University of rizonaMining and Geological Engineering 415 - Rock Excavation Course Notes – Spring 2003 Module 1-5 Compiled by: Dr. Sean Dessureault MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 i Table of Contents MODULE 1: INTRODUCTION ......................................................................................................... 1 1.1 INTRODUCTION.............................................................................................................................. 1 ADDITIONAL RESOURCES I .................................................................................................................. 1 1.2 GEOLOGICAL PROPERTIES IN ROCK EXCAVATION......................................................................... 1 1.2.1 Minerals and Rocks .................................................................................................................. 1 1.2.2 The properties of rocks ............................................................................................................. 3 1.2.3 Rock classification.................................................................................................................... 6 1.2.4 Rock Strength............................................................................................................................ 8 ADDITIONAL RESOURCES II............................................................................................................... 16 1.3 ROCK BREAKING PROCESSES....................................................................................................... 16 1.3.1 Primary................................................................................................................................... 17 1.3.2 Secondary ............................................................................................................................... 20 1.3.3 Tertiary................................................................................................................................... 21 1.3.4 Miscellaneous ......................................................................................................................... 23 ADDITIONAL RESOURCES III ............................................................................................................. 24 MODULE 2: DRILLING AND BLASTING COMPONENTS......................................................... 1 2.1 DRILLING MECHANISMS ................................................................................................................ 1 2.1.1 Percussive Drilling................................................................................................................... 2 ADDITIONAL RESOURCES IV............................................................................................................... 5 2.1.2 Rotary Drilling ......................................................................................................................... 6 ADDITIONAL RESOURCES V............................................................................................................... 28 2.2 DRILLING EQUIPMENT.................................................................................................................. 28 ADDITIONAL RESOURCES VI ............................................................................................................. 28 2.3 EXPLOSIVES INTRODUCTION........................................................................................................ 28 2.3.1 Detonation Theory.................................................................................................................. 28 2.3.2 Comparative Explosive Properties ......................................................................................... 33 2.3.3 Thermochemistry of Explosives and the Detonation Reaction. .............................................. 38 ADDITIONAL RESOURCES VII............................................................................................................ 39 2.4 EXPLOSIVE PRODUCTS................................................................................................................. 40 2.4.1 History of Explosives .............................................................................................................. 40 2.4.2 Energetic material .................................................................................................................. 42 2.4.3 Explosives and Propellants..................................................................................................... 43 2.4.4 Single Molecule and Composite Explosives ........................................................................... 43 2.4.5 Primary, Secondary, and Tertiary explosives......................................................................... 46 2.4.6 Commercial Explosives .......................................................................................................... 46 2.4.7 Nitroglycerin-based Explosives .............................................................................................. 47 2.4.8 Dry Blasting Agents................................................................................................................ 48 2.4.9 Wet Blasting Agents................................................................................................................ 49 2.4.10 Primers and Boosters-........................................................................................................ 52 2.4.11 Initiators and Initiation Systems......................................................................................... 52 2.4.12 Electric Caps...................................................................................................................... 52 2.4.13 Non-electric caps ............................................................................................................... 53 2.4.14 Electronic Detonators ........................................................................................................ 54 2.4.15 Detonating cord ................................................................................................................. 54 ADDITIONAL RESOURCES VIII .......................................................................................................... 55 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 ii MODULE 3: BLAST DESIGN.......................................................................................................... 56 BLAST DESIGN BASICS ............................................................................................................................. 56 System Approach ................................................................................................................................. 56 Idealized fragmentation curves............................................................................................................ 59 Preliminary guidelines for blast layout ............................................................................................... 62 Ratios for initial design ....................................................................................................................... 70 Powder Factor..................................................................................................................................... 73 Determination of K B ............................................................................................................................. 74 Effects to pattern design in changing explosives ................................................................................. 76 Effects to pattern design in changing rock types ................................................................................. 77 Numerical Examples............................................................................................................................ 78 Stemming & Decking........................................................................................................................... 80 Concluding Notes on D, B, and PF ..................................................................................................... 82 ADDITIONAL RESOURCES IX............................................................................................................. 83 GEOLOGICAL IMPACTS ON BLAST DESIGN................................................................................................ 83 Rock Properties ................................................................................................................................... 83 Structure .............................................................................................................................................. 85 ADDITIONAL RESOURCES X............................................................................................................... 92 PATTERNS AND SEQUENCING.................................................................................................................... 93 Effective Burden and Spacing.............................................................................................................. 94 Designing the Timing........................................................................................................................... 96 Generalized Timing ............................................................................................................................. 97 Skipping a Period ................................................................................................................................ 99 Pattern Types..................................................................................................................................... 100 ADDITIONAL RESOURCES XI ........................................................................................................... 104 MODULE 4: BLASTING APPLICATIONS.................................................................................. 105 DRIFT DESIGN......................................................................................................................................... 105 Cuts.................................................................................................................................................... 107 Blast Layout ....................................................................................................................................... 109 ADDITIONAL RESOURCES XII.......................................................................................................... 110 PRODUCTION BLASTING.......................................................................................................................... 111 Short-hole Production ....................................................................................................................... 111 Longhole Production Blasting........................................................................................................... 112 ADDITIONAL RESOURCES XIII ........................................................................................................ 118 CONTROLLED BLASTING......................................................................................................................... 119 Line Drilling ...................................................................................................................................... 120 Pre-splitting....................................................................................................................................... 121 Smoothwall blasting........................................................................................................................... 124 Trim Blasting ..................................................................................................................................... 125 Blasting Ornamental Rock................................................................................................................. 126 Underwater Rock Excavation............................................................................................................ 129 Controlled Blasting Rules of Thumb - Supplemental......................................................................... 131 ADDITIONAL RESOURCES XIV ........................................................................................................ 132 ENVIRONMENTAL ISSUES........................................................................................................................ 132 Airblast .............................................................................................................................................. 133 Fly-Rock............................................................................................................................................. 135 Disturbance of the Natural Ground Profile....................................................................................... 137 Dust.................................................................................................................................................... 137 Fumes ................................................................................................................................................ 138 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 iii Ground Vibrations............................................................................................................................. 138 Vibration reduction............................................................................................................................ 142 Cooperating Charges – Ground vibration......................................................................................... 143 The effect of Delayed Explosion ........................................................................................................ 144 Effect of Geological Factors.............................................................................................................. 145 Risk Assessment ................................................................................................................................. 145 Sustainable Development Issues in Rock Excavation........................................................................ 146 ADDITIONAL RESOURCES XV.......................................................................................................... 146 EVALUATION OF BLAST RESULTS ........................................................................................................... 146 Fragmentation and Swelling of the Muckpile.................................................................................... 147 Geometry of Muckpile, Its Height and Displacement ........................................................................ 152 Condition of the Remaining Mass...................................................................................................... 153 Analysis of the Bench Floor............................................................................................................... 154 Boulders............................................................................................................................................. 154 Vibrations and Airblast...................................................................................................................... 155 Blast Evaluation in Underground Mining ......................................................................................... 155 ADDITIONAL RESOURCES XVI ........................................................................................................ 155 MODULE 5: MECHANICAL EXCAVATION............................................................................. 156 5.1 MECHANICAL EXCAVATION BASICS.......................................................................................... 156 5.1.1 Synopsis ................................................................................................................................ 162 ADDITIONAL RESOURCES XVII....................................................................................................... 162 5.2 LONGWALL................................................................................................................................ 162 5.2.1 Mining System Description................................................................................................... 164 5.2.2 Selection of Cutting Machine................................................................................................ 165 5.2.3 Dimensions of Shearer.......................................................................................................... 165 ADDITIONAL RESOURCES XVIII ..................................................................................................... 167 5.3 CONTINUOUS MINING MACHINES.............................................................................................. 167 5.3.1 Underground continuous mining machines.......................................................................... 168 5.3.2 Continuous Surface Mining Technology............................................................................... 189 ADDITIONAL RESOURCES XIX ........................................................................................................ 196 5.4 DOZERS AND RIPPING................................................................................................................ 196 5.4.1 Determining Rippability. ...................................................................................................... 200 5.4.2 Operational issues:............................................................................................................... 205 5.4.3 Estimating Ripping Production ............................................................................................ 205 5.4.4 Ripping vs. Blasting.............................................................................................................. 208 ADDITIONAL RESOURCES XX.......................................................................................................... 209 Table of Figures FIGURE 1-1: ORIGINS OF VARIOUS ROCKTYPES ............................................................................................... 2 FIGURE 1-2: DRILL RATE INDEX FOR A VARIETY OF ROCKS ............................................................................. 7 FIGURE 1-3: UNCONFINED COMPRESSIVE STRENGTH VARYING BY TIME. ...................................................... 8 FIGURE 1-4: CRACK PROPAGATION ............................................................................................................... 10 FIGURE 1-5: CURVES OF STRESS-DEFORMATION FOR DIFFERENT TYPES OF ROCKS. ....................................... 12 FIGURE 1-6: CLASSIFICATION OF THE ROCK MASSES. .................................................................................... 16 FIGURE 1-7: PRIMARY................................................................................................................................... 17 FIGURE 1-8: STRESS RELATING TO SIZE OF INDENTER ................................................................................... 18 FIGURE 1-9: INCREASING ENERGY WITH DEPTH............................................................................................. 19 FIGURE 1-10: PENETRATION DEPTH OF VARIOUS BUTTON SHAPES................................................................. 19 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 iv FIGURE 1-11: SECONDARY BREAKAGE PROCESS ........................................................................................... 21 FIGURE 1-12: TERTIARY BREAKAGE PROCESS .............................................................................................. 21 FIGURE 1-13: TERTIARY BREAKAGE A TENSILE EFFECT ................................................................................ 22 FIGURE 2-1: THE DRILLING PROCESS............................................................................................................... 2 FIGURE 2-2: COMPONENTS OF A PERCUSSION DRILL – PISTON AND STEEL....................................................... 3 FIGURE 2-3: PISTON MOVING TOWARD STEEL ................................................................................................. 3 FIGURE 2-4: COMPRESSIVE WAVES IN BOTH PISTON AND STEEL...................................................................... 3 FIGURE 2-5: REFLECTIONS AND WAVE TRANSMISSION. ................................................................................... 4 FIGURE 2-6: DYNAMICS OF SURFACE BLASTHOLE DRILL................................................................................. 6 FIGURE 2-7: FLOOR PLAN FOR ROTARY DRILL P&H 100XP ............................................................................ 7 FIGURE 2-8: SELECTION OF DRILLING METHOD............................................................................................... 7 FIGURE 2-9: COMPONENTS IN ROTARY DRILLING............................................................................................ 8 FIGURE 2-10: BLADE AND ROLLER STABILIZER............................................................................................... 9 FIGURE 2-11: SWIVEL MOUNT SHOCK SUB..................................................................................................... 10 FIGURE 2-12: VIBRATION REDUCING EFFECTS OF USING SHOCK SUBS ........................................................... 11 FIGURE 2-13: BLASTHOLE COMPONENTS WITH RESPECT TO DRILL RIG.......................................................... 12 FIGURE 2-14: PENETRATION RATE VS. ROCK STRENGTH ............................................................................... 13 FIGURE 2-15: RECOMMENDED PULLDOWN WEIGHTS PER INCH OF BIT DIAMETER.......................................... 14 FIGURE 2-16: BIT LIFE VS. PULLDOWN WEIGHT FOR 9 1/4 INCH DIAMETER ROTARY BITS IN HARD FORMATION ............................................................................................................................................................. 14 FIGURE 2-17: ROTARY DRIVE TYPES............................................................................................................. 15 FIGURE 2-18: PULLDOWN MECHANISMS........................................................................................................ 15 FIGURE 2-19: RACK AND PINION PULLDOWN................................................................................................. 15 FIGURE 2-20: AIR BAILING CHART ............................................................................................................... 17 FIGURE 2-21: TONS DRILLED PER OPERATING HOUR FOR ROTARY DRILLS FOR VARIOUS HOLE DIAMETERS. .. 22 FIGURE 2-22: ROTARY DRILLING COSTS ESTIMATES BY METER DRILLED AND DIAMETER OF HOLE. .............. 23 FIGURE 2-23: COST PER TON OF MATERIAL EXCAVATED............................................................................... 23 FIGURE 2-24: DM-1 PRODUCTION MONITORING SYSTEM............................................................................. 24 FIGURE 2-25: DM-2 MATERIAL RECOGNITION SYSTEM................................................................................ 24 FIGURE 2-26: DM-3 DRILL CONTROL SYSTEM ............................................................................................. 25 FIGURE 2-27: OPERATOR’S VIEW WHEN NAVIGATING ................................................................................... 26 FIGURE 2-28: GPS RECEIVERS....................................................................................................................... 26 FIGURE 2-29: DETONATION PROCESS OF EXPLOSIVE CARTRIDGE.................................................................. 29 FIGURE 2-30: DETONATION PROCESS FOR CYLINDRICAL EXPLOSIVE ............................................................ 31 FIGURE 2-31: PRESSURE SHAPE FOR A-HIGH EXPLOSIVE AND B-COMMERCIAL EXPLOSIVE CONTAINING HIGH GAS VOLUME ........................................................................................................................................ 31 FIGURE 2-32: SHOCK WAVE PROPAGATION ................................................................................................... 32 FIGURE 2-33: GENERALIZED RELATIONSHIP BETWEEN VOD AND DIAMETER............................................... 35 FIGURE 2-34: EXPLOSIVES FAMILY TREE...................................................................................................... 42 FIGURE 2-35: EXPLOSIVE CLASSIFICATION SCHEME..................................................................................... 43 FIGURE 2-36: ENERGY OUTPUT VS. PERCENT FUEL OIL ADDED TO AMMONIUM NITRATE............................... 48 FIGURE 2-37: VOD VS. BOREHOLE DIAMETER FOR SELECTED INDUSTRIAL EXPLOSIVES ............................... 49 FIGURE 2-38: GENERIC DESIGN DETONATORS............................................................................................... 54 FIGURE 3-1: SYSTEMS APPROACH................................................................................................................. 57 FIGURE 3-2: CASCADING CONSTRAINTS AND GOALS ..................................................................................... 58 FIGURE 3-3: CONTROLLABLE AND UNCONTROLLABLE INPUT VARIABLES AND OUTPUT GOALS..................... 59 FIGURE 3-4: LOADING COST CURVE............................................................................................................... 60 FIGURE 3-5: HAULING COST CURVE ............................................................................................................. 60 FIGURE 3-6: CRUSHING COST CURVE............................................................................................................ 60 FIGURE 3-7: DRILLING AND BLASTING COST CURVES .................................................................................. 61 FIGURE 3-8: OVERALL COST CURVE. ............................................................................................................. 62 FIGURE 3-9: ISOMETRIC VIEW SHOWING NOMENCLATURE............................................................................. 62 FIGURE 3-10: PLAN VIEW OF BENCH SHOWING FIRST ROW. ........................................................................... 63 FIGURE 3-11: EFFECT OF HOLE DIAMETER ON BURDEN ................................................................................. 65 FIGURE 3-12: TOE CONFINEMENT.................................................................................................................. 66 FIGURE 3-13: RUN-UP DISTANCE TO ACHIEVE STEADY STATE VOD.............................................................. 66 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 v FIGURE 3-14: SECTION VIEW COMPARING THE SPHERICAL CHARGE (A) AND CYLINDRICAL CHARGE (B) MINIMUM DISTANCE FROM THE COLLAR ............................................................................................... 67 FIGURE 3-15: VIEW FACING BENCH COMPARING EQUIVALENT CYLINDRICAL AND SPHERICAL CHARGES. ..... 68 FIGURE 3-16: LANGEFORS & KIHLSTROM’S TOE BREAKING EQUIVALENCE OF SPHERICAL AND CYLINDRICAL CHARGES .............................................................................................................................................. 68 FIGURE 3-17: EQUIVALENT SPHERICAL AND CYLINDRICAL CHARGES............................................................ 69 FIGURE 3-18: BURDEN TO DIAMETER RELATIONSHIP..................................................................................... 70 FIGURE 3-19: LIMITING THE CHARGE DIAMETER AND BURDEN. .................................................................... 70 FIGURE 3-20: SQUARE PATTERN.................................................................................................................... 71 FIGURE 3-21: STAGGERED PATTERN.............................................................................................................. 71 FIGURE 3-22: BLASTHOLE PLACEMENT IN RELATION TO JOINT PLANES. ........................................................ 85 FIGURE 3-23: USE OF POCKET CHARGES....................................................................................................... 86 FIGURE 3-24: STAB HOLE .............................................................................................................................. 86 FIGURE 3-25: IDEALIZED RADIAL CRACKING SURROUNDING A SINGLE HOLE. ............................................... 87 FIGURE 3-26: EXTENT OF CRACKING FOR TWO PATTERNS WITH DIFFERENT HOLE DIAMETERS AND SAME PF 88 FIGURE 3-27: MAXIMUM BLOCK DIMENSIONS FOR HOLE PATTERNS. ............................................................. 89 FIGURE 3-28: FRAGMENTATION ENHANCEMENT ACHIEVED BY REDUCING THE SPACING .............................. 89 FIGURE 3-29: EFFECT OF JOINTING ON FRAGMENTATION .............................................................................. 90 FIGURE 3-30: POSSIBLE BLAST PATTERNS IN JOINTED ROCK ........................................................................ 90 FIGURE 3-31: SHOOTING WITH THE DIP ......................................................................................................... 91 FIGURE 3-32: SHOOTING AGAINST THE DIP.................................................................................................... 92 FIGURE 3-33: SHOOTING ALONG STRIKE ....................................................................................................... 92 FIGURE 3-34: INCREASING NUMBERS OF ROWS INCREASES UPWARD MOVEMENT DUE TO LOWER VELOCITY OF PREVIOUSLY BROKEN ROCK. ................................................................................................................. 93 FIGURE 3-35: MUCKPILES ILLUSTRATING THE DIFFERENCE BETWEEN MILLISECOND TIMED AND INSTANTANEOUS................................................................................................................................... 94 FIGURE 3-36: CORNER AND FACE BLASTS. .................................................................................................... 94 FIGURE 3-37: ECHELON OR HALF CHEVRON. ................................................................................................. 95 FIGURE 3-38: “V” (VEE) OR CHEVRON INITIATION PATTERN ON A SQUARE DRILL PATTERN. ........................ 96 FIGURE 3-39: THE EVENTS AND TIMING IN A BLASTHOLE.............................................................................. 98 FIGURE 3-40: BLAST TIMING SEQUENCE SKIPPING PERIODS.......................................................................... 99 FIGURE 3-41: CHEVRON OR V PATTERN WITH DOUBLE HOLE INITIATION FOR SLIGHTLY MORE FORWARD MOVEMENT......................................................................................................................................... 101 FIGURE 3-42: ECHELON DELAY PATTERN................................................................................................... 101 FIGURE 3-43: CHANNEL PATTERN ............................................................................................................... 102 FIGURE 3-44: FLAT FACE PATTERN.............................................................................................................. 102 FIGURE 3-45: ALTERNATING DELAY PATTERN ............................................................................................ 103 FIGURE 3-46: SINKING SHOTS...................................................................................................................... 103 FIGURE 3-47: SEQUENTIAL FIRING.............................................................................................................. 104 FIGURE 4-1: SYSTEMS FOR ADVANCING LARGE FACES ................................................................................ 106 FIGURE 4-2: DIFFERENT ZONES IN A DRIFT ROUND...................................................................................... 106 FIGURE 4-3: ORIENTATION OF THE CONTOUR BLASTHOLES TO MAINTAIN TUNNEL PROFITE-LOOKOUTS..... 106 FIGURE 4-4: VARIOUS TYPES OF BURN CUTS (SOLID DOTS ARE LOADED HOLES) ......................................... 108 FIGURE 4-5: ANGLE CUTS (VEE OR WEDGE, TOP LEFT; DOUBLE VEE OR BABY CUT, TOP RIGHT; THREE-HOLE PYRAMID, BOTTOM LEFT; AND A DRAW CUT, BOTTOM RIGHT) ............................................................ 109 FIGURE 4-6: CYLINTRICAL CUT ................................................................................................................... 109 FIGURE 4-7: SEQUENCING DRIFTS ............................................................................................................... 110 FIGURE 4-8: SHORT-HOLE PRODUCTION BLAST. .......................................................................................... 112 FIGURE 4-9: RING BLASTING....................................................................................................................... 113 FIGURE 4-10: SPACING CONSTRUCTION FOR RING BLAST DESIGN................................................................ 114 FIGURE 4-11: STEMMING LENGTH FOR RING BLASTING............................................................................... 115 FIGURE 4-12: BENCH BLASTING................................................................................................................. 116 FIGURE 4-13: VCR...................................................................................................................................... 117 FIGURE 4-14: VCR LOADED EXPLOSIVE COLUMN ....................................................................................... 118 FIGURE 4-15: DIAGRAMMATIC REPRESENTATION OF THE BDT OF A FULLY CHARGED HOLE....................... 119 FIGURE 4-16: DBT FOR A TWO ROW PRODUCTION BLAST............................................................................ 120 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 vi FIGURE 4-17: BDT AFTER BLASTING TWO PRODUCTION ROWS. .................................................................. 120 FIGURE 4-18: LINE DRILLING POSITIONED ALONG THE PLANNED FINAL PERIMETER.................................... 121 FIGURE 4-19: PRESPLIT BLAST COUPLED TO A 250MM PRODUCTION BLAST................................................ 122 FIGURE 4-20: RELATIONSHIPS BETWEEN HOLE DIAMETER AND SPACING FOR PRESPLITTING FROM VARIOUS RESEARCHERS..................................................................................................................................... 123 FIGURE 4-21: SMOOTHWALL DAMAGE ZONES ............................................................................................. 124 FIGURE 4-22: TRIM HOLE ROW TRIMMING THE FRACTURED AND INFLUENCED ROCK. ................................. 125 FIGURE 4-23: ORNAMENTAL STONE MATERIALS HANDLING – CRANES........................................................ 127 FIGURE 4-24: SEQUENCE OF OPERATIONS IN UNDERWATER ROCK EXCAVATION. ........................................ 130 FIGURE 4-25: AIRBLAST WAVES IN VARIOUS CONDITIONS........................................................................... 135 FIGURE 4-26: CAUSES OF FLY ROCK. ........................................................................................................... 136 FIGURE 4-27: OBTAINING SITE FACTORS FOR GROUND VIBRATIONS............................................................ 140 FIGURE 4-28: GROUND VIBRATIONS RESULTS FROM 2500 BLASTS IN 40 DIFFERENT OPERATIONS............... 141 FIGURE 4-29: COOPERATING CHARGE RELATED TO DISTANCE. ................................................................... 144 FIGURE 30: GIRD METHOD OF EVALUATING FRAGMENTATION AND FIELD ENVIRONMENT........................... 148 FIGURE 31: HIGH SPEED CAMERA BENCH BLAST ......................................................................................... 150 FIGURE 32: DIFFERENT GEOMETRIES OF THE MUCKPILE.............................................................................. 153 FIGURE 5-1: LONGWALL SHEARING MACHINE WHOSE CUTTING BITS ATTACK THE ROCK IN COMPRESSION. 156 FIGURE 5-2: PICK SHAPES - POINTED TOOL, SIMPLE CHISEL, AND POINTED TOOL. (| = BACK CLEARANCE ANGLE, o = RAKE ANGLE).................................................................................................................. 157 FIGURE 5-3: MEAN FORCE VS. DEPTH OF CUT (F C VS. D)............................................................................. 158 FIGURE 5-4: SPECIFIC ENERGY VS. DEPTH OF CUT (E S VS. D) ....................................................................... 158 FIGURE 5-5: MEAN FORCE VS. RAKE ANGLE (F C VS. |) ................................................................................ 159 FIGURE 5-6: MEAN FORCE VS. BACK CLEARANCE ANGLE (F C VS. |)............................................................ 159 FIGURE 5-7: MEAN FORCE VS. PICK WIDTH (F C VS. W)................................................................................ 160 FIGURE 5-8: COMPRESSIVE AND TENSILE STRENGTHS UPON VARYING MEAN CUTTING FORCE (FC VS. o C AND FC VS. o T )........................................................................................................................................... 161 FIGURE 5-9: CHISEL VS. PICK BITS............................................................................................................... 161 FIGURE 5-10: TYPICAL LONGWALL RETREAT METHOD ............................................................................... 163 FIGURE 5-11: LONGWALL SHEARING SYSTEM............................................................................................. 164 FIGURE 5-12: SHEARER DIMENSIONS........................................................................................................... 165 FIGURE 5-13: NOMOGRAPH FOR SHEARER DIMENSIONING .......................................................................... 166 FIGURE 5-14: FOUR-ROTOR BORING MACHINE........................................................................................... 168 FIGURE 5-15: MILLING HEAD CONTINUOUS MINING MACHINE..................................................................... 169 FIGURE 5-16: MILLER-HEAD CONTINUOUS MINER. ...................................................................................... 170 FIGURE 5-17: MINING OF STEEP SEAMS WITH ROADHEADDER AND MILLER HEAD (DRUM-TYPE) MINERS.... 171 FIGURE 5-18: MINING IN VARIABLE HEIGHT SEAMS.................................................................................... 172 FIGURE 5-19: DOUBLE HEAD ROADHEADDER............................................................................................. 172 FIGURE 5-20: MD1100 WITH TRAVERSE BOOM.......................................................................................... 173 FIGURE 5-21: ROCK MASS CUTTABILITY INDEX VS. CUTTING RATE............................................................ 174 FIGURE 5-22: ROADHEADER PERFORMANCE VS. ROCK CLASS. .................................................................... 175 FIGURE 5-23: TUNNEL BORING MACHINE .................................................................................................... 175 FIGURE 5-24: COMPONENTS OF A TBM....................................................................................................... 176 FIGURE 5-25: DISC CUTTERS ON TBM........................................................................................................ 176 FIGURE 5-26: CROSS SECTION OF DISC CUTTER............................................................................................ 177 FIGURE 5-27: KERF AND PINAPPLE CUTTERS............................................................................................... 177 FIGURE 5-28: SPACING VERSUS DEPTH RELATIONSHIP................................................................................. 178 FIGURE 5-29: EMPIRICALLY DERIVED EXCAVATION RATES OF TBMS BY DIAMETER FOR IGNEOUS AND METAMORPHIC ROCK TYPES ............................................................................................................... 179 FIGURE 5-30: EMPIRICALLY DERIVEN EXCAVATION RATES OF TBMS BY DIAMETER FOR SEDIMENTARY ROCK TYPES ................................................................................................................................................. 180 FIGURE 5-31COST COMPARISON BETWEEN TUNNEL BORING AND CONVENTIONAL DRILL AND BLAST. ........ 181 FIGURE 5-32: STEPS IN RAISEBORE MINING. ................................................................................................ 182 FIGURE 5-33: RAISEBORING IN OPERATION TOP & BOTTOM........................................................................ 182 FIGURE 5-34: RAISE BORING DIAGRAM – UP AND DOWNWARD BORING. ..................................................... 183 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 vii FIGURE 5-35: BLIND BORING OPERATION.................................................................................................... 184 FIGURE 5-36: BLIND BORING BIT HEAD. ...................................................................................................... 184 FIGURE 5-37: DIRECTIONAL BORING IN URBAN ENVIRONMENT. .................................................................. 185 FIGURE 5-38: HORIZONTAL BOREHEAD ....................................................................................................... 185 FIGURE 5-39: CROSS SECTION OF THE BOUBLY POTASH MINE STRATA........................................................ 186 FIGURE 5-40: REMOTE OPERATION ALLOWS DISTANCE MINING AND THE OPERATOR TO VIEW FROM DIFFERENT ANGLES............................................................................................................................................... 187 FIGURE 5-41: MARRIETA MINER IN ROCANVILLE........................................................................................ 188 FIGURE 5-42: MARRIETA DIGGING A TRENCH, BEING MAINTAINED............................................................. 188 FIGURE 5-43: CROSS SECTIONAL VIEW OF FOUR-ROTOR BORING IN POTASH............................................... 189 FIGURE 5-44: VARIOUS ASPECTS OF A SURFACE MINER.............................................................................. 190 FIGURE 5-45: SURFACE MINER IDEAL FOR THIN STRATA ............................................................................. 191 FIGURE 5-46: CUTTING CAPACITIES FOR VARIOUS WIRTGENAMERICA PRODUCTS ...................................... 191 FIGURE 5-47: SURFACE MINER IDEAL FOR DEFINED THIN VERTICAL OR FLAT SEAMS .................................. 191 FIGURE 5-48: COMPONENTS OF SURFACE MINER......................................................................................... 192 FIGURE 5-49: BUCKETWHEEL NOMENCLATURE........................................................................................... 193 FIGURE 5-50: BUCKET CUTTING AND DISCHARGE ....................................................................................... 193 FIGURE 5-51: RIPPER IN ACTION................................................................................................................. 197 FIGURE 5-52: (A) RADIAL RIPPER, (B) PARALLELOGRAM RIPPER, (C) ADJUSTABLE PARALLELOGRAM RIPPER ........................................................................................................................................................... 197 FIGURE 5-53: RIPPER NOMENCLATURE....................................................................................................... 198 FIGURE 5-54: NOMENCLATURE FOR RIPPERS (SEE TABLE 5-2) .................................................................... 199 FIGURE 5-55: ROCK QUALITY CLASSIFICATION IN RELATION TO EXCAVATION PROCESSES. ........................ 200 FIGURE 5-56: SEISMIC VELOCITIES IN RELATION TO RIPPING....................................................................... 201 FIGURE 5-57: RIPPER PRODUCTION CHART................................................................................................. 206 FIGURE 5-58: PERFORMANCE CHART FOR D11............................................................................................ 207 FIGURE 5-59: D11 WITH RIPPER .................................................................................................................. 207 List of Tables TABLE 1-1: IGNEOUS ROCK............................................................................................................................. 5 TABLE 1-2SEDIMENTARY ROCKS.................................................................................................................... 6 TABLE 1-3: METAMORPHIC ROCKS.................................................................................................................. 6 TABLE 1-4: FRACTURE TOUGHNESS.............................................................................................................. 11 TABLE 1-5: RELATIVE QUARTZ (COMMON ABRASIVE) CONTENT................................................................... 13 TABLE 1-6: COMMON ROCK NAMES AND THEIR GEOLOGICAL DEFINITIONS................................................... 14 TABLE 1-7: ROCK PROPERTIES ACCORDING TO ORIGIN ................................................................................. 14 TABLE 1-8: SUMMARIZING THEORY OF FORCES AND ENERGY IN BREAKING PROCESSES ............................. 24 TABLE 2-1: COMPARISON OF HYDRAULIC AND PNEUMATIC DRILLS................................................................ 5 TABLE 2-2: COMPARATIVE TABLE................................................................................................................... 5 TABLE 2-3: AVERAGE DRILL BIT FOOTAGES (FROM LARGE IRON ORE MINE IN CANADA) ............................ 16 TABLE 2-4: STANDARDS FOR FUME CLASS ................................................................................................... 33 TABLE 2-5: THERMODYNAMIC DATA FOR SOME EXPLOSIVE COMPONENTS AND GASES. ................................ 38 TABLE 2-6: SINGLE CHEMICAL EXPLOSIVE SUBSTANCES. .............................................................................. 44 TABLE 2-7: MORE SINGLE CHEMICAL EXPLOSIVES ...................................................................................... 45 TABLE 2-8: TYPICAL COMPOSITIONS OF SELECTED SLURRIES AND EMULSIONS............................................ 50 TABLE 2-9: TYPICAL DELAY TIMES FOR DONATORS...................................................................................... 53 TABLE 3-1: RATIO SUMMARY ....................................................................................................................... 73 TABLE 3-2: POWDER FACTOR........................................................................................................................ 82 TABLE 3-3: EFFECT OF DRILLING PATTERNS AND S/B RATIOS ON THE AREA COVERED BY FRACTURE CIRCLES (ENERGY DISTRIBUTION)....................................................................................................................... 98 TABLE 4-1: POWDER FACTORS FOR RING BLASTING.................................................................................... 114 TABLE 4-2: BDT CHARACTERIZATION........................................................................................................ 119 TABLE 4-3: FACTORS FOR DETERMINING HOLE SPACING............................................................................. 121 TABLE 4-4: RECOMMENDED CHARGE LOADS AND BLAST GEOMETRIES FOR TRIM BLASTING....................... 126 TABLE 4-5: AIRBLAST EFFECT ON STRUCTURES ......................................................................................... 134 TABLE 4-6: FACTORS THAT INFLUENCE GROUND MOTION........................................................................... 141 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 viii TABLE 4-7: RECOMMENDED MAXIMUM PEAK PARTICLE VELOCITIES. ......................................................... 142 TABLE 4-8: EFFECT OF DELAYED CHARGES ON PPV FOR A TOTAL CHARGE OF 2900 KG AT A DISTANCE OF 220M. ................................................................................................................................................. 144 TABLE 9: LEVELS OF DAMAGE TO PIT WALLS .............................................................................................. 154 TABLE 5-1: APPLICATIONS FOR BORING...................................................................................................... 183 TABLE 5-2: RIPPER SELECTION FOR LARGE CAT DOZERS. ......................................................................... 199 TABLE 5-3: EXCAVATION CHARACTERISTICS IN RELATION TO ROCK HARDNESS AND STRENGTH................ 201 TABLE 5-4: EXCAVATION CHARACTERISTICS IN RELATION TO JOINT SPACING............................................ 201 TABLE 5-5: VARIABLES IN EXCAVABILITY INDEX ....................................................................................... 202 TABLE 5-6: MASS STRENGTH NUMBER FOR ROCKS (MS) ........................................................................... 202 TABLE 5-7: JOINT COUNT NUMBER (J C ) AND JOINT SET NUMBER (J N )........................................................... 203 TABLE 5-8: RELATIVE GROUND STRUCTURE NUMBER (J S )........................................................................... 203 TABLE 5-9: JOINT ROUGHNESS NUMBER (J R ) ............................................................................................... 204 TABLE 5-10: JOINT ALTERATION NUMBER (J A )............................................................................................ 204 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 1 Module 1: Introduction 1.1 Introduction The information necessary to pass this course will be presented in these notes and through other medium such as lectures and field trips. To guide the student to additional resources either out of interest or mandated by the course requirements, a list of additional resources are indicated after every major topic. The appearance of this list will be as seen below: ADDITIONAL RESROUCES [NUMBER] The above information is additionally supplemented by: - Readings (Mandatory or Optional) - Field trip (Always mandatory) - Assignment - Lecture Important note: These class notes borrow heavily from the material assembled by: - Dr. Paul Lever, Mr. William Cummings, - Persson, Per-Anders, Holmberg, Roger, and Lee, Jaimin. Rock Blasting and Explosvies Engineering. CRC Press: New York. 2000. 534 p. ADDITIONAL RESOURCES I The above information for is additionally supplemented by: - Module 1.1 Lecture 1.2 Geological Properties in Rock Excavation A brief review of rock properties is required to provide students without this background or as a review. 1.2.1 Minerals and Rocks 1 The earth's crust consists of a variety of rocks, formed under different circumstances. Rocks consist of one or more composite minerals. A mineral is a substance formed by nature. A mineral may be an element or may consist of chemical compounds containing several elements. There are more than 3,000 different minerals. Of 103 known elements, oxygen is by far the most common, making up about 50 per cent of the earth's crust. Silicon, about 25 per cent, aluminum, iron, calcium, sodium, potassium, magnesium and titanium, together with oxygen, account for about 99 per cent. 1 From SECOROC’s Geology Drilling Theory. Document S-96174 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 2 Silicon, aluminum and oxygen occur in our most common minerals, quartz, feldspar and mica. They form the large group known as silicates, a silicate being a compound formed of silicic acid and other elements. Also included are amphiboles and pyroxenes, which contain aluminum, potassium and iron. Some of the earth's most common rocks, granite and gneiss, are composed of silicates. Figure 1-1: Origins of various rocktypes Oxygen often occurs in combination with metallic elements and forms our oxidic ores (the iron ores magnetite and hematite). Sulphur readily combines with metallic elements and forms sulphide ores (galena, sphalerite, molybdenite, arsenopyrite. Halogenides (fluorite, halite), carbonates (calcite, dolomite, malachite), sulphates (barite), tungstates (scheelite) and phosphates (apatite) are other large mineral groups. Gold, silver, copper and lead are elements that can occur as native metals. Feldspar accounts for MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 3 almost 50 per cent of the mineral composition of the earth's crust. Next come the pyroxene and amphibole minerals, closely followed by quartz and mica. These minerals make up about 90 per cent of the composition. Some of the characteristics of the minerals are hardness, density, color, streak, lustre, fracture, cleavage and crystalline form. Hardness can be graded according to the Moh's 10-point scale. (example followed by test) 1. Talc - Easily scratched with the fingernail 2. Gypsum - Just barely scratched with the fingernail 3. Calcite - Very easily scratched with a knife 4. Fluorite - Easily scratched with a knife 5. Apatite - Can be scratched with a knife 6. Orthoclase - Hard to scratch with a knife, can be scratched with quartz 7. Quartz - Scratches glass, can be scratched with a hardened file 8. Topaz - Scratches glass, can be scratched with emery 9. Corundum - Scratches glass, can be scratched with a diamond 10.Diamond - Scratches glass Molybdenite, hardness 1.5, blackens a thumb that is rubbed against it. The density of light -colored minerals is for the most part below 3.0. Exceptions are barite or heavy spar (BaSO4), density 4.5, scheelite(CaWO4), density 6.0, and cerussite (PbCO3), density 6.5 Dark-colored minerals with some iron, silicates, have densities between 3.0 and 4.0. Ore minerals have densities over 4.0. Gold has a very high density at 19.3, and tungsten at 19.4. The highest density is shown by osmium and iridium, 22.5. Streak is the color of the mineral powder produced when the mineral is scratched against unglazed, white porcelain (e.g. an ordinary electric fuse). Fracture is the surface produced by breaking off a piece of mineral, not following a crystallographically defined plane. Fracture is usually uneven in one way or another. Cleavage denotes the properties of a crystal whereby it allows itself to be split along flat surfaces parallel with certain formed or otherwise crystallographically defined surfaces. 1.2.2 The properties of rocks In order to be able to forecast the result of drilling in respect of penetration rate, hole quality, drill-steel costs, etc., we must be able to make a correct appraisal of the rock concerned. In doing so we distinguish between microscopic and macroscopic properties. A rock is composed of grains of various minerals, and among the microscopic properties are mineral composition, grain size, and the form and distribution of the grains. Taken together, these factors decide important properties of the rock, such as hardness, abrasiveness, compressive strength and density. These properties, MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 4 in their turn, determine the penetration rate that can be achieved and how heavy the tool wear will be. The drillability of a rock depends on, among other things, the hardness of its constituent minerals and on the grain size and crystal form. Quartz is one of the commonest minerals in rocks. Since quartz is a very hard material, a high quartz content (SiO 2 ) makes the rock very hard to drill and causes heavy wear, particularly on the drill bits. We say that the rock is abrasive. Conversely, a rock with a high content of calcite is easy to drill and causes little wear on the drill bits. As regards crystal form, minerals with high symmetry, e.g. cubic (galena) are easier to drill than minerals with low symmetry, e.g. fibrous (amphiboles and pyroxenes). A coarse-grained structure is easier to drill in and causes less wear than a fine-grain structure. Consequently, rocks with essentially the same mineral content may be quite different as regards drillability. For example, quartzite may be fine-grained, grain size 0.5-1 mm, or dense, grain size 0,05 mm, while a granite may be coarse-grained, grain size> 5 mm, medium-grained (grain size 1-5 mm) or fine-grained (grain size 0.5-1 mm). A rock can also be classified on the basis of its structure. If the mineral grains are mixed in a homogeneous mass, the rock is massive (e.g. granite). In mixed rocks the grains are arranged in layers. A slaty rock also has the minerals arranged in different layers, but in this case pressure and heat have compacted each layer in plates. Among the macroscopic properties are slatiness, fissuring, contact zones, layering, veining and inclination. These factors are often of great significance in drilling. For example, cracks or inclined and layered formations can cause hole deviation and can occasionally cause drilling tools to get stuck. Soft rocks make it difficult to achieve good hole quality, since the walls often cave in and in extreme cases the flushing air disappears into cracks in the rock without reaching the surface. The enormous variety of rocks and rock formations makes it impossible to give the subject adequate treatment in just a few pages. We must therefore content ourselves with giving a summary description of the more important rocks and referring interested readers to the literature of the subject for further study. Rocks are classified into three main groups on the basis of their origin and the way in which they were formed: 1. Igneous or magmatic rocks (formed from solidified lava or "magma"). 2. Sedimentary rocks (formed by deposition of broken material or by chemical precipitation). MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 5 3. Metamorphic rocks (formed by the transformation of igneous or sedimentary rocks, in most cases by an increase in pressure and heat). 1.2.2.1 Igneous rock Igneous rocks are formed when a magma solidifies deep down in the earth's crust (plutonic rock), or as it rises towards the surface (dyke rock) or on the surface (volcanic rock). The most important constituents (minerals) are quartz and silicates of various composition, chiefly feldspars. Plutonic rocks solidify slowly and are therefore coarse-grained, while volcanic rocks solidify quickly and become fine grained. Depending on whether the magma solidifies at depth, or as a dyke rock, or on the surface, the rock is given different names even if the composition is the same. This is evident from the table below, which also shows that it is customary to classify the igneous rocks by their silicon content (SiO 2 ). The greater the silicon content, the larger the amount of quartz in the rock will be. Table 1-1: Igneous Rock SiO 2 Plutonic Dykes Volcanic Basic <52% SiO 2 Gabbro Diabase Basalt Intermediary 52-65% SiO 2 Diorite Syenite Porphyrite Syneite porphyry Andesite Trachyte Acid >65% SiO 2 Quartz diorite Granodiorite Granite Quartz porphyrite Granodiorite porphyry Quartz porphyry Dacite Rydodacite Rhyloite 1.2.2.2 Sedimentary rocks Sedimentary rocks are formed by the deposition of material by mechanical or chemical action and a consolidation of this material under the pressure of overlying layers. It frequently occurs that the rock formation is broken down by mechanical action (weathering), carried away by running water and deposited in still water. Thus the original rock will determine the characteristics of the sedimentary rock. The weathering or erosion may proceed at different rates depending partly on climate and partly on how easily the rock breaks up. Sedimentary rocks can also arise as a result of the chemical precipitation of minerals or by proliferation of organic organisms, as for example in coral reefs or carbon deposits. Since formation takes place by deposition, several distinct layers can often be observed in a sedimentary formation. Sedimentary rocks make up a very heterogeneous family with widely varying characteristics. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 6 Table 1-2Sedimentary Rocks Rock Original material Conglomerate Greywacke Sanstone Argillaceous schist Aluminous slate Limestone Gravel, stones, boulders Clay, gravel Sand Clay Clay plus organisms Calcium carbonate or various organisms 1.2.2.3 Metamorphic rocks The effects of chemical action or increased pressure and/or temperature on a rock formation can sometimes be so great that it produces a transformation, which the geologist calls metamorphism. For example, pressure and temperature might increase under the influence of up-welling magma, or because the formation has sunk down deeper into the earth's crust. This results in the recrystallization of the mineral grains or the formation of new minerals. A characteristic of the metamorphic rocks is that they are formed without any complete melting. They are also frequently hard or very hard, and also compact and fine-grained, and are therefore often difficult to drill. The earth's crust is in a constant state of flux, and the various rock formations may be subjected to very powerful forces. The result is deformation in one way or another, and in nature we may therefore observe, for example, pronounced folding, crushed zones, faults and other phenomena that can have a telling effect on drilling. Table 1-3: Metamorphic rocks Rock Original Rock Degree of metamorphism Amphibolite Basalt, diabase, gabro High Mica-schist Mudstone, greywacke, etc. Medium/high Gneiss Various igneous rocks High Greenschist Basalt, diabase, gabbro Low Quartize Sandstone Medium/high Lepitte Dacite Medium Slate Shale Low Veined gneiss Silicic-acid-rich silicate rocks high 1.2.3 Rock classification Many attempts have been made to classify rocks on the basis of their drillability, and various measuring methods have been evolved with the aim MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 7 of making it possible to predict productivity and tool wear by carrying out a laboratory test before a rock job begins. The term "drillability" is used here to mean primarily the rate at which the tool penetrates into the rock, but in the wider sense it extends to the resulting hole quality, the straightness of the hole, the risk of tool jamming etc. Tool wear is often proportional to drillability, though it also depends on how abrasive the rock is. The drillability of a given rock is determined by a number of factors. Foremost among these are the mineral composition, the grain size and the brittleness. Attempts are often made to describe drillability by stating the compressive strength or hardness; these are crude methods but they will often suffice for rough calculations. The Norwegian Technical University has developed a more sophisticated method for calculating the DRI and BWI. DRI, the drilling-rate index, describes how fast the drill steel can penetrate; it also includes on the one hand a measurement of brittleness and on the other hand drilling with a small rotating bit into a sample of the rock. The higher the DRI, the higher penetration rate. Figure 1-2 shows how the DRI can vary from one rock to another. Figure 1-2: Drill rate index for a variety of rocks MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 8 BWI stands for "bit-wear index", and gives an indication of how fast the bit wears down. It is determined by an abrasive test. The higher the BWI, the faster will be the wear. In most cases the BWI and the DRI are inversely proportional to each other, so that a high DRI will give a low BWI and vice- versa. However, the presence of hard minerals may produce heavy wear on the bit despite relatively good drillability. This is particularly true in the case of quartz. The quartz content has been shown to exert great influence on wear, which explains why relatively easily drilled sandstone, for example, can cause very heavy wear on the bits. In other cases, metamorphic, compact, quartz-bearing rocks may also prove to be very difficult to drill. Other examples of hard minerals that impair drillability are certain sulphides in orebodies. 1.2.4 Rock Strength 2 The compressive strength of many rock materials is a factor of 5 or more greater than their tensile strength. The scatter of the strength values from a series of test specimens of the same rock material is considerable. This is a result of the randomly distributed weak planes, microcracks, or flaws in the rock which greatly influence the rock strength. The flaws are often so small and the microcracks so fine that they are difficult to detect by the naked eye. Figure 1-3: Unconfined Compressive Strength Varying by Time. There is also an element of creep in the strength and deformation characteristics of rock, as evidenced by the ability of rock to flow and deform plastically under tectonic stress over long periods of time (hundreds of 2 Persson, Per-Anders, Roger Homberg, and Jaimn Lee. Rock Blasting and Explosives Engineering. New York: CRC Press. 1994. 50p. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 9 thousands to millions of years). Experiments have shown that the strength of rock is time dependent, so that the compressive strength when the load is applied in 1 msec is a factor of 2 or 2.5 greater than when the load is applied in 10 sec. (Figure 1-3) With confinement, that is when lateral expansion is restricted, rock in compression becomes stronger. This is because deformation to failure takes place as a shearing of weak planes. The action of confinement is to resist shearing, partly by creating lateral forces that resist the shear motion, and partly by increasing the friction on potential shearing surfaces by increasing the normal load thereby also increasing the rock strength. Shear strength is composed of two parts: the friction between sliding crack surfaces characterized by a friction coefficient ,µ; and the fracturing or plastic deformation of the crystal grains, which approach a limiting shear strength t i when the deformation is entirely plastic. 1.2.4.1 Microcracks in Rock The reasons for the strength behavior of rock and other brittle materials can be sought in the presence of microscopic cracks and flaws in the base material that is in itself strong. Most rock materials are aggregates in which separate crystal grains of different strength, different elastic and thermal moduli, and different size are cemented or grown together. Any deformation of sufficient magnitude will lead to local cracking or the development of microscopic flaws, pores, or weakened regions. Such flaws are also nearly always present in most natural rock materials because of the deformation the rock has undergone under the influence of tectonic forces and temperature changes. In tension, microcracks grow, join, and ultimately lead to fracture at a low load. In compression, the friction on such microcracks that are stressed in shear leads to increased strength. With confinement, the crack growth is further restricted and friction is increased. This leads to a further increase in strength. In the limit, with increased hydrostatic confining pressure, we approach the real strength of the aggregate base material. It is conceivable, with a sufficiently high hydrostatic pressure, that the deformation of the weaker part of the aggregate grains will be plastic, while the hard grains still only deform elastically. The random distribution in space, size, and direction of the micro-cracks or flaws is the reason for both the scatter of experimental strength measurement data and the dependence of strength on the size of the specimen. In a large specimen, it is more probable than in a small specimen that a sufficiently large flaw will have a direction favoring fracture in a given stress situation. Therefore, a large specimen or rock volume has a lower strength than a small specimen. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 10 1.2.4.2 Fracture Mechanics The most important aspect of the strength of brittle materials is their ability to break by crack propagation. Because the tensile strength of these materials is so much lower than the compressive strength, and possibly also because they already contain micro-cracks, cracks form easily and, once formed, expand because of the concentration of tensile stresses at the crack tip (Figure 1-4). We will limit this discussion to cracking under biaxial stresses, that is, stress situations where two principal stresses are equal and the third is zero. The stress concentration in front of the crack tip can be represented by the expression: x K x I t o 2 ) ( ~ where the stress intensity factor K I is a function of the crack length and the load o o . The critical value of K I when the crack just starts moving is a material constant K IC . As the crack propagates, energy is absorbed by deformation work by the material at the crack tip or dissipated as elastic wave energy radiating out through the material from the crack tip. The work done per unit new crack surface is G IC which is coupled to K IC through the relation: 2 2 1 IC IC K E v G ÷ = Figure 1-4: Crack propagation MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 11 Where v is Poisson’s ratio 3 and E is Young’s modulus. 4 G IC is called fracture toughness and is the fundamental material constant. Some authors refer to K IC as fracture toughness. Table 1-4 shows fracture toughness of various materials Table 1-4: Fracture Toughness 1.2.4.3 Elasticity The majority of rock minerals have an elastic-fragile behavior, which obeys the Law of Hooke, and are destroyed when the strains exceed the limit of elasticity. Depending upon the nature of deformation, as function of the stresses produced by static charges, three groups of rocks are taken into consideration: 1. The elastic-fragile or those which obey the Law of Hooke, 2. The plastic- fragile, that have plastic deformation before destruction, 3. The highly plastic or very porous, in which the elastic deformation is insignificant. The elastic properties of rocks are characterized by the elasticity modulus 'E' and the Poisson coefficient 'v'. The elasticity module is the proportionality factor between the normal stress in the rock and the relative correspondent deformation, its value in most rocks varies between 0.03 x 10 4 and 1.7 x 10 5 MPa, basically depending upon the mineralogical composition, porosity, type of deformation and magnitude of the applied force. The values of the elasticity modules in the majority of sedimentary rocks are lower than those corresponding to the minerals in their composition. The texture of the rock also has influence on this parameter, as the elasticity module in the direction of the bedding or schistosity is usually larger than when perpendicular. 3 A material stretches when pulled under tensile load and usually contracts transversely. The Poisson's ratio is the ratio of transverse strain to axial strain during axial load. For example, if a bar is pulled in the axial direction then the deformed bar (besides being longer in the axial direction) also contracts in the transverse direction by the percentage amount indicated by the Poisson's ratio. 4 Young's modulus is the ratio of longitudinal stress to the resultant longitudinal strain (stress/strain). Stiffness of the material MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 12 Poisson's coefficient is the factor of proportionality between the relative longitudinal deformations and the transversal deformations. For most rocks and minerals it is between 0.2 and 0.4, and only in quartz is it abnormally low, around 0.07. Figure 1-5: Curves of stress-deformation for different types of rocks. 1.2.4.4 Plasticity As indicated before, in some rocks the plastic deformation precedes destruction. This begins when the stresses exceed the limit of elasticity. In the case of an ideally plastic body, that deformation is developed with an invariable stress. Real rocks are deformed and consolidated at the same time: in order to increase the plastic deformation it is necessary to increase the effort. The plasticity depends upon the mineral composition of the rocks and diminishes with an increase in quartz content, feldspar and other hard minerals. The humid clays and some homogeneous rocks have plastic properties. The plasticity of the stony rocks (granites, schistoses, crystallines and sandstones) becomes noticeable especially at high temperatures. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 13 1.2.4.5 Abrasiveness Abrasiveness is the capacity of the rocks to wear away the contact surface of another body that is harder, in the rubbing or abrasive process during movement. This property has great influence upon the life of drill steel and bits. In Table 1-5, the mean amounts of quartz for different types of rock are indicated. The factors that enhance abrasive capacities of rocks are the following: - The hardness of the grains of the rock. The rocks that contain quartz grains are highly abrasive. - The shape of the grains. Those that are angular are more abrasive than the round ones. - The size of the grains. - The porosity of the rock. It gives rough contact surfaces with local stress concentrations. - The heterogeneity. Polymineral rocks, although these are equally hard, are more abrasive because they leave rough surfaces with hard grains as, for example, quartz grains in a granite. Table 1-5: Relative quartz (common abrasive) content 1.2.4.6 Texture The texture of a rock refers to the structure of the grains or minerals that constitute it. The size of the grains are an indication, as well as their shape, porosity etc. All these aspects have significant influence on drilling performance. When the grains have a lenticular shape, as in a schist, drilling is more difficult than when they are round, as in a sandstone. The type of material that makes up the rock matrix and unites the mineral grains also has an important influence. As to porosity, those rocks that have low density MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 14 and, consequently, are more porous, have low crushing strength and are easier to drill. In Table 1-6 the classification of some types of rocks is shown, with their silica content and grain size. Table 1-7 shows characteristics properties of different types of rocks according to their origins Table 1-6: Common rock names and their geological definitions Table 1-7: Rock Properties according to origin MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 15 1.2.4.7 Rock Mass Strength and Structure A rock mass, as distinguished from a rock strength test specimen, is a body of rock with its naturally occurring network of flaws and discontinuities, cracks, joints, and planes of weakness. These are important for a proper understanding of the real ability of a rock mass volume to withstand load, of how and why it fails, and of the resulting fragment size and shape. Recurring discontinuities are easily identifiable in the form of bedding planes, foliation partings, cracks, fissures, or joints. Intersecting groups are common such as parallel planes, or random, irregular structures. A description of the three-dimensional network of intersecting planes, has to include the compass bearing (called the strike) of its intersection with a reference plane (normally the horizontal), and the slope angle (dip) between the plane and the horizontal. For regular or recurring cracks there are two further descriptors, namely the average crack length and the average distance between parallel cracks. The strength of joints is normally considerably less than that of the adjacent rock. It is described by two simple measures, the tensile or adhesive strength (often zero) at right angles to the plane, and the shear strength or friction angle along the plane. For a detailed understanding of the rock mass behavior under stress and vibration, we also need a measure of the elastic or MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 16 plastic deformability of the joint (its "spring constant") and the way the shear strength or friction angle varies with shear deformation and crack separation. Figure 1-6: Classification of the rock masses. ADDITIONAL RESOURCES II The above information is additionally supplemented by: - Lecture module 1.2. (mandatory) 1.3 Rock Breaking Processes 5 This section is concerned with the basic principles of breaking brittle hard rock. The main considerations in breaking rock are the forces required to induce fractures in the rock and the energy consumed in breaking rock. Force is important because it determines the limitation on the type of machinery that can be used to break the rock and on the materials of construction that can be used in the machinery. As the breaking mechanism of the machine changes, so would the energy required to break the rock since the strength of rock varies depending on the type of stress induced on the material. Energy is important because it determines the rate at which rock breaking can be carried out. All machines are limited in the power that can be applied to the rock and hardness of the manufactured components of the machine. Therefore a process that demands substantial energy will result in a slow rock breaking rate. The rock breaking process is classified into three major groups: primary, secondary, and tertiary. Each process is described below. 5 From Dr. Paul Lever’s 415 course notes. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 17 1.3.1 Primary This is the application of a force by means of a hard indenter to a free rock face much larger than the indenter. This generates chips which are of a size similar to that of the indenter at the sides of the indenter and a pulverized zone immediately below the indenter. Figure 1-7: Primary Primary breakage processes would include the following: 1. Impact or hammering. Dynamic forces are applied 2. Percussive drilling. Application of a hard indenter to the bottom of a hole. The force is applied from one side only and the bottom of the hole is the free face. The force applied dynamically and after each application the hard indenter is moved slightly to break out more chips on the next application 3. Button type cutters for raise and tunnel borers. The buttons are loaded slowly (quasi-statically) and are moved away to be re-applied elsewhere, that is, indexing occurs by rolling to the next button. Repeated applications over a large surface area maintain the flat face 4. Disc type cutters for raise and tunnel borers. Hard indenter indexed by rolling. Forces at a point in the rock rise very slowly. 5. Drag –bit. A hard indenter forced onto the rock and indexed by dragging across the surface. 6. Diamond bits. A very hard surface and very small indenter dragged across the surface. The real breaking is done by the force thrusting the diamonds against the rock. Diamonds produce very small fragments because they are small indenters. 1.3.1.1 Theory The theory of the forces and energy in the primary breakage process relates to the confinement of the rock and the energy of the indenter. In the primary breakage process, the rock surrounding the area immediately under the indenter provides confinement for the rock so that stresses are very much greater than the uniaxial compressive strength (confined stresses for rock is higher than uniaxial). Figure 1-8shows how the size of the indenter is directly related to the size of the indenter. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 18 F a Figure 1-8: Stress relating to size of indenter The stress to penetrate the rock depends on the size of the indenter (note that the nomenclature of this process is such that a is the width or diameter of the implement striking the rock): a constant = p o The force to penetrate is therefore is equal to: 1.5 2 a a a constant indentor of area stress p p p p k F F F = × = × = The energy for primary breaking is directly proportional to the stress multiplied by the strain. However, considering an elastic behavior, stress is directly proportional to the strain (Young’s modulus). Therefore the primary breaking energy is directly proportional to the square of the stress. a erefore th a constant finally, so or strain, stress since however, strain stress 2 2 p p p p p p K E E E E E = | | . | \ | · · · · · × · o c o oc Note that these formulae are only valid when tools are driven to form the first chip. As penetration increases, forces and therefore required energy also increase. This may be due to the increasing confinement of the rock. Practically speaking, this effect would essentially cause the button/indenter to be buried. Indexing (moving through rotation of the bit or roller) the indenter to a new free surface would allow the indenter to penetrate with a force or stress closer to 1 st chip formation. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 19 o or Force Penetration depth 1 st chip formed 1 st chip formed 2 nd chip formed 2 nd chip formed 3 rd chip formed 3 rd chip formed Figure 1-9: Increasing energy with depth Side note on indenter shape: From the above, it can be seen that indenter shape directly impacts the energy required to penetrate the rock. Some drill bit manufacturers sell drill button bits with ballistic or circular bits. This can be seen in Figure 1-10 which compares a ballistic shaped bit with a circular or dulled bit. Figure 1-10: Penetration depth of various button shapes 1.3.1.2 Example 1 Determine the boring rate of a 75 kW raise boring machine using roller cutters with buttons of 1 cm diameter and boring head of 2 m diameter. Assume correct thrusting, that is, a chip forms with every pass. With buttons of 1 cm the mean chip size will also be about 1 cm. 3 2 - 6 -2 MJ/m 150 10 10 5 . 1 a energy fracture m 10 a = × = = = p p K E MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 20 m/hour 57 . 0 m) (1 hr / m 8 . 1 r hr / m 8 . 1 area hole rate breaking rate Boring hr / m 8 . 1 hour seconds 3600 s / m 2000 1 MJ/m 150 J/s 10 75 s / m energy fracture Power ume energy/vol Power rate Breaking J/s 10 75 kW 75 Power 2 3 2 3 3 3 3 3 3 3 = = = = = × = × = = = × = = t t 1.3.1.3 Example 2 Determine the boring rate of a 75 kW raise boring machine using disc cutters on a 2m diameter head. Assume correct thrusting. The shape of the indenter is not very well defined with the disc cutter. However, a few centimeters of the disc are in contact with the rock as the disc rolls and chips are long and narrow with a width of about 4 cm. Take the width of the chip as the value for a. m/hour 3 . 2 m) (1 hr / m .2 7 r hr / m .2 7 area hole rate breaking rate Boring hr / m 2 . 7 hour seconds 3600 s / m 1000 2 MJ/m 5 . 37 J/s 10 75 s / m energy fracture Power ume energy/vol Power rate Breaking MJ/m 5 . 37 10 4 10 5 . 1 a energy fracture m 10 4 a 2 3 2 3 3 3 3 3 3 3 2 - 6 -2 = = = = = × = × = = = = × × = = × = t t p p K E Note that from the previous 2 examples, the energy for boring with disc cutters is four times smaller and the boring rate is four times greater than with the button cutters. This is because the fragment size with the disc cutters is four times larger. 1.3.2 Secondary This is the application of forces inside a hole near to the rock face. The forces inside the hole generate tension at the sides of the hole to which produces cracks which ultimately run to the free surface. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 21 Rock Free face F Cracks Rock Free face F Cracks Figure 1-11: Secondary breakage process Secondary breakage processes includes: 1. Wedging. Wedge driven into a hole which produces crack 2. Blasting. Explosive generates a pulverized zone through compression but the real breaking process is by driving tensile cracks. Specific theoretical aspects of energy required for secondary breakage will be discussed in further blasting aspects of the course. 1.3.3 Tertiary This is the application of forces from more than one side to a free surface. F F F F Figure 1-12: Tertiary Breakage Process Tertiary breakage processes include: 1. Breaking boulders by impact or mud blasting 2. Crushing 3. Milling According to theory, the tertiary breakage process is closely related to breaking the rock in tension. From Figure 1-13 is can be seen that loading of a sphere by diametrically opposed forces causes a uniform tensile stress across the diametrical plane. This causes the sphere to split in tension, that is, at a stress very much lower than the uniaxial compressive stress. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 22 F F o t o t Figure 1-13: Tertiary Breakage a tensile effect It has been found that the tertiary stress, o t , is also dependent on the size of the rock, but not as important as the size of the indenter for primary breakage. Larger boulders break at lower stresses, hence o t can be represented as: 0.25 a constant ~ t o Therefore the splitting forces could be calculated as: 1.75 2 0.25 a a a constant a sphere of area stress t t t k F = × = × = × = o It should also be noted that F >>F t Finally, the energy for tertiary breaking is derived by: a a constant squared stress strain stress 2 0.25 2 t t t K E = | . | \ | = = = × = o 1.3.3.1 Example Milling reduces rock to 70% minus 75 µm. Milling typically consumes 25 kWh/t. Compare this value with that predicted by the simple formula for tertiary rock breaking processes. 3 3 3 MJ/m 243 m 7 . 2 kWh 3.6MJ kWh/t 25 milling t/m 2.7 rock of Density MJ 6 . 3 hr min 3600 1000 kWh 1 = × × = = = × = t E According to the formula, tertiary breaking energy requires: a K E t t 6 10 5 . 1 a × = = MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 23 If 70% of the rock is reduced to minus 75 µm, then the mean fragment size is about 50 µm. 3 6 6 -6 MJ/m 214 10 50 10 5 . 1 m 10 50 a = × × = × ~ ÷ t E The simplified E t formula gives a reasonably good estimate of energy consumed in milling. 1.3.3.2 Example A jaw crusher is driven by a 10 kW motor and is set to produce fragments of 1 cm. Determine its ‘crushing’ capacity in tons per hour. t/hour 6.5 2.7 2.4 rate crushing t/m 2.7 rock of density / m 4 . 2 3600 10 15 000 , 10 E Power rate crushing MJ/m 15 10 10 5 . 1 m 10 1 a 3 3 6 - t 3 2 6 -2 = × = = = × × = = = × = × = ÷ hr hr s E t 1.3.4 Miscellaneous Several other breakage processes exist, these include: 1. Thermal spalling. This depends on intense heat (flame) being applied to the rock and traversed so that a high temperature gradient is produce in the rock resulting in differential expansion which produces mechanical strains and ultimately breaking of the rock. It is used in taconite and certain quarrying operations, usually in cold climates. Thermal spalling is also used for finishes on rock surfaces and where high forces must be avoided during breakage. (commonly used ancient technique) 2. Water jets. The water jets create high stagnation pressures against the surface it impinges on. Used for drilling in porous hard rock where water goes into pores and breaks grains out. Water jets are known to be wasteful on energy and are used only for special applications. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 24 Table 1-8: Summarizing Theory of Forces and Energy in Breaking Processes Breakage process Primary Tertiary Fracture Force (in Newtons) 1.5 a p p k F = 1.75 a t k Fracture Energy (J/m 3 ) a p p K E = a t t K E = Typical value k k p =10 8 kt =10 7 Typical value K K p =1.5 x 10 6 K t =1.5 x 10 6 Only for a in meters and valid for a << 1.0 ADDITIONAL RESOURCES III The above information is additionally supplemented by: - Readings – Chapter 9.1 – Mining Engineering Handbook. (Mandatory, this covers modules 1 and 2) - Lecture module 1.3. (mandatory) - Assignment 1 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 1 Module 2: Drilling and Blasting Components 6 The following notes provide a summary of the tools and theory involved in drilling and blasting. This covers basically drills and explosives. This module is organized into four main sub-modules: 1. Drilling mechanisms 2. Drilling equipment 3. Explosives Introduction 4. Explosives Products 2.1 Drilling Mechanisms Drilling is used in several industries and purposes. These various sectors are briefly listed: - Mining : drilling and blasting of rock 6 These notes were assembled directly from the following references: - From Dr. Paul Lever’s 415 course notes Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2 nd ed. 1992 - Hartman, Howard L. and Jan M. Mutmansky, Introductory Mining Engineering, 2 nd ed. New York: John Wiley & Sons. 2002, 570p. - Atlas Copco, The Raise Boring Handbook 2 nd ed. - Stefanko, Robert. Coal Mining Technology Theory and Practice. Littleton CO.:Society of Mining Engineers. 1983 - McKercher, R.M., Ed. Potash Technology: Mining, Processing, Maintenance, Transportation, Occupational Health and Safety, Environment. Pergamon Press: Toronto. 1983 - Mining-Technology.com, search: continuous mining - Bell, F. G. ed. Engineering in Rock Masses. Butterworth-Heinemann: London. 1992. ISBN: 0 7506 1063 8 - Caterpillar Performance Handbook, 28th Ed. - 2001 Class notes, from Bob Cummings - 1997 Class notes, from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department.Wirtgen America Inc. Surface Mining Manual. Edition 2002 - Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2nd ed. 1992 - Kennedy, B.A. Editor. Suface Mining 2nd Edition. SME: Port City Press, ML. 1990. - Persson, Per-Anders, Holmberg, Roger, and Lee, Jaimin. Rock Blasting and Explosvies Engineering. CRC Press: New York. 2000. 534 p. - Cook, Melvin A. The Science of Industrial Explosives. IRECO Chemicals: Salt Lake City, Utah. 1974. 449p. - Sen, Gour C. Blasting Technology for Mining and Civil Engineers. Sydney: University of New South Wales Press. 1995. 146 p. - The History of Explosives. http://sis.bris.ac.uk/~dj9006/explosives/history.html - Hustrulid, William. Blasting Principles for Open Pit Mining. Vol. 2 – Theoretical foundations. Rotterdam: A.A. Balkema. 1999. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 2 - Exploration: diamond core drilling, percussive, churn, hammer, rotary (collect chips as samples) - Quarrying: including specialized, small diameter (to create slabs) - Construction: Intermediate drill sizes (typically to reduce excess damage, to maintain slope) In focusing on mining applications, several features can be remarked: - May need different drills for ore and waste, due to the variances in hardness and for grade control - May need different drills depending on the location of drilling (air pressures) - Underground typically uses smaller holes - Surface mines typically use very large diameter holes - Drill selection can depend on bench height (surface) or direction of drilling (underground) The four main components in drilling include: - Energy source (drill and compressed air) – where the drill converts energy (such as the potential energy in compressed air) into mechanical energy. - Energy transmitter – (steel) – the steel is the primary mover of the bit - Energy applicator – (bit) – the bit attacks the rock mechanically - Fluid – cleans hole, lubricates bit & hole wall, controls dust, cools bit, stabilizes hole. The fluid can be water, air, or mud. There are several losses of energy once the drill has converted the original energy into mechanical energy: - Compression of steel - Bending - Elastic strain @ couplings - Internal friction in the drill - Heat - Side friction of cuttings MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 2 The two primary rock drilling functions are: - Penetration – where the fracturing or breaking of the material from in-situ - Circulation – where the debris is ejected. Figure 2-1 shows the drilling process. Three components of the drilling process should exist at all times. If any one of these components is ineffective, the drilling process will deteriorate rapidly and the information obtained from the hole will be invalid. Figure 2-1: The drilling process 2.1.1 Percussive Drilling Percussive drill bits break rock by indentation. The peak stress applied to the bit by the drilling machine causes the bit to strike the rock at a stress calculated by: cv p µ o = Where: v - bit velocity at the rock face µ - density of the impacting body, steel = 7850 kg/m3 c - compression wave velocity of the impacting body (~5000 m/s) Hence, o p =39.25v MPa for units of m/s Figure 2-2 shows the mechanical components and the nomenclature of a typical percussion drill. Either compressed air or water pushes the piston in the drill to strike the steel in a repetitive manner. The energy to break the rock is imparted through the drill bit by the compressive stress wave imparted by the piston to the steel. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 3 Figure 2-2: Components of a percussion drill – piston and steel Issues of drill efficiency and energy losses associated to coupling, the mechanics of energy along the drill steel are presented by the following discussion. The process of imparting energy into the rock begins with the piston being forced at high speed into the drill steel as seen diagrammatically in Figure 2-3. Figure 2-3: Piston moving toward steel When the piston strikes the steel the kinetic energy is transformed into a compressive wave imparted to both the steel and piston as seen in Figure 2-4. Figure 2-4: Compressive waves in both piston and steel These waves travel at a velocity c, to the ends of the piston and drill steel where they are reflected. The nature of the reflected wave depends on the boundary condition at the ends. For a bar with a free end (piston), the wave is reflected as a tensile wave, whereas at a fixed end it is reflected as a compressive wave. The piston remains in contact with the steel until the tensile wave returns to the steel-piston interface and causes the two to separate. This ends the pulse upon which the hammer imparts energy onto the steel. Calculating the duration time of the pulse has been used to calculate the yield strength for the drill before the machine would break. The reflection and separation effect is shown diagrammatically in Figure 2-5. Since the drill steel is in contact with the rock (not directly fixed), some of the energy is imparted into the rock to break out chips and the remaining energy returns as a compressive or tensile wave. The type of wave depends on the rock type, and contact between the drill steel and the rock. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 4 Figure 2-5: Reflections and wave transmission. Consider the case where the piston has a length l 1 , and the drill steel has a length l 2 , and where the diameters of the piston and drill steel are the same. The length of the compressive wave set up in the steel would therefore be equal to 2l 1 divided by the speed of the piston. c l t s 1 2 = Some have observed that the yield strength of steel used for the piston and drill steel limits the stresses that can be applied to the rock by using the piston-type devices. The maximum stress in the piston o h and the drill steel o t are given by: ( ) h t t h A A A cv + = µ o ( ) h t h t A A A cv + = µ o Where: v - bit velocity at the rock face µ - density of the impacting body, steel = 7850 kg/m3 c - compression wave velocity of the impacting body ~5000 m/s A t is the cross-sectional area of the drill steel A h is the cross-sectional area of the piston Therefore, where A t and A h are equal, o h = o t = µcv/2 = 19.63v MPa. If it is generally desirable to limit the stresses in these steel components to about 212 MPa, the maximum piston velocity is equal to about 10.8 m/s. However, these velocity limits are only for hydraulic drills as pneumatic drills cannot generate such high velocities. Not all the energy in the fluid in drilling is expended in breaking the rock. Sources of energy loss include: - Friction at coupling and other contact points, - Bit wear - Noise MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 5 - Vibration. - Flushing (positive work) - Rotation of steel (positive work) Machine mounted drills are known as drifters but most commonly as jumbo drill rigs. Most of the drills in service are pneumatic, meaning compressed air provides the energy to move the piston. Hydraulic, using water as the energy transmission method, drills are also used but are technically more sophisticated and are typically found on jumbo drill rigs. The advantages and disadvantages of the hydraulic and pneumatic drills are listed below. Table 2-1: Comparison of Hydraulic and Pneumatic drills Advantages Disadvantages Pneumatic - Low cost - Traditional, well established technology - Simple mechanical components - Most mines have compressed air lines through the drifts - Very low efficiency (less than 5% of the input power to the compressor is delivered to the rock) - Poor environment for the machine operator (machine is noisy and lubrication causes a fog) Hydraulic - More efficient (25-30% of energy input power delivered to rock - 8-10 decibels less than pneumatic drills (quieter) - Components more complex - More capital outlay Table 2-2: Comparative table Typical Hole Sizes Typical Hole Lengths Drill type Inches mm Ft m Jacklegs / stopers 0.75 – 1.25 19 - 32 1 – 4 0.3 – 1.2 Drifter / Jumbo 1.50 – 2.25 38 – 57 4 – 100 1.2 – 30.5 ADDITIONAL RESOURCES IV The above information is additionally supplemented by: - Lecture module 3.1 (mandatory) - Readings, Chapter 2: Rotary Percussive Drilling, from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (Mandatory) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 6 2.1.2 Rotary Drilling In rotary drilling, the bit attacks the rock with energy supplied to it by a rotating drill stem. The drill stem is rotated while a thrust is applied to it by a pulldown mechanism using up to 65% of the weight of the machine. The bit breaks and removes the rock by either a ploughing-scraping action in soft rock, or a crushing and chipping action in hard rock, or a combination of the two. Compressed air or water is supplied to the bit via the drill stem. The air cools the bit and provides a flushing medium for the cuttings. The drills typically operate in the vertical position although many types can drill up to 25 to 30 degrees horizontal. Figure 2-6 provides the nomenclature and dynamic movement for rotary surface blasthole drills. Figure 2-7 provides the deck plan for a P&H 100 rotary drill. Figure 2-6: Dynamics of surface blasthole drill MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 7 Figure 2-7: Floor plan for rotary drill P&H 100XP Rotary drilling is one of the most popular drilling techniques for large surface mines where large diameter holes are used in blasting. Figure 2-8 shows how rotary tri-cone bits are the ideal application in hardrock surface mines. Figure 2-8: Selection of Drilling method MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 8 2.1.2.1 Rotary drill string components The drill string consists of three main systems: - The suspension and drive that connect the string to the rig and power system - The drill pipe that transmits power, fluid, and cuttings - The drill tools and bit that advance and shape the hole and provide samples. Figure 2-9: Components in rotary drilling 2.1.2.1.1 Stabilizers The efficiency of rotary cutting is lost when the drill pipe is allowed to bounce around or when it fails to run smoothly in the center of the hole. The bit is steadied and hole maintained on grade by the use of stabilizers. There are two primary types of stabilizers, blade and roller (Figure 2-10). Blade stabilizers are designed to center the drill collars in the hole. They are run between or above the drill collars. Blade stabilizers come in many forms, some have spiraled blade arrangements, other run blades vertically along the stabilizer length. Roller stabilizers have throwaway or replaceable rollers. They have the advantage over blade stabilizers by providing better stabilization, longer life, and lower torque requirements than blade types. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 9 Figure 2-10: Blade and roller stabilizer The advantages of using stabilizers should be considered when designing or selecting a drilling process. The primary reason is costs. These costs include reduced cost per ton of rock removed, the more efficient use of expendable items such as rock bits and drill steel can be achieved. Rock Bits are designed to rotate about their own center. Stabilization assures that the bit will do this and thus cause the energies and forces exerted on it to be most efficiently utilized in an axial direction. Lateral movement or stumbling is restricted and hole is produced in the direction intended. More footage per bit and an increased penetration rate is achieved by proper utilization of the forces applied to the bit. Dull conditions of bits give evidence of this effective stabilization. Gage wear is uniform and less severe. Shirt-tail wear problems are reduced. The inner row cutting structure is more uniformly worn and inner and outer flank wear is reduced. Stabilizers also affect drill stem life. Without stabilization, rough spiral bores, ledges and other unconformities are obtained. The possibilities of crooked hole are enhanced. Drill steel rotating in these rough and crooked bores scrubs and scrapes against the bore wall and thereby abrades. Stabilizers also affect drill availability. The smooth bore produced with adequate stabilization permits faster rock bit and drill steel retraction from the bore hole. Furthermore, the smooth bore sloughs less than a rough bore. This means that less rock particles fall to the bottom of the hole. Re-drill time required is normally eliminated or drastically reduced. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 10 In the selection of a stabilizer, one must take into account the adequate stabilization. Unless the guiding elements of a stabilizer are very near the bore wall, adequate stabilization of the bit is not achieved. Theoretically, the guiding elements should have the same diameter as the bit. Unfortunately this is not practical because of the normal attrition of rock bit gage wear surfaces. The stabilizer should therefore be held at the largest diameter practicable. Concentricity (the quality of having the same center - as circles inside one another) of guiding elements with the axis of the bit and steel is also quite important to proper stabilization. Eccentricities (different centers) of these elements tend to void any hope of reducing drilling costs with a stabilizer. The stabilizer which maintains guiding elements close to hole wall is most efficient. The conventional welded rib or cast rib stabilizer (also known as blade) does an adequate job when new and unworn, provided the ribs are concentric. The problems with this type of stabilizer are wear rate and maintenance cost. Rib-type stabilizers, due to their construction, necessarily drag and scrape against the bore wall and abrade rapidly. Due to the cost in maintaining close-to-hole wall contact these stabilizers are left in the drill string long after adequate stabilization has been lost. The most efficient stabilizer is one with true-rolling centralizers that are in rolling contact with the bore wall. Scraping and gouging of the guiding elements are eliminated and co-centricities are maintained. Roller-stabilizers provide adequate stabilization without imposing additional torque 2.1.2.1.2 Substitutes (Subs) Substitutes (known as subs) are used as adapters where threads of one size or type must be coupled with threads of another size or type. They are also used at points of heavy wear to provide a readily replaceable thread. Shock absorber subs are run above hammers or above the bit or rod in rough drilling situations, thus reducing drill string vibration and increasing bit life. A swivel mount shock sub is also used to dampen vibration effects on the rig structure thereby reducing maintenance on the rig and mechanisms for pulldown and rotation. Figure 2-11: Swivel mount shock sub MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 11 Shock and vibratory energy transferred between drill and bit are absorbed through a series of rubber bonded segments. The absorbed energy is released to the atmosphere in the form of heat. A. The pulldown or drilling weight to the bit is transferred through the segment rubber placed in shear. B. The rotational torque (or torsional loading) to the bit is transferred through the segment rubber placed in compressive loading. C. The accelerated changes in axial and torsional loadings are minimized by the segment rubber. Figure 2-12: Vibration reducing effects of using shock subs The shock absorber is most beneficial when drilling in fractured formations, intermittent hard and soft layers or hard formations. The benefits include: - Reduces drilling machine maintenance by dampening torsional and axial shock loads. - Increases drilling rates by keeping bit in more uniform contact with formation. Allows use of more weight and higher rotary speeds in rough drilling areas. - Increases bit life by dampening cyclical shock loading normally transmitted to the bit bearings and cutting structure. - Decreases operator noise level by eliminating the metal to metal contact between rotary drive and drill pipe. Normally, no modification of the drill is required and installation can be completed in a short time. On multi-pass drills where length is not critical, a saver sub is recommended for use between rotary shouldered connection and drill pipe. 2.1.2.1.3 Rotary Drill Stems Factors involved in selecting drill stems (aside from bailing requirements) includes: - Fabricated (welded) or integral (machined from single steel bar) drill steel; - Thread size and type - Wall thickness - Types of connections MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 12 Figure 2-13: Blasthole components with respect to drill rig 2.1.2.2 Rock drillability / Penetration rate Rock drillability is defined as the penetration rate of a drill bit into the rock. It is a function of several rock properties such as: - Mineral composition; - Texture; - Grain size; - Degree of weathering. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 13 As is obviously important to costs and productivity, penetration rate is one of the most important factors in drilling. Some empirical equations have been developed from extensive rotary tests in iron ore. The Bauer and Calder method states that penetration rate, P, can be calculated using the following equation: 300 ) log 28 61 ( 10 rpm W S P c - ÷ = | where p = penetration rate in ft/hr Sc = uni-axial compressive strength, in thousands of psi W/| = weight per inch of bit diameter, in thousands of pounds Rpm = revolutions of drill per minute The adjustable factors in the above equations are the variables controlled by the operators. Rock compressive strength is not but can be estimated using the graph in Figure 2-14. Figure 2-14: Penetration rate vs. Rock strength Several other indices exist for rock drillability which include: - Drilling Rate Index; - Classification of rock types based on the drillability of Barre granite - Mohr’s test - Indices (sensor-measured), e.g. AQUILA’s drill performance monitoring MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 14 2.1.2.3 Rotary Drill Pulldown Weight A portion of the machine weight is applied by the pulldown motor via the pulldown chain or chains, rotary head and drill stems to the drill bit. Figure 2-15 illustrates recommended bit loadings for different bit sizes. As the bit diameter increases, the bearing size increases thus allowing an increase in the tolerable load. Overloading the bit results in severe loss of bit life as illustrated in Figure 2-16. Figure 2-15: Recommended pulldown weights per inch of bit diameter. Figure 2-16: Bit life vs. pulldown weight for 9 1/4 inch diameter rotary bits in hard formation There are several method of rotary drive. The bit may be turned by: 1. rotating a rotary table which turns the pipe as it slides through. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 15 2. rotating the pipe directly by a drive unit which moves down with the pipe. 3. rotating the bit, using a down hole turbine drive. Figure 2-17: Rotary drive types Figure 2-18: Pulldown mechanisms Figure 2-19: Rack and pinion pulldown MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 16 General parameters for drill penetration can be seen in Table 2-3. Table 2-3: Average Drill Bit Footages (from large Iron Ore mine in Canada) 2.1.2.4 Flushing medium Air is used to bail the drill cuttings from the hole as well as cool the bit bearings and, when used, roller stabilizer bearings. Approximately 20% of the air is forced through the roller cones for cooling purposes by adjusting the air pressure across the bit using the bit nozzles. The air volume is the primary requirement for bailing cuttings from the hole. Air velocity up the hole is dependent on the air volume per minute as well as the hole annulus (ring-shape where hole and stem meet). The velocity of the drill cuttings in this air is dependent on the chip size, density, and shape. Experimentally, the balancing air velocity in feet per minute is given by: Um = 264p ½ d ½ Where: d = diameter of the chip in inches p = density of chip in Ib/ft3 At air velocities above this balancing value, the chips begin to move, their velocity being approximately one half the excess air velocity above the balancing value. A bailing velocity of 1800 mpm/s (6000 fpm) is usually adequate to bail 13 mm (1/2 in.) chips. Figure 2-20 illustrates a typical air requirements chart. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 17 Figure 2-20: Air Bailing Chart Factors involved with choosing the air velocity are that higher velocities: 1. give higher bailing velocities; 2. will bail larger chips; 3. tend to give higher bit life; 4. will help cater for hole cavities, etc.; 5. will help cater for drill stem wear; 6. may give higher penetration rates and possibly lower cost per ft; and 7. reduce the volume of cuttings in the hole for a given penetration rate; The drawbacks of increased bailing velocities include: 1. will give increased stabilizer and pipe wear; 2. increased dust deflector and deck bushing wear 3. may damage borehole walls in soft drilling MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 18 2.1.2.5 Operational tips The following are some operational tips on bit and rig for top performance when using air circulation bits. 1. Use straight drill steel with properly maintained threaded connections. 2. Use a good grade of thread grease and maintain connections properly. 3. Use care attaching and removing a bit from the drill pipe. 4. Open air valves before starting to drill with a bit and keep the air on until the bit is out of the hole. 5. Break in a new bit by drilling at reduced down pressure and rotational speed for a short period. 6. To collar or start new hole, reduce down pressure and rotation. 7. Re-establish bottom hole pattern with reduced down pressure and rotation when drilling is interrupted. 8. Never finish an old hole with a new bit. This can pinch the cones, damaging the bearings. and gauge teeth. 9. Always maintain drilling air pressure at appropriate levels. 10.Rotary speed should be decreased as down pressure increases. 11.Do not use more water than is necessary to control dust and maintain hole wall. 12.Maintain rotation while tripping into or out of a hole. 13.In wet holes, maintain as high an air pressure and volume as is possible. 14.Guard against dropping the bit and drill steel to the bottom of the hole at high speed. High speed impact is a common cause of damage. 15.Periodically inspect the bit for damage or impending failure. Non- uniform cone temperatures may indicate obstructed air courses and potential bearing failure. 16.Before an idle period, clean the bit by passing air or water through it while rotating the cutters by hand. After an idle period and before re- using a bit, make sure all cones turn freely by hand. 2.1.2.6 Safety Tips The following safety tips should be taken into account around large surface blasthole drills: 2.1.2.6.1 Moving the Drill As with all heavy equipment, care must be taken when blasthole drills are moved around the job site. Watch for people, overhead electrical power lines, other equipment, low or narrow clearances, ground-bearing limits, and steep hills or uneven terrain. Use a signal man when traveling and use an audible warning signal to alert people in the area before traveling. Know the height, width and weight of the machine. Always secure on-board equipment before traveling. Do not travel with people riding on the outside of the machine or inside the machinery house. 2.1.2.6.2 High Voltage dangers MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 19 Hazardous voltage from the machine contacting power lines can cause electrocution and burns to anyone near the blasthole drill. MSHA and OSHA regulations require at least ten (10) feet of clearance form overhead lines carrying 50,000 volts or less. Greater clearances are required for lines with higher voltages. Some local regulations require greater distances than OSHA or MSHA. 2.1.2.6.3 Trailing Cables Blasthole drills with electric tail cables can break the cable connection or crush the cable when traveling. Maintain cable slack while operating and traveling. Use signalmen during traveling to prevent damage to the cable or cable connection. Rotational Dangers During drill operation, keep all people off the drilling platform and drill mast, and away from drill stems. Moving components or rotating drill stems can entangle clothing and can crush, pinch or strangle personnel. 2.1.2.6.4 Tipping Dangers Exceeding the slope or grade limitations given for your specific machine and machine configuration can cause machine tip-over. Prior to propelling, always determine the slope limitations for your specific machine and never travel or operate on site slope conditions which exceed these limitations. When propelling with the mast elevated, always position the machine with the mast on the uphill side of the machine. When the mast is down, propel the machine with the operator's cab on the downhill side of the machine. Sudden tip-over can occur when raising, lowering, or leveling the machine with jacks. Always level the machine at the lowest practical height that will unload the crawler belts. Inspect the ground for lifting support and add cribbing or support mats before lowering jacks. When lowering the machine to the ground, always lower the machine slowly and in stages, maintaining a level condition until the crawler belts contact the ground and are supporting the machine. Use a signal person to assist in watching the jacks, crawlers and machine during the process. 2.1.2.6.5 Crushing Danger from Lowering Mast Before lowering the mast, notify personnel to evacuate the roof and drill deck and inspect mast storage area for obstructions. After storing the mast, check for secure mast attachment on the front jack caps. 2.1.2.6.6 Chips and Dust Drilling produces flying debris and dust which can cause serious respiratory disease. Always lower dust curtains before drilling and inspect the curtains to MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 20 make sure all dust curtains are installed and in the lowered position. Curtains should not be lifted to remove cuttings while the drill is drilling in the hole. Keep personnel off drilling platform and away from the drill hole while drilling is in process. Avoid dust contamination from work cloths, eating or drinking. Follow mine procedures for air monitoring, exposure limitations, and protection methods for crystalline silica exposure. Specific design/plan-related safety issues to consider - Where and when the blasting is scheduled for the day - The other equipment, powerlines, or structures which must be moved or avoided during movement of the drill. - The ground surface strength in terms of having adequate weight bearing ability for the drill mass - The grade of the slope which the drill will be working on (too sloped- will cause a tip-over) - Steps necessary to keep unnecessary people and equipment at a safe distance from drill area. 2.1.2.7 Chip Sampling Penetrating always involves" chip making". Usually penetration efficiency is improved when large chips are produced and cleared quickly. The best chips are large chips; large chips often make better samples. Large chips need: - bits with sharp cutting edges. - large amounts of energy per chip. - rapid clearing and transport up the hole to prevent regrinding of the chip. Although larger amounts of energy are required for each large chip, fewer chips are produced and the total energy used per metre of hole drilled is less (less surface area produced. Many rigs produce good chips at the bottom of the hole but fail to clear them or lift them out of the hole before they are broken up. For best chips, the operator should use the drill as follows: - Use sharp blade bits or long toothed roller bits with bottom clearing nozzles. - Use high thrust or feed. - Keep hydrostatic head to a minimum. - Use high flushing flow of low viscosity, low solids fluid (must be balanced against erosion of hole wall). - Use reverse circulation techniques for broken, fractured, cavernous or other formations prone to lost circulation problems. 2.1.2.8 Single pass Most drills in surface mines use single pass drilling, either by blast design or through purchasing equipment with taller masts. Single pass drilling is MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 21 where the entire length of the hole is drilled without having to add more drill stem. The advantage to single pass drilling includes: - Eliminates adding stems - Reduces associated thread damage - Reduces machine downtime for rod changing equipment and tool racks - Facilitates the cleaning of boreholes - Permits a continuous air flow through the bit at all times. Disadvantages of single pass are: - High masts make the drill more instable - Extra care is needed in moving drill long distances - Pulldown chains become long and may require special attention 2.1.2.9 Capacities. Key operating capacities for a drill rig includes: - Maximum mast loading capacity - Maximum hoist or hook load capacity - Head torque capacity. 2.1.2.10 Productivity Estimate Figure 2-21 shows a generalized productivity estimate for various materials. Productivity calculations would be site-specific and should take into account: - set-up time, - movement time, - idle periods, - bit and stem changes - penetration rate - Burden and spacing - bit diameter (alters the penetration rate) - bit life (changes the frequency of bit and stem changes) - hole depth (changes the frequency of the movement time) - number of holes (changes the frequency of set-up time) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 22 Figure 2-21: Tons drilled per operating hour for rotary drills for various hole diameters. Figure 2-22 shows an estimate of the cost per meter. The drilling costs are made up of two different parts, the drill consumables per unit length of drilled hole and the balance of the drill cost. The balance of the drill cost is converted to the drill cost per unit length of hole using the drill penetration rate as the drill cost would be time dependent. The reason for this split is that the drill consumables cost, predominantly the drill bit cost, is independent from the drill penetration rate, assuming correct operating methods (poor operation can cause increased bit wear life), whereas all other costs are penetration rate dependent. Drilling records or statistics are therefore important to record so that accurate budgeting can be undertaken both in terms of costs that are incurred along with the output from the drill. Determining the drilling costs must take into account, the following factors: Ownership costs: include amortization and depreciation, interest on borrowed money (if not purchased out of cash-flow, and taxes and insurance. Operating costs: includes power, maintenance, direct and indirect labor, warehousing, and consumables. These consumables are dependant on penetration rate and wear rate and would include items such as stabilizers, drill stems, and drill bits (at times one of the most significant cost). MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 23 Figure 2-22: Rotary Drilling costs estimates by meter drilled and diameter of hole. Figure 2-23: Cost per ton of material excavated 2.1.2.11 Drilling technology Most advanced modern drilling technology focuses around the navigation and positioning of the drill. GPS based navigation systems are available from companies such as AQUILA (Caterpillar owned). The company has a both a drill and shovel product line that uses GPS and PLCs to enhance drilling productivity and provide susbstantical information from which better mine planning can be undertaken. AQUILA DM systems are available in two ways: as factory installs on new drills, or can be installed as retrofits. There are five main options for the drilling systems: DM-1: Production Monitoring System DM-2: Material Recognition System DM-3: Drill Control System DM-4: Guidance system for vertical drilling DM-6: Guidance system for inclined drilling MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 24 2.1.2.11.1 DM-1 Production Monitoring System DM-1 uses AQUILA's Advanced Monitoring Platform (AMP) with Graphical User Interface (GUI) software to give the operator immediate feedback on drilling productivity and performance. The DM-1 is designed to minimize the amount of operator effort, which means faster work with fewer errors. For example, the start of drilling can be automatically detected, so the operator doesn't have to zero the bit depth. Steel changes are also automatically sensed, eliminating errors in determining hole depth. Figure 2-24: DM-1 production Monitoring System 2.1.2.11.2 DM-2 Material Recognition System DM-2 uses a vibration sensor and pattern recognition software to automatically process and analyze drill variables, and determine hole geology while drilling. DM-2 pinpoints the location of ore and waste interfaces, then delivers real-time information via a color LCD screen. The result is immediate, accurate-to-the-centimeter information for both the operator and planning engineers in the mine office. Figure 2-25: DM-2 Material Recognition System MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 25 DM-1 and DM-2 functioning in tandem would enable the following: - Detailed, real-time geological information on a location-by-location basis. - Improved explosives usage, better fragmentation and reduced ore dilution. - Wireless connectivity to and from the mine office. - Comprehensive production reporting. 2.1.2.11.3 DM-3 Drill Control System DM-3 delivers consistent drill operation and performance in all kinds of conditions, for operators of all skill levels. By interfacing the DM-1 or DM-2 to Programmable Logic Controller (PLC) modules and electrohydraulic actuators, DM-3 makes automatic hole drilling possible - from collar to design depth. It regulates pulldown pressure and rotary speed within the most productive limits for torque and vibration. This means more consistent performance, with optimized penetration rates for changing downhole conditions. And the ultimate benefits of longer drill bit life and more productivity. Figure 2-26: DM-3 Drill Control System 2.1.2.11.4 DM-3 provides: - Reduced maintenance and consumables costs. - Optimized operation for overall productivity. - More consistent hole depths and more even benches. 2.1.2.11.5 DM-5 Guidance System The DM-5 Guidance System for Vertical Drilling combines high-resolution, Real-Time Kinematic (RTK) GPS receivers and an AMP platform. This allows the operator to place a blasthole within centimeters of his target, without the need for traditional surveying or staking. Once the drill is leveled and the hole started, the DM-5 automatically determines collar elevation, calculates the required drilling depth and displays the information on an easy-to-read LCD screen. And all the information on blasthole positions is stored by the DM-5 and transmitted to the mine office for use in blast design and updating the geological model. This technology allows the operator to navigate to each MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 26 blasthole via a navigation screen. The figure below shows the view provided to the operator during the navigation phase of drill positioning and the second picture shows the operator’s view when within 1 meter of the designed hole location (the view automatically zooms-in). Figure 2-27: Operator’s view when navigating The location where the drill finally punched the hole, along with sensor data (that can be interpreted as geological information) can be recorded and sent back to the engineering office. The positioning information is provided by differential GPS, from two receivers on the head of the mast of the drill. Figure 2-28: GPS receivers 2.1.2.11.6 DM-6 Guidance System for Inclined Drilling MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 27 DM-6 has the same basic features as the DM-5, but is specifically designed for use on machines that routinely drill inclined blastholes with angles between five and 30 degrees from the vertical. For inclined drilling, the DM-6 uses a servo-controlled Automatic Leveling System (ALS) that automatically maintains the GPS antennas level with the horizon. The DM-6 also uses software designed to facilitate drill setup and improve alignment accuracy when drilling inclined holes. And like the DM-5, this system also stores all blasthole positions, which are transmitted to the mine office for updating the design file. Using both DM-5 and DM-6 allow: - Greater hole location and depth accuracy for both inclined and vertical drilling. - Time and money savings from no-stake marking and better blast results. - The elimination of pre- and post-blast pickups. 2.1.2.11.7 Case Study: Highland valley Copper mine A post-audit was undertaken in 1998 of the implementation of AQUILA technology on three B&E 49Rs at the Highland Valley Copper mines. The benefits achieved after only 1 year of utilization are listed below. It should be noted that further development or utilization of the the data provided by the drill were used in other initiatives. 1. By designing the blast pattern on an office computer (as opposed to on-site by the surveyors), engineering considerations can more readily be taken into account. 2. HVC was able to place up to five surveyors (who formerly would be marking blast patterns) on other revenue generating projects within engineering. 3. The creation of a pattern no longer requires a large amount of survey consumables. 4. Adverse climatic conditions no longer impede the blast design and implementation. 5. There is consistency in blast design (only one technician does all of the designs). 6. Design time has been reduced (no surveying is required). 7. Much less skilled drill operators can operate the drill, since positioning the drill is made easy through the ‘video game’ type navigation interface. This has also tended to increase the productivity of the drill through a reduction in tramming and positioning time. 8. Supervision was increased as the productivity and delays are recorded and delivered in the log file. 9. Any need to re-survey the drilled pattern was eliminated (according to mine regulation, the position of each hole must be known so that bootlegs are avoided.) 10. The most important change incurred by the implementation of the technology was the increase in collaring accuracy (previously, errors of MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 28 up to a meter in an X & Y direction were common, the current average error is 10 cm in X & Y). The misalignment of the blastholes induced poor fragmentation. Engineering staff consider that such oversize caused increased wear on shovels, trucks, and crushers. It was also observed that poor fragmentation would adversely affect the mill. Throughput would fluctuate when processing such ore, resulting in decreased recovery. ADDITIONAL RESOURCES V The above information is additionally supplemented by: - Lecture module 3.1 (mandatory) - Readings, Chapter 4: Rotary Drilling with Tricone Bits, from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (optional) - Assignment 4: Drill selection 2.2 Drilling equipment This sub-module will familiarize students with the equipment that, as engineers, they will manage as part of a wider fleet to ensure system efficiency. ADDITIONAL RESOURCES VI The primary component of this sub-module is provided by: - Lecture module 3.2 (mandatory) - PowerPoint presentations: material developed for 2002’s 415 Rock Excavation course. Available upon request 2.3 Explosives Introduction The physics and chemistry behind explosives and their detonation are introduced in this first sub-module on explosives. The second sub-module will cover the properties of explosives products. 2.3.1 Detonation Theory Blasting theory is one of the most controversial topics in the rock excavation industry. No single concept has been developed and accepted that fully explains the mechanisms of rock breakage in every situation, yet a vast amount of research work has contributed valuable information and insight into blasting theories. Some of the theories and findings are discussed in this section. An explosion is a self-propagating, exothermic reaction. The stable end products are gases that are compressed, under elevated temperature and very high pressures. It is the sudden rise in temperature and pressure from MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 29 ambient conditions that results in a shock wave, or a detonation traveling through the un-reacted explosive. The velocity of detonations (VOD) lies in the approximate range of 5000 to 30,000 fps (1500 to 9000 m/s), well above the speed of sound in the explosive material. Deflagration is the chemical burning of explosive ingredients at a rate well below the sonic velocity. It is associated with heat only and carries no shock due to its much slower reaction rate. Deflagration occurs when less than ideal hole-loading conditions or explosive formulation are involved. The maximum energy release upon detonation occurs when the explosive mix is formulated for oxygen balance. An oxygen-balanced mixture is one in which there is no-excess or deficiency in oxygen, such that the gaseous products formed are chiefly H 2 O (water vapor), CO 2 (carbon dioxide), and N 2 (nitrogen). In actual blasting practice, small amounts of noxious gases such as NO (nitric oxide), CO (carbon monoxide), NH 4 (ammonia), CH 4 (methane), and solid carbon, are formed resulting in nonideal detonations and somewhat less than ideal pressures and energies. Commercial explosive formulation attempts to achieve an oxygen-balanced mixture. The work done by chemical explosives in the fragmentation and displacement of rock depends on the shock energy as well as the energy of the expanding gases. The self-sustained shock-wave produced by a chemical reaction in a gaseous medium was described by D.L. Chapman and E. Jouquet as a space. This space of negligible thickness is bounded by two infinite planes - on one side of the wave is the un-reacted explosive and on the other, the exploded gases, as seen in Figure 2-29. Figure 2-29: Detonation Process of explosive cartridge There are three distinct zones: (a) the undisturbed medium ahead of the shock wave; (b) a rapid pressure rise at Y leading to a zone in which chemical reaction is generated by the shock, and proceeds until complete at X; and (c) a steady state wave where pressure and temperature are maintained. This condition for stability exists at hypothetical X, which is commonly referred to as the Chapman-Jouquet (C-J) plane. Between the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 30 two planes X and Y there is a conservation of mass, momentum, and energy. A simplified and approximate velocity of detonation (VOD) can be obtained from the following empirical relation: ) 3 . 1 1 ( µ + = J C d Where: C d is the VOD in m/s J is heat of reaction in MJ/kg µ is the specific gravity The detonation pressure P d in N/m 2 which exists at the C-J plane can be estimated with comparable accuracy using the equation below: 4 2 d d C P µ = Explosion pressure, p a , is the pressure when the product gases have the same density as the unreacted explosive. The explosion pressure is approximately half the detonation pressure. 2.3.1.1 C-J Plane A deeper exploration of the individual zones in the C-J theory should be undertaken. Note Figure 2-30 as being the idealized detonation wave traveling through a cylindrical explosive shape, producing an increase in pressure. The steady-state chemical reaction takes place behind the shock front within the reaction zone. At the end of this zone, a nonsteady-state region exists. It is created by a flow of expanding gases in a direction opposite to that of the traveling wave front. Once again, the C-J plane is seen as the boundary between the steady and non-steady state, where the reactions are considered complete. This is also the plane where all the thermodynamic properties are calculated. These are: p – pressure V – velocity T – temperature E – internal energy or Q – heat of formation and µ – density MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 31 Figure 2-30: Detonation Process for cylindrical explosive The maximum pressure and duration of a wave pulse is directly proportional to the shock energy and gas pressure of the explosive, respectively. High explosives such as military explosives or highly sensitive commercial explosives are characterized by an intense shattering effect upon detonation (known as brisance). They liberate gaseous products very quickly. The distance between the shock front and the C-J plane is very short and results in a pressure pulse of high amplitude and short duration. The pressure pulse for less-sensitive commercial explosives shows a decreased pressure amplitude and a longer pulse length. In this case, the reaction is slower and the gas volume is greater, as seen in Figure 2-31. Figure 2-31: Pressure shape for A-high explosive and B-commercial explosive containing high gas volume MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 32 2.3.1.2 Quick note on Shock Wave Propagation As the stress wave front generated from a blast travels outwards, it has a tendency to compress the material at the wave front through a volume change. There is a tangential or hoop stress at right angles to the compression wave front. If this tangential stress is strong, then radial failure from the explosive source is evident. When the compression wave front travels from one material to another, there could be three possibilities at the interface as seen in Figure 2-32. These possibilities depend on the material's acoustic impedance I r which is defined as the product of the material's density µ and sonic velocity V. As seen in the figure, the material in which the energy wave originates is labeled A, and the material into which the wave travels, B. When the ratio of acoustic impedance of material A to material B is less than 1, some of the energy is transferred to material B as compression waves, whereas the rest is reflected back as compression waves as well. When the acoustic impedance ratio is 1, all the energy is transmitted into material B as compression waves. When the impedance ratio is greater than 1, some of the energy travels through the interface as compression waves and the rest is reflected back as tension waves. When a compression wave traveling through a rock medium encounters an interface such as a free face, nearly all of the energy will be reflected back as a tension wave. If the distance between the free face and the explosive charge is relatively small, most of the energy will be spent in peeling off layers of rock at the free face. Figure 2-32: Shock wave propagation MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 33 2.3.2 Comparative Explosive Properties Explosives and blasting agents are characterized by various properties that indicate how they will perform under field conditions. These properties include fume class, density, water resistance, temperature effects, detonation velocity, detonation pressure, borehole pressure, sensitivity, and strength. Each of these properties will be covered. 2.3.2.1 Fume Class Class-Fumes are noxious gases that are produced from the detonation of explosives. The production of these gases is most critical in underground and other confined workings. Many factors affect the volume of poisonous gas produced including oxygen balance and adverse loading of explosives. The fume class is a measure of the toxic gases in cubic feet per 0.44 lb (200 g) of un-reacted explosive. Institute of Makers (IME) has developed a fume class classification scheme as seen in Table 2-4. The now-disbanded US Bureau of Mines (USBM) limits the volume of poisonous gases produced by permissible explosives (those used in underground coal and other gaseous mines) to 2.5 lb (1.14 kg). Table 2-4: Standards for Fume Class Class Volume of poisonous gas per 200g of explosive, in ft 3 1 0.16 2 0.16-0.33 3 0.33-0.67 2.3.2.2 Density The density of an explosive is defined as the weight per unit volume or the specific gravity. Commercial explosives range in density from 0.5 to 1.7. Explosives with a density less than 1 will float in water. Therefore, in water- filled holes, an explosive with a density greater than 1 is required. For certain granular explosives such as dynamite, density correlates to the energy released in a given borehole volume. However, for water-based explosives, this is not the case, and often the reverse is true. Density is most useful in determining the loading density or the weight of explosives one can load per unit length of borehole (in pound per foot or kilogram per meter). Note that knowledge of loading density is required for blast-design calculations, and is calculated in English units as: LD = 0.3405 µ D 2 Where: µ is density D is explosive column diameter in inches. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 34 2.3.2.3 Water Resistance The ability of an explosive to withstand exposure to water for long periods of time without loss of strength or ability to detonate defines the water resistance. A numerical rating is used based on the results of tests performed on the explosive. However, explosive manufacturers individually rate products based on a relative basis as good, fair, or poor rating. The presence of moisture in amounts greater than 5% dissolves chemical components in dry blasting agents and alters the composition of gases produced, contributing to the formation of noxious fumes and lower energy output. Gelled granular products have good water resistance, and certain water- based mixtures have an excellent rating. 2.3.2.4 Temperature Effects Extreme low temperatures affect the stability as well as the performance of explosives. The sensitivity and detonation velocity are hampered for certain water-based explosives at low temperatures while dynamites can become dangerously unstable below freezing temperatures. Explosives manufacturers recommend the appropriate range of temperature for storage and use. 2.3.2.5 Detonation Velocity The detonation velocity is the speed at which the detonation front moves through a column of explosives. For high explosives such as dynamite, the strength of an explosive increases with detonation rate. For dry blasting agents and water-based explosives, field loading conditions greatly affect detonation velocity. Such conditions include (not exclusive list): - borehole diameter - density - confinement within the borehole - the presence of water The speed of detonation is important when blasting in hard, competent rock where a brisance effect is desired for good fragmentation. For most explosives, there is a minimum diameter D min below which detonation velocity increases nonlinearly with increasing borehole diameter as can be seen in Figure 2-33. Above D min the explosive has reached its steady-state velocity. At this point, all thermodynamic properties are at a maximum as the reaction front approaches a plane shock front. At diameters less than D min , complete reactions do not take place, and less than ideal energy and pressure evolve from the slower detonation rates. This represents a loss in terms of dollars spent on explosive energy. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 35 Figure 2-33: Generalized relationship between VOD and Diameter 2.3.2.6 Detonation Pressure The detonation pressure is the maximum theoretical pressure achieved within the reaction zone and measured at the C-J plane in a column of explosives. The actual pressure achieved is somewhat less than this maximum due to non-ideal loading conditions always present in practice and due to certain explosive formulation. Most commercial explosives achieve pressures in the range of 0.29 to 3.48 x 10 6 psi (2 to 24 GPa). Although detonation pressure is related to the temperature of the reaction, a number of simplifying formulas are available for estimating detonation pressure for granular explosives based on detonation velocity and density, for example (in English units): P = 0.00337 µ V 2 where P is detonation pressure in psi, µ is density XXX V is detonation velocity in fps. 2.3.2.7 Borehole Pressure Borehole pressure is the maximum pressure exerted within the borehole upon completion of the explosive reaction measured behind the C-J plane. Such measurements cannot be made directly and are done during underwater tests performed for energy and strength determinations. With the use of hydrodynamic computer models, theoretical calculations of borehole pressures are made. There is little agreement in the literature regarding specific estimates of actual borehole pressures. In general, pressures after detonation within the borehole are estimated to be less than 30% of the theoretical detonation pressure. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 36 2.3.2.8 Sensitivity The definition of explosive sensitivity is two-fold. It includes sensitivity against accidental detonations in addition to the ease by which explosives can be intentionally detonated. From the standpoint of safety and accidental detonations, the sensitivity of an explosive to shock, impact, friction, and heat determine its storage and handling characteristics. Standardized tests for high explosives have been adapted for commercial explosives that include the friction (pendulum), impact (fallhammer), and projectile tests, among others. The term properly used to define the propagating ability of an explosive is sensitiveness. In this respect, tests such as the No.8 strength blasting cap test, air-gap test, and the minimum critical diameter test are used. The cap sensitivity test measures the minimum energy required for initiation and is used to classify explosives (e.g., cap sensitive vs. noncap sensitive products) or the ability to initiate an explosive directly with a standard cap. The No.8 cap is an industry standard cap of specific dimensions and charge characteristics. The air-gap test measures the distance between the ends of adjacent cartridge explosives for which reliable initiation can be propagated from one cartridge to another. The critical diameter of an explosive is the smallest diameter at which an explosive will maintain a steady-state detonation. Below this critical diameter, explosives may deflagrate or "dead press." Dead pressing occurs when an explosive is densified to a point that no free oxygen is available to ensure the start or progression of detonation. 2.3.2.9 Strength The strength of an explosive is a measure of its ability to break rock. The terms "weight strength" and "bulk strength" were useful many years ago when explosives were primarily comprised of nitroglycerin cartridges, packaged in 50 lb (23-kg) boxes. In recent years, with the development of bulk blasting agents and less sensitive ingredients, new testing methods have been established to determine relative energies for all commercial products regardless of ingredients or packaging. The performance potential of an explosive is a function of the detonation velocity and density, as well as the volume of liberated gases and the heat of the reaction. A number of methods are used to establish this energy including the use of theoretical computer models and tests such as crater, ballistic mortar, and underwater tests. Of these methods, underwater tests give the best correlation to rock- breakage performance. Underwater tests were developed to measure both the shock energy and the gas (bubble) energy released during the detonation of standard test samples. These energy values have been useful in predicting the rock-breaking capabilities of explosives for comparative purposes. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 37 Other terminology widely used by manufacturers is based on the theoretical heat of reaction determined by explosive formulation. Absolute bulk strength (ABS) in calories per cubic centimeter and absolute weight strength (AWS) in calories per gram are computed from the heat liberated during the detonation and formation of gaseous end products. Note ABS and AWS can be computed from one another if density is known, and it is the volumetric basis of reaction heat which correlates with energy. Most manufacturers of explosives will include either value with technical product literature. A mixture of ammonium nitrate and fuel oil (ANFO) is by far the most widely used commercial blasting product. Depending on the proportions of the mix, the heat of reaction is approximately 850 cal/g. As a dry, free-running blasting agent, ANFO is capable of being loaded or packaged at varying densities. For a typical density of 0.85 and an AWS of 850 cal/g, the ABS = (850 cal/g) (0.85) = 723 cal/cm3. Other common strength terms are the relative weight strength (RWS) and relative bulk strength (RBS) in which the relative measure of energy available per unit weight or volume of an explosive is compared to an equal weight or volume of the standard commercial explosive ANFO. The RWS and RBS are computed as a percentage of that available from ANFO. Example 1 Determine the relative strengths of explosives A (ABS 645 cal/cm 3 and density of 0.8) and B (ABS 980 cal/cm 3 and density 1.25) Relative Bulk Strength % 2 . 89 723 645 3 3 = = = cm cal cm cal ABS ABS RBS ANFO A A % 5 . 135 723 980 3 3 = = = cm cal cm cal ABS ABS RBS ANFO B B Relative Weight Strength % 9 . 94 850 ) 8 . 0 ( 645 3 3 = = = cm cal cm cal AWS AWS RWS ANFO A A % 2 . 92 850 ) 25 . 1 ( 980 3 3 = = = cm cal cm cal AWS AWS RWS ANFO B B Therefore on a volume basis, explosive A is less powerful than ANFO while explosive B is more powerful. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 38 2.3.3 Thermochemistry of Explosives and the Detonation Reaction. In order to achieve maximum rock-breaking efficiency from an explosive, an oxygen-balanced mixture is formulated to ensure the formation of nonnoxious gases upon detonation. With an oxygen-balanced formula, it is assumed that optimum values of thermodynamic properties such as energy, temperature, and pressure are achieved, and that these values are not affected by changes in the reactants (explosive mixture). This is the case for ideal reactions. Unfortunately, commercial explosives are non-ideal materials. Changes in the physical nature rather than the chemistry of the explosive mixture, such as particle size and borehole diameter, vary the rate of detonation and hence affect thermodynamic variables. Furthermore, many granular explosives, such as ANFO and nitroglycerin, do not completely react, whereas water-based blasting agents react more efficiently, releasing optimum energy as predicted by formulation. It is possible, however, to estimate the thermodynamic properties of an explosive reaction, assuming an oxygen balance. The following calculations are made to illustrate the methods used to estimate explosive properties of interest. Although there are a number of methods available, the procedures selected herein are the least difficult to apply. Thermodynamic data used for these calculations are found in Table 2-5. Table 2-5: Thermodynamic data for some explosive components and gases. 2.3.3.1 Oxygen Balance As previously explained, an oxygen-balanced explosive formula is one in which the amount of oxygen O 2 is sufficient to form the desired detonation gases CO 2 , HO, and N 2 but not noxious fumes such as CO, NO, and CH 4 (if the mix is oxygen deficient), or free oxygen (if the mix has an excess of oxygen). Oxygen balance for an explosive component or a mixture of components is usually reported in percentages or as moles of monatomic oxygen. It is computed using the mass balance relationship for a reaction of 1 kg of explosive, assuming the formation of ideal gaseous products. Example 2 Determine the oxygen balance for nitroglycerin, C 3 H 5 0 9 N 3 . MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 39 The mass-balance equation is: C 13.21 H 22.02 O 39.62 N 13.21 13.21 CO 2 + 22.02/2 H 2 O + 13.21/2 N 2 + (39.62/2 – 13.21 – 11.01/2) O 2 from the above equation the O 2 is (19.81 – 13.21 – 5.505) = 1.095, therefore the reaction has an unbalance total of 2.19 moles of oxygen. Hence, the oxygen balance (OB) can be calculated by: OB = (O excess ) / (O available) = 2.19 / 39.62 = 0.0553 = 5.53% Note that the OB for an oxygen deficient component is: OB = (O deficient) / (O deficient – O required) Note that some authors (Persson) consider that oxygen balance is given as the mass of oxygen which needs to be removed or added to the composition to achieve oxygen balance. In this case, it is expressed as a fraction or percentage of the explosive formula mass, for example grams O 2 /100 grams of explosives. A simpler method of calculating oxygen balance for explosives that contain only carbon, oxygen, nitrogen, and oxygen, is given by: weight molecular explosive 100 2 4 y x 32 balance oxygen | . | \ | ÷ + ÷ = w Where C x H y N z O w Therefore, consider the previous example where nitroglycerin = C 3 H 5 0 9 N 3 . . N 0 H C of gO/g 0.0352 or % 5228 . 3 227.09 100 2 9 4 5 3 32 balance oxygen 3 9 5 3 + = | . | \ | ÷ + ÷ = 2.3.3.2 Explosive Energy Other parameters of explosive thermodynamics are best described in the readings assigned with this lecture. ADDITIONAL RESOURCES VII The material in the above submodule is further supplemented by: - Readings: Hustrulid, William. Blasting Principles for Open Pit Mining. Vol. 2 – Theoretical foundations. Rotterdam: A.A. Balkema. 1999. (mandatory) - Lecture Module 3.3: Introduction to Explosives. - Readings: “Chapter 9: Thermochemistry of explosives and the detonation process.” Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 40 translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (Optional) 2.4 Explosive Products Note that section 2.4.1 (below) is optional. Those interested in the history of explosives are welcome to become more informed. The remainder of this sub-section is mandatory readings. 2.4.1 History of Explosives The history of explosives is interesting from its profound influence on world development. The key developments are summarized and described below: Black Powder - Saltpeter or “Nitre” - Probably originated with the Chinese around the 10th Century (mostly for fireworks, rockets, etc.) - Roger Bacon published a formula for Black Powder around 1242 o In the 13th century Roger Bacon, a European, was interested in the new knowledge from far east. o He studied fireworks it and tested it over and over again. After many months he found the perfect ratio of saltpeter, sulfur, and a new ingredient, charcoal. o After he found out the perfect ratio he wrote the ingredients and the amounts in code in his diary. o Roger Bacon had made, and recorded, the first black powder (the early form of gunpowder). o Bacon did not get credit for the making of black powder because he didn't use his invention. Berthold Schwarts saw this and exploited it. - Berthold Schwartz invented the gun around 1300 which resulted in further refinement of black powder. o Schwarts used the black powder to launch a pebble at high speed out of a metal tube. o Gunpowder also sped up the very slow process of digging up stones. o With gunpowder they could blow the stones out of the ground. o Now even some poor people cold have a house of stone. This was a great technological step. - Blasting with powder replaced “fire setting” for loosening rock around the beginning of the 17th Century. Mercury Fulminate - Discovered by Howard in 1800 - Later used as detonator for dynamite by Alfred Nobel. Nitroglycerin MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 41 - Nitroglycerin (NG) and Nitrocellulose (NC) discovered by Ascanio Sobrero in 1846 in Switzerland - He was afraid of it, destroyed his notes and warned against its use Dynamite - Alfred Nobel and his father built a NG small factory in 1861 in Sweden - Loading and transporting nitroglycerin was dangerous (liquid poured in to holes and ignited with various types of black powder igniters. - NG proved to be very dangerous and resulted in the death of many people including his brother Emil. - Nobel discovers dynamites by accident (Dynamite is derived from the word Dynamis, meaning power): o When ‘blasting oil’, NG spilled into kieselguhr (NG was packed in it), Nobels saw that the kieselguhr abosrobed about 3 times its weight of NG. o Nobel began to sell the 75/25 NG/kies., the first of the dynamites. Eventually went to wood pulps which increased the energy output of the NG. o This development allowed the relatively safe transport transportation of NG. o Patented in 1867 - Nobel knew the amount of destruction his invention would cause and he did not want to be associated with thousands of deaths, so he left a large amount of money to the awarding of prizes in science, literature, politics, etc… every year. - Pros: more powerful than black powder, higher detonation velocity and more effective in breaking rock (not only moved, but broke them apart) - Cons: Would freeze, nitro fumes and nitro headaches Safety Fuse - William Bickford of England devised the safety fuse, originally a textile- wrapped cord with a black powder core, which for the first time enabled safe, accurately timed detonations. - In 1865 Nobel invented the blasting cap, providing the first safe and dependable means for detonating nitroglycerin and thereby considerably expanding its use for industrial purposes. - Electrical firing, first used successfully in the late 19th century, allows greater control over timing. Ammonium Nitrate - First synthesized in 1659 by J. R. Glauber by combining ammonium carbonate and nitric acid. - Two major uses: Fertilizer and Blasting - Initially used to replace a portion of nitroglycerin in dynamite - Dupont introduced NITRAMON in 1935 - Dupont called his products “Blasting Agents” due to their safe handling, low cost and non-headaches MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 42 - Two disastrous shipload explosions resulted in development of ANFO (Ammonium Nitrate & Fuel Oil) Texas City, Brest France - In 1955, Ammonium nitrate/fuel oil (ANFO) was discovered to greatly increase the energy output of AN prills. - Watergel slurries invented simultaneously by Dupont USA and CIL Canada. - Further sensitized by powered aluminum - ANFO eventually replaced dynamite in dry holes - Pros: easier to load, cheaper and safer than dynamite - Cons: desensitized by water - What can be concluded about the history of explosives are the key aspects in the design of commercial explosives: o Explosive power o Safety in transportation and handling o Cost o Controllable and predictable detonation. 2.4.2 Energetic material Energetic materials are all materials that can undergo exothermal chemical reaction releasing a considerable amount of thermal energy. An explosion is basically any rapid expansion of matter into a volume much larger than the original. An explosive is a material that can undergo an exothermal chemical reaction resulting in a rapid expansion of the reaction products into a volume larger than the original. Figure 2-34 shows a simplified version of an explosives family tree. Figure 2-34: Explosives Family Tree MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 43 Figure 2-35: Explosive Classification Scheme 2.4.3 Explosives and Propellants The difference between an explosive and a propellant is often functional rather than fundamental. Explosives are intended to function by detonation following shock initiation by a detonator or a booster charge. Propellants are intended to burn steadily at a rate determined by the design pressure of the rock or gun breech, and they are ignited to burning by a flame that provides a spray of hot burning particles. 2.4.4 Single Molecule and Composite Explosives There are two fundamentally different kinds of explosive materials, namely single explosive substances and composite explosive mixtures. Single explosives are chemical substances that contain in one well-defined molecule all that is needed for an explosion. The molecule decomposes into mainly gaseous reaction products, such as CO 2 , N 2 , and H 2 O. The solid explosive trinitrotoluene, TNT (C 7 H 5 (NO 2 ) 3 ), and nitroglycerine, are examples of single explosive substances. A composite explosive can be mixture of two single explosive substances, a mixture of a fuel and oxidizer, or an intermediate mixture containing one or more single explosive substances together with a/or oxidizer ingredients. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 44 Table 2-6: Single chemical explosive substances. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 45 Table 2-7: More Single Chemical Explosives MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 46 2.4.5 Primary, Secondary, and Tertiary explosives. Single molecule explosives range with respect to the strength of the stimulus required for initiation or a self-supporting chemical decomposition reaction: from primary explosives, such as lead azide (PbN 6 ), which are used as igniting charges in detonators; through the secondary explosives, of which NG, NM, and TNT are examples; to tertiary explosives, of which ammonium nitrate, AN (NH 4 NO 3 ), is an example. The primary explosives are able to transit from surface burning to detonation within very small distances. A 0.2 mm thick grain of lead azide when ignited will transit from burning to detonation within a distance less than the grain thickness. This is because the lead azide molecule is very simple, decomposing in a very simple two-step reaction, and also because the reaction products have a high molecular weight. Reaction products are generated at the surface faster than they can expand away from the surface, which results in a quick build-up of pressure at the burning surface. The secondary explosives, too, can burn to detonation, but only in relatively large quantities. For example, a stick of dynamite can burn as a candle, slowly, if ignited with a flame (it is strongly advised against performing a demonstration of this without due precautions against the chance event of a detonation), whereas a truckload of dynamite may burn to detonation. Under normal conditions, tertiary explosives are extremely difficult to explode and are in fact officially classed as non-explosives provided that certain conditions are observed (such as that an oxidizer not be mixed with fuels or sensitizers, and that the grain size exceeds a certain minimum size.) They are nonetheless explosive, as demonstrated by some of the largest accidental explosions in history, such as the (April 16, 1947) Texas City explosion of ammonium nitrate. A more detailed classification scheme for explosive material can be seen in Figure 2-35 which shows primary, secondary, and tertiary explosives and the differences between single and composite explosives. 2.4.6 Commercial Explosives Figure 2-34 shows the breakdown of commercial explosive categories. This classification scheme shows how military explosives are often segregated from military applications. Examples of military explosives include TNT, PETN, RDX (cyclonite), Tetryl, and compositions such as A3, B, B4, C4, which are mixtures of RDX, TNT, and additives to provide a moldable consistency. These explosives are chiefly used in the weapons industry. However, small amounts of HE are added to commercial explosives to increase strength and sensitiveness. Many commercial or industrial explosives are classified as HE because they contain critical amounts of military explosives or nitroglycerin, and usually they are cap sensitive. Others, such as dry blasting agents, are not classified as HE, and require boosters or primers of HE for initiation. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 47 Industrial explosives are classified as one of the following: - nitroglycerin-based, - dry blasting agents, - water gels, - emulsions, - permissibles, (for underground coal and will not be discussed) - primers, and boosters. Two-component explosives, a common category, actually contain mixtures or characteristics that fall in other classifications. Often the difference among these products is formulation; however, product packaging and consistency can also change a classification Explosive components are referred to as: - Oxidizers: Oxidizers contribute oxygen for oxygen balance, and include nitrated salts such as ammonium nitrate (AN), sodium nitrate (SN), and calcium nitrate (CN). - Fuels: include fuel oil, carbon, granular aluminum, TNT, black powder, or any carbonaceous material that produces heat. Many of these components are also referred to as sensitizers and can also act as absorbents. - absorbents, are products, such as wood pulp, sawdust, cotton, and cellulose, that incorporate liquid explosive components such as nitroglycerin. - Stabilizers: include flame retardants, gelatins, densifiers, water, gum, emulsifying agents, and thickeners. 2.4.7 Nitroglycerin-based Explosives Dynamite is a trade name introduced by Alfred Nobel. It is comprised primarily of a stable yet powerful mix of nitroglycerin (nitrostarch). Since its invention, a number of nitroglycerin (NG)-based products have been developed of three basic types: granular, gelatin, and semigelatin, which are all considered HE. Gelatins and semi gelatins contain nitrocotton that combines with NG to form a gel structure whose consistency is controlled by the percentage of nitrocellulose. Dynamites are packaged in cylindrical cartridges from 'Is in. (22 mm) in diameter and 8 to 24 in. (203 to 610 mm) in length. The quality of the waxed paper wrapping is important for: - water resistance, - fume production, - ease and safety of loading. Straight dynamite derives its energy source from NG, SN, and AN, including absorbants such as wood pulp and flour that also act as combustibles. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 48 - Ammonia dynamite (or "extra" dynamite) is a granular mix that contains a smaller quantity of NG mixed with AN and SN. - Gelatin dynamites are either straight gelatin or ammonium (extra) gelatin. Each has similar mixtures as straight and ammonia (extra) dynamites with the addition of nitrocellulose for a gel consistency. - Semigelatins are ammonia gelatin with a small amount of nitrocellulose and a 65% weight strength. They are also used as primers and boosters. 2.4.8 Dry Blasting Agents Dry blasting agents are one form of a general category of blasting agents. A blasting agent is, by definition, a mixture of fuel and oxidizer. It is not classified as an explosive, and cannot be detonated with a No.8 blasting cap. A dry agent is a granular, free-running mix of a solid oxidizer (usually AN), prilled into porous pellets onto which a liquid fuel oil or propellant is absorbed. ANFO is the most widely used blasting product, with approximately 94.5% industrial-grade ammonia nitrate and 5.5% fuel oil. Figure 2-36 shows the varying effect of the addition of fuel oil. No.2 grade diesel fuel oil is used for a nearly oxygen-balanced mix. Figure 2-36: Energy output vs. percent fuel oil added to ammonium nitrate Typical values of specific gravity range from 0.75 to 0.95. The properties of dry blasting agents vary significantly with borehole diameter, density, confinement, particle size, water conditions, and size of primer used for initiation. Figure 2-37 shows the varying effect of diameter confined on various explosives. The steady state detonation velocity of ANFO is over 15,000 fpS (4500 m/s) and is achieved in borehole diameters greater than 15 in. (381 mm). The critical diameter of ANFO is between 2 and 4 in. (51 and 102 mm) and is a topic of controversy among blasters and, in particular, those who blast underground using small-diameter holes. The exact values of MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 49 the critical diameter depends on the loading conditions; however, ANFO does not detonate reliably within the range cited above. Figure 2-37: VOD vs. borehole diameter for selected industrial explosives Aluminum in granular form can be added to ANFO to increase the heat or energy output. For increasing percentages of aluminum by weight up to 6%, a measurable increase in fragmentation energy is noted. The cost of additional aluminum beyond 6% does not result in proportionally increased work output and, therefore, is not cost effective. 2.4.9 Wet Blasting Agents Blasting agents that contain more than 5% water by weight are referred to as wet blasting agents. Within this category are: - water gels or slurries, - emulsions, and - heavy ANFO. Heavy ANFO is a combination of prilled ANFO and emulsion. The development of wet blasting agents, led by slurries in the 1950s, came about in response to the disadvantages of ANFO in certain applications. These were: - lack of water resistance - low bulk strength due to low density. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 50 Table 2-8: Typical Compositions of selected slurries and Emulsions Table 2-8 shows some typical compositions of water gels and emulsions. The critical diameter of wet blasting agents is often less than 1 in. (25 mm). Three varieties of wet blasting agents are in common use in the mining industry: slurries, emulsions, and Heavy ANFO. 2.4.9.1 Slurries Slurries, or water gels, are a colloidal suspension of solid AN particles suspended in a liquid AN solution that is gelled, using cross-linking agents. The gels (guar gum) effectively surround the solid AN, rendering the oxidizer water resistant while thickening the explosive mix. Fuels and sensitizers such as TNT, nitrostarch, Composition B, ethyl alcohol, fuel oil, and glass bubbles (microspheres) are dissolved or added to the liquid phase. Granular aluminum, added as a sensitizer, increases weight and bulk strength. Up to 18% aluminum by weight has been found to provide increased energy output. In general, 20% water is used. Certain mixtures, containing high- explosive sensitizers, are cap sensitive and hence should not be classified as a blasting agent, but rather as a slurry explosive. Slurries are characterized by: - excellent water resistance, - high density and bulk strength, - good oxygen balance, confinement, and coupling within the borehole. 2.4.9.2 Emulsions Definition of emulsion: A suspension of small globules of one liquid in a second liquid with which the first will not mix. Emulsion explosives and MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 51 blasting agents are the latest development away from ingredients that are in themselves explosive substances. Emulsions are a two-liquid phase containing microscopic droplets of aqueous nitrates of salts (chiefly AN) dispersed in fuel oil, wax, or paraffin using an emulsifying agent. The water- in-oil structure depends on entrapped air or microspheres for sensitivity, thereby eliminating the need for expensive explosive compounds. Microspheres, microscopic glass, or plastic air-filled bubbles and the AN droplets form the oxidizer, while the fuel oil exists as the oil phase. . In the resulting margarine- or Vaseline-like, smooth mixture, the AN solution stays as a super cooled liquid without crystallizing even upon cooling to sub-zero temperatures. By distributing in it finely dispersed voids in the form of glass microballoons or gas bubbles, that can act as hot spots to initiate the chemical reaction upon shock compression, a variety of emulsion explosives or blasting agents of different sensitivity levels can be produced. As they contain no ingredient that is an explosive in itself, and also because of the desensitizing effect of the water content, all such emulsions have a high degree of inherent safety. In contrast to ANFO which cannot be used in water-filled drillholes because of the high water solubility of AN, emulsion explosives have excellent water resistance since each AN/water droplet is surrounded by a thin film of oil which repels water. The extremely small droplet size, and the sub micron thickness of the oil film gives a very large contact area between the fuel and the oxidizer solution; the intimacy of mixing of the fuel and oxidizer approaches that of a solution. Over time, the oil phase migrates, and droplets combine to form larger particles sizes whose bulk surface area is reduced. Less fuel is in contact with the oxidizer, and less than ideal explosive properties are achieved. Emulsions provide high detonation pressures of 1.45 to 1.74 x 10 6 psi (10 to 12 GPa). Densities range from 1.15 to 1.45. Emulsions have excellent water- resistant properties regardless of packaging. The cost of emulsion products is within the range for slurries. Emulsions can be mixed on site and pumped from bulk trucks. Premixed emulsions are available in plastic tubes in a variety of diameters and lengths. Depending on product diameter and sensitizers used, emulsions can be cap sensitive 2.4.9.3 Heavy ANFO Heavy ANFO is a product comprised of up to 45 to 50% ammonium nitrate emulsion mixed with prilled ANFO. It was developed in an attempt to increase the bulk density of ANFO. The only fuel component is in the ANFO (or a liquid fuel), while the emulsion contains no solid fuel, making the mixture a "repumpable" consistency. The final product has improved strength and provides good water resistance in comparison to ANFO, with a price MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 52 range between that of ANFO and emulsions. Research has shown, however, that the ability of emulsion to prevent ANFO from being dissolved in the presence of water, thereby reducing blasting efficiency, is questionable. 2.4.10 Primers and Boosters- A primer charge is an explosive ignited by an initiator, which, in turn, initiates a noncap-sensitive explosive or blasting agent. A primer contains cap-sensitive high explosive ingredients. Often cartridges of dynamites, highly sensitized slurries, or emulsions are used with blasting caps or detonating cord. Other primers are cast into specific shapes and weights, using TNT and PETN, designed with wells for initiator acceptance. Boosters are highly sensitized explosives or blasting agents, used either in bulk form or in packages of weights greater than those used for primers. Boosters are placed within the explosive column where additional breaking energy is required. Often-times, cartridge or plastic-bagged dynamites or sensitized wet blasting agents are used as primers as well as boosters. Boosters are often used near the bottom of the blasthole at the toe level as an additional charge for excessive toe burden distances. They are also placed within the explosive column adjacent to geological zones that are difficult to break or intermittently within the main explosive charge to ensure continuous detonation. 2.4.11 Initiators and Initiation Systems Initiators are devices containing high explosives that, upon receiving an appropriate mechanical, or electrical impulse, produce a detonation or burning action. Initiators are used as components within a system of explosives and other devices to start the detonation of all other components. Initiation systems are either electric or nonelectric, and include blasting caps, safety fuse, detonating cord, or non-electric shock tubes. 2.4.12 Electric Caps Electric blasting caps are a commonly used method of initiation. Electrical energy (ac or dc) is sent through copper or iron legwires to heat an internal- connecting bridgewire. This heat, in turn, starts a chain reaction of explosives burning within the metal cap shell, through a powder delay train. This process detonates a high-explosive base charge, igniting a cap-sensitive explosive. They are manufactured with an instantaneous (no delay train) time of initiation, or time delays (in milliseconds) used in delayed blasting practices. Time delays with intervals of 25, 50, 100, 500, and 1000 ms are available for short- (ms) or long-period (LP) delays. Table 2-9 summarizes typical delay time intervals available for electric and non-electric initiating systems. Short delays are used in surface blasting operations, while longer delays are used underground where blasting conditions are more confined. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 53 The use of time delays in blasting enhances fragmentation and the control of ground vibrations. In recent years, improvements have been made in the manufacturing of blasting caps that increase the accuracy in detonation time. The next generation of high-precision detonators will contain an electronic circuit instead of pyrotechnical delay elements. The integrated circuits will permit microsecond rather than millisecond timing accuracy and allow programmability for onsite selection of each cap detonation timing. Ac power lines and capacitor-discharge dc power sources approved for blasting are used to energize caps. Precise calculations are needed to determine the entire blasting circuit resistance, including all accessory connecting wires. This is to ensure that the power source supplies the correct current to each cap in the circuit. Safe blasting practices dictate that precautions are used to avoid blasting in the vicinity of extraneous electricity such as stray current, static electricity, electrical storms, and radio frequency energy when using electric caps. Table 2-9: Typical Delay times for donators 2.4.13 Non-electric caps Non-electric initiation systems include a cap similar to that of an electric cap, but they are connected to plastic tubing or a transmission line that carries an initiation (shock and heat) to initiate the cap. The energy source in the tubing is either a gas mixture or an internal coating of special explosive. Non-electric tubing is not used in underground coal or gassy mines as it carries an open flame. The plastic tube itself does not detonate; therefore, the only noise source is the cap itself. Caps and tubes of varying lengths are connected with special connectors between holes to configure unique blast pattern arrays. Surface delay elements, when used in conjunction with in- MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 54 hole delays, provide nearly infinite numbers of delays in blasting patterns. Delays are available in short and long periods as well as in-hole and surface delays. The advantage of non-electric systems over electric systems is the ability to design blasts with a greater number of holes than traditional electric blasting. In addition, concerns about the effects of accidental detonations of electric caps due to stray currents are eliminated with the use of non-electric systems. Figure 2-38: Generic Design detonators 2.4.14 Electronic Detonators Electronic detonators have an electronic counter on a microchip in place of the pyrotechnic delay charge, and a capacitor to supply the discharge energy for ignition. Advantages compared to regular electric or nonel detonators: - Higher timing precision (10 µs instead of 1-10 ms delay scatter) - Same high timing precision at long delay times (10 µs at 5 second delays) - Increased control over time delay - Greater safety against accidental ignition (coded firing signal) Current disadvantages include: - Higher price because of chip and capacitor cost - Back to electric wiring – risk of ground faults or poor contacts 2.4.15 Detonating cord Detonating cord is a flexible but strong continuous detonator that can be several hundred meters long. A detonator is required to initiate a length of MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 55 detonating cord which cannot be normally initiated by fire. Detonating cord has two functions: - to provide simultaneous detonation of several interconnected blasthole charges, thus avoiding the need for multiple electric or plain detonators - to provide continuous initiation of the full length of an explosive column in a blasthole, as distinct from point initiation with individual detonators. Detonating cord (sometimes known as Cordtex) consists of a core of PETN enclosed in a tape wrapping that is further bound by counter-laced textile yarns. The cord is either reinforced or completely enclosed by strong waterproof plastic. Detonating cords are available with a variety of charge weights, tensile strengths and protective coatings, depending on the application. Their energy release depends on the amount of PETN in the core, which generally varies from 1.5 g/m to 70 g/m. However, 10 g/m is the PETN weight of standard detonating cord whose VOD is about 7000 m/s. ADDITIONAL RESOURCES VIII The material for this sub-module is additionally supplemented by: - Lecture module 3.4: Explosive products. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 56 Module 3: Blast Design 7 The section will cover the basic aspects of blast design in terms of geometry, explosives, geology, sequencing, and initiation. Note that these are the key theory, tools and techniques used for most blasting applications. More unique applications of blasting will be covered in the next module, along with underground blast design. Blast Design Basics Blast design is a semi-empirical systematic method that involves balancing numeric and qualitative assessments of rock properties, explosives, and desired products. System Approach The first step in the process of design is to determine the goal or purpose of the intended design. From a historical perspective, it has been seen that 7 These notes were assembled directly from the following references: - From Dr. Paul Lever’s 415 course notes Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2 nd ed. 1992 - Hartman, Howard L. and Jan M. Mutmansky, Introductory Mining Engineering, 2 nd ed. New York: John Wiley & Sons. 2002, 570p. - Atlas Copco, The Raise Boring Handbook 2 nd ed. - Stefanko, Robert. Coal Mining Technology Theory and Practice. Littleton CO.:Society of Mining Engineers. 1983 - McKercher, R.M., Ed. Potash Technology: Mining, Processing, Maintenance, Transportation, Occupational Health and Safety, Environment. Pergamon Press: Toronto. 1983 - Mining-Technology.com, search: continuous mining - Bell, F. G. ed. Engineering in Rock Masses. Butterworth-Heinemann: London. 1992. ISBN: 0 7506 1063 8 - Caterpillar Performance Handbook, 28th Ed. - 2001 Class notes, from Bob Cummings - 1997 Class notes, from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department.Wirtgen America Inc. Surface Mining Manual. Edition 2002 - Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2nd ed. 1992 - Kennedy, B.A. Editor. Suface Mining 2nd Edition. SME: Port City Press, ML. 1990. - Persson, Per-Anders, Holmberg, Roger, and Lee, Jaimin. Rock Blasting and Explosvies Engineering. CRC Press: New York. 2000. 534 p. - Cook, Melvin A. The Science of Industrial Explosives. IRECO Chemicals: Salt Lake City, Utah. 1974. 449p. - Sen, Gour C. Blasting Technology for Mining and Civil Engineers. Sydney: University of New South Wales Press. 1995. 146 p. - The History of Explosives. http://sis.bris.ac.uk/~dj9006/explosives/history.html - Hustrulid, William. Blasting Principles for Open Pit Mining. Vol. 2 – Theoretical foundations. Rotterdam: A.A. Balkema. 1999. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 57 drilling and blasting has been a process by which the mine has undertaken full responsibility. Its key goal was to deliver to the mill the least expensive product. This often limited the blast design to enable the efficient loading and hauling of the material from the stope to the crushing circuit. However, the more recent ‘systems design’ approach whereby the scope changes from simply the mine to the overall mine-mill system, an optimum design for simply the mine will most likely change when the mill is included. Figure 3-1: Systems Approach In visualizing the system, the variables upon which an engineer is to optimize based on lowest overall cost can be seen in Figure 3-2. For example, as can be seen, the internal and external environments specify a minimization of wall damage, flyrock, noise, and other detrimental aspects of working with explosives. These aspects of the design are facilitated in that they are often specified by law. The internal environment is also less limiting as rock damage to walls that are not permanent are also not to dominate the design. Other constraints not shown are the limitations posed by pit design, limited operating room, bench sizes, equipment, etc… The proper fragmentation to optimize the loading& hauling, crushing, and grinding processes are to be taken into account. All these factors must be balanced with cost. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 58 Figure 3-2: Cascading constraints and goals Figure 3-3 shows the numerous controllable variables that can be manipulated to achieve the optimum design. The uncontrollable variables and output requirements should be taken into account in the design. This focus of design assumes that the optimization is to be taken at the blasting stage, however, it should be noted that changes in downstream processes may reduce some of the cost upstream where an expensive aspect of the design can be removed. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 59 Figure 3-3: Controllable and uncontrollable input variables and output goals. Idealized fragmentation curves Some 30 years ago, MacKenzie presented his now classic conceptual curves showing the cost dependence of the different mining unit operations on the degree of fragmentation. They are presented in their original form in Error! Reference source not found. to Figure 3-8. As can be seen some of the costs decrease with increasing fragmentation while others increase. By adding the curves together one obtains the overall cost versus degree-of- fragmentation curve presented in Figure 3-8. It has the form of a saddle indicating that there is a certain degree of fragmentation for which the overall cost is a minimum. In the particular case shown, the base of the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 60 saddle is quite broad suggesting that the overall costs change little over a wide fragmentation range. Before discussing the development and application of these curves it is important to understand the logic behind them. Beginning with the loading, hauling and crushing curves the logic, as presented by MacKenzie, is as follows: Figure 3-4: Loading cost curve Loading An increase in the degree of fragmentation will give the shovel a higher rate of productivity. At standard operating costs per hour (for all practical purposes independent of the production rate) this will result in lower costs per ton or cubic yard moved. The effect of wear and tear will also decrease, giving lower operating cost per hour. Figure 3-5: Hauling Cost Curve Hauling Under similar conditions of haul, lift, size and type of truck and haul road conditions, truck production per hour will increase with greater degree of fragmentation due to faster shovel loading rates and a decrease in bridging (and hence waiting time) at the crusher. There will be a consequent decrease in cycle time. At a standard operating cost per hour, this increase in truck speed or productivity will result in lower unit operating costs. Figure 3-6: Crushing Cost Curve Crushing An increase in the degree of fragmentation gives lower crushing costs as more material passes through as undersize. Liner costs, repair and maintenance, and bridging time will decrease and the crushing rate per hour will increase. As indicated decreased bridging time also cuts down on truck delay time at the crusher which in turn gives higher truck and shovel MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 61 productivity. Any increase in degree of fragmentation means less work for the crusher. The % bridging time is one indicator, along with shovel loading rate of this degree of fragmentation. These have been the easiest to explain since the unit costs always decrease with increasing fragmentation. The same is not true for the drilling and blasting costs. There are many possible combinations which can occur depending upon the particular design. Figure 3-7: Drilling and Blasting Cost Curves Drilling & Blasting For a given rock type, geologic structure, and firing sequence, an increase in the degree of fragmentation may be achieved by: 1. increasing the consumed quantity of a given explosive 2. changing to an explosive having greater energy content per unit hole volume (higher energy content! density) 3. combinations of both. For blasting case (1.) the associated drilling cost would increase if the explosive quantity were to be increased by simply drilling the same diameter drill holes but on a tighter pattern. Thus there would be more drill holes required to blast a given volume. If larger diameter drill holes were substituted and the increased hole volume (explosive quantity) achieved in this way then the rate of increase or decrease would depend upon the comparative drilling cost per foot of hole. For case (2.), presuming that the same hole diameter and pattern is used, the drilling costs would remain constant independent of the fragmentation. For case (3.) the drilling cost could: remain constant, increase, or decrease depending upon the situation. If the same fragmentation is desired and a more energetic explosive is substituted for the one currently in use, then the unit drilling cost could decrease due to the possibility of increasing the hole spacing (spreading the pattern). In his original presentation MacKenzie has explained the drilling dependence as follows: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 62 Generally speaking, for a given type of drilling and of explosive, the cost per cubic yard or ton will remain constant or increase with the degree of fragmentation. If higher energy explosives are substituted, the drilling cost per yard will decrease. The rate of increase or decrease or decrease will be dependent upon the drilling cost per foot. Therefore summing all cost curves together, the overall optimum fragmentation can be assumed to be the lower section of the saddle. Figure 3-8: Overall cost curve. Preliminary guidelines for blast layout The preliminary guidelines for blast layout were taken directly from the Hustrulid (1999) which in turn were taken directly from Ash (1967). These preliminary guidelines provide the reasoning (proofs) for the five key relationships for blast design. The readings prescribed in the additional readings provide additional theory and equations. Figure 3-9 shows the basic nomenclature for the key variables discussed here. Note that the basic guidelines shown here apply to surface mines. However, the same key variables are also applicable to underground mine blasts, as will be shown in future modules. Figure 3-9: Isometric view showing nomenclature The key variables for blast design include: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 63 - drilled burden (B) - is defined as the distance between the individual rows of holes. It is also used to describe the distance from the front row of holes to the free face. When the bench face is not vertical the burden on this front row of holes varies from crest to toe. - spacing (S) - is the distance between holes in any given row. - Subgrade (J) - Generally the holes are drilled below the desired final grade. This distance is referred to as the subgrade drilling or simply the sub-drill - Stemming (T) - A certain length of hole near the collar is left uncharged. This will be referred to as the stemming length (T) whether or not it is left unfilled or filled with drill cuttings/crushed rock. - Bench height (H) – is the vertical height from the toe to the crest. - drilled length (L) - is equal to the bench height plus the sub-drill. - length of the explosive column (L e ) - is equal to the hole length minus the stemming. This column may be divided into sections (decks) containing explosives of various strengths separated by lengths of stemming materials. Sometimes the explosive strength is varied along the hole, i.e. a higher strength bottom charge with a lower strength column charge. As will be seen in the next section, the different dimensions involved in a blast design are not arbitrary but closely related to one another. The selection of one, for example the hole diameter, fixes within rather strict limits, many of the others. Spacing to burden relationship As can be seen in Figure 3-10, the hole spacing (S) and burden (B) can be directly related through the following relationship: B K S S = Where K s is a constant relating spacing to the burden. Figure 3-10: Plan view of bench showing first row. Burden to hole diameter relationship Note Figure 3-10, where it is shown that for a particular burden and spacing, each hole diameter is expected to break a particular volume of rock. The volume of rock to be broken by A UNIT of hole length is: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 64 1 × × = S B V R A particular amount of energy (E V ) is required to break the a unit volume of rock. The total energy for a unit of hole length is therefore V S R S V V R R E B K E B K S E S B E V E 2 since however, = = × × = × = Since K S and E V are constants, the required amount of explosive energy is directly proportional to B 2 . 2 B E R · The amount of explosive energy available is determined by the explosive volume that can be loaded into that unit of length of borehole: 2 4 e e D V t = Where D e is explosive diameter and E e is the explosive bulk strength (strength by unit volume). Therefore the explosive available is determined using: e e A E D E 2 4 t = Since E e is a constant related to explosive type, the D e can be seen as directly proportional to E a 2 e A D E · If using packaged explosives, as is sometimes the case at pit perimeters where perimeter blasting techniques are employed, the charge diameter (D e ) may be less than the diameter (D) of the hole. However, where bulk blasting agents are used, the entire cross-sectional area of the hole is filled with explosive. Therefore hole diameter (D) is equal to explosive diameter (D e ). This would make the explosives available also directly proportional to hole diameter: 2 D E A · MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 65 Furthermore, this would also require that the required explosives and available explosives to be equal since additional explosives cannot be added beyond the hole diameter. R A E E = however, remember that: B D B D B E R · · · e therefor re therefo 2 2 2 therefore the diameter is can be determined from a proportionality constant K B , relating to the hole diameter: D K B B = Therefore as the burden increases, the diameter should also increase proportionally, as seen in Figure 3-11. Figure 3-11: Effect of hole diameter on burden Subdrill to burden relationship The toe region is a highly confined volume. Therefore extra explosive energy must be applied to assure adequate fragmentation. This extra explosive power is generally provided by extending the drill hole below the toe elevation and filling the so-called subdrill length J with explosive. There are several different rationales used for selecting the appropriate length. Here an explanation based upon explosive run-up distance will be presented. The results are essentially the same with all techniques. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 66 Figure 3-12: Toe confinement There is a certain distance (called the run-up distance) characteristic of the initiating system/explosive which the shock wave must travel away from the point of initiation before steady state conditions are reached in the explosive column. To break the confined toe, the borehole pressure should be as high as possible. As seen in previous classes, the explosion (borehole wall) pressure (P e ) is proportional to the square of the detonation velocity: 2 VOD P e · The elevation in the hole at which steady state velocity is reached should not be higher than the bench toe elevation. To be conservative the minimum run-up distance will be assumed to be 6D. Figure 3-13: Run-up distance to achieve steady state VOD. In addition, the primer is seldom placed directly at the bottom of the blasthole due to the presence of cuttings and water. A normal offset is of the order of 2D. Therefore, the distance from the drilled end of the hole to the toe elevation (the subdrill distance J) should be: B K J B J B D D J J = · · ~ ip relationsh ality proportion a in resulting since ore theref 8 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 67 Stemming to burden relationship Near the hole collar, the rise of the explosive should be controlled so that the possibility of breaking upward toward the horizontal free surface should be ‘as difficult’ or more difficult than breaking, as desired, toward the vertical free face. This can be seen conceptually, by the placement of a spherical charge having the same distance from the collar as the burden. Therefore the “as or more difficult” relationship can be summarized as: B V > a) b) Figure 3-14: Section view comparing the spherical charge (a) and cylindrical charge (b) minimum distance from the collar An initial assumption is made in that the degree of equivalence of the charges will depend on the proximity to the charge and that this relationship is linear and expressed as: B K T TC C = From Figure 3-14, the following relationship is evident: 2 1 where to simplified be can h whic 2 1 2 : how g considerin and where ip relationsh difficult" as " the g considerin 2 TC T T TC TC TC C C K K B K T K B B K B T B K T B V T V T ÷ = = | . | \ | ÷ = ÷ = = = ÷ = Determining K TC is difficult. The other relationships are derived in a later section however, this relationship proof is explained here. In bench blasting, Langefors & Kihlstrom (1963) have empirically derived the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 68 spherical/cylindrical charge equivalence is as shown in Figure 3-16. To explain the significance of the curve, consider a bench containing two vertical side-by-side blastholes. The burden is the same for both. Rather than discussing the collar region which is the subject of this portion, this example will involve the toe region. The reason for this is that the explanation is easier and the principle is the same. Consider a spherical charge of quantity Q o placed at the toe elevation in one of the holes. In the second blasthole a cylindrical charge with a linear charge concentration of 1 kg/m of hole is emplaced. The bottom of the charge is at toe elevation and then the column extends upward towards the collar, as seen in Figure 3-15. Figure 3-15: View facing bench comparing equivalent cylindrical and spherical charges. The length of the elongated charge is expressed in multiples of the burden B. For a cylindrical charge of length B, the total charge would be B x 1 (remember that the explosive volume per length emplaced is 1 kg/m). From the Figure 3-16 one can see that at the toe this elongated charge has only the equivalent breaking power of a spherical charge of weight 0.6 x 1 x B. This is understandable since the energy contained in that part of the elongated charge near the collar must travel a much longer distance to reach the toe and in the process the energy is spread over a much larger volume of rock. The energy density by the time it reaches the toe is much less than that produced by energy which has traveled a shorter distance. Figure 3-16: Langefors & Kihlstrom’s Toe breaking equivalence of spherical and cylindrical charges MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 69 For a linear charge of length 0.3B the total charge has a mass of 0.3 x 1 x B. From the curve it is seen that this has the same effect at the toe as a spherical charge placed directly at the toe elevation with a mass 0.3 x 1 x B. For charges shorter than 0.3B this relationship holds as well, i.e. the elongated charge of a given weight has the same effect at the toe as a spherical charge of the same weight. For elongated charges with lengths greater than 0.3B, the effect at the toe diminishes rapidly with increasing length. The same effect could be achieved by considering the elongated charge extending from the toe elevation downward. Thus an elongated charge extending from 0.3B below the toe to 0.3B above the toe elevation (for a total explosive weight of 0.6 x 1 x B) would, according to the curve have the same breaking capacity as a spherical charge with a weight of 0.6 x 1 x B placed directly at the toe elevation. This can be seen in Figure 3-17. Figure 3-17: Equivalent spherical and cylindrical charges In transferring this concept to the collar region one finds that 7 . 0 2 4 . 1 2 6 . 0 1 , 2 1 where and 6 . 0 that noting and 6 . 0 = = ÷ = ÷ = s = s T TC T TC TC C C K K K K B K T B T Bench height to Burden relationship To this point in the discussion there has been no specific mention of the bench height. If one continues to increase the scale (hole diameter) as shown in Figure 3-18, the center of the charge progresses further and further down the hole. The limiting condition is when the center of charge reaches the toe elevation (Figure 3-19). This occurs for a hole diameter which yields a burden just equal to the bench height. The fifth and last of the fundamental relationships is: B K H H = where K H is a constant relating bench height to the burden. The value of KH is therefore: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 70 1 > H K As a rule of thumb, to derive a method of limiting the choice of hole diameter, note how: B B H B H B K H D D K K D K K H D K B s > > = = limiting a in resulting H derive can we , 1 since and re therefo Figure 3-18: Burden to diameter relationship. Figure 3-19: Limiting the charge diameter and burden. Ratios for initial design In the previous section, five relationships were derived for preliminary blast design. This section will discuss the general values of the constants which is used as base designs from which to further optimize based on other design constraints. Table 3-1 provides the summary of this section. The remainder provides the reasoning and proof behind these initial values. Ratio K s As will be covered in a later lecture, the optimum burden and spacing ratio depends on the energy coverage of the bench (among other variables). When using a square pattern, the best energy coverage is with K s =1 however, empirically, there is little difference between K s =1 to K s =1.5. For a staggered drilling pattern, the best energy coverage is with K s = 1.15. Note that a staggered pattern provides more uniform energy coverage. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 71 Figure 3-20: Square pattern Figure 3-21: Staggered pattern Ratio K B The most critical and important dimension in blasting is that of the burden. There are two requirements necessary to define it properly. To cover all conditions, the burden should be considered as the distance from a charge measured perpendicular to the nearest free face and in the direction in which displacement will most likely occur. Its actual value will depend on a combination of variables including the rock characteristics, the explosive used, etc. But when the rock is completely fragmented and displaced little or not at all, one can assume the critical value has been approached. Usually, an amount slightly less than the critical value is preferred by most blasters. There are many formulae that provide approximate burden values but most require calculations that are bothersome or complex to the average man in the field. Many also require knowledge of various qualities of the rock and explosives, such as tensile strengths and detonation pressures, etc. As a rule, the necessary information is not readily available, nor is it understood. A convenient guide that can be used for estimating the burden, however, is the K B ratio. Experience shows that when K B = 30, the blaster can usually expect satisfactory results for average field conditions. To provide greater throw, the K B value could be reduced below 30, and subsequently finer sizing is also expected to result. Light density explosives, such as field-mixed ANFO mixtures, necessarily require the use of lower K B ratios (20 to 25), while dense explosives, such as slurries and gelatins, permit the use of K B near 40. The final value selected should be the result of adjustments made to suit not only the rock and explosive types and densities but also the degree of fragmentation and displacement desired. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 72 To estimate the desired K B value one should know that densities for explosives are rarely greater than 1.6 or less than 0.8 g/cm3. Also, for most rocks requiring blasting, the density in g/cm3 rarely exceeds 3.2 nor is less than 2.2 with 2.7 far the most common value. Thus, the blaster can, by first approximating the burden at a K B of 30 make simple estimations toward 20 (or 40) to suit the rock and explosive characteristics, densities for the latter exerting the greater influence. As a rule of thumb consider: - For light explosives in dense rock use K B = 20, - For heavy explosives in light rock use K B = 40, - For light explosives in average rock use K B = 25, - For heavy explosives in average rock use K B = 35. Ratio K J The most common value of K J is 0.3. In certain sedimentary deposits with a parting plane at toe elevation subdrilling may not be required. In very hard toe situations, the subdrilling may be increased over that indicated by using K J = 0.3. However it is probably better to consider using a more energetic explosive. It must be remembered that the subdrill region generally forms the future crest/bench top for the bench below. Unwanted damage done at this stage may have a long and costly life. In addition excessive subdrill results in: - A waste of drilling and blasting expenditures - An increase in ground vibrations - Undesirable shattering of the bench floor. This in turn creates drilling problems, abandoned blastholes and deviations for the bench below. - It accentuates vertical movement in the blast. This increases the chances for cutoffs (misfires) and overbreak. Ratio K T The minimum recommended value for K T for large hole production blasting is K T = 0.7. Some specialists suggest the use of K T = 1.0. Collar and stemming are sometimes used to express the same thing. However, stemming refers to the filling of blast holes in the collar region with materials such as drill cuttings to confine the explosive gases. But stemming and the amount of collar, the latter being the unloaded portion of a blasthole, perform other functions in addition to confining gases. Since an energy wave will travel much faster in solid rock than in the less dense unconsolidated stemming material, stressing will occur much earlier in the solid material than compaction of the stemming material could be accomplished. Thus the amount of collar that is left, whether or not stemming is used, determines the degree of stress balance in the region. The use of stemming material then assists in confining the gases by a delayed MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 73 action that should be long enough in time duration to permit their performing the necessary work before rock movement and stemming ejection can occur. For stress balance in bench-blasting of massive material, the value of T should equal the B dimension. Placing the charge too close to the collar can result in backbreak, flyrock and early release of the explosive gases with resulting poor fragmentation. On the other hand, increasing the length of stemming may reduce the energy concentration in the collar region to the point where large boulders result. Usually a K T value of less than 1 in solid rock will cause some cratering, with back break and possible violence, particularly for collar priming of charges. However, if there are structural discontinuities in the collar region, reflection and refraction of the energy waves reduce the effects in the direction of the charge length. Thus the K T value can be reduced under such circumstances, the amount depending upon the degree of energy reduction at the density or structural interfaces. Field experience shows that a K T value of 0.7 is a reasonable approximation for the control of air blast and stress balance in the collar region. Ratio K H Currently most open pit operations have K H values which are approximately 1.6 or more. Too small a K H value will result in substantial cratering. Summary of Ratios The table below simply summarizes some of the basic ratio values and equations. Table 3-1: Ratio Summary Name of Relationship Equation Value Spacing - Burden B K S S = 1-1.5 Burden – Diameter D K B B = 25 Subdrill – Burden B K J J = =0.3 Stemming – Burden B K T T = 0.7 Bench height - Burden B K H H = 1.6 Powder Factor The following example is provided to illustrate the relationships developed in the last two sub-sections. A key indicator of blast design is the specific charge, also known colloquially as the ‘powder factor’. Design a blast considering these initial variables: - Rock = syenite porphyry (SG=2.6) - Explosive = ANFO (µ=0.8, S ANFO =1) - Bench Height (H)=15m MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 74 - Hole Diameter (D)=381mm (15”) - Staggered drilling pattern, vertical holes - 1 blast = 4 rows of holes each containing 6 holes from the relationship, we see: - D K B B = , where K B 25 (assumed), therefore B=(25)(0.381)=9.5 m - B K S S = , where K S =1.15 (staggered), therefore S=(1.15)(9.5)=11 m - B K J J = , where K J =0.3, therefore J=(0.3)(9.5)=3 m - B K T T = , where K T =0.7, therefore T=(0.7)(9.5)=6.5 m - L=H+J, hole length therefore equals to 15+3=18 m As an assurance, in verifying K H , we find 6 . 1 5 . 9 15 = = = B H K H , considering the value is usually equal to or greater than 1.6, we find these parameters to be acceptable. To calculate the powder factor, the volume and weight of the explosive used is calculated: kg 25,176 4 6 kg 1049 explosive of weight total hole per kg 1049 m / kg 800 m 31 . 1 m 31 . 1 ) 5 . 6 18 ( ) 81 . 3 ( 4 ) ( 4 3 3 3 2 2 = × × = = × = = = ÷ = ÷ = EXP e e e T V W T L D V µ t t The weight of the rock that will be broken is calculated: tons 812 , 97 ) t/m 6 . 2 )( 15 )( 11 )( 5 . 9 ( 24 holes of number 3 = = × × × × = × = µ µ H S B n V T R R Hence, the powder factor using the explosive ANFO is found to be: ton kg 26 . 0 812 , 97 716 , 25 = = = R EXP ANFO T T PF Determination of K B As seen in the above discussion, K B is by far the most important constants in the design of blasts. Selecting the proper burden is therefore a key step in blast design, and the factor K B allows the selection of an appropriate burden and diameter as seen in the equation: D K B B = As mentioned above, the best initial estimate for K B is 25, when using ANFO in rocks with an average density. However, consider the requirements for selecting a new K B when explosives or rock type varies. The approach described below can be used as an approximation until field results are available to guide the designed toward more optimal solutions. Note that the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 75 proof below is valid for the metric system. Consider this blast design has must be designed to other variables, including: SG E = specific gravity of explosive used SG R = specific gravity of the rock PF EXP = powder factor (kg/ton) TF = tonnage factor (m 3 /ton) As seen in a previous example, the total weight that one borehole is expected to break can be calculated using the following: R H S rock R R SG K K B H S B V T 3 = × × × = × = = µ µ Once the total amount of rock material to be removed is known, the amount of energy required is calculated using the powder factor: EXP R H S EXP R qyured PF SG K K B PF T E 3 Re = × = The amount of explosive available is defined by the size of the borehole, which can be given as: E T J H e E T J H e E e Avail SG K K K D B SG BK BK BK D SG T J H D E ) ( 4 ) ( 4 ) ( 4 2 2 2 ÷ + = ÷ + = ÷ + = t t t Setting the amount of explosives to the amount required results in: | | . | \ | ÷ + | | . | \ | = = | | . | \ | ÷ + | | . | \ | = ÷ + = H S T J H EXP R E B B H S T J H EXP R E e E T J H e EXP R H S K K K K K PF SG SG K D K B K K K K K PF SG SG D B SG K K K D B PF SG K K B ) ( 4 used, being is agent bulk a assuming and since and , ) ( 4 ) ( 4 2 3 t t t The powder factor based on the explosive used is replaced with the equivalent ANFO powder factor, denoted by the variable PF ANFO : MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 76 lbs/ton 2000 at lbs/ton in factor powder equivalent ANFO where ) ( 4 2000 to equivalent is units english in equation that this note ) ( 4 : in results which cal/gm), (usually t unit weigh per energy the is Q where , S to equivalen is S explosive, an of ANFO o strength t weight relative where , ANFO ANFO = | | . | \ | ÷ + | | . | \ | | | . | \ | = | | . | \ | ÷ + | | . | \ | | | . | \ | = = = ANFO H S T J H ANFO ANFO R E B H S T J H ANFO ANFO R E B ANFO EXP ANFO ANFO EXP PF K K K K K PF S SG SG K K K K K K PF S SG SG K Q Q S PF PF t t The above formula can be used for several purposes, as will be seen in the following examples. Effects to pattern design in changing explosives One of the major ways the equation can be used is to study the effect of changes in the explosive on the blasting pattern while keeping other factors constant. Note that the bench height, K H , depends on the burden which also depends on K B . Therefore the burden can change. The approach can be seen as follows: Explosive 1: | | . | \ | ÷ + | | . | \ | | | . | \ | = 1 1 1 1 1 ) ( 4 H S T J H ANFO ANFO R E B K K K K K PF S SG SG K t Explosive 2: | | . | \ | ÷ + | | . | \ | | | . | \ | = 2 2 2 2 2 ) ( 4 H S T J H ANFO ANFO R E B K K K K K PF S SG SG K t In most blast design alternatives, a constant powder factor is maintained. A ratio between the two K B values, where the powder factor ratio is eliminated, can be expressed by: | | . | \ | ÷ + ÷ + | | . | \ | | | . | \ | = ) ( ) ( ) 1 ( ) 2 ( 1 2 2 1 1 2 1 2 T J H H S T J H H S ANFO ANFO E E B B K K K K K K K K K K S S SG SG K K Taking this ratio and for now, ignoring the changes in K H with changing burden, the equation is reduced to: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 77 ( ) ( ) 1 2 1 2 ANFO E ANFO E B B S SG S SG K K × × = The above is a first approximation of the K B2 and is equivalent to the square root of the bulk strength ratio for the explosives in question. To refine the K B2 value, an iterative process is used where: 1) Initial value of K B2 is substituted in the formula: e B D K B 2 2 = 2) Then the new K H2 is derived from the equation 2 2 B H K H = 3) This new value of K H2 is then inserted into the equation ( ) ( ) | | . | \ | ÷ + ÷ + | | . | \ | × × = ) ( ) ( 1 2 2 1 1 2 1 2 T J H H S T J H H S ANFO E ANFO E B B K K K K K K K K K K S SG S SG K K 4) If the resulting K B2 is then compared to the original estimate. If they are the same, the process stops. If not, then the new K B2 is used in step 1 and the process continues, until the value of K B2 converges. Effects to pattern design in changing rock types The effect to the pattern design when rock types change is very similar to that discussed in the previous subsection. Here, the initial approximation can be made by the ratio: 2 1 1 2 R R B B SG SG K K = Once again, the iteration process involving three equations is undertaken: 1) Initial value of K B2 is substituted in the formula: e B D K B 2 2 = 2) Then the new K H2 is derived from the equation 2 2 B H K H = 3) This new value of K H2 is then inserted into the equation | | . | \ | ÷ + ÷ + | | . | \ | = ) ( ) ( 1 2 2 1 2 1 1 2 T J H H S T J H H S R R B B K K K K K K K K K K SG SG K K MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 78 4) If the resulting K B2 is then compared to the estimate at the beginning of the iteration. If they are the same, the process stops. If not, then the new K B2 is used in step 1 and the process continues, until the values of K B2 converge. Numerical Examples Consider the mine where the current design has the following parameters: - Hole diameter = 12 ¼ inches - Bench height = 40 ft - Burden = 25 ft - Spacing = 29 ft - Subdrill = 7ft - Stemming = 17ft - ANFO: S ANFO = 1 - SG ANFO = 0.82 - Q = 912 cal/gm - Rock: SG = 2.65 - PF ANFO = 0.5 lbs/ton | | . | \ | ÷ + | | . | \ | | | . | \ | = H S T J H ANFO ANFO R E B K K K K K PF S SG SG K ) ( 4 2000 t using imperial units Example 1: Changing diameter of holes What would the pattern be if changing to 15 inch diameter holes? The first step in this problem is to derive K B . First, the other ratios need to be derived: K H = 40/25 = 1.6 K J = 7/25 = 0.3 K T = 17/25 = 0.7 K S = 29/25 = 1.15 These values are plugged into the following equation: 2 . 25 ) 15 . 1 )( 6 . 1 ( ) 7 . 0 3 . 0 6 . 1 ( 5 . 0 1 65 . 2 82 . 0 4 2000 ) ( 4 2000 = | | . | \ | ÷ + | . | \ | | . | \ | = | | . | \ | ÷ + | | . | \ | | | . | \ | = t t H S T J H ANFO ANFO R E B K K K K K PF S SG SG K This is about what is expected from the recommended ranges discussed previously. For the 15 inch diameter holes, the first approximation for the burden would be: ft 5 . 31 foot inches 12 inches 15 2 . 25 = | | | . | \ | = = E B D K B MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 79 However, this new value for the burden would result in a change in K H . Therefore the new K H is found to be: 27 . 1 5 . 31 40 = = = B H K H This new value is then input back into the original K B equation: 1 . 24 ) 15 . 1 )( 6 . 1 ( ) 7 . 0 3 . 0 6 . 1 ( 5 . 0 1 65 . 2 82 . 0 4 2000 ) ( 4 2000 = | | . | \ | ÷ + | . | \ | | . | \ | = | | . | \ | ÷ + | | . | \ | | | . | \ | = t t H S T J H ANFO ANFO R E B K K K K K PF S SG SG K this process is repeated through several iterations until a stable K B is found, which is: K B = 24.3 This results in a pattern with the following dimensions: B = 30 ft S = 34.5 ft T = 21 ft J = 9ft The resulting powder factor may be slightly different from the original on account of the rounding off, however note that this pattern would result in increased coarseness in fragmentation. To maintain fragmentation, the powder factor would have to be increased. Example 2: Changing Explosives What would be the change in the pattern in changing the explosive from ANFO to heavy ANFO with the following properties (original hole diameter of 12 ¼ inches): SG = 1.10 Q = 815 cal/gm We will need to use the weight strength of this explosive product with respect to ANFO, which is calculated: 89 . 0 912 815 = = ANFO S We use the initial estimate, as derived in a previous section, and using the initial K B value from the previous question: ( ) ( ) 5 . 27 00 . 1 82 . 0 89 . 0 10 . 1 ) 2 . 25 ( 1 2 1 2 = × × = × × = ANFO E ANFO E B B S SG S SG K K MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 80 The new burden would therefore be: ft 1 . 28 foot inches 12 inches 25 . 12 5 . 27 2 2 = | | | . | \ | = = E B D K B The new K H is then calculated to be: 42 . 1 1 . 28 40 2 2 = = = B H K H This value is then substituted into the equation: ( ) ( ) ( ) ( ) ( ) ( ) ( ) ( ) 95 . 26 ) 7 . 0 3 . 0 6 . 1 ( 42 . 1 ) 7 . 0 3 . 0 42 . 1 ( 6 . 1 1 82 . 0 89 . 0 10 . 1 ) 2 . 25 ( ) 7 . 0 3 . 0 ( ) 7 . 0 3 . 0 6 . 1 ( 6 . 1 1 82 . 0 89 . 0 10 . 1 ) 2 . 25 ( ) ( ) ( ) ( ) ( 2 2 2 2 2 2 1 1 1 2 1 2 1 2 2 1 1 2 1 2 = | | . | \ | ÷ + ÷ + | | . | \ | × × = | | . | \ | ÷ + ÷ + | | . | \ | × × = | | . | \ | ÷ + | | . | \ | ÷ + | | . | \ | × × = | | . | \ | ÷ + ÷ + | | . | \ | × × = B H H B H T J H T J H H ANFO E ANFO E B B T J H H S T J H H S ANFO E ANFO E B B K K K K K K K K K K K K S SG S SG K K K K K K K K K K K K S SG S SG K K The new burden is calculated to be: ft 51 . 27 foot inches 12 inches 25 . 12 95 . 26 2 2 = | | | . | \ | = = E B D K B this process is repeated until a stable value of K B2 results, which for this example, the final K B2 = 27.0 The blast pattern would therefore be (in feet): B = 27.6 S = 31.7 J = 8.3 T = 19.3 Stemming & Decking Many blast designs use decked charges formed by dividing the explosive column into two or more individual charges, initiated on the same or different delays, separated by inert stemming material. Decking is employed to: - Reduce explosive use adjacent to weak rock zones, faults or clay seams MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 81 - Reduce charge quantity detonated at one time delay, lowering ground vibrations - Bring the powder column up higher in the hole to assure good breakage near the collar Decked charges should be separated by stemming materials at a length beyond which two adjacent decks do not affect one another. If interdeck stemming is too small, the deck designed to initiate on the earlier time delay may prematurely initiate the second deck. This situation is referred to as sympathetic detonation and may lead to excessively high ground vibrations or flyrock and a loss of fragmentation due to confinement from improper timing. A rule of thumb for the design of interdeck stem length is to employ the hole radius dimension in feet. The following example gives the design procedures for a blast design in which the explosive charge is limited to control ground vibrations. Example of Deck design Determine the change in layout if decking is needed in the layout by limiting the weight of explosive to 275 kg/delay, keeping the same powder factor. Note that the collar stem, bench height, and subgrade drilling remain unchanged. The basic data is: - Rock = syenite porphyry (SG=2.6) - Explosive = ANFO (µ=0.8, S ANFO =1) - Bench Height (H)=15m - Hole Diameter (D)=381mm (15”) - Staggered drilling pattern, vertical holes from the relationship, we see: - D K B B = , where K B 25 (assumed), therefore B=(25)(0.381)=9.5 m - B K S S = , where K S =1.15 (staggered), therefore S=(1.15)(9.5)=11 m - B K J J = , where K J =0.3, therefore J=(0.3)(9.5)=3 m - B K T T = , where K T =0.7, therefore T=(0.7)(9.5)=6.5 m - L=H+J, hole length therefore equals to 15+3=18 m - A following calculations from section 0, above, we see that the powder factor is 0.26 kg/ton and that the amount of explosives per hole is: 1049 kg. Solution: The amount of explosives per unit length is calculated as: length unit kg/per 91 ) 8 . 0 ( ) 381 . 0 ( 4 4 2 exp 2 exp = = = t µ t lo lo D L W Charge decks per hole is found to be: 275 kg / 91 kg = 3 meters per deck. Therefore assuming if three decks are used, the overall explosive column MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 82 length is 9 meters leaving 1.5 meters of stemming length available meaning 0.75 meters of stemming between each deck. 9.7m ) (1.15)(8.4 S new the therefore 4 . 8 ) / ( 11.7 825 re therefo 825 ) / ( 11.7 : determined is following the B, for solving / 26 . 0 ) / 6 . 2 ( 15 ) 15 . 1 ( kg) )(275 3 ( deck) per decks)( ( 2 2 2 3 = = = = = = × × × = × × × = m m kg (kg) B (kg) m kg B ton kg m ton B B H B K B W n PF rock S µ Concluding Notes on D, B, and PF Drillhole diameter, burden, and powder factor are the most important variables in blast design. Although this lecture provided the preliminary blast design layout equations, be aware that blasting is an never-ending process of fine-tuning and modifications. This approach is necessary due to the many factors that cannot be controlled such as geology and explosive loading conditions. Burden should be selected based on geology and explosive energy output. Hole diameter is usually set by the drill rig capacity which is matched to the range of hole depths anticipated for the job. It is desirable to select a size that will provide an adequate powder factor (the ratio of explosive distributing the explosive evenly throughout the hole depth). Fragmentation and particle size distribution are a function of hole diameter and burden. The capacity of the excavation equipment or requirements of downstream processing dictates the required fragmentation. Higher powder factors result in fine fragmentation and are therefore required for small capacity removal equipment such as a front-end loader. Higher powder factors result in coarser fragmentation and are typically used for rock removal using draglines and large shovels. Note that since the powder factors are lower, the drillhole diameters and burden are typically lower for quarries or mines using smaller equipment in weaker rock. Table 3-2: Powder Factor MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 83 ADDITIONAL RESOURCES IX Additional learning resources include: - Readings, Chapter 19: Controllable parameters of blasting and Chapter 20: Bench Blasting, from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (Mandatory for assignment 5) - also considered readings for Module 4.3 – Patterns and sequencing. - Assignment 5 (you will learn much when doing this assignment. - Lecture Module 4.1 – Basics of Blast Design Geological Impacts on Blast Design Prior to developing each blast design, care should be taken to review existing geology and geologic structure. This can be accomplished by simply review existing outcrops in , around, or adjacent to the blast site. A detailed drill log indicating discontinuities at various depths, may also be used. Rock Properties Blasting performance is usually influenced more by rock properties than by the properties of the explosive. Rocks show numerous planes of weakness, natural fissures, and cracks formed as a result of previous blasting. Thus there are planes of preferential fracture orientation in anyone or more of an infinite number of directions relative to the blasthole's axis. The following properties of rock may have a significant influence on blasting results. Dynamic Compressive Strength If the explosive's outgoing strain wave exceeds the dynamic compressive breaking strength of rock an annulus of crushed rock is formed around the charge. This crushed zone is detrimental to the transmission of strain waves in the surrounding rock. Hence it is important that the dynamic compressive strength of the rocks in situ be determined, and an explosive of low density and velocity should be used in rocks with low compressive strength values. Elastic Moduli The elastic moduli give the behavior of rocks under stress and should be determined by sonic techniques (dynamic moduli) rather than by the use of mechanical tests (static moduli). It has been found that the explosion pressure should not exceed 5% of the dynamic Young's modulus in order to obtain optimum results from blasting. The post-detonation gas pressure exerted in the cracks between the blasthole and the free face pushes the burden forward and produces heave. The bulk modulus of a rock has to be known in order to calculate heave. Density The density of rock is closely correlated with its strength. An increase in rock density often results in a decrease in the displacement of a rockmass MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 84 fragmented by blasting. Adequate displacement of higher-density rock can be achieved by following one of three courses: - increasing the blasthole diameter, - reducing the blasthole pattern, or - changing to an explosive which has stronger heave energy. Porosity Porosity tends to reduce the efficiency of blasting operations. The lengths of strain-wave-induced cracks in a highly porous rock are calculated to be only about 25% of those in a non-porous rock of identical mineralogy. This implies that highly porous rocks are fragmented mainly by heave energy. Hence, post-detonated gases have to be kept trapped at high pressure until they have performed their task. This can be achieved by bottom priming and by having adequate stemming to prevent premature venting of gases. Internal Friction Internal friction is a relative measure of a rock's ability to attenuate strain waves by the conversion of some of the mechanical energy into heat. It increases with a high degree of porosity, permeability, and jointing of the rock mass. Generally, internal friction values for igneous or metamorphic rocks are lower than for sedimentary rocks, which require high energy explosives for satisfactory blasting. However, if the rock pores are filled with water, the internal friction factor reduces considerably, giving easier passage of the strain wave and improved fragmentation. Water Content Water saturation considerably increases the velocity of propagation of strain waves, owing to the filling of pores with water, which is a good medium for elastic wave transmission. However, fluids in a porous rock reduce both the compressive and tensile strengths, owing to the lower friction characteristic between grain surfaces. If water is present in discontinuities adjoining a block of rock which is being blasted, strain waves may have a greater ability to weaken that rock mass by means of water being jetted considerable distances through interconnected fissures. This has a wedging action which will have a considerable influence on overbreak; and hence slope instability. It is therefore advisable in open-pit mining to dewater a rock mass where a 'permanent' slope, which is intended to remain for several years with minimum maintenance, has to be formed. In Situ Static Stress High in situ static stresses often exist well within the rock body and blasting results can be affected by these stresses. A typical example of such a phenomenon is where some of the radial cracks from the blasthole tend to curve off into the direction of the static stress field. There is also a strong possibility of the closing of microcracks by the static stress in the rock mass when the confining pressure is above 100-300 MPa. Again, when a stress field exists in a direction normal to pre-existing radial cracks around a MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 85 blasthole, it can be sufficiently strong to prevent extension of these cracks. Moreover, it may induce the formation of new cracks in the direction of the stress field. These types of unexpected blast results are often due to in situ stresses in the rock mass. Structure The bedding planes and joints in a rock mass tend to dominate the nature of the blast-induced fracture pattern. Maximum fragmentation is generally achieved where the principal joint planes are parallel to the free face. Where the angle between joint planes and the face is within a region between 30° and 60°, the blastholes may produce an irregular new face, owing to the formation of wide cracks behind the blastholes. When the joint planes are at right angles to the face, each block requires at least one blasthole in order to obtain satisfactory fragmentation (Figure 3-22). Figure 3-22: Blasthole placement in relation to joint planes. If there is a high density of joint planes normal to the face, it is worth while to consider adopting smaller diameter blastholes at closer spacings. Pronounced bedding planes can also inhibit the explosive's strain energy transmission from one stratum to the next. This has particular importance where the stratum near the collar of the blasthole is hard and is filled with a prescribed length of stemming material to avoid fly-rock or early escape of the explosive's gas energy. This situation can be dealt with in two ways. The first is to place a small 'pocket' charge centrally within the stemming column (Figure 3-23). The second method is to drill 'stab holes' half-way between the drilled burden and spacing, and lightly charge them although this may not be practical from an operational standpoint. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 86 Figure 3-23: Use of Pocket Charges Figure 3-24: Stab hole The physical distance between the structures and how that would affect blasting should be considered. Pre existing cracks in this area will direct or even dictate the fragmentation size. The three most negative geological effects (structurally related) on blast performance include: - Rock that has open structures - Zones of incompetence within the rock, in which structures are unpredictable - Rock with alternate zone of competent and incompetent rock. Close and tight rock structure are preferable as the explosive energy is not lost or vented. The problems arise when the energy is not confined or when the transmission of stress waves within the rock mass are interrupted. Open or widely separated structures can result in poor fragmentation due to: - Interruption of the explosive generated stress waves. Causes an inconsistent formation of cracks - Disruption of confinement resulting in oversize. Venting and airblast can also occur in weak seams or open layers of rock. This can be corrected through stemming and decking. These areas can be identified in the drilling process when the drill experiences slower penetration rates related to poor hole flushing. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 87 Some solutions to addressing these problems include: - Closer initiation intervals (achieves desired rock breakage before allowing existing cracks to open further) - Altering design (burden, spacing, and hole diameter) - Selecting more dense explosives or blasting agents (detonate at higher velocity) - Use of multiple decks or cartridged explosives smaller in diameter into the stemming zone. Structurally induced Interruption of stress waves Figure 3-25 shows the type of radial cracking which one might expect when blasting a single hole in a brittle, massive rock formation. There will be a relatively few long cracks (6-8) spaced uniformly around the hole. As one approaches the hole the cracks will be shorter and more numerous. Figure 3-25: Idealized radial cracking surrounding a single hole. The maximum length (Re) of the radial cracks for a given explosive and rock type can be shown to be directly dependent on the hole radius. Thus as the hole diameter is increased from 150 mm to 310 mm the length of the longest cracks would be expected to about double. This is consistent with the design relationship B = K B D presented earlier since the burden should be related to the lengths of the cracks generated e R B · If the strength of the explosive used in a hole of given diameter is increased or decreased, the outer crack radius should change accordingly. This is reflected in the value of K B chosen. Since in general, a larger diameter hole is less expensive to drill than one of smaller diameter (on a cost/volume basis) the natural conclusion would be to drill as large diameter holes as possible. Unfortunately fragmentation considerations would suggest just the opposite, MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 88 i.e. the holes should be smaller to better distribute the explosive throughout the rock mass. To illustrate this some simple geometric reasons will be given. Figure 3-26 shows two possible blast patterns using different size holes but the same explosive. The specific energy (powder factor) is the same for both. A simplified representation of the radial cracks after blasting is shown in Figure 3-27 for each pattern. As the hole diameter is increased and the pattern expands, the distance between adjacent crack tips becomes greater. For the case shown L>L' Thus even though the energy density is the same, the fragmentation is more coarse. Generally as the pattern is spread, the powder factor (energy factor) must be increased to maintain acceptable fragmentation. Figure 3-26: Extent of cracking for two patterns with different hole diameters and same PF MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 89 Figure 3-27: Maximum block dimensions for hole patterns. One way of maintaining proper fragmentation is to increase the powder factor by limiting the pattern spread to some proportion of the theoretical value. As shown in Figure 3-28, there is now an overlap of the longest fractures. Another way of accomplishing this would be to increase the energy of the explosive being used. Figure 3-28: Fragmentation enhancement achieved by reducing the spacing Therefore, even in massive rocks, because of the point introduction of energy into the rock and the fracture geometry, there are limiting hole diameters/burdens/spacings which yield acceptable fragmentation. It is well known that an actual rock mass generally contains many discontinuities of different types. If such structures (joints in particular) are now introduced, such as is shown in Figure 3-29, the story becomes even more complex. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 90 Figure 3-29: Effect of jointing on fragmentation The radius of influence for any given hole is significantly reduced since - The radial cracks will not cross the gaps formed by the joints. - The high pressure gasses can be short-circuited by the less resistant joints compared with the fresh cracks. Therefore the primary fracturing effectiveness is reduced as well as that produced by a sustained heave of the fractured material. Although these pre-existing cracks limit the formation of new cracks and provide avenues of escape for the explosive gases, mobilization of these is a major reason why the specific breakage energy in blasting is much lower than other processes which must attack the intact rock. Figure 3-30 shows two potential drilling patterns in the jointed rock. The smaller diameter, closely spaced holes yield almost one hole per block and the fragmentation would be expected to be good. On the other hand, the larger holes on wide spacings could yield a large number of substantial blocks largely isolated from the effect of the explosive by the joints. Pattern (a) would have higher associated drilling and blasting costs than Pattern (b). By assigning costs to the degree of fragmentation and knowing the overall ideal fragmentation requirements, an 'optimum' pattern can be determined. Figure 3-30: Possible blast patterns in Jointed rock Structure Orientation The orientation of the major structures can have a significant effect on blasting results. There are three cases to be considered: - Shooting with the dip - Shooting against the dip - Shooting along the strike MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 91 In shooting with the dip (Figure 3-31) one finds - a tendency to get more back break - less toe problems - a smoother pit floor - more movement away from the face and therefore a lower muckpile profile. Note: less sub-drilling may be required in this case owing to the fact that the explosive energy may follow the strata downward, eliminating toe problems. Furthermore, inclination of the drillholes in the direction of dip may reduce backbreak but will also tend to ‘cast’ (throw) and spread the muckpile. A slight addition of delay time in the back row may provide relief to the back of the shot resulting in a more stable highwall. Figure 3-31: Shooting with the dip When shooting against the dip (Figure 3-32) one finds - less backbreak since the strata is dipping into the wall. - the toe would be more difficult to pull. - a rougher floor condition. - the muckpile may be higher with less movement from the face. In this situation, the rock tends to move upward, parallel to the plane of the joints. As a result of explosive energies migrating into the strata, a rock unit may cause back break. This situation may result in the creation of an unstable highwall. Also the muckpile would tend to be poorly displaced (creating more work during excavation). If a rock unit is massive, a saw- toothed effect along the floor and overhand at the crest of the slope could result. To eliminate toe problems, the blaster may consider: - a blast using angled drill holes, - high energy explosives in the toe area and/or additional subdrilling. - Decking, satellite holes or small diameter explosive charges in he crest area might help to alleviate an overhang. - Pre-splitting may also be a highwall stabilization option. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 92 Figure 3-32: Shooting against the dip Finally, when shooting along the strike (Figure 3-33) one finds that: - the floor can be highly sawtoothed due to the different rock types intersecting the floor. - for the same reasons the back break is irregular. - these are some of the worst conditions for those involved in drilling and blasting. To overcome this, the working face may be reoriented to a more favorable conditions. Figure 3-33: Shooting along strike ADDITIONAL RESOURCES X Additional learning resources include: - Readings, Chapter 17: Rock and rock mass properties and their influence on the results of blasting, from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (Mandatory for assignment 5) - also considered readings for Module 4.3 – Patterns and sequencing. - Lecture Module 4.2 – Geological impacts on Blast Design. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 93 Patterns and Sequencing The formations should be examined to identify the strike and dip direction of the most prominent joints. In igneous and metamorphic rock formations, one should consider aligning the rows of holes parallel to the alignment of the dominant joint system. In sedimentary rocks, the drill holes should be placed in rows drilled parallel to the formation strike line. When considering the option of altering the blasthole pattern by increasing the spacing (parallel to joints) and reducing burdens (perpendicular to joints) or using a staggered pattern, the direction of movement or angle of movement controlled by delay intervals should be considered. A diagram of blast hole location and firing time intervals will assist in predicting the true burden and spacing firing angles relative to existing geologic structure. Millisecond (MS) delay blasting was introduced in open pit quarry blasting many years ago. Even when blasting to a free face, the rock movement time can be an important factor. This is particularly true in multiple row blasts. For a typical quarry with 15-foot (4.6 m) spacings, the initial movement at the free face may occur in 10 to 12 milliseconds, but the burden only moves about 0.5 foot (15 cm) in 10 milliseconds. With one or two rows of holes, the prime movement is directly out from the face. As the number of rows increases, the rock movement will tend toward the vertical. This is caused by the low velocity of the broken rock successively reducing the relief toward the free face. This can contribute to "tight" bottom as well as flyrock Figure 3-34: Increasing numbers of rows increases upward movement due to lower velocity of previously broken rock. Delay Blasting techniques are employed to improve fragmentation, control of rock movement, overbreak, and to reduce ground vibrations. Delays are incorporated into the blast design using electric or nonelectric caps or delay connectors with detonating cord. The delay patterns used in design will determine the sequence of hole or deck initiations, thereby, dictate the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 94 overall direction of blasted rock movement and resulting fragmentation. Depending on the S/B ratio, the actual timing (in milliseconds) between detonating charges will determine muck pile displacement height and distance from the bench. Figure 3-35: Muckpiles illustrating the difference between millisecond timed and instantaneous. Figure 3-35 illustrates the difference between two adjacent blasts. The blast on the left used millisecond timing and the blast on the right instantaneous firing. It may also be important to provide additional nomenclature for the various types of blasts in terms of degree of fixation. Figure 3-36 shows a corner blast and face blast. Figure 3-36: Corner and face blasts. Effective Burden and Spacing Depending on initiation sequence, an effective burden B e and effective spacing S e result as shown in Figure 3-37. The figure shows a variation of timing used for surface blasting called echelon or half chevron. The effective spacing is the distance between holes in a row defined by adjacent time MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 95 delays (e.g., delays by rows). Effective burden is the distance in the direction of resultant rock mass movement. Figure 3-37: Echelon or half chevron. The “V” (Vee) pattern, also known as chevron, (Figure 3-38) is applicable to most types of formations. It can be readily adapted to the square or rectangular pattern. When a V/MS delay pattern is used in conjunction with a square drill pattern, the angle of movement is 45 degrees to the open face. Therefore, a 10 by 100 foot (3 by 3 m) square pattern becomes a rectangular pattern with a 7.07-foot (2.1 m) burden and a 14.14-foot (4.3 m) spacing or the burden is only one-half the spacing. If a rectangular drill pattern is used, the angle of movement will vary in relation to the relative (also known as ‘effective) burden and spacing dimensions. The formula to determine the angle of movement in relation to the open face for a rectangular pattern is: S B A = tan where B = burden S = spacing A = angle of movement Therefore effective spacing S e and effective burden B e would be calculated by: A S B A B S e e sin and sin = = Example: Determine the effective burden and spacing for a V initiation pattern with an 8 ft drilled burden and 13 foot spacing on a square drilled pattern. Arctan (8/13) = 31.6˚ Se = 8 / sin(31.6) = 15.26 ft Be = 13* sin(31.6) = 6.8 ft MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 96 Figure 3-38: “V” (Vee) or Chevron initiation pattern on a square drill pattern. Designing the Timing The design of initiation timing for multiple-hole blasting is critical to the blasting effectiveness. If the interhole delay is too short, the movement of row burdens is restricted and fragmentation is poor. High ground vibrations result, and backbreak along the new high wall may persist, jeopardizing the stability of the slope. If interhole delays are too long, cutoffs of surface delays may occur. The minimum time for design is controlled by the stress wave travel distance (= 2 B e ) in order for radial cracking to begin to develop, contributing to the detachment of the rock mass in the vicinity of the hole. This detachment forms an internal free face (or relief) to which successive detonations will interact with the reflection of stress waves. The minimum timing is, therefore: 3 10 2 × = p e C B t where t = stress wave travel time in milliseconds (ms) B e = effective burden or distance form the hole to the free face in feet C p is velocity of sound for the rock in fps The maximum timing is that at which the burden is fully detached and accelerating as gas pressures build. Research has shown that stress wave travel time is a fraction of the time required to develop radial cracks. Furthermore, studies using high-speed photography indicate that the burden moves within a timeframe which is between 2 to 10 times the wave travel MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 97 time to the face. Other research has shown that the time to burden movement ranges from 5 to 50 ms, and suggests an optimum range of timing for design between 1.5 to 2.5 ms/ft of B e . Timing studies have been performed to investigate resulting fragmentation and muck pile shapes. Reduced-scale research using a variation in delay ratios suggests improved fragmentation for timing between 11 to 17 ms/ft of B e . Even more research has demonstrated improved fragmentation for S/B ratios of two at timing ratios of 1 ms/ft of B e or greater. Subtantial research has been undertaken in production-scale, multiple-row blasting resulting in recommended timing to improve various aspects of the resulting muckpile.: - For optimum fragmentation, some suggest delays of 1 to 5 ms/ft within rows and 2 to 15 ms/ft of B e between rows (or on the echelon). - For optimum breakage and forward movement, from measured flyrock velocity, and gas venting, through the collar stemming, it was established that 3.4 ms/ft of hole spacing and 8.4 ms/ ft of B e . - Forward throw and muckpile shape: similar work in which muck pile profiles were mapped indicates that optimum forward throw and muck pile height reduction occur for delay ratios of 4.2 ms/ft of S e and 10 ms/ft of B e while forward throw is minimized, resulting in high muck piles, with ratios of 1.5 to 2 ms/ft of S e and 5 to 6 ms/ft of B e . - For single-row production shooting and S/B of 1.2 to 1.6 that timing ratios greater than 1.2 ms/ft of B e are ideal. - Rock Types: One researcher recommended 1.2 ms/ft of B e for multiple- row production blasting in hard rock, while using high powder factors and short stem lengths. A 2.4 ms/ft of B e was recommended for soft rock with long stem lengths and low powder factors. - Ground vibrations: to control ground vibrations, it has been recommended that a 1.3 ms/ft of S e and 1.2 to 4.3 ms/ft of B e is to be used. The timing ratios cited are found to vary over a wide range. A great deal of research on the effects of initiation timing cannot be compared due to the lack of similar variables such as geology, scale, and explosive type. Some researchers have recognized the need to qualify delay ratios, in a general way, based on existing fracture density. Competent dense rock requires lower delay ratios to achieve fine fragmentation, while weak fractured rock fragments best with higher delay ratios. Generalized Timing When taking both pattern shape and timing into account, the most desirable overall drilling/initiation pattern would the drilling pattern with the best energy distribution. Table 3-3Error! Reference source not found. provides the effect of drilling patterns and S/B ratios on the area covered by fracture circles. However, the existence of fractured ground where existing MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 98 fracture planes limited the development of new radial cracks would reduce the efficiencies involved in maximizing the energy distribution, as discussed in previous lectures. Overall initiation patterns are best where each blasthole is initiated separately and in sequence. Alternatively, simultaneously initiated blastholes should be far enough apart to prevent mutual interaction between their stress fields. This is usually best achieved through the “chevron” or V patterns. Table 3-3: Effect of drilling patterns and S/B ratios on the area covered by fracture circles (energy distribution) Figure 3-39 provides a summary of events and their timing for a single blasthole allowing the reader to visualize the chain of events. Figure 3-39: The events and timing in a blasthole MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 99 Skipping a Period It is a common practice of many blasters to double the delay time on the last row. This provides additional time for the rock ahead of the last row to move forward so that the relief on the last row will be increased. This practice, called "skipping a period", will also reduce the upward ripping action and materially reduce the backbreak on the face. When the blast consists of as many as eight or nine rows, the timing on MS delays should provide the additional time without skipping a period. As shown in Figure 3-40, the No.1 through No.8 periods (25 through 200 MS) provided a nominal 25 milliseconds between each period. No.8 through No.15 (200 MS through 500 MS) provided a nominal 50 milliseconds between each period. No. 15 through No. 19 (500 MS through 1,000 MS) provided a nominal 100 milliseconds between each period. This sequence is provided only as an example for discussion: the actual sequences and intervals of detonator timing vary with manufacturer. Figure 3-40: Blast timing sequence skipping periods Always base timing designs on the limitations of detonator accuracy since delays of a given period have a range of actual firing times. Check with the suppliers to avoid overlap or crowding. Even with additional time between rows, the tendency still exists for the rock to stack if the number of rows is excessive. The hole diameter, burden and spacing, and height of face all have a pronounced effect on the number of rows that can be fired successfully without excessive stacking or without encountering high bottom. When the rock is broken, it will occupy on the average 30 percent more volume (swell factor) than it did in the solid. Remember that swell factor will vary with the type of rock. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 100 In most cases the material has only two directions to move, to the front and vertically. Obviously, excessive movement in either direction will result in dangerous flyrock. If the number of rows is excessive, forward movement is limited, thus additional space for forward expansion cannot be provided. Pattern Types The next few sections will provide general guidelines for a few blast designs which can be altered by mine planners to meet needs and conditions. Remember that there are three basic drill pattern types into which these patterns can be designed: square; staggered; and rectangular. Note that various initiation systems can be used to achieve the various timing options. The specific numerical dimensions and timining of the figures reflects only suggestions or possible options. The shape, direction of movement and relative timing are what the figures are intended to represent. “V” (Vee) Pattern The V pattern (Figure 3-38) is applicable to most types of formations. It can be readily adapted to the square or rectangular pattern. It may be used with a staggered pattern, but is not as practical for ease of loading under field conditions. If a rectangular drill pattern is used, the angle of movement will vary in relation to the relative burden and spacing dimensions. Note that this pattern results in a high concentrated and centralized rockpile. Another advantage may be that the impacts between rock in different rows may additionally fragment the muck. The forward movement is controlled within reasonable limits and the broken muck is deposited in a windrow 90 degrees to the open face. Depending on the formation and the number of holes per row, it is sometimes desirable, as seen in Figure 3-41, to open the blast using two holes on the first delay. This will result in slightly more forward movement in the muckpile. Note that his type of drill pattern also results in an effectively staggered initiation sequence. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 101 Figure 3-41: Chevron or V pattern with double hole initiation for slightly more forward movement. Echelon Delay Pattern When the blast area is open on two adjacent sides in an external corner, the blast should be designed to take advantage of the reduced degree of fixation. Note that these patterns would result in a lower, flatter muckpile. Other advantages include a minimization of tight toe problems and an easier connect-up (pattern tying) since the paths can be easily seen. Figure 3-42 is an example of an echelon delay pattern and is also known as a half-chevron. Figure 3-42: Echelon Delay Pattern MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 102 Channel Delay Pattern Operations are frequently opened on side hill terrain where there is not sufficient area opened in front of the blast for forward movement associated with the V or echelon patterns. In these cases a channel delay pattern can be used to confine the broken rock to the blast area. The design of the blast must be relatively narrow, in certain formations it may be desirable to skip a delay period to allow additional time for rock movement because only the direction of movement for the initial holes is vertical. The channel pattern will not hold all the broken rock within the blast area if the terrain is steep. Figure 3-43: Channel pattern Flat-face pattern The flat face pattern will move the broken material farther from the face and usually will result in rock of larger size. This pattern should only be used for special conditions. It is almost always used with a staggered pattern unless large blocky material is desired. Figure 3-44: Flat face pattern MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 103 Alternating hole delay pattern This pattern has been used with limited success in thinly laminated formations on very wide spacings. It is not recommended for most formations. Figure 3-45: Alternating delay pattern Sinking Blast Pattern When opening a new pit or starting a new lift in an existing pit, it may be necessary to make a sinking blast, also known as a sinking shot or sink. A sinking blast varies from most blasts because there is no open face or relief, and the direction of blasted rock movement must be vertical. Since the entire blast will be ‘in the tight’, which levels of vibrations and the generation of flyrock are more likely. In most formations it is necessary to decrease the burden and spacing of the initial holes in the delay pattern in order to open an area of relief to which the remaining holes may break. Figure 3-46: Sinking shots MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 104 Sequential Pattern Variant Note that considering that various initiation products (det. cord, det. cord delays, shock tubes, electric detonators, programmable detonators, etc…) can provide any timing sequence possible, variants of the timing can be created allowing movement planes not directly aligned along drill holes. Note each hole in Figure 3-47 is initiated at different times yet is similar to the movement plane of an echelon blast. Blasting each hole individually may also aid in reducing vibrations from the blast as less explosive material is initiated simultaneously. Figure 3-47: Sequential Firing ADDITIONAL RESOURCES XI Additional learning resources include: - Readings, Chapter 27: Initiation Sequence from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (optional) - Lecture Module 4.3 – Patterns and Sequencing MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 105 Module 4: Blasting Applications 8 The section will cover the more advanced aspects of blast design in terms of its various applications. Drift Design The blasts in tunnels and drifts are characterized by the initial lack of an available free surface towards which breakage can occur; only the tunnel heading itself. Note that for tunnel rounds that are too large to be drilled in a single pass (as seen in many civil applications), can be developed in multiple stages as seen in Figure 4-1. 8 These notes were assembled directly from the following references: - From Dr. Paul Lever’s 415 course notes Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2 nd ed. 1992 - Hartman, Howard L. and Jan M. Mutmansky, Introductory Mining Engineering, 2 nd ed. New York: John Wiley & Sons. 2002, 570p. - Atlas Copco, The Raise Boring Handbook 2 nd ed. - Stefanko, Robert. Coal Mining Technology Theory and Practice. Littleton CO.:Society of Mining Engineers. 1983 - McKercher, R.M., Ed. Potash Technology: Mining, Processing, Maintenance, Transportation, Occupational Health and Safety, Environment. Pergamon Press: Toronto. 1983 - Mining-Technology.com, search: continuous mining - Bell, F. G. ed. Engineering in Rock Masses. Butterworth-Heinemann: London. 1992. ISBN: 0 7506 1063 8 - Caterpillar Performance Handbook, 28th Ed. - 2001 Class notes, from Bob Cummings - 1997 Class notes, from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department.Wirtgen America Inc. Surface Mining Manual. Edition 2002 - Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2nd ed. 1992 - Kennedy, B.A. Editor. Suface Mining 2nd Edition. SME: Port City Press, ML. 1990. - Persson, Per-Anders, Holmberg, Roger, and Lee, Jaimin. Rock Blasting and Explosvies Engineering. CRC Press: New York. 2000. 534 p. - Cook, Melvin A. The Science of Industrial Explosives. IRECO Chemicals: Salt Lake City, Utah. 1974. 449p. - Sen, Gour C. Blasting Technology for Mining and Civil Engineers. Sydney: University of New South Wales Press. 1995. 146 p. - The History of Explosives. http://sis.bris.ac.uk/~dj9006/explosives/history.html - Hustrulid, William. Blasting Principles for Open Pit Mining. Vol. 2 – Theoretical foundations. Rotterdam: A.A. Balkema. 1999. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 106 Figure 4-1: Systems for advancing large faces The principle behind tunnel blasting is to create an opening by means of a cut and then stoping is carried towards the opening. The opening usually has a surface of 1 to 2 m 2 , although with large drilling diameters it can reach up to 4 m 2 . In fan cuts, the cut and cut spreader blastholes usually occupy most of the section. Stoping can be geometrically compared to bench blasting although it requires powder factors that are 4 to 10 times higher. This is due to drilling errors, the demand made by swelling, the absence of hole inclination, the lack of co-operation between adjacent charges and, in some areas, there is a negative action of gravity as happens in lifter holes. Figure 4-2: Different zones in a drift round Contour holes are those which establish the final shape of the tunnel and are placed with little spacing and directed towards the interior of the mass to make room for the drills in collaring and advance, as seen in Figure 4-3. Figure 4-3: Orientation of the contour blastholes to maintain tunnel profite- lookouts MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 107 The position of the cut has influence on rock projection, fragmentation and also on the number of blastholes. Of the three positions, comer, lower center and upper center, the latter is usually chosen as it avoids the free fall of the material, the profile of the broken rock is more extended, less compact and better fragmented. Cuts. The blasts in tunnels and drifts are much more complex than bench blastings owing to the fact that the only free surface is the tunnel heading. The powder factors are elevated and the charges are highly confined. On the other hand, burdens are small, which requires sufficiently insensitive explosives to avoid sympathetic detonation and at the same have a high enough detonation velocity, above 3000 ms, to prevent channel effects in the cartridge explosives placed in large diameter blastholes. This phenomena consists of the explosion gases pushing the air that exists between the column charge and the wall of the blasthole, compressing the cartridges in front of the shock wave, destroying the hot spots or excessively increasing the density of the explosive. Cuts can be classified in two large groups: - Parallel hole cuts - Angled hole cuts. The first group is most used in operations with mechanized drilling, whereas those of the second have fallen in disuse due to the difficulty in drilling. They are only applied in small excavations. In the following, the different types of cuts are explained in their order of importance, as well as calculation of the patterns and charges in the rest of the sections which are, generally speaking, independent from the type of cut applied. The primary function of the cut remains the same regardless of the type of cut or its variations. To be successful, it must break the rock and move it forward. This creates a void which provides additional relief for the remaining holes to be fired later in a predetermined sequence. Burn Cut (Hopler) This is a closely spaced group of boreholes drilled parallel to the direction of advance and perpendicular to the existing face. They are blasted at or near the center of the face to break a roughly cylindrical opening to the intended depth of the round. The boreholes that surround this cut area are sequenced to fire later and break to this newly created opening. It is important that the bum cut holes be drilled accurately and parallel to each other. Improper location of the burn cut holes may result in "bootlegs," (lengths of borehole left after the blast that may contain un-exploded explosives. Bootlegs may be caused by: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 108 - sympathetic detonation from propagation between holes that are too close, thereby destroying the delay sequence - excessive burdens between the bottom of the holes. Figure 4-4: Various types of burn cuts (solid dots are loaded holes) The varied rock types and structures determine the drill spacings, detonator delays, and types of explosive that will successfully fragment and remove the rock to the full depth of these boreholes. Angle Cut An angle cut is a group of boreholes drilled at various angles inclined to the free face to provide as much freedom of movement for the rock as possible. Types of angle cuts include the Vee, the pyramid, and the drnw or hammer. Angle cuts generally require fewer holes and less powder-per-round than burn cuts. However, they are generally more difficult to drill and require more experienced miners. Usually the angle cut will move the rock further down the heading and produce coarser fragmentation out of the cut area. Longer drill steel is required in the cut holes to achieve the same depth as the surrounding boreholes in the round. In narrow headings it is difficult to drill a sufficiently wide angle to insure "pulling" the cut. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 109 Figure 4-5: Angle cuts (Vee or wedge, top left; double Vee or baby cut, top right; three-hole pyramid, bottom left; and a draw cut, bottom right) Cylindrical cut The cylindrical cut can be considered a parallel hole cut, and is the most frequently used in tunnelling and drifting, regardless of their dimensions. It is considered to be an evolution or perfection of the bum cuts which will be discussed later on. This type of cut consists of one or two uncharged or relief blastholes towards which the charged holes break at intervals. The large diameter blastholes (65 to 175 mm) are drilled with reamer bits which are adapted to the same drill steel which is used to drill the rest of the holes. All the blastholes in the cut are placed with little spacing, in line and parallel, which explains the frequent use of jumbos which come with automatic parallelism. The type of cylindrical cut most used is the four section, as it is the easiest one to mark out and execute. The calculation method for patterns and charges of this cut and for the rest of the tunnel zones, uses the Swedish theories and empirical updates. Figure 4-6: Cylintrical cut Blast Layout Designing the actual blast is considerably difficult. Beginning with an initial design, an experienced miner would modify the pattern as ground conditions MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 110 change or as assessing blast results point to necessary changes. Design specifications are typically made for a ‘standard’ for each rock type in a mine, to be modified depending on the cut location, ground conditions, or other factors that may improve the results. Note that in most countries, drilling into bootlegs is not permitted. Therefore cuts usually alternate from side to side. There are two mechanisms used to sequence the holes, some blast the back last, other the floor (ostensibly, to ‘fluff-up’ the muck that has compressed from downward blasting). Figure 4-7a shows the general areas that are initiated in order based on blasting the floor last while Figure 4-7b shows a layout when blasting the back last. Figure 4-7: Sequencing drifts Follow the mandatory readings to determine the charge lengths, densities, hole spacing, and lengths of the blast. ADDITIONAL RESOURCES XII Additional learning resources include: - Readings, Chapter 22: Blasting for tunnels and drifts from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (mandatory) - Lecture Module 5.1 – Drift Development (mandatory) - Assignment 6 – in class (part of lecture) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 111 Production Blasting Metal-bearing orebodies are extracted in underground mines by various methods, depending on the size, orientation, depth and geological characteristics of the deposit. However, the excavation work is usually divided into two broad categories; development and production. Development involves tunnelling, shaft sinking, cross cutting, raising, etc., so that the ore bodies are easily accessible and transportable after excavation. The production work can be subdivided into two categories: short-hole and long-hole blasting. Short-hole Production In short-hole blasting the diameter and length of shotholes are usually limited to 43 mm and 4 m respectively. Short-hole blasting can be used in both stope and room-and-pillar mining. Short-hole 'breast stope' blasting is most commonly applied in narrow, tubular orebodies such as gold or platinum reefs. The usual blast pattern in a South African narrow reef of about 1.0 to 1.2 m stope width (or height) comprises staggered rows of 35 mm diameter blastholes. The length of the blastholes is about 1.2 m. Both rows are drilled at 70° to a line parallel to the stope face and close to the hanging wall and footwall, the burden of each hole being 0.5 to 0.6 m. The most common types of explosive used include a cartridged nitroglycerine-based semi-gelatine or emulsion explosive with a composition density of 1.25 g cm-3 and a column of pneumatically loaded ANFO with a density of between 0.8 and 0.9 g cm-3. In an average stope, there are about 120 shotholes per panel and, to prevent the hanging wall and footwall being damaged, the present practice is not to detonate two or more shotholes simultaneously. Many mines use a pyrotechnic system of initiating explosives in the shotholes. Capped safety fuse and igniter cord (burning speed of about 18 s m-1) make up the carrier of the initiating system. The ends of the safety fuses protruding from the charged blastholes are connected sequentially with a trunkline of igniter cord, as seen in Figure 4-8. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 112 Figure 4-8: Short-hole production blast. The igniter cord itself is usually fired by an electric starter for igniter cord (ESIC). The ESIC consists of a plastic capsule into which a standard electric fusehead (fitted with short lead wires to energise the system with an exploder) and a pigtail of fast igniter cord which has been crimped. If a correct combination of burning speeds of safety fuse and igniter cord is achieved the shotholes are detonated in sequence at 4 s to 8 s intervals. The distance separating a detonating shothole and the flame front in the igniter cord is known as the burning front, and is usually between 3 m and 5 m. Experience has established that the likelihood of igniter cord trunkline cut offs from rock projectiles decreases as the burning front increases. The powder factor varies between 1 to 2 kg/m 3 . Longhole Production Blasting Basically there are three long-hole blasting systems: ring blasting, bench mining and vertical crater retreat. Ring Blasting Ring blasting has wide application in massive are bodies with their high rate of extraction at low unit costs. The method requires three distinct operations - The formation of a tunnel, called the ring drive, from a sublevel along the axis of the proposed excavation. - The excavation of an empty space, called the slot, at the end of the ring drive, to the full width of the excavation. - The drilling of sets of radial holes, called Rings, parallel to the slot at appropriate spacing and burden. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 113 Figure 4-9: Ring Blasting In ring drilling, the distance between two consecutive rings is called the burden, whereas the term spacing refers to the ends of the adjacent holes in the same ring, measured at right angles to one, and straddling the outline of the are block, using construction lines (Figure 4-10). Normally, the spacing/burden ratio is about 1.3, but it can be as high as 1.5. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 114 Figure 4-10: Spacing construction for ring blast design. When blasting into a confined slot, an allowance of at least 30% expansion from the solid is recommended, in order to achieve good fragmentation. Also the stemming length should not be more than two-thirds of blasthole length. A guide to the powder factor for some types of rocks is given in Table 4-1. Table 4-1: Powder factors for Ring blasting Since the blastholes in a ring radiate from a centre point of the ring drive, the collars of these blastholes will be fairly close to each other. Hence the blastholes need to have a variable stemming length (Figure 4-11) in order to avoid serious overcharging in the ore body close to the ring drive. It should be noted that the degree of success in ring blasting depends on the degree of accuracy in designing and drilling the blasting holes. The blasthole MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 115 in each ring could be drilled upward as well as downward. If conditions permit, it is preferable to use one delay per ring, but this may generate a shattering effect in the adjacent rock. This phenomenon can be avoided by using two or even three different delays per ring. These delays are not alternated between holes but are apportioned to whole sections of the ring. Figure 4-11: Stemming length for ring blasting Bench Blasting Bench blasting is essentially similar to surface excavation. First a development heading is excavated at the top sublevel to provide drilling space. Then, depending on the thickness of the ore body and/or the availability of drilling machinery, either vertical or horizontal blastholes are drilled to increase the height of the excavation (Figure 4-12). The blastholes can be from 32 mm up to 250 mm in diameter, depending on the amount of pull (which could be 6.0 m or more) and other factors such as quality of rock, fragmentation requirement, etc. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 116 Figure 4-12: Bench Blasting Vertical crater retreat (VCR) Essentially, vertical or subvertical blastholes are drilled downward from the top level to the bottom level. A cuboid of ore body can be excavated from the lower level upward by a number of horizontal slices using the same blastholes. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 117 Figure 4-13: VCR It is imperative that the first set of charges in the blastholes breaks through into the undercut. Theoretically, spherical charges should be placed, to obtain maximum cratering effect. In practice, however, this is achieved when the deviation from the true spherical charge is not greater than a 1:6 diameter to length of explosive column ratio. In the vertical crater retreat method, gravity enlarges the crater dimensions by excavating the whole rupture zone. The size of this cavity can exceed the optimum distance of the charge from the back many times and its extent depends on rock properties and the local structural geology. The optimum distance for positioning the explosive charge should be determined through small-scale crater tests using the same explosive-rock combination. The blasthole charging details have been well described in a case study where the blastholes were of 159 mm diameter and average length 35 m. First a square section wooden plug (100 X 100 X 200 mm) was lowered down the hole, using a 6 mm diameter polypropylene cord, to the desired depth of blockage, about 1.8 m above the free face (Figure 4-14). Then a small amount of 16 mm crushed rock was poured into the hole to obtain a seal. The average explosive charge length was 0.8 m, in the middle of which was a primer attached to a 10 g/m detonating cord down line. The explosive charge was first stemmed by 1 m of sandfill, followed by 1 m of crushed rock. Finally, two detonators of the same delay were used to initiate the down line. The average depth of slice achieved was 3.4 m. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 118 Figure 4-14: VCR loaded explosive column The main advantage of the VCR technique is the decreased possibility of damaging the surrounding rock which in turn reduces the risk of dilution. This technique has been applied successfully in blasting out pillars between cemented cut and fill stopes. It is important to use an explosive of high energy and high-detonation velocity in the VCR method to achieve maximum advance with minimum drilling costs. ADDITIONAL RESOURCES XIII Additional learning resources include: - Readings, Chapter 24: Underground production blasting in mining and underground tunneling, from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (mandatory) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 119 - Lecture Module 5.2 – Production blasting (mandatory) Controlled Blasting The following discusses more of the theoretical aspects of blasting design in avoiding damage control. To facilitate the discussion, consider an area in the blast called the blast damage transition zone (BDT), as seen in Figure 4-15. The extent of each zone is characterized by a radius from the center of the production change. The zones, their extent (expressed as hole diameter) and the corresponding peak particle velocity (PPV values resulting when ANFO in medium strength rock are assumed can be similar to the value in Table 4-2. Figure 4-15: Diagrammatic representation of the BDT of a fully charged hole. Table 4-2: BDT Characterization These zones have been drawn on Figure 4-16 for a blast containing 2 rows of production blast holes (9-7/8" diameter). As seen, it is a square pattern with 5 holes in each row, a burden of 20' and spacing of 20'. In this example it has been assumed for simplicity that R c =5D=4ft R f = 12 D = 10 ft R i = 55 D = 45 ft MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 120 Figure 4-16: DBT for a two row production blast After blasting, the situation is shown in Figure 4-17. In the BDT zone there exists a crushed and a fractured zone surrounding each production hole. Although the crushed zone as well as a small portion of the fractured zone may have been loaded out along with the rock from production rows 1 and 2, here the rock will be assumed to still be in place. The zone lying between the fracture zone and the boundary of the BDT consists of an inner portion which has been influenced by both rows of blastholes (to a distance of about 25') and an outer portion (20' in extent) influenced by just one row of production holes. Figure 4-17: BDT after blasting two production rows. Therefore in the design and implementation of any blast damage control techniques, one should take into account the BDT limit into account. Line Drilling Line drilling, as the name implies, involves the drilling of closely spaced holes along the limit of the excavation. This is shown with respect to the example case in Figure 4-18. The object is to create an artificial plane of weakness which serves to limit the extent of the fracture and influence zones from both the production holes and any buffer (helper) holes placed between the final production row and the perimeter. Generally, these line-drilled holes are not charged with explosive but, if charged, it is with detonating cord or a highly decoupled charge. The purpose for lightly charging the holes is to destroy the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 121 integrity of the rock web. As can be appreciated, close drilling control is essential for the method to succeed. The holes must be drilled so that they all lie in one plane corresponding with the dip of the final pit wall. Some recommendations for hole spacing as provided in Table 4-3. To get hole spacing one multiplies the values in the table by the hole diameter expressed in the same units. When line drilling 6" diameter holes in copper ore, the hole spacing (c-c) should be 12". Figure 4-18: Line drilling positioned along the planned final perimeter Table 4-3: Factors for determining hole spacing In the example there is 45 ft wide zone between hole row 2 and the perimeter. To excavate this rock one might consider using another row of production holes at normal burden and spacing and then one or two rows of smaller diameter buffer holes. For line drilling to be most effective: - It must be used in conjunction with a buffer row - The main excavation charges should be 1 to 3 rows from the pit limit. Line drilling produces one of the best final surface - a smooth, clean face with no backbreak or crest fracture. However because of its high drilling cost, the method has not been commonly used in open pit work. Pre-splitting The pre-splitting technique also involves the careful drilling of relatively closely spaced parallel holes along the final perimeter. Now however the holes are lightly charged and shot instantaneously. The objective is to generate a line of cracks connecting the holes. In this way, it is intended to achieve nearly the same effects: - Terminate the growth of the radial cracks - Act as a barrier to the shock wave - Provide an escape route for the explosive gases. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 122 It attempts to produce the same effect as with line drilling but at a significantly lower cost. Obviously to be of any use the presplit line must be created prior to the blasting of any holes lying closer than I-BOT distance away. Figure 4-19 illustrates typical presplit blast layout using 102-mm (4-in) diameter presplit holes for 250-mm (9y'-in) diameter production holes. For this type of blast, presplit holes would normally be drilled first, ahead of main production holes. The choice can then be made between loading and firing the presplit line or infilling the main blast. In the latter case, the presplit line would be fired instantaneously 100 to 150 millisecond before the main blast. As shown in the figure, the presplit line is formed ahead of the main blast and allows the gas being driven back from the buffer row through the radial cracks to terminate at the presplit line. Figure 4-19: Presplit blast coupled to a 250mm production blast The presplit row in Figure 4-19 has a spacing of 2 m for a 102-mm (4-in.) diameter hole and is inclined at 15° to the vertical. The presplit angle is somewhat dictated by rock structure although a slight angle is preferred regardless of structure for long-term stability as well as for best initial results with large production holes. The figure illustrates the upper bench where two benches will finally run together to form the final face between berms. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 123 Presplit drill requirements become clear when presplit holes needed for the next bench are considered. The drill must be capable of drilling close to the previously produced bench face at an angle of 15° beneath itself so the face can be continued to depth. Currently, this means some form of drifter drill is required limiting the hole size to 102 to 127 mm (4 to 5 in.) diameter. The back row of the main production blast, termed the buffer row, must also be carefully designed with respect to standoff distance from the presplit row and spacing as well as explosives load. The inset sketch on the right side of Figure 4-19 shows how the top portion of the buffer row hole charge acts as a spherical crater charge breaking to the bench surface. Subsequently, main blastholes after the buffer row are designed at regular spacing, burden, and loading for the type of material blasted. One further point to note from Figure 4-19 is the subgrade or, more accurately, lack of subgrade used on the presplit and buffer row holes. This is to prevent damage to the bench below or to the wall at that point. Figure 4-20 is a graph of recommended hold spacings used in presplitting as a function of hole diameter. These data cover various types of material, but are not fully specified in the open literature. Figure 4-20: Relationships between hole diameter and spacing for presplitting from various researchers. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 124 Smoothwall blasting In smoothwall blasting, as opposed to line drilling and pre-splitting, the final pit perimeter lies in the zone of influence from the final row of production holes. This is shown diagrammatically in Figure 4-21. Since the final row of holes lies in the influenced zone, some minor crest fracturing or backbreak may result but the amount of damage is much less than would be produced by the main production blast if no control blasting was used. Figure 4-21: Smoothwall damage zones There are five general followed in the design of the smoothwall row: Rule 1 The burden, spacing and charge concentration of the smoothwall line of holes are selected so that the extent of the associated influence region does not exceed that of the production holes. The hole size for the smoothwall and buffer row holes may be the same as in the production round with the required reduction in influence zone dimension occurring through pattern adjustments and decoupling or smaller diameter holes may be used with or without decoupling. Rule 2 The hole spacing is less than the burden. Often the relationship S/B= 0.8 is used Rule 3 The holes in the smoothwall row are shot on the same delay with detonating cord downlines to assure as simultaneous detonation as possible. Rule 4 The delay time between the helper row (that adjacent to the smoothwall row) and the smoothwall row should be chosen so that the smoothwall holes can shoot to a free face. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 125 Rule 5 All of the smoothwall and buffer row holes are shot together with the main production round. Today, smoothwall blasting is much more common to use than most other methods in mining operations because the method involves less drilling and less complicated blasting. Trim Blasting Trim blasting, as the name implies, involves trimming away some of the fractured and influenced rock from the pit perimeter after the production blast has been shot and cleaned up. The trimming may be accomplished using one or several rows of blast holes depending on where in the BDT zone it is desired that the final crest should fall. The design process differs from the smoothwall technique in that the layout begins at the actual pit perimeter and works outward toward the desired final pit limit rather than vice versa. This is shown diagrammatically in Figure 4-22. Some rules for selecting the burden, spacing and charge concentration as a function of hole diameter are presented in Table 4-4. As in all other types of perimeter control, accurate drilling is important. To achieve the best results, the holes should be drilled at the final pit slope angle. The boreholes are drilled in a line along the planned excavation limits, loaded lightly, and blasted to remove the undesired material. As noted earlier, a reduced explosive load can be obtained in various ways. The use of low density, bulk-loaded explosives in larger diameter holes is one way of improving the economics of the method. Figure 4-22: Trim hole row trimming the fractured and influenced rock. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 126 Table 4-4: Recommended charge loads and blast geometries for trim blasting Blasting Ornamental Rock Ornamental rock is all stone that is used, in blocks or slabs, for its aesthetic characteristics such as color, texture, shine, grain, etc. and technical such as strength, facility of elaboration, polish, etc. The most common types of rock can be generically classified in three large groups: granites, marbles and marmoreal limestones. Granite is cut from the "bed" of the quarry with a jet piercing machine that produces a flame burning at approximately 3,000 degrees Fahrenheit. This high-velocity flame, created by burning oxygen and fuel oil, is directed at the granite to be removed, causing a continuous flaking action. As the flame nozzle is moved up and down, a channel is created around large sections in the quarry. In some quarries, diamond wire saws are used. A long loop of small steel cable, impregnated with industrial diamond segments, cuts the sections free from the bed of the quarry. After a section has been completely wiresawed or channeled by the burner, it is separated from the bottom by explosives. Likewise, when high-speed drills are used, rows of drilled holes are loaded with explosives. The explosives are detonated to free the sections of granite on all sides and on the bottom by explosives. The large sections are then broken into workable sizes by wedging. In this process, steel wedges are driven manually into holes previously drilled along the desired line of cleavage. The sections are readily forced apart and cross-wedged into rectangular blocks. Large cranes, or derricks, lift these blocks to the quarry's rim (Figure 4-23). Requirements for monumental granite are exacting, and only about 50 percent of the granite removed from the quarries finds its way into finished monuments. The remainder is consigned to commercial applications such as street curbing and gravel, or is sent to "grout piles" as waste products. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 127 Figure 4-23: Ornamental stone materials handling – cranes. Case studies show that the cutting methods consist in primary separation from the rock mass of a large block (100 to 4000 m 3 ), in parallel piped form, which is subdivided afterwards to achieve sizes that are easily handled and within the ranges that the transformation industries require, generally lengths of 1.8 to 3.5 m, widths of 1 to 1.50 m and heights of 0.9 and 1.2 m. The cutting technique is usually with explosives, although not exclusively, because cutting systems with helicoidial and diamond wire, with mounted rock cutters, with flame torching and with water jet kerfing are often applied. The blasting techniques are a special type of presplitting, but with slight variations as it is of maximum importance not to damage the rock and at the same time take into account the properties: strength, homogeneity, schistocity, fissurization, etc. Although it is difficult to give general recommendations for design in this type of blasting, as there are many different rock types and exploitation conditions, the following criteria should be of use: Drilling diameters. They depend upon the phase of excavation and the type of rig used, but generally around 25 to 45 mm. Spacing. It is established as a function of the rock properties and explosive charge characteristics. The usual interval is between 4 and 8 D. To be able to make an analytic calculation, the formula suggested by Berta can be applied: D D RT d PE S e S + × × × × = 2 2 µ where: PE s = Specific pressure (MPa), µ e = Density of the explosive (g/cm3), d = Diameter of the explosive charge (m), D = Diameter of the blasthole (m), RT = Tensile strength (MPa). Explosives. In the vertical benching planes detonating cords with a core of pentrite are usually used, while for the horizontal planes explosives of low detonation velocity are also used, as they generate a large volume of gases. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 128 In these last planes of the cut, the structural properties of the rock mass should be used to advantage. In some countries, there is extensive use of charges prepared in connecting plastic tubes that contain powdery explosives with low density and detonation velocity, made up of nitroglycerine, sodium nitrate and other ingredients. Powder factors. These vary greatly depending upon the type of rock, explosive and extraction phase. Charge configuration. The explosive columns are generally designed to be continuous and decoupled with an air chamber although, in some cases such as in hard rock, to increase the energy transmitted to the rock by the detonating cords, the blastholes are filled with water. Also, if a blackening by explosion smoke of the cut surfaces is to be avoided, the holes can be filled with sand or drilling waste. Distribution of the charge in the borehole. In order to eliminate breakings or fracturation in the corners of the blocks, it is suggested that empty holes be used at the end of the line or next to the free surfaces. Apart from this, in vertical blastholes there is no subdrilling and they are usually drilled to a few centimeters above the horizontal plane. Stemming. They are necessary to use the maximum pushing power of the gases. As the rock characteristics become poorer, the heights are usually shortened to assure that the pressure of the gases do not act upon the rock for a long period and therefore produce damage. In general, with the detonating cords the stemmings are small, whereas with powder a larger confinement is necessary. Initiation. As in contour blasting, instantaneous initiation of all the blastholes with detonating cord is recommended. Example A block of granite is to be extracted by drilling blastholes and blasting with detonating cord. What should the spacing between boreholes be when the initial data is: - Tensile rock strength RT = 10 MPa. - Drilling diameter D = 0.032 m. - Diameter of the detonating cord core of pentrite - d = 0.0034 m. - Density of the pentrite charge µ e = 1.3g/cm3. - Specific pressure PE s = 1200 MPa. m S 14 . 0 032 . 0 032 . 0 10 0034 . 0 3 . 1 1200 2 2 = + × × × × = MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 129 The ratio S/D is equal to 4.37, which is within the practical interval of 4 to 8 D. If the rock were of worse quality with a tensile strength of 5 MPa, the spacing should be increased to S = 0.26 m. Underwater Rock Excavation Underwater blasting of rock is usually done to deepen the sea bed in and around a harbor to allow the passage of larger vessels, or excavate trenches for pipelines. The technique can also be extended to winning underwater mineral deposits. The most important factor is the placement of the charges in the designated area. This is accomplished in a number of ways: 1. Using divers to drill and to charge the blastholes. 2. Using shaped explosive charges fastened in a predetermined pattern onto a metal frame or high explosives packed in boxes and connected to a line of detonating cord and then lowered onto the sea bed. 3. Using a barge or platform to drill and charge the shots. Method 1 is only economical if the water is shallow and the area of excavation is small. The procedure is then the same as for land excavation. Method 2 is fairly labor intensive but economical for excavating selected small areas. Its success depends on the intimacy of the contact between the explosives and the rock, whose thickness to be removed by blasting in one pass is generally limited to one meter. In deep water (say >10 m) this is the only method that can be applied economically in the present state of technical knowledge. Moreover, in deeper water, a charge becomes effectively tamped by the pressure of the water, and consequently the efficiency of the explosive in breaking rock is increased. Individual charges are normally 10 to 25 kg, depending on rock type, and the corresponding charge ratio varies between 3.5 and 7 kg/m 3 . Most underwater blasting is done by method 3, that is, from a platform, which may be of either the floating or the jack-up type. Overburden drilling (0D) rigs are used by both types and the blastholes are normally between 51 and 102 mm in diameter. 0D rigs are fitted with chucks so that the equipment can drive a hollow outer casing of steel tubing through soft material, as well as drill rock through the casing. Strict control on positioning each blasthole at its designed place is extremely important for the success of this method. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 130 Figure 4-24: Sequence of operations in underwater rock excavation. The sequence of operations is shown in Figure 4-24 and can be described as: - The outer casing, fitted with a hollow ring bit with serrated edges, is driven through the water until the ring bit, by rotating, grips firmly into the top part of the bedrock (a). - Then the casing is uncoupled from the drill chuck and the normal drill string is inserted through the hollow casing for blasthole drilling (b). - When the required depth is reached (Note: normally, the subgrade length is the same as the burden), the drill string is withdrawn (c) and the charging procedure takes place through the hollow casing as follows: the primer cartridge with detonating cord or signal tube is pushed through the hole first by using a string of stemming rods with a flat-ended wooden coupling for easy handling of explosives (d). - The length of the cord or tube should be adequate to allow for rise and fall of the tide when it is brought up to deck level inside the casing. The required cartridges are then added. When charging is complete, a slip- ring attached to a hemp rope is lowered to the bottom. Then as the outer casing is withdrawn, the tube or cord is retrieved from inside the tube (e) and brought up to deck level. - A weight is attached to the end of the cord or tube, and the latter is placed on a roller away from the drilling activity. The gantry is then moved to the next line of holes and the operation repeated. The explosive used in this type of operation should have high velocity of detonation and high density. It should have a good degree of water resistance so that 24-hour immersion does not affect its performance, yet its properties should be such that it would be made inert by a long period (say MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 131 one month) of immersion. The powder factor for this type of operation varies between 0.5 and 2.0 kg/m 3 depending on rock conditions. The waterborne blast-induced overpressure P (kPa), can be calculated approximately from R m P 3 1 3 10 55× = where m is mass of explosives in kg, and R is the distance from the charge to the point affected by pressure, in meters. It is estimated that a peak waterborne overpressure not exceeding 40 kPa is safe for humans and animals. Controlled Blasting Rules of Thumb - Supplemental The following points are controlled blasting – presplitting and trim blasting – additions that blasting consultants consider important. 1. Critical geometric parameters are hole diameter, decoupling ratio (d powder /d h ), spacing/diameter ratio (S/d h , charge density (charge/unit length) and drill accuracy. 2. Critical geologic parameters are rock brittleness, rock strength, fracture density, and fracture orientation. 3. Typical S/d h = 10-14 for pre splitting, 16 (to 18) for trim blasting, with the lower values necessary under less ideal conditions. S/d h is more likely to be too large than too small. If the ratio is too high a more rugged, more damaged slope will develop, and there is essentially no way to compensate through loading or timing adjustments. 4. Typical decoupling ratios = 0.1-0.3, the higher ratios being necessary in rock of higher tensile strength. Decoupling avoids shattering and reduces gas penetration. 5. Typical charge densities 0.2-0.4 lb/ft in-hole, face coverage 0.10 - 0.18 lb/sq ft, the higher values associated with stronger rock and/or larger hole diameters. 6. Use linear, distributed charges with cord (50 to 400 grain most common), presplit powder (skinny segmental or linkable cartridges), cartridges taped to cord at intervals, air decked loads, or mass decked loads, in order of increasing potential for rock damage. 7. Presplit works best in brittle, homogeneous rock, trim blasting in fractured, weaker rock. 8. Rock with closed fractures spaced less than 0.3S, or open fractures spaced 0.6S or less, will dissipate presplit elastic stresses. Compensate by closer hole spacing. 9. Fracturing striking less than 15 degrees to a presplit line is very difficult to overcome. Fracturing striking 15-45 degrees is less difficult, but may require decreasing spacing by 25-50%. Unless the fractures are very tight and the rock is very weak, even closely-spaced presplit will tend to break to such fractures. Trim blasting may be more cost-effective in such cases. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 132 10.Presplit shot together with production must be timed at least 50 ms ahead of small-diameter (<4") production holes, at least 100 ms ahead of medium-diameter (4"–8”) production holes, and at least 200 ms ahead of large diameter (>8") production holes. More time is better. The shot width should be at least 3 S (spacing lengths) for presplit confinement. 11.Presplitting: simultaneous firing is ideal. However, if the resulting charge weight exceeds vibration/scaled distance criteria, presplit holes can be delayed as much as 25 ms hole-to-hole, but with less effectiveness. To meet vibration concerns, start by grouping presplit in as many simultaneously-fired holes as possible, delaying 9 ms between holes or groups of holes, and then add time as necessary. Trim blasting: hole-to- hole timing is best; add an extra delay period or two for trim hole relief. 12.Standoff between presplit and buffer (if used) or production holes should be 0.3-0.5 x B aProduction (0.5-0.75 if trim blasting). The back break from production or buffer holes is needed for additional fragmentation in front of controlled holes. Refine standoff according to the buffer/production arrangement. 13.Presplit holes ideally are not stemmed, to permit release of gases. To control noise, they can be stemmed, 0.5S < T < S. Stem trim holes 0.7S - 1.0S. 14.Angled controlled holes perform best, but seldom are drilled flatter than 50 degrees or so. ADDITIONAL RESOURCES XIV Additional learning resources include: - Lecture Module 5.3 – Controlled Blasting (mandatory) - Guest Lecture – Bob Cummings Environmental Issues The use of explosives is probably the most widely used means of fracturing rock. Rock can also be excavated by methods other than blasting such as the use of rippers, hydraulic bursters, plugs and feathers, etc. Techniques involving explosives differ from other available systems in that the energy applied is released in a matter of milliseconds. If the energy release process is not adequately controlled, there is a potential danger of environmental disturbance. Environmental disturbance includes the effects of airblast, fly- rock, changes in the natural profile of the ground, dust, fumes, and ground vibration. The blasting engineer should be aware of the need for defense against allegations of damage caused by the above factors. Before any explosives are used, it is good practice to carry out a detailed survey of all properties that might conceivably be considered at risk of damage, and then to keep meticulous records of all blasts. It is now appropriate to analyze the energy distribution after the detonation of a quantity of explosives. The explosive's potential energy can be MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 133 manifested as heat, rock fragmentation, fly-rock, airblast, and ground vibration. Depending on the placement of the explosive charge, the percentage distribution of the above factors may vary, except for heat emission, which will be constant in all situations. For example, if a quantity of explosive is detonated on the ground surface, the major part of the explosive's energy will be converted into airblast, whereas if the same quantity of explosive is buried very deep in the rock, the majority of the explosive's energy will be manifested as ground vibration. Each aspect which can cause environmental problems is now discussed. Airblast Air overpressure due to an explosive charge (for a given amount) increases rapidly as the confinement decreases. Severe airblast is caused either by inaccurate charging, with the consequent wastage of explosive energy in the air, or by the firing of exposed detonating cord in a sensitive area. The ideal blasting condition for minimum airblast effect is where temperature decreases with increasing altitude, causing a decrease in the sound wave velocity, inducing the waves to bend upward away from the ground. On the other hand, if the air becomes warmer with increasing altitude, an increase in sound wave velocity results, causing the waves to be returned to ground by refraction. Moreover, if there is a strong wind in a particular direction the refracted sound waves are channeled into the wind path. This phenomenon greatly increases the airblast pressure at focal points. The most damaging cases of airblast are caused by unconfined surface charges. The resulting airblast overpressure may be estimated from the following formula: 2 . 1 3 1 185 ÷ | | | . | \ | = m R p where p = pressure (kPa), m = explosive charge (kg) R = distance from charge (m). For confined borehole charges, airblast overpressure may be estimated from 2 . 1 3 1 3 . 3 ÷ | | | . | \ | = m R p When assessing the effect of airblast it is usual to correlate this with the possibility of cracking panes of glass. The pressure waves generated by MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 134 explosives consist of energy over a wide range of frequencies, some of which are audible, but most of which occur below 20 Hz frequency in the concussion range, which is not detectable by the human ear, but can damage structures. While the most frequently mentioned complaint resulting from airblast is cracked plaster, research has shown that window panes fail before any structural damage results. Certain well-known symptoms caused by the peak overpressure levels are given in Table 4-5. Table 4-5: Airblast Effect on Structures Symptoms Peak (kPa) Overpressure (dB) Rattle of loose windows 0.03 140 Failure of poorly-mounted window pane 0.69 150 Damage to strongly-mounted window pane 6.9 170 Cracking plaster >6.9 >170 Noise from drilling and blasting can be annoying, and sometimes even painful. The technology for reducing noise and vibration from pneumatic and hydraulic drill operations without losing performance does not yet exist. On the other hand, close attention to the design of the canopy for compressors can bring about almost silent machines. The engine should be fitted with twin silencers and the inside of the canopy should be coated with sound-absorbing plastic foam to eliminate natural panel resonances. The sources of objectionable blast noise are: - the use of lay-on or plaster charges often used in secondary blasting; - poorly stemmed holes; - blown-out shots resulting from poorly designed blast patterns; - exposed surface lines of detonating cord. In sensitive areas, a covering of 20 cm of stemming material over detonating cord will bring about a significant reduction in noise. Atmospheric conditions such as temperature inversion and wind can lead to the phenomenon of focusing. Normally, temperature decreases with height; correspondingly, the velocity of sound in air also decreases with height, causing the sound waves to curve away from the surface. However, in a temperature inversion (air temperature increasing with height), sound velocity increases with height, causing the sound waves to bend downward towards the surface. The above conditions apply on the assumption that there is no appreciable wind. A combination of wind and temperature inversion can take place several kilometres from the blast and the overpressure due to airblast may increase by a factor of 100. Usually ideal atmospheric conditions for blasting exist in the early afternoon. These airblast wave effects can be seen from Figure 4-25. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 135 Figure 4-25: Airblast waves in various conditions Fly-Rock Fly-rock is the term for undesirable projectiles of blasted material. In a particular type of rock there is a compatible relation between the height of the explosives column in the holes, drilling pattern, and charge ratio. When this is compromised, the explosive's gas energy is vented violently into the atmosphere and propels rocks in front of it. Moreover, deviation of blastholes in surface mining operations can effectively reduce the burden, causing fly- rock. Proper stemming has an important role in trapping explosive gases in the blasthole to do useful work such as rock fragmentation and throw. Crushed angular rock is recommended as the ideal stemming medium and the stemming length should not be less than the burden distance. The major causes of fly-rock in mining are shown in Figure 4-26. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 136 Figure 4-26: Causes of fly rock. Moreover, an out-of-sequence shot has the same effect as an overburdened hole and is liable to cause fly-rock. In excavation work close to buildings, special precautions may be taken such as placing a thick rubber mat (such as heavy conveyor belting. Some very limited field studies reported by various researchers that suggest that for granite the maximum throw (L) as a function of the hole diameter (d) and specific charge is as shown in Figure 9.12. When the specific charge (q) is q 0.2 kg/m3 there is no throw. For other values of q the maximum throw is expressed by L=143 d (q-0.2) where d = hole diameter (inches), q = specific charge (kg/m3), L = maximum throw (m). A typical specific charge in bench blasting is 0.6 kg/m3. In this case the maximum throw expression becomes: L=57d For a 10 inch hole diameter the maximum expected throw would be L = 570 m. There are a number of different situations in which the actual conditions depart markedly from the ideal: - The explosive extends too high in the hole so that cratering to the upper surface occurs. - An irregular face brings the explosive column too close to the free face resulting in cratering. The lack of confinement offered by both of these situations provides a weak link for the gas to exploit. The rock plug involved is pushed out in an early MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 137 stage of the gas expansion process and the expansion energy is expended in propelling a relatively small volume of rock at high velocity. Hence the throw distance can be very great. Of all various types of flyrock, the most dangerous situation occurs when cratering occurs on the top part of the bench near the collar. If the weakest link in the system is the column of stemming and not the collar rock, this can be ejected much like a projectile from a cannon barrel. Some empirical and computer simulation have examined the maximum throw and boulder size as a function of hole diameter. It was found that for granite with a specific gravity of 2.6, the relationships for the maximum throw (L) involving rocks of diameter | L max = 260 d 2/3 |= 0.1 d 2/3 where d = hole diameter (inches), | = boulder diameter (m). Thus for a 10 inch diameter borehole the boulder size would be | = 0.1 (10) 2/3 = 0.47 m. Disturbance of the Natural Ground Profile In an open-cut mining operation by blasting, a turmoil of the natural surroundings is often experienced with the end result of residual benches of bare rock. This can resemble a deep scar in the immediate environment. If the mine working is below the natural profile of the surrounding country the void can be infilled with waste material, or can be filled with water to form a recreation area. However, if the workings are in elevated ground where older workings are actively eroding with rock fall and spalling onto the mine floor or benches beneath, the problem requires different treatment. A method of restoration blasting has recently been adopted in various areas, and has proved to be successful in providing an aesthetically pleasing landscape. This is the application of a carefully designed blast pattern in the berm (bench), initiated to reproduce a predicted sloping profile with infilled materials. The screen blast piles will mask the earlier scorch marks, and this will be the medium for selective establishment of vegetation. Dust Dust is mainly produced while drilling and could be a considerable nuisance in high winds. Apart from incorporating a dust collecting unit in the drill rig itself, or using wet drilling, very little can be done to suppress dust from this operation. Clearly the driller should wear a protective anti-dust mask. In a properly designed shot for rock blasting, very little dust is scattered. However, in demolition work, particularly of a tall structure, dispersal of fine dust is inevitable. This could be partly controlled by sprinkling a jet of water, and thereby creating an artificial curtain of fine mist to contain the falling MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 138 dust in a limited area. Jets of water are also sprayed in some surface mining operations where dust has caused problems in nearby habitats. Fumes The detonation of a commercial explosive in ideal conditions produces water vapor, carbon dioxide and nitrogen. However, in addition a small amount of undesirable poisonous gases, such as carbon monoxide and oxides of nitrogen, known as fume or toxic gases, are hazardous in confined spaces such as underground mines, tunnels, etc. Careful thought must be given to mechanical ventilation of such areas with fans so that fumes are diluted to a harmless level. This aspect is adequately safeguarded by legal requirements laid down by the relevant authorities. The fume characteristics of a cartridge explosive relate to the conditions where the explosive is fired in its cartridge. Removal of explosive from its cartridge will upset the oxygen balance and unfavorably affect the explosive's fume qualities. The presence of water in the blasthole may also adversely affect the fumes produced by a blast. Ground Vibrations When an explosive charge is detonated in a borehole, a pressure wave will be generated in the earth surrounding the hole. As the pressure wave travels away from the borehole it forms a seismic or vibration wave by displacing the particles around it. This excursion or oscillation of the individual particles is measured to determine the magnitude of the blasting vibration. For elastic wave transmission the strain is directly proportional to particle velocity. Since structural damage, in general, is strain related, the use of peak particle velocity is accepted as the principal parameter for vibration measurement. The longitudinal, vertical, and transverse components of the seismic waves caused by blasting are generally measured by instruments. Calculation of the resultant of these three components is sometimes preferred and countries have different standards as regards the suitability of measuring planes, or the resultant. The same amount of charge at a fixed distance does not necessarily produce the same magnitude of vibration. The vibration is largely influenced by the prevailing constriction and by certain physical site factors. The peak particle velocity (ppv) of ground motion can be related to distance and instantaneous explosive charge by the following equation: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 139 P m R K ppv ÷ | | | . | \ | = 2 1 where R = distance in meters, m = explosive charge in kg, K = site factor constant, and P = site exponent constant. The ratio P m R ÷ | | | . | \ | 2 1 is commonly known as scaled distance (SD). SD is a useful parameter for comparing one set of vibration results to others. Obviously, as the value of SD increases, the magnitude of ppv decreases. When blasting has to be done close to a sensitive structure, the following procedure determines the optimum charge. In a given situation when the same blasting is further away (the SD is more than 12), several (at least 6) vibration recordings are taken. On the other hand, if there is any possibility of causing damage to the structure by normal production blasts, a test shot (using a small amount of explosives so that the SD value is well over 12) has to be monitored at various distances. The simplest method of establishing the site constants is to use log-log graph and put the measured values straight onto the axes, that is, vertical (ordinate) for ppv and horizontal (abscissa) for SD. Such a graph is known as a regression curve. Allowing for a certain amount of scatter due to the variation in blasting constriction, the graph is a straight line. The constants K and P are obtained, respectively, from the intercept of the regression line on the ppv axis, and by calculating the slope of the regression line. Figure 4-27 is such a graph obtained by plotting ppv against SD for 10 monitored results. After drawing the best fitting line, two other lines parallel to the first one are also drawn to encompass all points in the graph. Example Consider that there is a sensitive structure where maximum ppv is stipulated as 5 mm s -1 , and this structure is 300 m from the blast area, what will be the maximum instantaneous charge? By drawing a horizontal line from ppv = 5, three values of SD can be obtained: 30, 40, 50. From this, it can be assumed determined that SD = 2 1 m R = 50 will be the best solution. The safest maximum charge per delay becomes 36 kg from: MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 140 36 6 50 300 50 2 2 2 = = | . | \ | = | . | \ | = R m Figure 4-27: Obtaining site factors for ground vibrations. It is essential for the blasting operator to know and to be able to control what will happen when an explosive charge is detonated. The factors listed in Table 4-6 identify the phenomena important in blasting. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 141 Table 4-6: Factors that influence ground motion Figure 4-28: Ground vibrations results from 2500 blasts in 40 different operations. A collection of scaled distance data for determination of peak particle velocity of some 2500 blasts from 40 mines, quarries and construction sites in the USA is shown in Figure 4-28, the graph being drawn from a considerable amount of scatter. The factors responsible for this scatter are the changes of geological conditions, differences between types of explosives, difference in the geometry of the blasts, and experimental error. The line labeled 50 is the best-fit line, whereas the other two are bounds below which 84% and 95% of MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 142 the data fall. The maximum ppv recommended by Standards Australia is given in Table 4-7. Table 4-7: Recommended maximum peak particle velocities. Type of building or structure Particle velocity mm s -1 Houses and low-rise residential buildings; commercial buildings not included in description below 10 Commercial and industrial buildings or structures of reinforced concrete or steel construction 25 This table does not cover historic buildings, particularly if they are in an indifferent state of repair, or some sensitive structures such as reservoirs and buildings with long-span or suspended floors. In the absence of particular site-specific data, a working recommendation is a maximum ppv of 5 mm/s. Vibration reduction By adapting the blasting method, drilling pattern, charging scheme, and ignition pattern, the size of the vibrations can be controlled. Among other things, the size of the vibrations depends on: - Cooperating charge - Confinement conditions - The character of the rock - The distance from the blasting site - The geology, e.g., overlay soil types Practical methods to reduce the ground vibrations by limiting the cooperating charging weight per interval are: - Adapt the ignition pattern so that the charging level is spread over more intervals and the scattering in the delay elements of the detonators is utilized - Reduce the number of holes and the hole diameter - Use decked charges by dividing up the necessary charge level in a drillhole into more ignition intervals by means of sand plugs - Use decoupled charges; charge diameter smaller than hole diameter - Divide the bench into more benches Do not blast to the final depth at once. At the moment of detonation, there should be as little confinement as possible. This can be obtained by: - A carefully adapted ignition pattern, so that all the holes will break the burden in the easiest way - Increased hole inclination (of the drillhole) - A voiding too large burdens and choke blasting For blasting at a shorter distance than 100 m, the risk for interaction between the different intervals is small. The risk for cooperation between the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 143 intervals increases with large blast in quarries: for instance, where any structures sensitive to vibrations are situated a large distance away. The size of the vibrations is then influenced by: - Charging level - Interval times - The resonance frequency of the ground (which depends on the depth and the character of the ground) - The local geology The cost for careful blasting near built-up areas increases very rapidly with decreasing permissible vibration level. The increase in costs primarily depends on the following factors: - Drilling - smaller or greater number of drillholes. - Charging - more detonators and higher cost of labor - Blasting - more rounds and longer stand-up time The costs of planning and control work will also increase in: - Blasting - Visual inspection - Vibration measurement - Blasting record - Insurance administration Cooperating Charges – Ground vibration In the preceding section, it was shown how the maximum detonating weight of charge can be estimated when the vibration level and the distance are known. The weight of the charge in question is the maximum total weight of charges that can be initiated at the same time. Using delay detonators, it is possible to blast rounds with considerably higher total charge weights per delay interval. The higher the interval number, the larger a total charge weight per interval can be used because the scatter of the delay time increases with the nominal delay time. The cooperating charge is defined as the total charge per interval, multiplied by the reduction factor appropriate for the interval used. In the USA, the USBM states that the delay interval should be equal to, or greater than, 8 ms in order that the two charges be considered separate charges. The expression cooperating charge is somewhat inappropriate as it is only applicable at certain distances. When blasting two separate charges with detonators having the same interval number (Figure 4-29), if the vibrations are observed at a short distance (A), the charges do not cooperate; while at a long distance (B), interference and reinforcement of the two vibrations may occur. Whether two charges, initiated one after the other, cooperate or not depends on the following factors: - Time interval between initiations MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 144 - Velocity of propagation of the vibrations in the bedrock - The decay time of the vibrations - Distance from the observation point to the charge - The geometry of the charge - Velocity of detonation - Confinement. Figure 4-29: Cooperating charge related to distance. The effect of Delayed Explosion Reduction of explosive weight per delay is perhaps the greatest factor in reducing the probability of structural damage due to ground vibration. Table 4-8 highlights this point for a situation where the total charge is 2900 kg, the ground vibration monitoring distance is 220 m, and the expected ground vibration is given in three different situations. Table 4-8: Effect of delayed charges on ppv for a total charge of 2900 kg at a distance of 220m. Mode of Detonation ppv in mm s -1 Instantaneous 75 Two delays with equal charges 43 Four delays with equal charges 25 Each delayed charge generates its own seismic wave which is separated from the subsequent delayed wave. However, precise timing for the detonation of each hole is imperative for effective delay blasting. The cardinal point for successful delay blasting is that the seismic waves from any detonating blasthole shall pass all other blastholes before any of them are initiated. If two seismic waves resulting from two blastholes meet at a point, the resultant motion will be the sum of the two motions and the vibration level will be significantly increased. Moreover, the seismic wavelength of the composite motion varies from a single wavelength to nearly twice the length of a single wave. At the point of maximum overlap, the period and frequency are those of the single wave. Since the period may approach twice that of a single wave, the frequency will be reduced by half. This condition may MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 145 produce a region of high seismic risk due to increased motion and reduced frequency of vibration. Effect of Geological Factors The intensity of ground vibration is often influenced by the following properties of rock: - the elastic properties of the medium, which determine the propagation velocity of the seismic waves (predominant range and type of waves); - the moisture content of the medium and the ground-water level; - the topography and the geological structure, which may have an influence in focusing seismic waves; - the absorption characteristics of the medium. Risk Assessment Near most sites where blasting is necessary, there are structures whose sensitivity will limit the maximum permissible vibrations. Client or contractor must then decide on the maximum charge that can be detonated without causing damage in the neighborhood. Where small-diameter holes are used, this maximum charge may be the charge having the same vibration effect as that from several holes detonated with detonators having the same nominal time delay. It is called the maximum cooperating charge. Because the cost increases considerably if the maximum cooperating charge has to be reduced, the economy of the blasting job is greatly influenced by that decision. Too large a cooperating charge may result in damages to neighboring buildings, damage claims, and even court disputes. Too conservative a decision on the size of the maximum cooperating charge will result in excessively increased costs and project time. To optimize the blasting work, it is necessary to carry out risk analysis in order to determine, first, what size vibrations the environment will accept and second, how large a charge can be blasted at a certain distance without exceeding that vibration limit. Before the blasting operations can begin, a risk analysis should be made involving a careful examination of the factors that can affect the blasting operations. The probability that a correct decision will be made increases as more information is made available. The decision data should be based on as many points as possible in the list below. Checklist for risk analysis: 1. Has a geological examination been made regarding working site as a risk area? 2. Is there a potential risk for lowering the groundwater level? 3. What is the nature of foundations and underground parts of buildings in the area? 4. What is the type of construction and the condition of buildings within the risk zone? MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 146 5. Is there any equipment (such as computers, electron microscopes, laser equipment, relays…) which are sensitive to vibrations in the neighborhood? 6. Are there any underground objects (tunnels, cable trenches, telegraph cables, oil cisterns, district heating culverts,...) that might be damaged by blasting? 7. What connections are there between vibration values, cooperating charges, and distances? 8. How are the inhabitants in the neighborhood influenced? 9. Has information about the blasting job been distributed to the neighbors? Sustainable Development Issues in Rock Excavation Sustainability issues in rock excavation, aside from esthetics of the excavation and implications of the operation itself, deal primarily with communities around the explosives consumers and manufacturers/transporters. Communities can be exposed to noise, fly-rock, ground vibration, fumes, transportation or storage accidents, and dust. As the importance of the community’s concerns continue to rise, engineers should be more sensitive and knowledgeable of the needs of the community. Adequate information is key. In some regions, the issue of explosive theft is important as criminal elements can obtain explosives for illicit purposes. Adequate storage and compliance with all licensing and transportation safety protocols should be taken. ADDITIONAL RESOURCES XV Additional learning resources include: - Readings, Chapter 33: Land Vibration, air blasts and their control from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (optional) - Lecture Module 5.4 – Vibrations (mandatory) - Assignment 6 – in class (part of lecture) Evaluation of Blast Results Once the blast has been carried out, it is necessary to analyze the obtained results, as its interpretation will give the successive modifications of the design parameters for the following rounds. This is the basis for the optimization process. To achieve a global evaluation, the following aspects must be analyzed: - Fragmentation and swelling of the muckpile. - Geometry, height and displacement of the muck pile. - State of the remaining rock and bench floor. - Presence of boulders in the pile. - Vibrations, fly rock and airblast produced by the blast. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 147 Fragmentation and Swelling of the Muckpile Apart from the classification of size distribution or screening of the muckpile in treatment plants, there is no method which enables a quantitative evaluation of fragmentation in conditions that would be trustworthy. Size distribution is the basic tool within the optimization process of blasting, as it is the only means of comparing the fragmentation obtained when a study is to be done on the sensitivity of the design parameters. Due to high costs and necessary time to achieve the complete size distribution curve, in mining operations the following approximate methods are used: - Qualitative visual analysis - Photographic methods - Photogrammetry methods - High-speed photography - Study of loading equipment productivity - Volume of material that requires secondary blasting - Bridging delays at the crusher - Partial screening - Image analysis by computer. Qualitative visual analysis This is the most widely used and usually the only analysis applied. The muckpile and the general aspect of the blast is observed immediately afterwards and the engineer responsible for the evaluation makes a subjective assessment. However, changes in fragmentation can only be distinguished when the differences are great, even if the person in charge has vast experience. The application of this technique is not very precise; it does not allow an exact distribution of the sizes and very frequently there is no written report on the results. In general, it is only good for an initial contact with the blast results so that the specialists can later make a complete study. Photographic method This technique has been used in various ways and methods of analysis. Some researchers used photographs of the muckpile and chose at random around 15% of the total to analyze sizes and number of the fragments with a superimposed grid. After studying five blasts under similar conditions, the results showed a deviation of more or less 9.6% in the mean fragment size. The most important source of error was in assuming that the surface fragment distribution was representative of the total mass. Reid used a series of photographs of the muckpile at different moments of digging, placing scaled targets in the field of vision for dimensional purposes, as seen in Figure 30. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 148 Figure 30: Gird method of evaluating fragmentation and field environment This system is one of the most useful 'and also gives a graphic documentation for analysis and comparison of different rounds. The only inconveniences are: - Time consuming in preparation and study, and - It is difficult to quantify the small sizes. Photogrammetric method The photogrammetric methods give more precision than in conventional photographs. As the investment in equipment and accessories is relatively high, it is only used as a complement to the main applications such as topographic control of the cuts and muckpiles, geological studies, etc. The advantage of the photogrammetry is that it allows a tri-dimensional study of the muckpile, thereby aiding in the calculation of each fragment size and the volume and swelling of the pile. The drawbacks, apart from cost, are the need for qualified personnel to use and interpret the method. High-speed photography The use of high-speed photography in the evaluation, design and, above all, control of the blasting has been widely undertaken. Some have considered it as a technique for the evaluation of fragmentation. The main problem is that the gases and dust in the environment obscure the vision of crack formation and muckpile displacement. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 149 The information taken from the analysis of the high peed photographs can be classified as: - Qualitative: o First rock movements o Confinement of the stemming o Trajectory of the muckpile movement. - Quantitative: o Exit time of the blast accessories o Time and efficiency of the gas confinement o Acceleration, direction and velocity of the fragments o Velocity of stemming ejection o Projection and displacement of the muckpile. The study of these data, along with detailed information of the design parameters of the blast and of the whole of the operation, is very useful for the detection and definition of: - The existence of misfires and their causes - Incorrect explosive charges - The effect of sub-drilling, of the presence of water and of the stemming - Determination of the best initiation sequence - Yield of the chosen initiation system - Global movement of the muckpile - Source of oversize at the face - Muckpile displacement and - Profile geometry of the muckpile. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 150 Figure 31: High speed camera bench blast MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 151 Digital processing of images Computer technology has opened the door to image analysis of the evaluation of muckpile fragmentation. The modem methods of image analysis quantify geometric aspects with images in two dimensions, such as the area, number, perimeter, shape, size and orientation. The procedures take in the following stages: 1. Image input. The image is captured by a camera, usually video, and subjected to an automatic digitization process. This means the conversion of the optic image to a digital format of picture points pixels, and giving each one a certain brightness, or grey level value from 0 (black) to 255 (white). 2. Scaling. The scale of the image is defined, normally using a marker placed on the muckpile as a reference. 3. Image enhancement. This stage uses digital filters which permit an enhanced viewing of the fragments. For example, low pass or Gaussian filters which eliminate noise, a shading filter which corrects illumination defects, etc. 4. Image segmentation. At this stage, the fragments are separated from the rest of the background to produce a binary (black and white) image. A grey level is defined, then the pixels with values above this level are turned to white (fragments), and they will be taken into account whereas the pixels below this level will be darker (back-ground) and are turned to black. 5. Binary image manipulation. The segmentation process is never perfect, as the contours of some segments will bridge together and others will be confused with the background. To correct this, an iterative process of dilation, erosion and line thinning is used. 6. Measurement. This system, after identifying each object in the binary image as an independent fragment, measures the diameter of an equivalent area circle and classifies them. 7. Stereometric interpretation. In this step the distribution of two dimensional size is converted into volumetric sizes or tri-dimensional. This conversion demands the use of stereometric principles along with some empirical relationships. Studies of loading equipment productivity This technique of fragmentation evaluation is based upon the assumption that the digging rates are an inverse function of the coarseness of the muckpile and a direct function of the swelling of the same. The presence of oversize, reduced swelling and poor toe condition will be immediately reflected in productivity. If the technique is applied correctly, a precise evaluation can be obtained. The lost time that is not directly related to the condition of the muckpile, such as waiting for the trucks, mechanical breakdowns, shifting the shovel or clean-up operations, should all be taken into account. The studies should be MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 152 made with the same machines and operators to eliminate the experience factor or erroneous estimates. Studies have been undertaken to monitor the continuous current motors of the rope shovels, recording the signals on magnetic tape and processing the data on the computer. The call for extra power can be caused by the following: - Defective floor breakage with toe appearance. - Insufficient swelling of the muckpile. - Excessive displacement of the pile, and - Coarse size distribution. Boulder count and secondary breakage Any fragments of rock produced by blasting that cannot be handled by the mining equipment is referred to as boulders or oversize. The sizes of these blocks depend upon each operation and during the same should be set aside for fragmentation, with the procedures described in the following chapter. The relative volume of the oversize should be maintained at a minimum, not only because of the high fragmentation cost, but because they affect the whole of the operation by giving low digging yield due to lost time dedicated to taking them aside and also delays at the crusher. Bridging delays at the crusher The production of any crusher depends basically upon the coarseness of the material that enters, and indirectly this can be an index of fragmentation taken from the yield and the energy consumed per treated ton. The interruptions at the crusher due to oversize should be controlled, as well as the wear of the steel plates. Screening This is the only precise method of quantitative fragmentation evaluation. In small operations this technique can be used with representative samples, but in large open pit mines, it would be impractical because of the high cost and time involved. Geometry of Muckpile, Its Height and Displacement The configuration of the muckpile is governed by: - The geometric design parameters: height of bench, blasthole inclination, burden, spacing and stemming. - Powder factors. - Initiation sequences and delay timing. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 153 The optimum geometry depends upon, in each case, the digging and haulage system used. - Profile I represents the ideal situation for front end loaders, but if the available equipment are rope shovels, the yield will be low and time consuming in collecting the material and cleaning up around the shovel. - Profile 2 requires minimum cleaning labor and the productivity is high, but there can be safety problems involved for the operators due to the fall of rock from great height. o Excessive clean up area, low productivity for rope shovel, high safety conditions. o Low clean up area, high productivity, dangerous. o Low clean up area, good productivity, safe. - Profile 3 reflects optimum conditions for rope shovels. This control procedure can be done with topographic equipment, aided by transversal photographs. Figure 32: Different geometries of the muckpile Condition of the Remaining Mass Once the muck pile has been loaded, it is possible to observe if there has been overbreak, and the quantity of damage to the remaining mass. The evaluation of the damage caused by the blasts to the remaining rock can be carried out by any of the methods of geomechanic characterization of the rock mass, but for what is wanted, the system in table below is one of the most used for its simplicity and pragmatism MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 154 Table 9: Levels of damage to pit walls Analysis of the Bench Floor Once the blast has been evacuated, the bench floor can show the following: - Toe in front of holes, - Toe between holes, - High floor, - Low floor. The toe between blastholes is due to overspacing and this should be decreased. The appearance of a high floor systematically could be due to a plane of weakness or to insufficient subdrilling and bottom charge. When the floor level is lower than was designed, the bottom charge and subdrilling should be diminished and the loading operation observed. Boulders Large sized blocks can appear in the muckpile in the following areas: - On the top or crest, - On the floor (usually occurs with toe problems), - Within the pile, and - In front of the pile. The boulders on top could be caused by a hard level or poor fragmentation of the top part of the bench. This is corrected by increasing the explosive column or by placing a small pocket charge in the stemming. The boulders at floor level are due to a weakness plane in the lower part of the bench. They are eliminated by the same method as for toes: decreasing the burden and increasing the bottom charge, the subdrilling and the between-row delay. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 155 The boulders in the interior of the pile are due to incorrect drilling, a poor blast yield because of defective priming or an alteration of the explosive such as local dampening of ANFO, and to a pernicious effect of the inner joints with respect to the opening of the grid. The boulders in front of the pile could originate in an excessive breakage of the previous blast in the same zone. Vibrations and Airblast If the delay timing and the initiation sequence of a blasting are not adequate, the following results will come about: - Poor fragmentation and insufficient swelling of the muckpile - Uncontrolled fly rock - High vibration levels, and - Low frequency vibrations. By analyzing the vibrations registered with a seismograph and later modifying the indicated parameters, a better use of the explosive energy can be obtained, giving a better size distribution of the muckpile and lower vibration intensity and frequency, which are potentially less dangerous. As to airblast, there are different mechanisms that cause it, but one of the most important is a premature escaping of the gases to the atmosphere after stemming ejection. For this reason, the studies with high-speed cameras can help define the optimum stemming height and the ideal type of material so as to achieve a larger confinement of the explosive without negatively affecting the fragmentation in the zone where the inert material is placed. Blast Evaluation in Underground Mining In underground mining, characterizing the quality of the resulting excavation is a key method of evaluating the blast results, especially for development. Overbreak is commonly calculated by comparing the planned excavation with the resulting excavation. Hard toes, bootlegs, and oversize is also a common method of characterizing the quality of a blast. Dilution, a factor of both the design and execution of a blast may also be used as a sign of blast evaluation. In certain methods, such as VCR, cratering at the top of the blasthole is a common indicator of a lack of stemming. Damage to ventpipe or other utilities can also be a sign of inadequate stemming or improper blast preparation procedures. ADDITIONAL RESOURCES XVI Additional learning resources include: - Lecture Module 5.5 – Blast Results (mandatory) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 156 Module 5: Mechanical Excavation This module on mechanical excavation considers the mechanical excavation basics and theoretical aspects of bits for excavation. Mechanical excavation can be considered the removal of rock from in-situ by mechanical means. Several groups of equipment will be considered in this section including continuous mining machines, roadheadders, longwalls, and rippers. 5.1 Mechanical Excavation Basics The two main types of mechanical breakage for production purposes include indenters and drag bits (also known as picks). Both are considered to break the rock in the primary breaking process. The main difference between indenters and drag bits is that an indenter breaks rock by applying a force that is predominantly in a direction normal to the rock surface. Comparatively, a sharp drag bit applies the main force in a direction predominantly parallel to the rock surface. The breaking mechanism for both is actually a tensile fracture. Intenders are used more widely than drag bits primarily due to the shear loading of the rock by the drag bit. The cutting edge of a drag bit is subjected to bending resulting in tensile stresses along the bit edge. As most cutting tools are made of tungsten carbine, a brittle material weak in tension, this makes the possibility of catastrophic failure likely. However the tools can be mounted in a fashion so that the bit contacts with the rock in a mainly compressive fashion, as seen in the Figure 5-1. Figure 5-1: Longwall shearing machine whose cutting bits attack the rock in compression Various theoretical models have been developed that can compare drag bits with indenters. Several will be presented with an emphasis on comparison between drag bits and indenters. This will allow the two modes of the primary breakage process to be compared. Various theoretical models have been developed as presented in the SME Mining Handbook (Chapter 9.1) yet few have admittedly been proven empirically. Furthermore, results from lab-based rock cutting experiments worldwide have reported variances in cutting conditions and variables such as pick shape and rock type. However, generalities do emerge from these experiments and are regarded as fundamental features of rock pick cutting. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 157 Important variables in understanding these generalities include: F C = Mean force F N = Peak force E S = Specific Energy (relates to the cutting force to amount of rock produced Various picks have been tested, such as; chisel, round bottom, v-front, v- bottom, and pointed. However, these many picks can be categorized into three main groups (as seen in Figure 5-2: Pick Shapes); - Pointed tool. - Simple chisel, - Round bottomed tool, Figure 5-2: Pick Shapes - Pointed tool, simple chisel, and pointed tool. (| = Back clearance angle, o = rake angle) Both F C and F N increase with d (depth of cut) for all pick shapes. From Figure 5-3, it can be seen that this relationship is approximately linear. (not linear when using wide picks at depths much less than their width but machines such as these do not exist.) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 158 Figure 5-3: Mean force vs. depth of cut (F C vs. d) From Figure 5-4, it can be seen that Specific energy varies inversely with depth of cut for all pick shapes, therefore: d E S constant = Figure 5-4: Specific energy vs. depth of cut (E S vs. d) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 159 Cutting and normal forces decrease monotonically 9 with increasing rake angle as seen in Figure 5-5. Most of the benefit to pick forces has been achieved at a rake angle of 20°, beyond which further marginal improvement is at an increasing penalty to pick strength and its potential to survive. Figure 5-5: Mean force vs. rake angle (F C vs. |) Figure 5-6 shows that increasing back clearance angle reduces pick forces up to about 5°, beyond which forces are independent of this angle. Figure 5-6: Mean force vs. back clearance angle (F C vs. |) 9 monotonic defined for mathematics: Designating sequences, the successive members of which either consistently increase or decrease but do not oscillate in relative value. Each member of a monotone increasing sequence is greater than or equal to the preceding member; each member of a monotone decreasing sequence is less than or equal to the preceding member. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 160 Higher productivity will be achieved by cutting slowly and increasing depth than when cutting at a higher speed yet more shallow cut since: - Cutting speeds of up to 5 m/s and beyond have no discernible affect on pick forces or specific energy. - Hence speed has no direct effect on pick forces - Cutting efficiency improves with cutting depth For a chisel shaped pick, F C and F N increase linearly with pick width as seen in Figure 5-7. Figure 5-7: Mean force vs. pick width (F C vs. w) Pick forces increase with rock strength, and it has yet to be determined what characteristic of rock strength is the predominant influence. Figure 5-8 compares the mean cutting force with both the tensile and compressive strength of various rocks. A further issue related to cutting forces is the degree of rock saturation as increased saturation results in a decreasing mechanical strength of the rock (the reason for this is not clearly understood). MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 161 Figure 5-8: Compressive and tensile strengths upon varying mean cutting force (FC vs. o c and FC vs. o t ) In terms of pick shape, when operating at the same rake and clearance angles and depth of cut, the pointed pick requires the least cutting and normal force. The chisel pick requires the greatest forces. However, considering that the chisel pick cuts a considerably larger volume of rock than the other two shapes, it cuts with the lowest specific energy and therefore is the most efficient shape (pointed pick least efficient). Due to the increased penetrating capability of the pointed pick, for a given available normal force, pointed picks operate more efficiently than the chisel bit. (Hence, pointed picks can but deeper for a given level of force, whereas chisel bit cut more material for a given depth of penetration. Operationally, picks are better.) Figure 5-9: Chisel vs. pick bits. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 162 5.1.1 Synopsis Drag bits are an efficient means of breaking rock. Although most are unable to survive in the hard-rock environment, they require substantially less energy than other mechanical cutters. Two main problems remain to be solved: - Drag bit materials that have far improved wear resistance and strength - Machines to provide sufficient thrust force to keep drag bits at an effective depth of cut, especially as bit wear proceeds. ADDITIONAL RESOURCES XVII The above information is additionally supplemented by: - Readings – Chapter 9.1 – Mining Engineering Handbook. (Mandatory, but this is for modules 1 and 2) - Lecture module 2.1. (mandatory) - Readings – Speight, H. E., “Obersvations on drag tool excavation and the consequent performance of roadheaders in strong rock” AusIMM Proceedings, No. 1, 1997, pp.17-37 (mandatory). 5.2 Longwall These lecture notes review the design and operational aspect of the longwall mining method. Particular focus will be on the cutting mechanism and design. Important issues related to longwall, such as conveyor design, ventilation and roof support, are left to materials handling and rock mechanics courses. Longwall mining has one of the simplest layouts, provides continuous production and can provide a safer work environment. The panel layout is conducive to good ventilation. Therefore this method is considered to be better than room and pillar methods used to mine similar coal deposits. Since this system uses full caving, fewer residual pillars remain hence coal recovery is higher and surface subsidence is relatively uniform and complete. Longwalls are used to mine flat horizontal coal seams of thicknesses ranging from 1.1 to 4 meters. The depth of overburden can range from anywhere between 60 to 820 meters. Figure 5-10 shows a typical plan view of a longwall layout and the nomenclature of the key design variables. Longwall panels are flanked by panel entries on both sides of the main entries, typically excavated by a continuous mining machine. The immediate entries on both sides of the panel are called the head entry and tail entry. The head entry is used for air intake and the transportation of coal on conveyors, personnel, and supplies. The tail entry is used as the air exhaust. Panel sizes and length are generally determined by: - Experience - size and shape of the coal seam - geologic conditions MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 163 - location of surface structures (buildings can subside horizontally when in the middle of a panel but would subside differentially if it straddles a pillar), - capacities of the transportation system - ventilation - power capacity (of equipment) Panel widths vary between 120 to 293 meters (centre to centre) and have lengths between 1220 to 1830 meters. The economic factors pushing for longer and wider longwalls include: - Reduction of the development cost as fewer panel entries would be required - Increases in recovery and recovery rate (few longwall assemblies, down- time between panels) Other issues that begin to surface in longwalls include: - Downtime (maintenance & repair) in large panels - Capital cost of equipment in a single panel Figure 5-10: Typical Longwall retreat method MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 164 5.2.1 Mining System Description Figure 5-11 shows a diagram of a coal operation. Coal at the face is cut by a shearer or plough and loaded onto an armoured flexible conveyor (AFC) and conveyed to the head entry T- junction. The coal is loaded onto a stage loader which empties onto the entry belt conveyor. Powered sectional supports are used to support the roof along the whole face. The AFC and powered supports are advanced hydraulically after each cutting cycle of the shear and the roof behind the supports are allowed to cave. The area between the rear edge of the supports and faceline is called the face area while behind the supports is called the gob. Panel entries are where the regular conveyors are maintained by roof bolting or other methods. However, at the tail entry where ventilation circuits are set up, cribs are usually required to strengthen the support. Figure 5-11: Longwall shearing system Technical factors of longwall mining technology include: - The longwall width is limited to the power and structural strength of the face conveyor. - Increasing panel width increases roof exposure time (when the shearer or plow moves across the face, an area of roof is opened. This small area may cave prior to the supports moving forward to secure the back). Now that the basics of longwall mining have been established, a focus on the actual rock excavation aspects will be explored in more detail. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 165 5.2.2 Selection of Cutting Machine The two main types of longwall machine are the shear loader and the plough. Shearers are used more widely in the US. Ploughs are simply a large blade that is thrust into the rock and dragged across the face resulting in the removal of a slice of coal. Key variables in the selection of the type of cutting machine are: - Mining height - Seam structure - Roof bonding strength of coal - Cutting/ploughing resistance of coal. Since shearers are the most common machine, only shearers are discussed. Major parameters in shear design are: - Type of shearer, - Dimensions, - Haulage speed, - Power capacity. There are several types of shearers: - Double-ended ranging-drum (DERD): two shearing drums mounted on cutter, articulated so that coal seams of varying height can be mined. Can mine thicknesses of 1.47 – 3.96 meters. Is by far the most employed shearer. - Single-ended ranging-drum (SERD) for thicknesses of 1.52 meters. - Single-ended fixed drum (SEFD) for thicknesses of 1.26 to 1.37 meters 5.2.3 Dimensions of Shearer Equipment suppliers provide various models with varying dimensions and power available. The key variables in selecting the shearer can be seen in Figure 5-12: Figure 5-12: Shearer dimensions Where: 2 sin 2 D L B H H a b c + + ÷ = o And where: H c - mining height D - diameter of the drum MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 166 H b – height of ranging arm La – length of ranging arm B - body depth o– angle of the ranging arm The maximum cutting height should not exceed 2 D. A nomograph for determining shearer dimensions and mining height is shown in Figure 5-13. Note that the conversion factor is 1 inch = 25.4 mm. Figure 5-13: Nomograph for shearer dimensioning MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 167 Example 1: Shearer dimensioning Use the nomograph in Figure 5-13 to select the drum diameter D, given the mining height and major dimension of a shearer: where: H c =90 in. (2.29m), H b =50 in (1270mm), La =70in (1778mm), B = 20 in. (508mm) and o=30˚ The dotted line on the nomograph can be followed from point a through to f in Figure 5-13. This results in a value of D being 30 inches (762 mm). In practice, in order to reduce the loading resistance of the drum and to increase loading efficiency, the drum diameter is selected larger than the minimum value, typically 75-80% of the mining height, in this example, D would therefore be equal to 60 inches. ADDITIONAL RESOURCES XVIII The above information is additionally supplemented by: - Lecture for module 2.2 – Longwall (mandatory) - Readings – Chapter 20.1 – Longwall Mining, from SME handbook (optional) 5.3 Continuous Mining Machines 10 Module 5.3 will cover underground continuous mechanical mining technology and, continuous surface mining equipment. These notes were assembled directly from the following references: - Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2 nd ed. 1992 - Hartman, Howard L. and Jan M. Mutmansky, Introductory Mining Engineering, 2 nd ed. New York: John Wiley & Sons. 2002, 570p. - Atlas Copco, The Raise Boring Handbook 2 nd ed. - Stefanko, Robert. Coal Mining Technology Theory and Practice. Littleton CO.:Society of Mining Engineers. 1983 - McKercher, R.M., Ed. Potash Technology: Mining, Processing, Maintenance, Transportation, Occupational Health and Safety, Environment. Pergamon Press: Toronto. 1983 - Mining-Technology.com, search: continuous mining - Wirtgen America Inc. Surface Mining Manual. Edition 2002 10 These notes were assembled directly from the following references: - Hartman, Howard L. Ed. SME Mining Engineering Handbook. 2 nd ed. 1992 - Hartman, Howard L. and Jan M. Mutmansky, Introductory Mining Engineering, 2 nd ed. New York: John Wiley & Sons. 2002, 570p. - Atlas Copco, The Raise Boring Handbook 2 nd ed. - Stefanko, Robert. Coal Mining Technology Theory and Practice. Littleton CO.:Society of Mining Engineers. 1983 - McKercher, R.M., Ed. Potash Technology: Mining, Processing, Maintenance, Transportation, Occupational Health and Safety, Environment. Pergamon Press: Toronto. 1983 - Mining-Technology.com, search: continuous mining - Wirtgen America Inc. Surface Mining Manual. Edition 2002 MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 168 5.3.1 Underground continuous mining machines Underground continuous mining machines typically operate in room and pillar mines in coal, salt, Trona, or potash (essentially soft material in horizontal orebodies). The machines are an alternative to drilling and blasting operations. Underground blasting however, is undesirable in coal operations due to the risk of secondary explosions caused by suspended coal dust or methane gas. Other advantages of continuous mining is combining the operations of cutting, drilling, blasting, and loading into one operation by a single machine. The continuous nature of this mining method eliminates the problems involved in the cycling of equipment. As the number of equipment is reduced, so are manpower requirements. However, these advantages are offset in some part by concentrating the reliance of production on a few headings. Maintenance issues therefore become more critical. Continuous mining equipment has evolved significantly over the last 100 years. While most of this equipment attack the solid face, ripping out and loading the coal in one step, sometimes using very different processes. Only those machines still in use today are reviewed. Figure 5-14: Four-Rotor Boring machine 5.3.1.1 Boring Machines Figure 5-14 shows a boring machine. These machines have rotating arms equipped with bits that bore out the coal. The boring machine has a cutting chain on the bottom that creates a flat working bottom and an upper trimming chain, which when working together, produce a roof configuration that is arched or ovaloidal opening. The arms revolve relatively slowly, produce a much coarser product than that of a ripper miner, minimizing gas and dust problems. Since the machine uses its tracks to provide the thrust, it cannot be used to bolt simultaneously as advancing. The machine can have drills mounted on a slide carriages and have met with success at bolting and mining simultaneously. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 169 The main disadvantage of this type of machine is its size. It typically occupies the entire stope. The tips of the arms can be extended or contracted hydraulically over certain limits to provide some machine clearance during tramming, however, the machine must move considerably slowly. Ventilation is also an issue when dealing with its size; therefore these machines are typically used in potash mines, not coal mines. Further discussions on potash mining will be discussed later. Figure 5-15: Milling head continuous mining machine 5.3.1.2 Milling head miner Milling head miners are the most popular continuous miner in operation in US coal mines. The machine is a ripper with disks or wheels located on arms. The disks rotate in a vertical plane. To mine the coal between the disks, the arms to which the disks are attached oscillate or are provided with a side splay or horizontal movement. While the machine can be equipped with a full-face head about 4.6 m (15 ft) wide, usually the width of the head is less so that two passes are required to mine the width of a room or entry. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 170 Figure 5-16: Miller-head continuous miner. The usual mode of operation is to push the machine into the face on its crawler while the head is in an elevated position and the disks are rotating. The milling machine sumps (pushes) into the top and shears downward. Loading arms located under the head load the coal. While the basic machine had an oscillating feature when it was introduced, most miners use the same bit attack with a rigid head and the machine is usually referred to as the Hardhead miner. In this type, bits are spirally wound on a fixed shaft so that each bit cuts an infinite number of vertical planes during a single revolution, accomplishing the same purpose without oscillating and with subsequently fewer maintenance problems. This machine is cheaper as well as requiring less maintenance, it is, therefore, quite popular today. In addition, its head can be tapered to provide an arched opening where desired, although the optimum shape with regard to ground control should be heeded. The usual procedure is to mine an opening cut the width of the machine [usually 3 m (10 ft)] then mine the other side of the face. Since this machine cuts by advancing the crawler to the face, drills for roof-bolting would have to be attached to the chassis in a flexible manner, negating its widespread use for concurrent bolting. However, some equipment allows both bolting and mining simultaneously. 5.3.1.3 Boom-Type Miner A boom-type miner or road-header consists of a rotary cutter-head mounted on an articulated boom, a crawler assembly, and a gathering arm system. The design of the machine permits a concentration of power at a single cutting bit at a time, enabling it to cut harder rock than drum-type continuous miners can excavate. While it is used for mining coal and other minerals, its primary use is for driving mine entries and tunnels in civil engineering projects. Roadheaders were introduced in the post World War II era, and at present there are over 2800 employed worldwide, with the majority outside the US. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 171 However, about 140 road-headers are in use in North America with approximately 75% of the machines working in mines. Roadheaders can excavate any size and shape of cross section: circular, horseshoe, arched, and rectangular. Unlike conventional drilling and blasting, the smooth cutting action of the cutterhead does not weaken the rock due to overbreak. Thus lighter support may be used or less concrete, aiding cost. The machine is being used to enlarge mine entries in a bituminous mine in central Pennsylvania by removing bottom rock. The wide flexibility in cutting at various angles also increases the utility of the roadheader when compared to the continuous miner, as seen in Figure 5-17. Figure 5-17: Mining of Steep seams with roadheadder and miller head (drum-type) miners MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 172 Figure 5-18: Mining in variable height seams Figure 5-19: Double head Roadheadder MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 173 Figure 5-20: MD1100 with traverse boom 5.3.1.4 Performance Analysis Predicting the cutting rate of these machines requires information on: - rock properties, both in terms of the rock and rock mass; - machine properties, including rock-tool interaction models. A few key measures include: - ICR: instantaneous cutting rate - OCR: operational cutting rate (includes utilization) Typically, the OCR is 0.45–0.60 x ICR, while 0.30(ICR) when trimming Utilization is affected by: - Support installation - Surveying - Pick replacement, - Maintenance/repairs - Haulage delays - Shift changes In the calculation of utilization, boom repositioning and final profiling are not included. Advance rate is calculated by: rate n Penetratio n utilizatio area face OCR rate Advance × × = Be warned: theoretical cutting models for roadheaders vary in applicability for machine types, and are not good. TBM models work better, however, TBM performance remains to be calculated empirically, as before: SE=specific energy HP=headpower MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 174 Where: SE = HP / ICR Most predictions use RMR, RQD and UCS. The rock cuttability index, in kg/cm 2 = 100 3 2 RQD c o ; can be used in conjunction with Figure 5-21. Figure 5-21: Rock Mass Cuttability Index vs. cutting rate Note that Figure 5-21 can also be used for predicting road-header productivity as a function of rock mass properties. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 175 Figure 5-22: Roadheader performance vs. rock class. 5.3.1.5 Tunnel Boring Machine This type of machine can be used to drive circular tunnels from 5.7 ft (1.75 m) to more than 36 ft (11 m) in diameter in rock types that range from weak, loosely consolidated to very strong and abrasive. In almost all cases breakage is effectuated by roller cutters mounted on the cutting head, as seen in Figure 5-23. Because these cutters break the rock by indentation, these machines are characterized by very high thrust requirements. This thrust is provided by hydraulic rams that press the cutterhead into the rock face. The thrust reaction force is reacted through gripper pads that are pressed hydraulically against the tunnel walls. The rock broken from the face by the cutters falls to the floor where it is scooped into buckets mounted around the gage of the cutterhead. This debris is lifted in the buckets to the tunnel crown, whereupon it is tipped onto a belt conveyor that runs through the center of the machine. Figure 5-23: Tunnel boring machine MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 176 Figure 5-24: Components of a TBM Figure 5-25: Disc cutters on TBM The most common type of cutting tool employed on these machines is the disk cutter. In some cases the cutting edge of this tool is a hardened steel surface and in other cases it is a row of cemented tungsten carbide buttons that are press-fitted into the disk rim. The most common cross section for a MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 177 hardened steel disk cutter is seen in Figure 5-26. The advantage of this design is that the tool area presented to the face is maintained resulting in relatively constant penetration rate. As the cutters become blunted, it is necessary to increase both the machine thrust force and torque. Other cutter types include the kerf cutter, which is simply multiple disks mounted on the same hub, and the pineapple cutter, which is a frustrum with cemented tungsten carbide buttons press-fitted onto the surface. Figure 5-26: cross section of disc cutter Figure 5-27: Kerf and Pinapple cutters A ranking of cutting efficiency of these tool types, in terms of specific energy, places the steel disk cutter as the most efficient, the disk-button cutter next, and the pineapple cutter as least efficient. However, the wear resistance, and therefore the capability of cutting strong abrasive formations. is the reverse of this efficiency ranking. Consequently, steel disks tend to be used for cutting weaker, less abrasive rocks, and pineapple cutters are used for machining the most abrasive and toughest formations. Several other comments can be made about the cutting behavior of disk cutters. First, in contrast with drag bits, the efficiency of the rock breakage process does not decrease when disk cutters are used in a groove deepening (when multiple passes are made by a tool taking a series of shallow cuts in a kerf before producing major rock chips) mode. This is fortunate because in practice, groove deepening is the cutting procedure most commonly employed with tunneling and with other boring machines. Second, similar to the findings for drag bits, an optimum spacing exists between an array of disk cutters working a rock face. The value of this optimum spacing depends on the depth of cut taken and on the rock type. However, whereas with drag bits an optimum s/d value of 2 to 3 is typical (see Figure 5-28), with disk cutters this value is more typically in the range 5 to 10. Third, the efficiency of the rock MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 178 breakage process is independent of whether the grooves are cut simultaneously, with multiple disks on a single hub, or sequentially, with independent disks. Figure 5-28: spacing versus depth relationship Skidding has a major influence on cutter longevity (more will reduce longevity). This dragging or ploughing action is particularly acute for rollers as the center of the cutting head. TBMs are widely used in civil applications. The primary advantage is that this machine is quasi-continuous, compared with the drill-blast-muck cyclical method. Therefore tunnel rates are somewhat more constant and higher where TBMs are used. Roughly, the speed of advance for a drill-blast-muck crew would be 10 feet per shift (ideal), whereas on a TBM, the advance rate would be about 33 feet per shift. A key environmental aspect is the lack of ground vibrations when compared to blasting. A further considerable advantage is the lack of secondary damage caused by the machine when compared to the damage cause by blasting. The disadvantages are: - High capital costs - Tunnel cross section is circular - Large turning radius (100 meters) - Cumbersome machine and system installation. Considering these disadvantages, a rule of thumb is that a tunnel of at least 2 km in length to justify the cost of a machine setup. The lack of flexibility, size, and cost of the equipment has made TBM not fully accepted in mining. Some attempts have been made to develop hardrock continuous mining machines that are more flexible, however, the bit wear costs remain prohibitive. A prototype oscillating disc cutter appears to show promise but has yet to be fully developed into a working production unit. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 179 The TBM is considered to be the primary equipment in the field of rapid excavation. Progress in rapid excavation can be measured by several performance parameters: (1) hardness of the rock, (2) time percentage of machine availability, (3) diversity of application, (4) rate of advance, (5) specific excavation rate, and (6) cost of advance. Several of these parameters deserve to be discussed further here. First, hardness of rock is an area in which some progress is being made. Tunneling through rocks up to 30,000 psi (207 MPa) has become fairly common; however, some tunnels have been driven through rock approaching 50,000 psi (350 MPa) in compressive strength TBM specific excavation rates have also gone up over the period of rapid excavation usage. Case studies dating back to 1980 have resulted in empirical tables of rock types with corresponding excavation rate. Note that the metamorphic, granitic, and volcanic rocks have TBM specific excavation rates of 8 to 360 X 10 -6 ft 3 of rock extracted per lb of force per cutter per revolution (0.00005 to 0.0023 m 3 of rock extracted per kN per cutter per revolution) and the rates in sedimentary rocks are 8 to 720 X 10 -6 ft 3 of rock per lb of force per cutter per revolution (0.00005 to 0.0046 m 3 of rock per kN per cutter per revolution). Figure 5-29: Empirically derived excavation rates of TBMs by diameter for igneous and metamorphic rock types MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 180 Figure 5-30: Empirically deriven excavation rates of TBMs by diameter for sedimentary rock types Although improvements have been made in these parameters, machine availability is still a problem area. Studies have shown that TBM availability is 35 to 50%. There are many good reasons that the TBM is not engaged in the excavation process a greater percentage of the time. Maintenance, backup equipment, ground control, cutter replacement, and other delays all contribute to the problem. One of the major deterrents to applying rapid excavation in underground mining is the rather massive dimensions of the typical TBM with its trailing gear. This is evident in the application of a TBM to development in an underground mine (Stillwater). In this project the mining company had to work with a TBM manufacturer to reduce the turning radius of its TBM to about 200 ft (60 m) in order to effectively use the device in mine development. The original turning radius of the machine (350 ft or 106 m) was simply too great to be able to effectively maneuver the machine underground. Another important limitation of the TBM in mining projects is the economics of conventional versus rapid excavation development, see Figure 5-31). For tunnels of less than about 22,000 ft (6 km), the TBM normally cannot provide a cost that is lower than the costs of conventional development practices. This eliminates its use for many mine development projects. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 181 Figure 5-31Cost comparison between tunnel boring and conventional drill and blast. Although TBMs have been used for development at the Stillwater Complex in Montana and at the Ray Mine in Arizona, the rapid excavation revolution has still not established this process as commonplace in providing horizontal development openings in mines. In the advance of vertical openings, rapid excavation has greatly improved mine development and civil works, particularly in the area of raise boring. In many mines, raise borers for development of stoping operations are routinely used and have replaced conventional development in all but unusual circumstances. 5.3.1.6 Raise / Blindhole / Shaft Boring Machines These machines use the same button roller cutter technology as TBMs, but they drive vertical or steeply inclined holes rather than tunnels. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 182 Figure 5-32: Steps in raisebore mining. Figure 5-33: Raiseboring in operation top & bottom Raise Borers are machines used to produce a circular excavation either between two existing levels in an underground mine or between the surface and an existing level. The process of raise boring is shown in Figure 5-32. First, the boring machine is set up on the upper level, and a small-diameter (of the order of 9 in. or 230 mm) pilot hole is drilled, usually with a tricone bit, down to the lower level. When this hole is completed, the drillbit is removed, at the lower level, and replaced by a reamer head having a diameter with the same dimension as the desired excavation. Some type of roller cutters are mounted on the reamer head. This head then is rotated and pulled back up towards the machine. The rock debris falls by gravity into the lower excavation where it is removed. These machines are very effective in driving raises, and they have become very popular, particularly in hard-rock underground mines. Frequently the direct costs of driving a raise, in terms of dollars per foot (dollars per meter), are reduced by using these machines. In addition, however, this raising system offers other significant advantages, such as: - safety--conventional drill-and-blast raising is notoriously dangerous. - improved excavation rates - improved productivity. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 183 - circular shape combined with the lack of blasting damage results in an excavation of greater strength and integrity than a hand-driven raise. - Reduced friction for ventilation raises. Applications of raise boring are in both civil and mining environments, specific applications are summarized in Table 5-1. Table 5-1: Applications for boring Mining Civil 1. Material Transfers (ore or rock passes) 2. Ventilation 3. Personnel access 4. Ore production 5. Slots for longhole 1. Penstocks and surcharge chambers in hydroelectric projects 2. Redirection and retrieval of water in hydroelectric projects 3. Storage of petroleum, pressurized gas, and nuclear waste 4. Road and rail tunnel ventilation 5. Storm water storage and drainage 6. Equipment access (pipes, hoses, cables) 7. Water inlets and outlets for fish farms. Figure 5-34: Raise boring diagram – Up and downward boring. Blindhole Borers are machines that will produce a circular excavation without the need for a pilot hole. The machine is set up on a level in an underground mine, and a steeply inclined hole, generally several feet (meters) in diameter, is excavated upward from this level. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 184 Figure 5-35: Blind boring operation Figure 5-36: Blind boring bit head. Boring Site Preparation. Correct site preparation eliminates major delays and adds noticeable efficiencies to the raise boring operation. Raise boring site preparation begins with a comprehensive plan of the site layout. Site planning considerations must include: - Derrick mounting systems - Selection of bailing fluid - Bailing fluid and cuttings discharge - Storage and positioning of drill string components - Overhead clearances - Floor space and equipment positioning - Compressed Air - Water - Electric power - Lighting MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 185 - Communications - Ventilation. Horizontal boring Use of horizontal raise boring in civil engineering projects is becoming increasingly popular as compared to alternatives since: - no trench is to be dug - no vibration or environmental side-effects of drilling and blasting - less set-up time and cheaper than TBMs. Combined with directional drilling (undertaken when drilling the pilot hole), a directional tunnel can be built. Figure 5-37: Directional boring in urban environment. Figure 5-38: horizontal borehead 5.3.1.7 Underground mining case studies The first case study of an underground continuous mining system is exploration of the Boulby potash mine in the UK. Note that the shaft had to be excavated by freezing the surrounding rock mass so that a minimum of water inflow would occur. A special impermeable lining was then constructed MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 186 long the shaft in the section under the aquifer. This can be seen in Figure 5-39. Figure 5-39: Cross section of the Boubly potash mine strata To cope with greater stresses in the ore below 1,200m and to raise output above 3Mt/y by 2001, CPL converted its continuous miners to remote- controlled operation (so they can be safely used in shaly sections without prior blasting). The mine is currently establishing a mine-wide digital micro- seismic monitoring system and has devised a two-road with stubs (rather than four-road) stress-relieving room-and-pillar technique which achieves more mining per shift. Jeffrey and Joy continuous miners discharge to Joy electric shuttle cars, which run to feeder breakers on the main conveyors to the hoisting shaft. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 187 Figure 5-40: remote operation allows distance mining and the operator to view from different angles. Case Study: Rocanville Saskatchewan Rocanville uses a long room-and-pillar method of mining. Ore is mined from rooms in three passes, separated by pillars supporting the overlying strata. Five automated Marietta two- and four-rotor continuous miners, each capable of extracting 650t/h of ore, form the production fleet. The run-of-mine ore is loaded on to extensible conveyors attached to the continuous miners. These connect to the main haulage conveyors, which move the ore to skip-loading pockets at the shafts, where it is hoisted to surface. In addition to automating its Marietta miners to measure ore grades and mine selectively, the company installed a central control room at Rocanville from which the entire mine, surface and underground, is controlled. The central control system, supplied by Allen-Bradley, oversees all aspects of the mine and mill operation, and has maximized ore extraction and recovery since it was commissioned. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 188 Figure 5-41: Marrieta miner in Rocanville Figure 5-42: Marrieta digging a trench, being maintained Several types of cuts are made in mining potash. These can be seen in Figure 5-43, which shows various boring type cuts. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 189 Figure 5-43: Cross sectional view of four-rotor boring in potash 5.3.2 Continuous Surface Mining Technology 5.3.2.1 Surface Miners Surface Miner combines the individual processes normally associated with traditional technology, such as drilling, blasting or tearing, extraction, transportation, crushing and loading. In particular, the fact that blasting is no longer required leads to a considerable reduction in environmental pollution from, gasses, dust, noise and flying rocks. Selective mining of thin or sloping layers results in a clear separation of material (overburden, minerals etc.). Eliminating dilution increases the efficiency of transportation and further processing. Surface Miners have reached a rated capacity of 1,250 bm³/h when engaged in lignite mining. This means that layers of up to approximately 1.2 m thickness can be mined in one step. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 190 Figure 5-44: Various aspects of a Surface miner Surface Miners are capable of continuously mining stone with 50 MPa or 100 MPa, irrespective of discontinuities and without blasting. This includes: - lignite - coal - anthracite - bauxite - phosphorite - marl - limestone - hard sediment stone (sandstone, - certain conglomerates, breccien - clays, loam MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 191 Figure 5-45: Surface miner ideal for thin strata Figure 5-46: Cutting capacities for various Wirtgenamerica products Figure 5-47: Surface miner ideal for defined thin vertical or flat seams Surface Miners are suitable for: - Mining for the extraction of minerals and overburden - Road construction for preparing roads, motorways and waterways - Earthworks for preparing construction sites etc. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 192 Figure 5-48: components of Surface miner 5.3.2.2 Bucket wheel excavators Originally designed for relatively easy digging materials (gravel, sand, loam, marl, clays, and lignites), bucket wheel excavators (BWEs) can now dig in relatively hard material. These have included compact sediments such as shales, black coal, some limestones, and tar sands. One of the most successful applications of BWEs has been in German and Australian lignite mines. The machine digs out the material using a large wheel with buckets that revolve as the wheel turns. The teeth on the individual buckets are the primary ground engaging tools that break out the material from the ground. BWEs are most often attached to a conveyor network where waste material is sent to a spreader or to an ore stockpile. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 193 Figure 5-49: Bucketwheel nomenclature Figure 5-50: Bucket cutting and discharge BWE considerations: - Hard consolidated materials, large boulders or blocky material cannot be handled - Sticky material build up in buckets and can gum up the conveyor (although, with modern systems, sticky material can be handled); MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 194 - Abrasive material produces excessive wear on the teeth (some modern tooth design can significantly increase tooth life). - The digging face should be stable - Actual output is usually only 45-110% of theoretical - Very limited flexibility (can be flexible in certain geological and equipment situations) - High capital costs but may be the most economical method of mining weak flat tabular deposits. - Some advantages in direct reclamation (environmental benefit); - Relatively little manpower - Must closely match downstream equipment (as in all mining operations) - Low total costs - Long life - There must be a matched system linking BWE, side-slewage conveyor belt system and spreaders Advantages of Continuous Excavators - Have lower impact loading than comparable single-bucket machines - Tends to reduce dynamic stresses, machine mass, maintenance costs, power consumption - Reduced slewing (swing) speed, reduced digging impacts, reduction in ground bearing pressure - BWE can mine both thin overburden and deeper overburden - Conditions possible where single bucket (dragline) is impossible Theoretical output of machine is: Factor Swell Fs Q th 60 = where Q is theoretical output in yd3/hr (or m3/hr) bank, F is capacity of single bucket S is number of bucket discharges per minute Swell factor is that of the material being excavated. The theoretical output of BWEs range between 200 to 20,000 m 3 /hr. Largest disadvantage is their size, which limits flexibility. Therefore the mine must have strict discipline in terms of the mine plan, such as: - Sufficient reserves to justify the capital cost - Relatively horizontal stratification over a wide area of the deposit - Uniform geologic conditions (i.e. absence of major faulting, severe undulations, large variations in overburden thickness Specific cutting force The digging resistance of the material being excavated is limited by: - The bucket wheel drive power - The mechanical strength of the bucket - Machine mass (service weight) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 195 Despite empirical research, no direct relationship mathematical model for prediciting specific cutting force between the key variables has been established. These key variables include: - Digging resistance - Intact rock strength - jointing - Bedding - Tooth shape and sharpness - Angle of attack - Wedge angle of the bucket cutting lip Hard Ground operation is possible with large, modern BWEs that have high wheel rotation speeds with slower slewing speeds. Material that cannot be dug out with a cable shovel such as hard clays, phosphates, sandstones, and frozen Tar Sands, can be excavated economically with BWEs. Provided that rock does not contain large hard boulders, and that the ground’s strength ranges between 15-18 MPa. Case Study: Big Brown Mine, Texas General - The Big Brown Mine, located near Fairfield in Freestone County, is composed of two active mining areas that supply fuel to the Big Brown Steam Electric Station. - Big Brown Mine fuel is used to generate approximately 7% of TU Electric's total electricity generation. - TXU Mining receives about 17% of its total lignite production from the Big Brown Mine. - About 12,000 acres have been mined at since mining began in 1971. - The Mine areas have a combined total pit length of approximately 7 miles. - Approximately 72 million cubic yards of overburden are moved per year. The total amount of overburden have been excavated since mining operations began is approximately 1.4 billion cubic yards. BWE Specifications: Height.................................41 feet Width..................................33 feet Length................................144 feet Weight................................374 tons Bearing Pressure .............14.5 P.S.I.* Wheel Diameter.................26 feet Wheel Speed.....................5.125 R.P.M.*** Number of Buckets...........12 Bucket Size.......................1.08 yd³ Belts..................................2 (1 receiving, and 1 discharge) Belt Width..........................60 inches Belt Speed.........................885 F.P.M.** Boom Length.....................49 feet (receiving) 82 (discharge) Crawler Size......................8 feet (width) 29 feet (length) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 196 Design Digging Height......49feet Design Digging Width........60 feet BWE will dig 3 feet below level of tracks Ground Speed.................. 32.8 F.P.M.** Personnel Required......... 1 operator and 1 oiler ADDITIONAL RESOURCES XIX The above information is additionally supplemented by: - Assignment 2 - Lecture for module 2.3 - Reading – Chapter 13.5 – Continuous Excavators, SME Handbook, (Optional) 5.4 Dozers and Ripping 11 Dozers can be utilized for ground preparation when using a ripper, which is mounted to the rear of the tractor. Rippers are used widely in civil engineering projects to excavate slopes and wide swaths of surface area. Rippers are used in quarries to excavate sand and gravel deposits. Rippers are also used to some extent in mining for surface preparation and in coal production. In the ripping process, the ripper shank is pulled through the soil or rock to loosen or fragment the material, which can then be loaded by a scraper or front-end loader, or handled by the dozer blade. Ripping is an inexpensive method of breaking discontinuous ground or soft rock masses. Figure 5-51 shows a ripper in action (back of dozer). In this operation, the dozer excavates the exposed coal and pushes the loosened material into windrows. Front-end loaders then load the material into haul trucks. Rippability is determined by: • Compressive strength • Bedding planes, joints and fractures; • Brittleness; • Softness (from weathering). There are three ripper types, as seen in Figure 2 through Figure 4 • Radial; • Parallelogram; • Adjustable parallelogram. 11 This lecture and its notes were complied using sources from: - Bell, F. G. ed. Engineering in Rock Masses. Butterworth-Heinemann: London. 1992. ISBN: 0 7506 1063 8 - Caterpillar Performance Handbook, 28th Ed. - Caterpillar Ripping Handbook, 12th Ed. - 2001 Class notes, from Bob Cummings - 1997 Class notes, from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department. Students are welcome to view these sources in my office. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 197 Figure 5-51: Ripper in Action (a) (b) (c) Figure 5-52: (a) Radial ripper, (b) Parallelogram ripper, (c) Adjustable Parallelogram ripper In a Hinge-type ripper, the linkage carrying the beam and shank pivots about a fixed point at the rear of the tractor. As the shank enters the ground and penetrates to maximum depth, the tooth angle is constantly changing. Hinge-type rippers offer the advantage of an aggressive entry angle, but cannot be adjusted to compensate for varying conditions. A Parallelogram- type ripper allows the linkage carrying the beam and shank to maintain an essentially constant tip-ground angle regardless of tooth depth. This type of ripper has advantages over the hinge-type when ripping above maximum depth, but does not provide the aggressive tooth angle necessary for hard- to-penetrate materials. The Adjustable Parallelogram ripper combines the features of both the hinge- type and parallelogram rippers. It can vary the tip angle beyond vertical for improved penetration and can be hydraulically adjusted while ripping to provide the optimum ripping angle in most MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 198 materials. The ripper may have a multishank arrangement where multiple shanks are pulled by a single dozer (typically up to three). Figure 5-53: Ripper Nomenclature Other ripper nomenclature includes: - Pryout, also known as breakout – the maximum sustained force upwards generated by the lift cylinders - Penetration force – the maximum sustained downward force, generated by the ripper lift cylinders measured at the ripper tip. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 199 Figure 5-54: Nomenclature for rippers (see Table 5-2) Table 5-2: Ripper Selection for Large CAT Dozers. Good penetration is essential for high production and depends on: - Material; - Down pressure; MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 200 - Point or tip angle with ground; - Ripping direction (material bedding planes, slopes) Cost comparison: Blasting vs. ripping: - Higher repair costs; - Ripper tip replacement should be included in hourly costs; - Machine life average 8000 hrs (continuous ripping) - May be 1/3 to ½ of blasting, in proper application. 5.4.1 Determining Rippability. Ripping is an inexpensive method of removing discontinuous ground or soft rock masses. Some weaker fragmented sedimentary rocks (less than 15 MPs compressive strength, example: mudstone) are not easily removed by blasting as they are pulverized in their immediate vicinity of the blasthole or may lift along bedding planes then fall back when gas pressure has been dissipated. Figure 5-55 shows a rough chart of various excavation methods considering several ground types. Table 5-3 provides excavation characteristics for a wide variety of material. Figure 5-55: Rock quality classification in relation to excavation processes. MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 201 Table 5-3: Excavation Characteristics in relation to rock hardness and strength Table 5-4: Excavation Characteristics in relation to joint spacing Figure 5-56: Seismic velocities in relation to ripping The most common method for determining if a rock mass is economically rippable is seismic refraction. Figure 5-56 shows the seismic velocity of MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 202 various rock types and their rippability. The upper limit of ripper operations is ground with a seismic velocity approximately 2 km/s. However, this generalized measurement only provides a very limited view of the issues in ripping operation. Seismic velocity can vary as much as 1 km/s in identical materials. Rippability assessments based on rock classification is far more accurate. An adapted Q system of rock classification can be used to define an excavability index, N, which ca be calculated by the equation below, and supplemented by Table 5-5. Table 5-6 through Table 5-10 provide comparative tables from which an excavability index can be calculated. a r s n S J J J J RQD M N × × = ) ( ) ( Table 5-5: Variables in excavability index M s mass strength number (amount of effort needed to excavate dry, homogeneous material with no discontinuities RQD Rock quality designation J n Number of joint sets J s Reducing effect which the block shape and orientation has on the force needed to break out the material J r Roughness of the most unfavorable joint sets J a Degree of alteration Table 5-6: Mass Strength Number for Rocks (Ms) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 203 Table 5-7: Joint count number (J c ) and Joint set number (J n ) Table 5-8: Relative ground structure number (J s ) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 204 Table 5-9: Joint roughness number (J r ) Table 5-10: Joint Alteration number (J a ) MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 205 5.4.2 Operational issues: 5.4.2.1 Ripping technique: Using hydraulic force to help fracture rock is an effective technique for ripping hard rock seams or pockets is to combine tractor drawbar power and hydraulic force. When the ripper has contacted the hard rock, the following procedure is undertaken: 1. The decelerator is used to control trackslip. 2. The ripper shank is controlled to adjust shank angle back slightly. 3. Engine speed is maintained high enough to allow the tractor to continue moving forward as the ripper shank angles back. 4. While the tractor moves forward, the ripper shank is returned to its original position (forward), combining ripper hydraulic force with tractor drawbar pull. 5.4.2.2 Ripping Downgrade Ripping downgrade can increase production. If the job layout permits, the downgrade approach can be helpful when working a hard spot or seam. There are several key issues to be aware of when ripping downgrade. - Traction on rock is less than on dirt. - Avoid ripping on or creating slopes the tractor cannot climb. - Avoid sideslopes. 5.4.3 Estimating Ripping Production Ripper production rates: can be determined empirically by: - Determining the average cycle time, measure average rip distance, rip spacing, and depth penetration, calculate production rate; or - Record time spent ripping, remove and weigh (through surveying & volume calculations), then calculate production rate. Since ripping is primarily compared to drilling and blasting, on a per ton or volume basis, an estimate of ripper production in these terms is needed. The simplest method is to record the time spent ripping, then remove the weight of the material. For example, consider an example where a CAT D10R dozer equipped with a single shank ripper excavates a path every 90 cm (width of tear) at an average speed of 1.6 km/h. After every 91 meters, the machine must turn (takes 0.25 minutes to raise shank, pivot dozer, and lower shank again). This is the equivalent of 1 pass. Consider that the ripper shank has a penetration depth of 61 cm. Solution: 1.6 km/h = 26.7 m/min. Therefore min 41 . 3 min / 7 . 26 91 = m m 3.41 + .25 (turn time) = 3.66 min/pass MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 206 - if operator works 45 min/h, is would be possible to make 45/3.66=12.3 passes per hour. - The volume ripped can be estimated by: 91m x 0.61m x 0.90m = 49.1 BCM per pass - Therefore the volume per pass is equivalent to 12.3 x 49.1 = 604 BCM/h Note that these estimates are approximately 10-20% higher than what would be expected on-site. Productivity can also be estimated using the ripper production charts, such as those seen in Figure 5-57. However, it should be recognized that in these charts: - Rip full time (does not doze simultaneously) - Power shift tractors with single shank rippers - 100% efficiency (60 min/hr) - For all classes of material. Figure 5-58 also provides estimates for ripper productivities for various material type. Figure 5-57: Ripper Production Chart MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 207 Figure 5-58: Performance chart for D11 Figure 5-59: D11 with ripper MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 208 5.4.4 Ripping vs. Blasting Improvements in ripping tractor capabilities have made ripping a viable alternative to blasting. As mentioned before, environmental factors will undoubtedly play an important role in reaching this decision. For example, in an urban area there may be restrictions prohibiting the use of explosives, making ripping a necessity. Political factors or the threat of terrorism severely restrict the use of explosives in some countries. But in most situations, where there is equal opportunity for the use of either method, the first consideration is probably one of cost - will it be cheaper to rip or blast? This initial-cost consideration must then be weighed with other influencing factors: the economics of fully utilizing equipment; the end use of the material; and transporting and loading methods. Full utilization of the equipment available or already on the job can help determine the best method of loosening the material. This is because many earthmoving jobs already involve track-type tractors and scrapers for a sizable portion of the total yardage. If this equipment can be used to finish the job - rather than bring in a rock crew with drills, explosives, loaders and hauling units - it's not difficult to appreciate the savings involved. It's soon apparent that considerable effort can be expended to rip the material in order to keep scrapers on the job. End-use of the material also influences the ripping vs. blasting decision. There are few size limitations when the rock is simply moved by a bulldozer and "wasted." If the material is used to form an embankment, however, very definite limitations are usually placed upon the size of the rocks to be accepted. Optimum compaction cannot be obtained if there are large rocks in the fill. Variations in ripping depth, spacing, and direction of passes usually can produce the desired material size. Blasting is at times unpredictable, as the desired rock fragmentation may be difficult to obtain and even require expensive secondary blasting (in effect, reblasting). Appreciable increases in crusher production have been realized by cement plants and aggregate quarries after switching from a blasting to a ripping operation. The final comparison of ripping vs. blasting can be made in terms of how the material is to be moved. As we stated, dozed material presents few problems. Material top loaded into hauling units cannot be larger than the loading bucket. Scrapers can inexpensively haul materials which are well broken up and loosened. Elevating loaders and conveyors are high capacity systems. Their greatest advantage - high production - can be achieved only if the material is in small pieces and easy to handle. Generally, ripping is the most cost- effective method to achieve these requirements. A cost analysis will indicate the economics of ripping over drilling blasting. This comparison indicates how ripper tip life is an important factor in deciphering the production needed for ripping to be cost effective. Ripper tips are the most expensive variable in the operating costs of ripping tractors, accounting for approximately 30 to 40 percent of total operating costs on the MGE 415 – Rock Excavation 2004 Dessureault 1/14/2004 209 largest tractors. In the final analysis, a ripping vs. blasting decision will depend on the total volume of material to be loosened and moved, on the production capabilities and costs of the ripping tractor(s) used, and on the size and relative efficiency of drilling and blasting techniques. ADDITIONAL RESOURCES XX The above information is additionally supplemented by: - Lecture for module 2.4 (Mandatory) - Assignment XX? MGE 415 – Rock Excavation 2004 Table of Contents MODULE 1: 1.1 INTRODUCTION ......................................................................................................... 1 INTRODUCTION .............................................................................................................................. 1 ADDITIONAL RESOURCES I .................................................................................................................. 1 1.2 GEOLOGICAL PROPERTIES IN ROCK EXCAVATION ......................................................................... 1 1.2.1 Minerals and Rocks .................................................................................................................. 1 1.2.2 The properties of rocks ............................................................................................................. 3 1.2.3 Rock classification.................................................................................................................... 6 1.2.4 Rock Strength............................................................................................................................ 8 ADDITIONAL RESOURCES II............................................................................................................... 16 1.3 ROCK BREAKING PROCESSES....................................................................................................... 16 1.3.1 Primary................................................................................................................................... 17 1.3.2 Secondary ............................................................................................................................... 20 1.3.3 Tertiary ................................................................................................................................... 21 1.3.4 Miscellaneous ......................................................................................................................... 23 ADDITIONAL RESOURCES III ............................................................................................................. 24 MODULE 2: DRILLING AND BLASTING COMPONENTS......................................................... 1 2.1 DRILLING MECHANISMS ................................................................................................................ 1 2.1.1 Percussive Drilling ................................................................................................................... 2 ADDITIONAL RESOURCES IV ............................................................................................................... 5 2.1.2 2.2 Rotary Drilling ......................................................................................................................... 6 DRILLING EQUIPMENT .................................................................................................................. 28 ADDITIONAL RESOURCES V............................................................................................................... 28 ADDITIONAL RESOURCES VI ............................................................................................................. 28 2.3 EXPLOSIVES INTRODUCTION ........................................................................................................ 28 2.3.1 Detonation Theory.................................................................................................................. 28 2.3.2 Comparative Explosive Properties ......................................................................................... 33 2.3.3 Thermochemistry of Explosives and the Detonation Reaction. .............................................. 38 ADDITIONAL RESOURCES VII............................................................................................................ 39 2.4 EXPLOSIVE PRODUCTS ................................................................................................................. 40 2.4.1 History of Explosives .............................................................................................................. 40 2.4.2 Energetic material .................................................................................................................. 42 2.4.3 Explosives and Propellants..................................................................................................... 43 2.4.4 Single Molecule and Composite Explosives ........................................................................... 43 2.4.5 Primary, Secondary, and Tertiary explosives......................................................................... 46 2.4.6 Commercial Explosives .......................................................................................................... 46 2.4.7 Nitroglycerin-based Explosives .............................................................................................. 47 2.4.8 Dry Blasting Agents................................................................................................................ 48 2.4.9 Wet Blasting Agents................................................................................................................ 49 2.4.10 Primers and Boosters-........................................................................................................ 52 2.4.11 Initiators and Initiation Systems......................................................................................... 52 2.4.12 Electric Caps...................................................................................................................... 52 2.4.13 Non-electric caps ............................................................................................................... 53 2.4.14 Electronic Detonators ........................................................................................................ 54 2.4.15 Detonating cord ................................................................................................................. 54 ADDITIONAL RESOURCES VIII .......................................................................................................... 55 Dessureault 1/14/2004 i MGE 415 – Rock Excavation MODULE 3: 2004 BLAST DESIGN .......................................................................................................... 56 BLAST DESIGN BASICS ............................................................................................................................. 56 System Approach ................................................................................................................................. 56 Idealized fragmentation curves............................................................................................................ 59 Preliminary guidelines for blast layout ............................................................................................... 62 Ratios for initial design ....................................................................................................................... 70 Powder Factor..................................................................................................................................... 73 Determination of KB............................................................................................................................. 74 Effects to pattern design in changing explosives ................................................................................. 76 Effects to pattern design in changing rock types ................................................................................. 77 Numerical Examples............................................................................................................................ 78 Stemming & Decking ........................................................................................................................... 80 Concluding Notes on D, B, and PF ..................................................................................................... 82 ADDITIONAL RESOURCES IX ............................................................................................................. 83 GEOLOGICAL IMPACTS ON BLAST DESIGN ................................................................................................ 83 Rock Properties ................................................................................................................................... 83 Structure .............................................................................................................................................. 85 ADDITIONAL RESOURCES X............................................................................................................... 92 PATTERNS AND SEQUENCING .................................................................................................................... 93 Effective Burden and Spacing.............................................................................................................. 94 Designing the Timing........................................................................................................................... 96 Generalized Timing ............................................................................................................................. 97 Skipping a Period ................................................................................................................................ 99 Pattern Types..................................................................................................................................... 100 ADDITIONAL RESOURCES XI ........................................................................................................... 104 MODULE 4: BLASTING APPLICATIONS.................................................................................. 105 DRIFT DESIGN ......................................................................................................................................... 105 Cuts.................................................................................................................................................... 107 Blast Layout....................................................................................................................................... 109 ADDITIONAL RESOURCES XII.......................................................................................................... 110 PRODUCTION BLASTING.......................................................................................................................... 111 Short-hole Production ....................................................................................................................... 111 Longhole Production Blasting ........................................................................................................... 112 ADDITIONAL RESOURCES XIII ........................................................................................................ 118 CONTROLLED BLASTING ......................................................................................................................... 119 Line Drilling ...................................................................................................................................... 120 Pre-splitting ....................................................................................................................................... 121 Smoothwall blasting........................................................................................................................... 124 Trim Blasting ..................................................................................................................................... 125 Blasting Ornamental Rock................................................................................................................. 126 Underwater Rock Excavation ............................................................................................................ 129 Controlled Blasting Rules of Thumb - Supplemental......................................................................... 131 ADDITIONAL RESOURCES XIV ........................................................................................................ 132 ENVIRONMENTAL ISSUES ........................................................................................................................ 132 Airblast .............................................................................................................................................. 133 Fly-Rock............................................................................................................................................. 135 Disturbance of the Natural Ground Profile....................................................................................... 137 Dust.................................................................................................................................................... 137 Fumes ................................................................................................................................................ 138 Dessureault 1/14/2004 ii MGE 415 – Rock Excavation 2004 Ground Vibrations ............................................................................................................................. 138 Vibration reduction............................................................................................................................ 142 Cooperating Charges – Ground vibration......................................................................................... 143 The effect of Delayed Explosion ........................................................................................................ 144 Effect of Geological Factors.............................................................................................................. 145 Risk Assessment ................................................................................................................................. 145 Sustainable Development Issues in Rock Excavation ........................................................................ 146 ADDITIONAL RESOURCES XV .......................................................................................................... 146 EVALUATION OF BLAST RESULTS ........................................................................................................... 146 Fragmentation and Swelling of the Muckpile .................................................................................... 147 Geometry of Muckpile, Its Height and Displacement ........................................................................ 152 Condition of the Remaining Mass...................................................................................................... 153 Analysis of the Bench Floor............................................................................................................... 154 Boulders............................................................................................................................................. 154 Vibrations and Airblast...................................................................................................................... 155 Blast Evaluation in Underground Mining ......................................................................................... 155 ADDITIONAL RESOURCES XVI ........................................................................................................ 155 MODULE 5: MECHANICAL EXCAVATION ............................................................................. 156 5.1 MECHANICAL EXCAVATION BASICS .......................................................................................... 156 5.1.1 Synopsis ................................................................................................................................ 162 ADDITIONAL RESOURCES XVII ....................................................................................................... 162 5.2 LONGWALL ................................................................................................................................ 162 5.2.1 Mining System Description................................................................................................... 164 5.2.2 Selection of Cutting Machine................................................................................................ 165 5.2.3 Dimensions of Shearer.......................................................................................................... 165 ADDITIONAL RESOURCES XVIII ..................................................................................................... 167 5.3 CONTINUOUS MINING MACHINES .............................................................................................. 167 5.3.1 Underground continuous mining machines .......................................................................... 168 5.3.2 Continuous Surface Mining Technology............................................................................... 189 ADDITIONAL RESOURCES XIX ........................................................................................................ 196 5.4 DOZERS AND RIPPING ................................................................................................................ 196 5.4.1 Determining Rippability. ...................................................................................................... 200 5.4.2 Operational issues:............................................................................................................... 205 5.4.3 Estimating Ripping Production ............................................................................................ 205 5.4.4 Ripping vs. Blasting.............................................................................................................. 208 ADDITIONAL RESOURCES XX .......................................................................................................... 209 Table of Figures FIGURE 1-1: ORIGINS OF VARIOUS ROCKTYPES ............................................................................................... 2 FIGURE 1-2: DRILL RATE INDEX FOR A VARIETY OF ROCKS ............................................................................. 7 FIGURE 1-3: UNCONFINED COMPRESSIVE STRENGTH VARYING BY TIME. ...................................................... 8 FIGURE 1-4: CRACK PROPAGATION ............................................................................................................... 10 FIGURE 1-5: CURVES OF STRESS-DEFORMATION FOR DIFFERENT TYPES OF ROCKS. ....................................... 12 FIGURE 1-6: CLASSIFICATION OF THE ROCK MASSES. .................................................................................... 16 FIGURE 1-7: PRIMARY ................................................................................................................................... 17 FIGURE 1-8: STRESS RELATING TO SIZE OF INDENTER ................................................................................... 18 FIGURE 1-9: INCREASING ENERGY WITH DEPTH ............................................................................................. 19 FIGURE 1-10: PENETRATION DEPTH OF VARIOUS BUTTON SHAPES................................................................. 19 Dessureault 1/14/2004 iii ............................................................................................................................................................................... 58 FIGURE 3-3: CONTROLLABLE AND UNCONTROLLABLE INPUT VARIABLES AND OUTPUT GOALS............................. 6 FIGURE 2-7: FLOOR PLAN FOR ROTARY DRILL P&H 100XP ......................................................................... BOREHOLE DIAMETER FOR SELECTED INDUSTRIAL EXPLOSIVES ............................................................... 66 Dessureault 1/14/2004 iv .................................................... 26 FIGURE 2-28: GPS RECEIVERS................... 62 FIGURE 3-10: PLAN VIEW OF BENCH SHOWING FIRST ROW......................... 24 FIGURE 2-26: DM-3 DRILL CONTROL SYSTEM ........................................................................................ 65 FIGURE 3-12: TOE CONFINEMENT ........................................................................................................... 21 FIGURE 1-12: TERTIARY BREAKAGE PROCESS . 12 FIGURE 2-14: PENETRATION RATE VS........ ROCK STRENGTH .............................................. PULLDOWN WEIGHT FOR 9 1/4 INCH DIAMETER ROTARY BITS IN HARD FORMATION .................................................................... 3 FIGURE 2-3: PISTON MOVING TOWARD STEEL .................. 2 FIGURE 2-2: COMPONENTS OF A PERCUSSION DRILL – PISTON AND STEEL ..................................................................................... 22 FIGURE 2-1: THE DRILLING PROCESS ....................... 61 FIGURE 3-8: OVERALL COST CURVE........................................................................................................................................................................................................................................................... 7 FIGURE 2-8: SELECTION OF DRILLING METHOD...................................................................................................................... 3 FIGURE 2-4: COMPRESSIVE WAVES IN BOTH PISTON AND STEEL ......................................................... 60 FIGURE 3-6: CRUSHING COST CURVE .............................................................................................................................................................................. 42 FIGURE 2-35: EXPLOSIVE CLASSIFICATION SCHEME .. 13 FIGURE 2-15: RECOMMENDED PULLDOWN WEIGHTS PER INCH OF BIT DIAMETER................. 14 FIGURE 2-16: BIT LIFE VS.......................... 21 FIGURE 1-13: TERTIARY BREAKAGE A TENSILE EFFECT ............................................................................................. 29 FIGURE 2-30: DETONATION PROCESS FOR CYLINDRICAL EXPLOSIVE ................................................. .................................................................................... PERCENT FUEL OIL ADDED TO AMMONIUM NITRATE ................................................MGE 415 – Rock Excavation 2004 FIGURE 1-11: SECONDARY BREAKAGE PROCESS .............................. ..................................... 26 FIGURE 2-29: DETONATION PROCESS OF EXPLOSIVE CARTRIDGE ................................................................................................... 8 FIGURE 2-10: BLADE AND ROLLER STABILIZER ............................................................................................ 11 FIGURE 2-13: BLASTHOLE COMPONENTS WITH RESPECT TO DRILL RIG ......................................................................................................................................................................................................... 63 FIGURE 3-11: EFFECT OF HOLE DIAMETER ON BURDEN ....................................................................................................................... ....... 66 FIGURE 3-13: RUN-UP DISTANCE TO ACHIEVE STEADY STATE VOD. 25 FIGURE 2-27: OPERATOR’S VIEW WHEN NAVIGATING ................................................................... 31 FIGURE 2-32: SHOCK WAVE PROPAGATION ........... 60 FIGURE 3-5: HAULING COST CURVE ......................................................................... ................................................................................................... 54 FIGURE 3-1: SYSTEMS APPROACH ........................... 43 FIGURE 2-36: ENERGY OUTPUT VS.................. 62 FIGURE 3-9: ISOMETRIC VIEW SHOWING NOMENCLATURE ............................... 15 FIGURE 2-20: AIR BAILING CHART .......................................................... 57 FIGURE 3-2: CASCADING CONSTRAINTS AND GOALS ................................................................................................................................................ 23 FIGURE 2-23: COST PER TON OF MATERIAL EXCAVATED ................................................................................................................................................. 35 FIGURE 2-34: EXPLOSIVES FAMILY TREE ................................................................................. 3 FIGURE 2-5: REFLECTIONS AND WAVE TRANSMISSION...................................................................................................................................................................................................................................................... 48 FIGURE 2-37: VOD VS.. 24 FIGURE 2-25: DM-2 MATERIAL RECOGNITION SYSTEM......................................................... 59 FIGURE 3-4: LOADING COST CURVE............ 7 FIGURE 2-9: COMPONENTS IN ROTARY DRILLING .............................................................................................................................................................. 32 FIGURE 2-33: GENERALIZED RELATIONSHIP BETWEEN VOD AND DIAMETER ...................................................................................... 22 FIGURE 2-22: ROTARY DRILLING COSTS ESTIMATES BY METER DRILLED AND DIAMETER OF HOLE.......................... 10 FIGURE 2-12: VIBRATION REDUCING EFFECTS OF USING SHOCK SUBS ...................................... 9 FIGURE 2-11: SWIVEL MOUNT SHOCK SUB........................................................................... 14 FIGURE 2-17: ROTARY DRIVE TYPES..................................................................................... 4 FIGURE 2-6: DYNAMICS OF SURFACE BLASTHOLE DRILL ......................................... 31 FIGURE 2-31: PRESSURE SHAPE FOR A-HIGH EXPLOSIVE AND B-COMMERCIAL EXPLOSIVE CONTAINING HIGH GAS VOLUME .............................................................................................. 15 FIGURE 2-18: PULLDOWN MECHANISMS .................................................................................... 17 FIGURE 2-21: TONS DRILLED PER OPERATING HOUR FOR ROTARY DRILLS FOR VARIOUS HOLE DIAMETERS........................................................................................................................................................................... 49 FIGURE 2-38: GENERIC DESIGN DETONATORS ................................................................................................................................. 23 FIGURE 2-24: DM-1 PRODUCTION MONITORING SYSTEM ............ 15 FIGURE 2-19: RACK AND PINION PULLDOWN ............ 60 FIGURE 3-7: DRILLING AND BLASTING COST CURVES .......................... ........................ 103 FIGURE 3-47: SEQUENTIAL FIRING ................................................................................. 118 FIGURE 4-15: DIAGRAMMATIC REPRESENTATION OF THE BDT OF A FULLY CHARGED HOLE.................................. 117 FIGURE 4-14: VCR LOADED EXPLOSIVE COLUMN ................. 94 FIGURE 3-36: CORNER AND FACE BLASTS.......................................................... 93 FIGURE 3-35: MUCKPILES ILLUSTRATING THE DIFFERENCE BETWEEN MILLISECOND TIMED AND INSTANTANEOUS............................................. 116 FIGURE 4-13: VCR............................................................................................................................................................................................................................................................................................................................ 120 Dessureault 1/14/2004 v ............................................................................................................ 69 FIGURE 3-18: BURDEN TO DIAMETER RELATIONSHIP................................................................................................ 90 FIGURE 3-30: POSSIBLE BLAST PATTERNS IN JOINTED ROCK ................................................................................................................................................................ BOTTOM LEFT................................................................................................................................................................................................................................................ AND A DRAW CUT.................................................. 71 FIGURE 3-22: BLASTHOLE PLACEMENT IN RELATION TO JOINT PLANES... 68 FIGURE 3-17: EQUIVALENT SPHERICAL AND CYLINDRICAL CHARGES............................ BOTTOM RIGHT) .. 89 FIGURE 3-29: EFFECT OF JOINTING ON FRAGMENTATION ......................................... 71 FIGURE 3-21: STAGGERED PATTERN.................................................................. 106 FIGURE 4-2: DIFFERENT ZONES IN A DRIFT ROUND ................................... 119 FIGURE 4-16: DBT FOR A TWO ROW PRODUCTION BLAST ................................ THREE-HOLE PYRAMID............................ ................ 113 FIGURE 4-10: SPACING CONSTRUCTION FOR RING BLAST DESIGN................................................................. 86 FIGURE 3-24: STAB HOLE .............................................................................................. 102 FIGURE 3-45: ALTERNATING DELAY PATTERN .............................................................................................................. 101 FIGURE 3-42: ECHELON DELAY PATTERN .............................................................. 99 FIGURE 3-41: CHEVRON OR V PATTERN WITH DOUBLE HOLE INITIATION FOR SLIGHTLY MORE FORWARD MOVEMENT........................................................................................................................................................................................... 85 FIGURE 3-23: USE OF POCKET CHARGES ............ .................................. 104 FIGURE 4-1: SYSTEMS FOR ADVANCING LARGE FACES .................................... ............ 92 FIGURE 3-33: SHOOTING ALONG STRIKE ................... TOP RIGHT.......................... 95 FIGURE 3-38: “V” (VEE) OR CHEVRON INITIATION PATTERN ON A SQUARE DRILL PATTERN................................................. 103 FIGURE 3-46: SINKING SHOTS....... 101 FIGURE 3-43: CHANNEL PATTERN ................................................. 109 FIGURE 4-7: SEQUENCING DRIFTS ......................................................................... 92 FIGURE 3-34: INCREASING NUMBERS OF ROWS INCREASES UPWARD MOVEMENT DUE TO LOWER VELOCITY OF PREVIOUSLY BROKEN ROCK............................................................................................ 115 FIGURE 4-12: BENCH BLASTING ................................................................ 70 FIGURE 3-19: LIMITING THE CHARGE DIAMETER AND BURDEN... 96 FIGURE 3-39: THE EVENTS AND TIMING IN A BLASTHOLE .............................................................................................. 114 FIGURE 4-11: STEMMING LENGTH FOR RING BLASTING ........................................................... 68 FIGURE 3-16: LANGEFORS & KIHLSTROM’S TOE BREAKING EQUIVALENCE OF SPHERICAL AND CYLINDRICAL CHARGES ..................................................................................... ...................................................................................... 91 FIGURE 3-32: SHOOTING AGAINST THE DIP ............................................................................................................................................................................... DOUBLE VEE OR BABY CUT. 106 FIGURE 4-4: VARIOUS TYPES OF BURN CUTS (SOLID DOTS ARE LOADED HOLES) .......................................................................................................... ........................................................................................................................................................... 106 FIGURE 4-3: ORIENTATION OF THE CONTOUR BLASTHOLES TO MAINTAIN TUNNEL PROFITE-LOOKOUTS........ TOP LEFT................................ 109 FIGURE 4-6: CYLINTRICAL CUT .................................................................................. .. 108 FIGURE 4-5: ANGLE CUTS (VEE OR WEDGE................................................ 110 FIGURE 4-8: SHORT-HOLE PRODUCTION BLAST..................................................................................... 94 FIGURE 3-37: ECHELON OR HALF CHEVRON........................................ 89 FIGURE 3-28: FRAGMENTATION ENHANCEMENT ACHIEVED BY REDUCING THE SPACING ............. 112 FIGURE 4-9: RING BLASTING.................................. 98 FIGURE 3-40: BLAST TIMING SEQUENCE SKIPPING PERIODS ...................................MGE 415 – Rock Excavation 2004 FIGURE 3-14: SECTION VIEW COMPARING THE SPHERICAL CHARGE (A) AND CYLINDRICAL CHARGE (B) MINIMUM DISTANCE FROM THE COLLAR ............................. 67 FIGURE 3-15: VIEW FACING BENCH COMPARING EQUIVALENT CYLINDRICAL AND SPHERICAL CHARGES.................................................................................................................. 102 FIGURE 3-44: FLAT FACE PATTERN............... 87 FIGURE 3-26: EXTENT OF CRACKING FOR TWO PATTERNS WITH DIFFERENT HOLE DIAMETERS AND SAME PF 88 FIGURE 3-27: MAXIMUM BLOCK DIMENSIONS FOR HOLE PATTERNS............................................................................................................................................................................ 90 FIGURE 3-31: SHOOTING WITH THE DIP ....... 86 FIGURE 3-25: IDEALIZED RADIAL CRACKING SURROUNDING A SINGLE HOLE..................................................................................................................................................................................................................................................................................... .............. 70 FIGURE 3-20: SQUARE PATTERN ............... ............. D)....... DEPTH OF CUT (ES VS. 171 FIGURE 5-18: MINING IN VARIABLE HEIGHT SEAMS ...................................... 157 FIGURE 5-3: MEAN FORCE VS............... 140 FIGURE 4-28: GROUND VIBRATIONS RESULTS FROM 2500 BLASTS IN 40 DIFFERENT OPERATIONS...... .............................................. ................ .............................. 141 FIGURE 4-29: COOPERATING CHARGE RELATED TO DISTANCE.................................................... 174 FIGURE 5-22: ROADHEADER PERFORMANCE VS................ 181 FIGURE 5-32: STEPS IN RAISEBORE MINING................................................................... 175 FIGURE 5-24: COMPONENTS OF A TBM..................................................... 168 FIGURE 5-15: MILLING HEAD CONTINUOUS MINING MACHINE....................................................................................................................................................................... 159 FIGURE 5-7: MEAN FORCE VS.................................................................................................................................................................................................................................................................................... 130 FIGURE 4-25: AIRBLAST WAVES IN VARIOUS CONDITIONS...................................................................... PICK WIDTH (FC VS................. D) ............................................................ .. SIMPLE CHISEL..... 182 FIGURE 5-34: RAISE BORING DIAGRAM – UP AND DOWNWARD BORING....................................................... 161 FIGURE 5-9: CHISEL VS................................................................................................................................................................................................................................................... 150 FIGURE 32: DIFFERENT GEOMETRIES OF THE MUCKPILE ........................................ AND POINTED TOOL............................................... )............................................... 172 FIGURE 5-19: DOUBLE HEAD ROADHEADDER................... 180 FIGURE 5-31COST COMPARISON BETWEEN TUNNEL BORING AND CONVENTIONAL DRILL AND BLAST......................................................................MGE 415 – Rock Excavation 2004 FIGURE 4-17: BDT AFTER BLASTING TWO PRODUCTION ROWS................................................... ..................... 163 FIGURE 5-11: LONGWALL SHEARING SYSTEM ....................... 177 FIGURE 5-27: KERF AND PINAPPLE CUTTERS ............................................................................................................................. 177 FIGURE 5-28: SPACING VERSUS DEPTH RELATIONSHIP ........ 175 FIGURE 5-23: TUNNEL BORING MACHINE ............. 164 FIGURE 5-12: SHEARER DIMENSIONS ..................................................................................................................................................................................................... .............................................................................. 159 FIGURE 5-6: MEAN FORCE VS....................................................... ....................................... PICK BITS..................................................................................................... = RAKE ANGLE)......... C AND FC VS......................................... 176 FIGURE 5-25: DISC CUTTERS ON TBM .......................... ..... 144 FIGURE 30: GIRD METHOD OF EVALUATING FRAGMENTATION AND FIELD ENVIRONMENT ...................................... .............. BACK CLEARANCE ANGLE (FC VS........................................ 123 FIGURE 4-21: SMOOTHWALL DAMAGE ZONES ............................................... 169 FIGURE 5-16: MILLER-HEAD CONTINUOUS MINER......................................... 158 FIGURE 5-4: SPECIFIC ENERGY VS........................................................................................................................ W).............................. 183 Dessureault 1/14/2004 vi ..... 153 FIGURE 5-1: LONGWALL SHEARING MACHINE WHOSE CUTTING BITS ATTACK THE ROCK IN COMPRESSION ....... 136 FIGURE 4-27: OBTAINING SITE FACTORS FOR GROUND VIBRATIONS...... 125 FIGURE 4-23: ORNAMENTAL STONE MATERIALS HANDLING – CRANES............... 173 FIGURE 5-21: ROCK MASS CUTTABILITY INDEX VS... 135 FIGURE 4-26: CAUSES OF FLY ROCK....................... 156 FIGURE 5-2: PICK SHAPES .............................. 170 FIGURE 5-17: MINING OF STEEP SEAMS WITH ROADHEADDER AND MILLER HEAD (DRUM-TYPE) MINERS ......................... 160 FIGURE 5-8: COMPRESSIVE AND TENSILE STRENGTHS UPON VARYING MEAN CUTTING FORCE (FC VS........................ RAKE ANGLE (FC VS...........................................................................................................................................................................POINTED TOOL....................... 122 FIGURE 4-20: RELATIONSHIPS BETWEEN HOLE DIAMETER AND SPACING FOR PRESPLITTING FROM VARIOUS RESEARCHERS........................................................ 158 FIGURE 5-5: MEAN FORCE VS.................................... 182 FIGURE 5-33: RAISEBORING IN OPERATION TOP & BOTTOM ..... DEPTH OF CUT (FC VS......................................................................................... 124 FIGURE 4-22: TRIM HOLE ROW TRIMMING THE FRACTURED AND INFLUENCED ROCK............. CUTTING RATE........... 166 FIGURE 5-14: FOUR-ROTOR BORING MACHINE............................................................................................................................ 148 FIGURE 31: HIGH SPEED CAMERA BENCH BLAST .. 172 FIGURE 5-20: MD1100 WITH TRAVERSE BOOM . T).......................... 165 FIGURE 5-13: NOMOGRAPH FOR SHEARER DIMENSIONING ................................................................................................................................................... ) ............................ 121 FIGURE 4-19: PRESPLIT BLAST COUPLED TO A 250MM PRODUCTION BLAST........ 120 FIGURE 4-18: LINE DRILLING POSITIONED ALONG THE PLANNED FINAL PERIMETER .............................................................................................................................................. 176 FIGURE 5-26: CROSS SECTION OF DISC CUTTER .................................. 178 FIGURE 5-29: EMPIRICALLY DERIVED EXCAVATION RATES OF TBMS BY DIAMETER FOR IGNEOUS AND METAMORPHIC ROCK TYPES .... ( = BACK CLEARANCE ANGLE............................................................................................................. ROCK CLASS....... 179 FIGURE 5-30: EMPIRICALLY DERIVEN EXCAVATION RATES OF TBMS BY DIAMETER FOR SEDIMENTARY ROCK TYPES ........................................................ 127 FIGURE 4-24: SEQUENCE OF OPERATIONS IN UNDERWATER ROCK EXCAVATION........ 161 FIGURE 5-10: TYPICAL LONGWALL RETREAT METHOD ....... ......... 191 FIGURE 5-46: CUTTING CAPACITIES FOR VARIOUS WIRTGENAMERICA PRODUCTS .............................................................................................................................................................................................................................................. 199 FIGURE 5-55: ROCK QUALITY CLASSIFICATION IN RELATION TO EXCAVATION PROCESSES.......................................................... 45 TABLE 2-8: TYPICAL COMPOSITIONS OF SELECTED SLURRIES AND EMULSIONS ................................. 73 TABLE 3-2: POWDER FACTOR......................................... 189 FIGURE 5-44: VARIOUS ASPECTS OF A SURFACE MINER...................................................................................................... 53 TABLE 3-1: RATIO SUMMARY .................................................... 33 TABLE 2-5: THERMODYNAMIC DATA FOR SOME EXPLOSIVE COMPONENTS AND GASES..... 82 TABLE 3-3: EFFECT OF DRILLING PATTERNS AND S/B RATIOS ON THE AREA COVERED BY FRACTURE CIRCLES (ENERGY DISTRIBUTION)............. 192 FIGURE 5-49: BUCKETWHEEL NOMENCLATURE......................................... 197 FIGURE 5-52: (A) RADIAL RIPPER...................................... 98 TABLE 4-1: POWDER FACTORS FOR RING BLASTING ........................................................................................ 134 TABLE 4-6: FACTORS THAT INFLUENCE GROUND MOTION ...... 207 FIGURE 5-59: D11 WITH RIPPER .................................................................................................................................................................................................................................................. 5 TABLE 1-2SEDIMENTARY ROCKS .................. 16 TABLE 2-4: STANDARDS FOR FUME CLASS .............................................. 13 TABLE 1-6: COMMON ROCK NAMES AND THEIR GEOLOGICAL DEFINITIONS ....................................... ............................................................................................. 188 FIGURE 5-43: CROSS SECTIONAL VIEW OF FOUR-ROTOR BORING IN POTASH................................ 207 List of Tables TABLE 1-1: IGNEOUS ROCK .................................................................................................................. 5 TABLE 2-2: COMPARATIVE TABLE ............................................................. 119 TABLE 4-3: FACTORS FOR DETERMINING HOLE SPACING .............................................. 44 TABLE 2-7: MORE SINGLE CHEMICAL EXPLOSIVES ........................... 5 TABLE 2-3: AVERAGE DRILL BIT FOOTAGES (FROM LARGE IRON ORE MINE IN CANADA) ........................................................ .................................................... 50 TABLE 2-9: TYPICAL DELAY TIMES FOR DONATORS ....................................................................................................................................................................... ..................................................................................................... 186 FIGURE 5-40: REMOTE OPERATION ALLOWS DISTANCE MINING AND THE OPERATOR TO VIEW FROM DIFFERENT ANGLES................................................................................................................................................................................ 193 FIGURE 5-50: BUCKET CUTTING AND DISCHARGE ........................................ 191 FIGURE 5-47: SURFACE MINER IDEAL FOR DEFINED THIN VERTICAL OR FLAT SEAMS ........................ 14 TABLE 1-7: ROCK PROPERTIES ACCORDING TO ORIGIN ..................................................................................................................................... 197 FIGURE 5-53: RIPPER NOMENCLATURE ......................... (B) PARALLELOGRAM RIPPER.............................................................................. 185 FIGURE 5-38: HORIZONTAL BOREHEAD ...................................... BEING MAINTAINED ............... 193 FIGURE 5-51: RIPPER IN ACTION .................................. (C) ADJUSTABLE PARALLELOGRAM RIPPER ........................................................... 126 TABLE 4-5: AIRBLAST EFFECT ON STRUCTURES ....................................................... 6 TABLE 1-4: FRACTURE TOUGHNESS ........................................................................................................................................................................................................................................................................... 198 FIGURE 5-54: NOMENCLATURE FOR RIPPERS (SEE TABLE 5-2) ...... 190 FIGURE 5-45: SURFACE MINER IDEAL FOR THIN STRATA ..... 114 TABLE 4-2: BDT CHARACTERIZATION ....................................................................................................................................................................................... 206 FIGURE 5-58: PERFORMANCE CHART FOR D11.............................................................. ................................................................................................................................................................................ 184 FIGURE 5-36: BLIND BORING BIT HEAD............................... 187 FIGURE 5-41: MARRIETA MINER IN ROCANVILLE . 201 FIGURE 5-57: RIPPER PRODUCTION CHART ................................... 38 TABLE 2-6: SINGLE CHEMICAL EXPLOSIVE SUBSTANCES.................................................................................................................... 188 FIGURE 5-42: MARRIETA DIGGING A TRENCH.......................................................................................................... 11 TABLE 1-5: RELATIVE QUARTZ (COMMON ABRASIVE) CONTENT ....................................................................................................... 14 TABLE 1-8: SUMMARIZING THEORY OF FORCES AND ENERGY IN BREAKING PROCESSES ................... 121 TABLE 4-4: RECOMMENDED CHARGE LOADS AND BLAST GEOMETRIES FOR TRIM BLASTING ....................................... 200 FIGURE 5-56: SEISMIC VELOCITIES IN RELATION TO RIPPING ..................................... 141 Dessureault 1/14/2004 vii ............. 191 FIGURE 5-48: COMPONENTS OF SURFACE MINER ............................................................................................................................................................................................................................................ 184 FIGURE 5-37: DIRECTIONAL BORING IN URBAN ENVIRONMENT............ 6 TABLE 1-3: METAMORPHIC ROCKS.............................................. 24 TABLE 2-1: COMPARISON OF HYDRAULIC AND PNEUMATIC DRILLS......................................... 185 FIGURE 5-39: CROSS SECTION OF THE BOUBLY POTASH MINE STRATA ...............................................................................MGE 415 – Rock Excavation 2004 FIGURE 5-35: BLIND BORING OPERATION ..................................................................................................................................... ............................................................... 202 TABLE 5-7: JOINT COUNT NUMBER (JC) AND JOINT SET NUMBER (JN).......................................................................................................................... 203 TABLE 5-9: JOINT ROUGHNESS NUMBER (JR) ........... 203 TABLE 5-8: RELATIVE GROUND STRUCTURE NUMBER (JS)...................................................................................................................................................................................... . 202 TABLE 5-6: MASS STRENGTH NUMBER FOR ROCKS (MS) ...........MGE 415 – Rock Excavation 2004 TABLE 4-7: RECOMMENDED MAXIMUM PEAK PARTICLE VELOCITIES............................................................ 154 TABLE 5-1: APPLICATIONS FOR BORING ............... 142 TABLE 4-8: EFFECT OF DELAYED CHARGES ON PPV FOR A TOTAL CHARGE OF 2900 KG AT A DISTANCE OF 220M............................... 201 TABLE 5-5: VARIABLES IN EXCAVABILITY INDEX ..................... 201 TABLE 5-4: EXCAVATION CHARACTERISTICS IN RELATION TO JOINT SPACING ............................... 199 TABLE 5-3: EXCAVATION CHARACTERISTICS IN RELATION TO ROCK HARDNESS AND STRENGTH............................................................................................................................................................................................................................................................... ........................................................ 204 TABLE 5-10: JOINT ALTERATION NUMBER (JA).................................................................................... 204 Dessureault 1/14/2004 viii .......... 183 TABLE 5-2: RIPPER SELECTION FOR LARGE CAT DOZERS................. 144 TABLE 9: LEVELS OF DAMAGE TO PIT WALLS ........................................... ..... A mineral may be an element or may consist of chemical compounds containing several elements. making up about 50 per cent of the earth's crust. A mineral is a substance formed by nature. iron. together with oxygen. oxygen is by far the most common. William Cummings. 1. magnesium and titanium. Silicon. Paul Lever. Document S-96174 1/14/2004 Dessureault 1 . formed under different circumstances. a list of additional resources are indicated after every major topic. Of 103 known elements. Rocks consist of one or more composite minerals.1 Lecture 1. Per-Anders. 2000.2.2 Geological Properties in Rock Excavation A brief review of rock properties is required to provide students without this background or as a review. aluminum. To guide the student to additional resources either out of interest or mandated by the course requirements.000 different minerals. 534 p. about 25 per cent. Holmberg. ADDITIONAL RESOURCES I The above information for is additionally supplemented by: Module 1. 1 From SECOROC’s Geology Drilling Theory. Rock Blasting and Explosvies Engineering. calcium.1 Introduction The information necessary to pass this course will be presented in these notes and through other medium such as lectures and field trips. Roger. sodium. There are more than 3. and Lee.MGE 415 – Rock Excavation 2004 Module 1: Introduction 1. The appearance of this list will be as seen below: ADDITIONAL RESROUCES [NUMBER] The above information is additionally supplemented by: Readings (Mandatory or Optional) Field trip (Always mandatory) Assignment Lecture Important note: These class notes borrow heavily from the material assembled by: Dr. Mr. potassium. account for about 99 per cent. Persson. Jaimin. CRC Press: New York.1 Minerals and Rocks1 The earth's crust consists of a variety of rocks. Also included are amphiboles and pyroxenes. Gold. molybdenite. carbonates (calcite. which contain aluminum. halite). Some of the earth's most common rocks. feldspar and mica. malachite).MGE 415 – Rock Excavation 2004 Silicon. silver. arsenopyrite. granite and gneiss. dolomite. tungstates (scheelite) and phosphates (apatite) are other large mineral groups. potassium and iron. quartz. copper and lead are elements that can occur as native metals. sulphates (barite). are composed of silicates. Figure 1-1: Origins of various rocktypes Oxygen often occurs in combination with metallic elements and forms our oxidic ores (the iron ores magnetite and hematite). Feldspar accounts for 2 Dessureault 1/14/2004 . sphalerite. Sulphur readily combines with metallic elements and forms sulphide ores (galena. aluminum and oxygen occur in our most common minerals. Halogenides (fluorite. a silicate being a compound formed of silicic acid and other elements. They form the large group known as silicates. A rock is composed of grains of various minerals. hardness 1.0. we must be able to make a correct appraisal of the rock concerned. Fluorite .Diamond .0. density 6. and the form and distribution of the grains.Easily scratched with the fingernail 2.. fracture. The density of light -colored minerals is for the most part below 3.Scratches glass. Calcite .Scratches glass. blackens a thumb that is rubbed against it. etc. have densities between 3.Can be scratched with a knife 6.Easily scratched with a knife 5. abrasiveness.2 The properties of rocks In order to be able to forecast the result of drilling in respect of penetration rate. Ore minerals have densities over 4. silicates.Just barely scratched with the fingernail 3. These minerals make up about 90 per cent of the composition. streak. scheelite(CaWO4). hole quality. Streak is the color of the mineral powder produced when the mineral is scratched against unglazed.4. Topaz . density 6.Scratches glass.Very easily scratched with a knife 4.3. Gold has a very high density at 19. Cleavage denotes the properties of a crystal whereby it allows itself to be split along flat surfaces parallel with certain formed or otherwise crystallographically defined surfaces.MGE 415 – Rock Excavation 2004 almost 50 per cent of the mineral composition of the earth's crust. and among the microscopic properties are mineral composition. density 4. Fracture is the surface produced by breaking off a piece of mineral. cleavage and crystalline form. Fracture is usually uneven in one way or another. can be scratched with a diamond 10.Scratches glass Molybdenite. can be scratched with quartz 7. Dessureault 1/14/2004 3 .5 Dark-colored minerals with some iron.5.g.2. Quartz . Hardness can be graded according to the Moh's 10-point scale. Talc . and tungsten at 19. These properties. drill-steel costs. such as hardness. 22. these factors decide important properties of the rock. Corundum . In doing so we distinguish between microscopic and macroscopic properties. Taken together. Exceptions are barite or heavy spar (BaSO4). white porcelain (e. closely followed by quartz and mica. an ordinary electric fuse). 1. and cerussite (PbCO3).0 and 4. can be scratched with a hardened file 8. Apatite . Next come the pyroxene and amphibole minerals.Hard to scratch with a knife. not following a crystallographically defined plane. (example followed by test) 1. The highest density is shown by osmium and iridium. grain size.0. Some of the characteristics of the minerals are hardness.5. Orthoclase . density. compressive strength and density. can be scratched with emery 9. color. Gypsum .0. lustre.5. rocks with essentially the same mineral content may be quite different as regards drillability. the rock is massive (e. particularly on the drill bits. Igneous or magmatic rocks (formed from solidified lava or "magma"). We say that the rock is abrasive. The drillability of a rock depends on. Since quartz is a very hard material. Conversely. fissuring. cracks or inclined and layered formations can cause hole deviation and can occasionally cause drilling tools to get stuck.5-1 mm). layering. Among the macroscopic properties are slatiness. These factors are often of great significance in drilling. but in this case pressure and heat have compacted each layer in plates.MGE 415 – Rock Excavation 2004 in their turn. grain size> 5 mm. A slaty rock also has the minerals arranged in different layers. The enormous variety of rocks and rock formations makes it impossible to give the subject adequate treatment in just a few pages. e. We must therefore content ourselves with giving a summary description of the more important rocks and referring interested readers to the literature of the subject for further study. A rock can also be classified on the basis of its structure. Sedimentary rocks (formed by deposition of broken material or by chemical precipitation). 2. since the walls often cave in and in extreme cases the flushing air disappears into cracks in the rock without reaching the surface. e. For example. medium-grained (grain size 1-5 mm) or fine-grained (grain size 0. granite). grain size 0. fibrous (amphiboles and pyroxenes). veining and inclination.g. Consequently.05 mm. Rocks are classified into three main groups on the basis of their origin and the way in which they were formed: 1. contact zones. cubic (galena) are easier to drill than minerals with low symmetry. among other things. or dense. minerals with high symmetry. As regards crystal form.g. Quartz is one of the commonest minerals in rocks. the hardness of its constituent minerals and on the grain size and crystal form. a rock with a high content of calcite is easy to drill and causes little wear on the drill bits.g. A coarse-grained structure is easier to drill in and causes less wear than a fine-grain structure. a high quartz content (SiO2) makes the rock very hard to drill and causes heavy wear. quartzite may be fine-grained.5-1 mm. determine the penetration rate that can be achieved and how heavy the tool wear will be. Soft rocks make it difficult to achieve good hole quality. In mixed rocks the grains are arranged in layers. Dessureault 1/14/2004 4 . For example. grain size 0. If the mineral grains are mixed in a homogeneous mass. while a granite may be coarse-grained. Sedimentary rocks can also arise as a result of the chemical precipitation of minerals or by proliferation of organic organisms. This is evident from the table below.2. which also shows that it is customary to classify the igneous rocks by their silicon content (SiO2). Metamorphic rocks (formed by the transformation of igneous or sedimentary rocks. as for example in coral reefs or carbon deposits. chiefly feldspars. 1. Plutonic rocks solidify slowly and are therefore coarse-grained.2. or as a dyke rock.2 Sedimentary rocks Sedimentary rocks are formed by the deposition of material by mechanical or chemical action and a consolidation of this material under the pressure of overlying layers. or as it rises towards the surface (dyke rock) or on the surface (volcanic rock). the rock is given different names even if the composition is the same. The greater the silicon content. the larger the amount of quartz in the rock will be. carried away by running water and deposited in still water. Since formation takes place by deposition.MGE 415 – Rock Excavation 2004 3. The weathering or erosion may proceed at different rates depending partly on climate and partly on how easily the rock breaks up. Dessureault 1/14/2004 5 . while volcanic rocks solidify quickly and become fine grained. It frequently occurs that the rock formation is broken down by mechanical action (weathering). Depending on whether the magma solidifies at depth. Sedimentary rocks make up a very heterogeneous family with widely varying characteristics. Thus the original rock will determine the characteristics of the sedimentary rock.1 Igneous rock Igneous rocks are formed when a magma solidifies deep down in the earth's crust (plutonic rock). or on the surface.2. The most important constituents (minerals) are quartz and silicates of various composition. several distinct layers can often be observed in a sedimentary formation.2. Table 1-1: Igneous Rock SiO2 Basic <52% SiO2 Intermediary 52-65% SiO2 Acid >65% SiO2 Plutonic Gabbro Diorite Syenite Quartz diorite Granodiorite Granite Dykes Diabase Porphyrite Syneite porphyry Quartz porphyrite Granodiorite porphyry Quartz porphyry Volcanic Basalt Andesite Trachyte Dacite Rydodacite Rhyloite 1. in most cases by an increase in pressure and heat). 2.2. crushed zones. The result is deformation in one way or another. The earth's crust is in a constant state of flux. faults and other phenomena that can have a telling effect on drilling.MGE 415 – Rock Excavation 2004 Table 1-2Sedimentary Rocks Rock Conglomerate Greywacke Sanstone Argillaceous schist Aluminous slate Limestone Original material Gravel. greywacke. pressure and temperature might increase under the influence of up-welling magma. gabbro Low Sandstone Medium/high Dacite Medium Shale Low Silicic-acid-rich silicate high rocks Original Rock of 1.2. for example. boulders Clay. For example. diabase. stones. and in nature we may therefore observe. gravel Sand Clay Clay plus organisms Calcium carbonate or various organisms 1. and also compact and fine-grained. This results in the recrystallization of the mineral grains or the formation of new minerals. Medium/high etc. and are therefore often difficult to drill. pronounced folding. They are also frequently hard or very hard. Various igneous rocks High Basalt. Table 1-3: Metamorphic rocks Rock Amphibolite Mica-schist Gneiss Greenschist Quartize Lepitte Slate Veined gneiss Degree metamorphism Basalt.3 Rock classification Many attempts have been made to classify rocks on the basis of their drillability. which the geologist calls metamorphism. A characteristic of the metamorphic rocks is that they are formed without any complete melting. diabase. and the various rock formations may be subjected to very powerful forces. gabro High Mudstone. or because the formation has sunk down deeper into the earth's crust. and various measuring methods have been evolved with the aim Dessureault 1/14/2004 6 .3 Metamorphic rocks The effects of chemical action or increased pressure and/or temperature on a rock formation can sometimes be so great that it produces a transformation. MGE 415 – Rock Excavation 2004 of making it possible to predict productivity and tool wear by carrying out a laboratory test before a rock job begins. describes how fast the drill steel can penetrate. these are crude methods but they will often suffice for rough calculations. the grain size and the brittleness. Figure 1-2 shows how the DRI can vary from one rock to another. The term "drillability" is used here to mean primarily the rate at which the tool penetrates into the rock. Figure 1-2: Drill rate index for a variety of rocks Dessureault 1/14/2004 7 . though it also depends on how abrasive the rock is. the higher penetration rate. Tool wear is often proportional to drillability. The drillability of a given rock is determined by a number of factors. the drilling-rate index. The Norwegian Technical University has developed a more sophisticated method for calculating the DRI and BWI. it also includes on the one hand a measurement of brittleness and on the other hand drilling with a small rotating bit into a sample of the rock. the risk of tool jamming etc. The higher the DRI. Attempts are often made to describe drillability by stating the compressive strength or hardness. the straightness of the hole. but in the wider sense it extends to the resulting hole quality. DRI. Foremost among these are the mineral composition. This is a result of the randomly distributed weak planes. There is also an element of creep in the strength and deformation characteristics of rock. This is particularly true in the case of quartz. which explains why relatively easily drilled sandstone. The flaws are often so small and the microcracks so fine that they are difficult to detect by the naked eye. The quartz content has been shown to exert great influence on wear. or flaws in the rock which greatly influence the rock strength. can cause very heavy wear on the bits. Dessureault 1/14/2004 2 8 . Figure 1-3: Unconfined Compressive Strength Varying by Time. and Jaimn Lee. and gives an indication of how fast the bit wears down. the faster will be the wear. compact. microcracks. In most cases the BWI and the DRI are inversely proportional to each other. New York: CRC Press. for example. so that a high DRI will give a low BWI and viceversa.4 Rock Strength2 The compressive strength of many rock materials is a factor of 5 or more greater than their tensile strength.2. The higher the BWI. the presence of hard minerals may produce heavy wear on the bit despite relatively good drillability. 1. quartz-bearing rocks may also prove to be very difficult to drill. Other examples of hard minerals that impair drillability are certain sulphides in orebodies. However. The scatter of the strength values from a series of test specimens of the same rock material is considerable. metamorphic. as evidenced by the ability of rock to flow and deform plastically under tectonic stress over long periods of time (hundreds of Persson. 1994. It is determined by an abrasive test. 50p. In other cases. Per-Anders. Rock Blasting and Explosives Engineering. Roger Homberg.MGE 415 – Rock Excavation 2004 BWI stands for "bit-wear index". which approach a limiting shear strength i when the deformation is entirely plastic. and direction of the micro-cracks or flaws is the reason for both the scatter of experimental strength measurement data and the dependence of strength on the size of the specimen. . microcracks grow. In a large specimen. the friction on such microcracks that are stressed in shear leads to increased strength. that the deformation of the weaker part of the aggregate grains will be plastic. (Figure 1-3) With confinement. with a sufficiently high hydrostatic pressure. The random distribution in space. that is when lateral expansion is restricted. Such flaws are also nearly always present in most natural rock materials because of the deformation the rock has undergone under the influence of tectonic forces and temperature changes. In tension. partly by creating lateral forces that resist the shear motion. with increased hydrostatic confining pressure.MGE 415 – Rock Excavation 2004 thousands to millions of years). so that the compressive strength when the load is applied in 1 msec is a factor of 2 or 2. pores. Therefore. and different size are cemented or grown together. In the limit. rock in compression becomes stronger. we approach the real strength of the aggregate base material. or weakened regions. and partly by increasing the friction on potential shearing surfaces by increasing the normal load thereby also increasing the rock strength. join. Any deformation of sufficient magnitude will lead to local cracking or the development of microscopic flaws. Dessureault 1/14/2004 9 . and the fracturing or plastic deformation of the crystal grains. a large specimen or rock volume has a lower strength than a small specimen. Experiments have shown that the strength of rock is time dependent. With confinement. the crack growth is further restricted and friction is increased. The action of confinement is to resist shearing. In compression. size. Most rock materials are aggregates in which separate crystal grains of different strength.5 greater than when the load is applied in 10 sec. while the hard grains still only deform elastically. and ultimately lead to fracture at a low load. 1. it is more probable than in a small specimen that a sufficiently large flaw will have a direction favoring fracture in a given stress situation. different elastic and thermal moduli. Shear strength is composed of two parts: the friction between sliding crack surfaces characterized by a friction coefficient .1 Microcracks in Rock The reasons for the strength behavior of rock and other brittle materials can be sought in the presence of microscopic cracks and flaws in the base material that is in itself strong.2. This leads to a further increase in strength. It is conceivable.4. This is because deformation to failure takes place as a shearing of weak planes. The stress concentration in front of the crack tip can be represented by the expression: ( x) KI 2 x where the stress intensity factor KI is a function of the crack length and the load o. that is. As the crack propagates. Because the tensile strength of these materials is so much lower than the compressive strength. The critical value of KI when the crack just starts moving is a material constant KIC. The work done per unit new crack surface is GIC which is coupled to KIC through the relation: GIC 1 v2 2 K IC E Figure 1-4: Crack propagation Dessureault 1/14/2004 10 . We will limit this discussion to cracking under biaxial stresses.4.2 Fracture Mechanics The most important aspect of the strength of brittle materials is their ability to break by crack propagation. cracks form easily and.MGE 415 – Rock Excavation 2004 1. stress situations where two principal stresses are equal and the third is zero. expand because of the concentration of tensile stresses at the crack tip (Figure 1-4). and possibly also because they already contain micro-cracks. energy is absorbed by deformation work by the material at the crack tip or dissipated as elastic wave energy radiating out through the material from the crack tip. once formed.2. 3.4 GIC is called fracture toughness and is the fundamental material constant. type of deformation and magnitude of the applied force. Depending upon the nature of deformation. The elastic properties of rocks are characterized by the elasticity modulus 'E' and the Poisson coefficient 'v'. The elasticity module is the proportionality factor between the normal stress in the rock and the relative correspondent deformation. basically depending upon the mineralogical composition. 2.MGE 415 – Rock Excavation 2004 Where v is Poisson’s ratio3 and E is Young’s modulus. A material stretches when pulled under tensile load and usually contracts transversely.2. and are destroyed when the strains exceed the limit of elasticity. Table 1-4 shows fracture toughness of various materials Table 1-4: Fracture Toughness 1. The values of the elasticity modules in the majority of sedimentary rocks are lower than those corresponding to the minerals in their composition. if a bar is pulled in the axial direction then the deformed bar (besides being longer in the axial direction) also contracts in the transverse direction by the percentage amount indicated by the Poisson's ratio. which obeys the Law of Hooke.03 x 104 and 1.4. in which the elastic deformation is insignificant. porosity. as function of the stresses produced by static charges.3 Elasticity The majority of rock minerals have an elastic-fragile behavior. Young's modulus is the ratio of longitudinal stress to the resultant longitudinal strain (stress/strain). The Poisson's ratio is the ratio of transverse strain to axial strain during axial load. as the elasticity module in the direction of the bedding or schistosity is usually larger than when perpendicular. The plastic. its value in most rocks varies between 0. The texture of the rock also has influence on this parameter. The highly plastic or very porous. For example. The elastic-fragile or those which obey the Law of Hooke. Some authors refer to KIC as fracture toughness. that have plastic deformation before destruction. three groups of rocks are taken into consideration: 1. Stiffness of the material Dessureault 1/14/2004 4 3 11 .fragile.7 x 105 MPa. MGE 415 – Rock Excavation 2004 Poisson's coefficient is the factor of proportionality between the relative longitudinal deformations and the transversal deformations. and only in quartz is it abnormally low.4 Plasticity As indicated before. The humid clays and some homogeneous rocks have plastic properties. crystallines and sandstones) becomes noticeable especially at high temperatures. The plasticity depends upon the mineral composition of the rocks and diminishes with an increase in quartz content.2 and 0. For most rocks and minerals it is between 0. in some rocks the plastic deformation precedes destruction.4. This begins when the stresses exceed the limit of elasticity. feldspar and other hard minerals. Real rocks are deformed and consolidated at the same time: in order to increase the plastic deformation it is necessary to increase the effort.07. The plasticity of the stony rocks (granites. 1.4. around 0.2. Figure 1-5: Curves of stress-deformation for different types of rocks. In the case of an ideally plastic body. that deformation is developed with an invariable stress. schistoses. Dessureault 1/14/2004 12 . As to porosity. The shape of the grains. as in a sandstone. The size of the grains. are more abrasive because they leave rough surfaces with hard grains as. The size of the grains are an indication.2.6 Texture The texture of a rock refers to the structure of the grains or minerals that constitute it. The heterogeneity. drilling is more difficult than when they are round. for example.2. All these aspects have significant influence on drilling performance. in the rubbing or abrasive process during movement. the mean amounts of quartz for different types of rock are indicated. This property has great influence upon the life of drill steel and bits.4. although these are equally hard. The porosity of the rock. those rocks that have low density Dessureault 1/14/2004 13 .4. When the grains have a lenticular shape. quartz grains in a granite. In Table 1-5. Those that are angular are more abrasive than the round ones. as well as their shape. The factors that enhance abrasive capacities of rocks are the following: The hardness of the grains of the rock.5 Abrasiveness Abrasiveness is the capacity of the rocks to wear away the contact surface of another body that is harder. as in a schist. The type of material that makes up the rock matrix and unites the mineral grains also has an important influence. Table 1-5: Relative quartz (common abrasive) content 1. It gives rough contact surfaces with local stress concentrations. Polymineral rocks. porosity etc. The rocks that contain quartz grains are highly abrasive.MGE 415 – Rock Excavation 2004 1. with their silica content and grain size. Table 1-7 shows characteristics properties of different types of rocks according to their origins Table 1-6: Common rock names and their geological definitions Table 1-7: Rock Properties according to origin Dessureault 1/14/2004 14 . are more porous.MGE 415 – Rock Excavation 2004 and. consequently. In Table 1-6 the classification of some types of rocks is shown. have low crushing strength and are easier to drill. and the shear strength or friction angle along the plane. For a detailed understanding of the rock mass behavior under stress and vibration. Recurring discontinuities are easily identifiable in the form of bedding planes. For regular or recurring cracks there are two further descriptors. or joints. the tensile or adhesive strength (often zero) at right angles to the plane.2. and the slope angle (dip) between the plane and the horizontal. fissures. cracks. irregular structures. cracks.7 Rock Mass Strength and Structure A rock mass.4. The strength of joints is normally considerably less than that of the adjacent rock. as distinguished from a rock strength test specimen. A description of the three-dimensional network of intersecting planes. and planes of weakness.MGE 415 – Rock Excavation 2004 1. is a body of rock with its naturally occurring network of flaws and discontinuities. or random. It is described by two simple measures. and of the resulting fragment size and shape. namely the average crack length and the average distance between parallel cracks. of how and why it fails. we also need a measure of the elastic or Dessureault 1/14/2004 15 . These are important for a proper understanding of the real ability of a rock mass volume to withstand load. joints. foliation partings. Intersecting groups are common such as parallel planes. has to include the compass bearing (called the strike) of its intersection with a reference plane (normally the horizontal). 1/14/2004 Dessureault 16 . (mandatory) 1. The rock breaking process is classified into three major groups: primary. Energy is important because it determines the rate at which rock breaking can be carried out. ADDITIONAL RESOURCES II The above information is additionally supplemented by: Lecture module 1.MGE 415 – Rock Excavation 2004 plastic deformability of the joint (its "spring constant") and the way the shear strength or friction angle varies with shear deformation and crack separation. As the breaking mechanism of the machine changes. Paul Lever’s 415 course notes. Each process is described below. Figure 1-6: Classification of the rock masses. The main considerations in breaking rock are the forces required to induce fractures in the rock and the energy consumed in breaking rock. secondary. All machines are limited in the power that can be applied to the rock and hardness of the manufactured components of the machine. so would the energy required to break the rock since the strength of rock varies depending on the type of stress induced on the material.2. Therefore a process that demands substantial energy will result in a slow rock breaking rate. 5 From Dr. and tertiary.3 Rock Breaking Processes5 This section is concerned with the basic principles of breaking brittle hard rock. Force is important because it determines the limitation on the type of machinery that can be used to break the rock and on the materials of construction that can be used in the machinery. 1 Primary This is the application of a force by means of a hard indenter to a free rock face much larger than the indenter. that is. Percussive drilling. The buttons are loaded slowly (quasi-statically) and are moved away to be re-applied elsewhere. Application of a hard indenter to the bottom of a hole. 6. Disc type cutters for raise and tunnel borers. Dynamic forces are applied 2. Forces at a point in the rock rise very slowly.3. indexing occurs by rolling to the next button. The force is applied from one side only and the bottom of the hole is the free face. 5. Diamonds produce very small fragments because they are small indenters.1. A very hard surface and very small indenter dragged across the surface. the rock surrounding the area immediately under the indenter provides confinement for the rock so that stresses are very much greater than the uniaxial compressive strength (confined stresses for rock is higher than uniaxial). The force applied dynamically and after each application the hard indenter is moved slightly to break out more chips on the next application 3. In the primary breakage process. Figure 1-8shows how the size of the indenter is directly related to the size of the indenter. Diamond bits. Drag –bit. A hard indenter forced onto the rock and indexed by dragging across the surface. 1. Dessureault 1/14/2004 17 . Repeated applications over a large surface area maintain the flat face 4. Hard indenter indexed by rolling. Button type cutters for raise and tunnel borers. This generates chips which are of a size similar to that of the indenter at the sides of the indenter and a pulverized zone immediately below the indenter.MGE 415 – Rock Excavation 2004 1. Figure 1-7: Primary Primary breakage processes would include the following: 1.3. Impact or hammering. The real breaking is done by the force thrusting the diamonds against the rock.1 Theory The theory of the forces and energy in the primary breakage process relates to the confinement of the rock and the energy of the indenter. As penetration increases. This may be due to the increasing confinement of the rock. Dessureault 1/14/2004 18 . E p a constant Kp a Note that these formulae are only valid when tools are driven to form the first chip. Ep stress strain strain. stress is directly proportional to the strain (Young’s modulus). forces and therefore required energy also increase. Indexing (moving through rotation of the bit or roller) the indenter to a new free surface would allow the indenter to penetrate with a force or stress closer to 1st chip formation. However. or 2 Ep Ep therefore E p 2 however.5 a2 The energy for primary breaking is directly proportional to the stress multiplied by the strain. this effect would essentially cause the button/indenter to be buried. Therefore the primary breaking energy is directly proportional to the square of the stress. considering an elastic behavior. since stress so finally. Practically speaking.MGE 415 – Rock Excavation 2004 F a Figure 1-8: Stress relating to size of indenter The stress to penetrate the rock depends on the size of the indenter (note that the nomenclature of this process is such that a is the width or diameter of the implement striking the rock): constant p a The force to penetrate is therefore is equal to: Fp Fp Fp stress area of indentor constant a k p a 1. 2 Example 1 Determine the boring rate of a 75 kW raise boring machine using roller cutters with buttons of 1 cm diameter and boring head of 2 m diameter. a chip forms with every pass. Some drill bit manufacturers sell drill button bits with ballistic or circular bits. Assume correct thrusting. With buttons of 1 cm the mean chip size will also be about 1 cm.MGE 415 – Rock Excavation 2004 3rd chip formed or Force 2nd chip formed 1st chip formed Penetration depth Figure 1-9: Increasing energy with depth Side note on indenter shape: From the above. that is. a 10 -2 m fracture energy E p Kp a 1. it can be seen that indenter shape directly impacts the energy required to penetrate the rock.1. Figure 1-10: Penetration depth of various button shapes 1.3.5 10 6 10 -2 150 MJ/m 3 Dessureault 1/14/2004 19 . This can be seen in Figure 1-10 which compares a ballistic shaped bit with a circular or dulled bit. Dessureault 1/14/2004 20 .3 m/hour hole area r2 (1 m)2 Note that from the previous 2 examples. the energy for boring with disc cutters is four times smaller and the boring rate is four times greater than with the button cutters.8 m 3 / hr 3 150 MJ/m 2000 hour 3 3 breaking rate 1.MGE 415 – Rock Excavation 2004 Power 75 kW 75 10 3 J/s Power energy/volume Power m3 / s fracture energy Breaking rate Boring rate 75 10 3 J/s 1 3600 seconds m3 / s 1.5 MJ/m 1000 hour 3 3 breaking rate 7. a 4 10-2 m Kp fracture energy E p a 1. a few centimeters of the disc are in contact with the rock as the disc rolls and chips are long and narrow with a width of about 4 cm.2 m / hr 7.8 m / hr 0.1.3 Example 2 Determine the boring rate of a 75 kW raise boring machine using disc cutters on a 2m diameter head.2 Secondary This is the application of forces inside a hole near to the rock face.5 MJ/m 3 Breaking rate Power energy/volume Power m3 / s fracture energy Boring rate 75 103 J/s 2 3600 seconds m3 / s 7. However.3. 1. Assume correct thrusting. The shape of the indenter is not very well defined with the disc cutter.2 m3 / hr 3 37.2 m / hr 2. The forces inside the hole generate tension at the sides of the hole to which produces cracks which ultimately run to the free surface.2 37.57 m/hour hole area r2 (1 m) 2 1.3. This is because the fragment size with the disc cutters is four times larger. Take the width of the chip as the value for a.8 m / hr 1.5 106 4 10. MGE 415 – Rock Excavation 2004 Free face Rock F Cracks Figure 1-11: Secondary breakage process Secondary breakage processes includes: 1. Wedging. Wedge driven into a hole which produces crack 2. Blasting. Explosive generates a pulverized zone through compression but the real breaking process is by driving tensile cracks. Specific theoretical aspects of energy required for secondary breakage will be discussed in further blasting aspects of the course. 1.3.3 Tertiary This is the application of forces from more than one side to a free surface. F F Figure 1-12: Tertiary Breakage Process Tertiary breakage processes include: 1. Breaking boulders by impact or mud blasting 2. Crushing 3. Milling According to theory, the tertiary breakage process is closely related to breaking the rock in tension. From Figure 1-13 is can be seen that loading of a sphere by diametrically opposed forces causes a uniform tensile stress across the diametrical plane. This causes the sphere to split in tension, that is, at a stress very much lower than the uniaxial compressive stress. Dessureault 1/14/2004 21 MGE 415 – Rock Excavation 2004 F t F t Figure 1-13: Tertiary Breakage a tensile effect It has been found that the tertiary stress, t, is also dependent on the size of the rock, but not as important as the size of the indenter for primary breakage. Larger boulders break at lower stresses, hence t can be represented as: t constant a 0.25 Therefore the splitting forces could be calculated as: Ft stress area of sphere t a constant a2 0.25 a k t a 1.75 It should also be noted that F >>Ft Finally, the energy for tertiary breaking is derived by: Et stress strain stress squared 2 t constant a 0.25 2 Kt a 1.3.3.1 Example Milling reduces rock to 70% minus 75 m. Milling typically consumes 25 kWh/t. Compare this value with that predicted by the simple formula for tertiary rock breaking processes. 1 kWh 1000 3600 min hr Density of rock 2.7 t/m 3 3.6MJ E milling 25 kWh/t kWh 3.6 MJ 2.7t m3 243 MJ/m 3 According to the formula, tertiary breaking energy requires: Et Kt a 1.5 106 a Dessureault 1/14/2004 22 MGE 415 – Rock Excavation 2004 If 70% of the rock is reduced to minus 75 m, then the mean fragment size is about 50 m. a Et 50 10 -6 m 1.5 10 6 50 10 6 214 MJ/m 3 The simplified Et formula gives a reasonably good estimate of energy consumed in milling. 1.3.3.2 Example A jaw crusher is driven by a 10 kW motor and is set to produce fragments of 1 cm. Determine its ‘crushing’ capacity in tons per hour. a Et 1 10 -2 m 1.5 10 6 10 2 15 MJ/m 3 Power Et 10,000 15 10 -6 3600 s hr 2.4 m 3 / hr crushing rate density of rock 2.7 t/m 3 crushing rate 2.4 2.7 6.5 t/hour 1.3.4 Miscellaneous Several other breakage processes exist, these include: 1. Thermal spalling. This depends on intense heat (flame) being applied to the rock and traversed so that a high temperature gradient is produce in the rock resulting in differential expansion which produces mechanical strains and ultimately breaking of the rock. It is used in taconite and certain quarrying operations, usually in cold climates. Thermal spalling is also used for finishes on rock surfaces and where high forces must be avoided during breakage. (commonly used ancient technique) 2. Water jets. The water jets create high stagnation pressures against the surface it impinges on. Used for drilling in porous hard rock where water goes into pores and breaks grains out. Water jets are known to be wasteful on energy and are used only for special applications. Dessureault 1/14/2004 23 MGE 415 – Rock Excavation 2004 Table 1-8: Summarizing Theory of Forces and Energy in Breaking Processes Breakage process Fracture Force (in Newtons) Fracture Energy (J/m3) Primary Tertiary Fp k pa 1.5 k t a 1.75 Et Kt a Ep Kp a Typical value k kp=108 Typical value K Kp=1.5 x 106 Only for a in meters and valid for a << 1.0 ADDITIONAL RESOURCES III Kt=1.5 x 106 7 kt=10 The above information is additionally supplemented by: Readings – Chapter 9.1 – Mining Engineering Handbook. (Mandatory, this covers modules 1 and 2) Lecture module 1.3. (mandatory) Assignment 1 Dessureault 1/14/2004 24 1995. Littleton CO. Howard L. 2 – Theoretical foundations. The Science of Industrial Explosives. This covers basically drills and explosives. Explosives Products 2. Environment.A. 1992. 28th Ed. New York: John Wiley & Sons. Sen. Cook. Ed. Gour C. Roger. 146 p. Pergamon Press: Toronto. Suface Mining 2nd Edition. 2nd ed.A.bris. ed. G. Holmberg. 1999.html Hustrulid. Ed. Blasting Technology for Mining and Civil Engineers. 1974. 570p. Explosives Introduction 4. Vol. Melvin A. 1992 Hartman. These various sectors are briefly listed: Mining : drilling and blasting of rock 6 These notes were assembled directly from the following references: From Dr. Persson. and Lee. Maintenance. 2000. SME Mining Engineering Handbook. Editor. Atlas Copco. Rock Blasting and Explosvies Engineering. 1990. R.ac.MGE 415 – Rock Excavation 2004 Module 2: Drilling and Blasting Components6 The following notes provide a summary of the tools and theory involved in drilling and blasting. Occupational Health and Safety. Potash Technology: Mining. Transportation. Balkema.1 Drilling Mechanisms Drilling is used in several industries and purposes. Processing. search: continuous mining Bell. Drilling equipment 3. F. Rotterdam: A. 449p. 2001 Class notes. Blasting Principles for Open Pit Mining. from Bob Cummings 1997 Class notes. SME Mining Engineering Handbook. Howard L. and Jan M. 1992 Kennedy. Howard L. Coal Mining Technology Theory and Practice. Edition 2002 Hartman. 2nd ed. William. The History of Explosives. This module is organized into four main sub-modules: 1.Wirtgen America Inc. The Raise Boring Handbook 2nd ed. Engineering in Rock Masses. 1983 McKercher.:Society of Mining Engineers. Jaimin. 1983 Mining-Technology. CRC Press: New York.M. http://sis. Sydney: University of New South Wales Press. Introductory Mining Engineering. Ed. Robert.. Dessureault 1/14/2004 1 . ML. Drilling mechanisms 2. from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department.com. ISBN: 0 7506 1063 8 Caterpillar Performance Handbook. SME: Port City Press. Butterworth-Heinemann: London. Utah. B. IRECO Chemicals: Salt Lake City. Surface Mining Manual. 2002. Per-Anders.uk/~dj9006/explosives/history. 534 p. Mutmansky. Paul Lever’s 415 course notes Hartman. Stefanko. 2nd ed. or mud. hammer. There are several losses of energy once the drill has converted the original energy into mechanical energy: Compression of steel Bending Elastic strain @ couplings Internal friction in the drill Heat Side friction of cuttings Dessureault 1/14/2004 2 . controls dust. small diameter (to create slabs) Construction: Intermediate drill sizes (typically to reduce excess damage. lubricates bit & hole wall. cools bit. Energy transmitter – (steel) – the steel is the primary mover of the bit Energy applicator – (bit) – the bit attacks the rock mechanically Fluid – cleans hole. due to the variances in hardness and for grade control May need different drills depending on the location of drilling (air pressures) Underground typically uses smaller holes Surface mines typically use very large diameter holes Drill selection can depend on bench height (surface) or direction of drilling (underground) The four main components in drilling include: Energy source (drill and compressed air) – where the drill converts energy (such as the potential energy in compressed air) into mechanical energy. rotary (collect chips as samples) Quarrying: including specialized. air. several features can be remarked: May need different drills for ore and waste. The fluid can be water. churn. stabilizes hole.MGE 415 – Rock Excavation 2004 Exploration: diamond core drilling. to maintain slope) In focusing on mining applications. percussive. 1. The peak stress applied to the bit by the drilling machine causes the bit to strike the rock at a stress calculated by: p cv Where: v .1 Percussive Drilling Percussive drill bits break rock by indentation. steel = 7850 kg/m3 c . Three components of the drilling process should exist at all times. If any one of these components is ineffective.25v MPa for units of m/s Figure 2-2 shows the mechanical components and the nomenclature of a typical percussion drill. Figure 2-1 shows the drilling process. the drilling process will deteriorate rapidly and the information obtained from the hole will be invalid.bit velocity at the rock face . The energy to break the rock is imparted through the drill bit by the compressive stress wave imparted by the piston to the steel. Dessureault 1/14/2004 2 .MGE 415 – Rock Excavation 2004 The two primary rock drilling functions are: Penetration – where the fracturing or breaking of the material from in-situ Circulation – where the debris is ejected.compression wave velocity of the impacting body (~5000 m/s) Hence.density of the impacting body. Figure 2-1: The drilling process 2. p=39. Either compressed air or water pushes the piston in the drill to strike the steel in a repetitive manner. the mechanics of energy along the drill steel are presented by the following discussion. and contact between the drill steel and the rock. The piston remains in contact with the steel until the tensile wave returns to the steel-piston interface and causes the two to separate. some of the energy is imparted into the rock to break out chips and the remaining energy returns as a compressive or tensile wave. Dessureault 1/14/2004 3 . This ends the pulse upon which the hammer imparts energy onto the steel. to the ends of the piston and drill steel where they are reflected. the wave is reflected as a tensile wave. Figure 2-3: Piston moving toward steel When the piston strikes the steel the kinetic energy is transformed into a compressive wave imparted to both the steel and piston as seen in Figure 2-4. The type of wave depends on the rock type. Calculating the duration time of the pulse has been used to calculate the yield strength for the drill before the machine would break. The process of imparting energy into the rock begins with the piston being forced at high speed into the drill steel as seen diagrammatically in Figure 2-3. whereas at a fixed end it is reflected as a compressive wave.MGE 415 – Rock Excavation 2004 Figure 2-2: Components of a percussion drill – piston and steel Issues of drill efficiency and energy losses associated to coupling. The reflection and separation effect is shown diagrammatically in Figure 2-5. Figure 2-4: Compressive waves in both piston and steel These waves travel at a velocity c. Since the drill steel is in contact with the rock (not directly fixed). For a bar with a free end (piston). The nature of the reflected wave depends on the boundary condition at the ends. Sources of energy loss include: Friction at coupling and other contact points. and the drill steel has a length l2.63v MPa.bit velocity at the rock face . these velocity limits are only for hydraulic drills as pneumatic drills cannot generate such high velocities. h = t = cv/2 = 19. where At and Ah are equal. Not all the energy in the fluid in drilling is expended in breaking the rock.compression wave velocity of the impacting body ~5000 m/s At is the cross-sectional area of the drill steel Ah is the cross-sectional area of the piston Therefore. the maximum piston velocity is equal to about 10.8 m/s. The length of the compressive wave set up in the steel would therefore be equal to 2l1 divided by the speed of the piston. Bit wear Noise Dessureault 1/14/2004 4 . and where the diameters of the piston and drill steel are the same. However. Consider the case where the piston has a length l1. steel = 7850 kg/m3 c . ts 2l1 c Some have observed that the yield strength of steel used for the piston and drill steel limits the stresses that can be applied to the rock by using the piston-type devices.MGE 415 – Rock Excavation 2004 Figure 2-5: Reflections and wave transmission.density of the impacting body. If it is generally desirable to limit the stresses in these steel components to about 212 MPa. The maximum stress in the piston h and the drill steel t are given by: h cv At At Ah t cv Ah At Ah Where: v . Carlos Lopez.5 ADDITIONAL RESOURCES IV The above information is additionally supplemented by: Lecture module 3.2 4 – 100 1. using water as the energy transmission method.25 19 .3 – 1. Balkema.32 1.A. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro.75 – 1. Hydraulic. Chapter 2: Rotary Percussive Drilling. meaning compressed air provides the energy to move the piston.1 (mandatory) Readings. The advantages and disadvantages of the hydraulic and pneumatic drills are listed below. well established 5% of the input power to the technology compressor is delivered to the Simple mechanical components rock) Most mines have compressed Poor environment for the air lines through the drifts machine operator (machine is noisy and lubrication causes a fog) More efficient (25-30% of Components more complex energy input power delivered to More capital outlay rock 8-10 decibels less than pneumatic drills (quieter) Table 2-2: Comparative table Drill type Jacklegs / stopers Drifter / Jumbo Typical Hole Sizes Inches mm 0. Table 2-1: Comparison of Hydraulic and Pneumatic drills Pneumatic Hydraulic Advantages Disadvantages Low cost Very low efficiency (less than Traditional. Most of the drills in service are pneumatic.MGE 415 – Rock Excavation 2004 Vibration. from: Jimeno. and Francisco Javier Ayala Carcedo. (Mandatory) Dessureault 1/14/2004 5 .2 – 30.25 38 – 57 Typical Hole Lengths Ft m 1–4 0. Emilio Lopez Jimeno. Flushing (positive work) Rotation of steel (positive work) Machine mounted drills are known as drifters but most commonly as jumbo drill rigs. drills are also used but are technically more sophisticated and are typically found on jumbo drill rigs.50 – 2. Rotterdam: A. the bit attacks the rock with energy supplied to it by a rotating drill stem. The drills typically operate in the vertical position although many types can drill up to 25 to 30 degrees horizontal. Figure 2-7 provides the deck plan for a P&H 100 rotary drill. Figure 2-6 provides the nomenclature and dynamic movement for rotary surface blasthole drills. or a crushing and chipping action in hard rock.MGE 415 – Rock Excavation 2004 2. Figure 2-6: Dynamics of surface blasthole drill Dessureault 1/14/2004 6 . The air cools the bit and provides a flushing medium for the cuttings.2 Rotary Drilling In rotary drilling. The bit breaks and removes the rock by either a ploughing-scraping action in soft rock. or a combination of the two. The drill stem is rotated while a thrust is applied to it by a pulldown mechanism using up to 65% of the weight of the machine. Compressed air or water is supplied to the bit via the drill stem.1. Figure 2-8: Selection of Drilling method Dessureault 1/14/2004 7 .MGE 415 – Rock Excavation 2004 Figure 2-7: Floor plan for rotary drill P&H 100XP Rotary drilling is one of the most popular drilling techniques for large surface mines where large diameter holes are used in blasting. Figure 2-8 shows how rotary tri-cone bits are the ideal application in hardrock surface mines. fluid. and cuttings The drill tools and bit that advance and shape the hole and provide samples. other run blades vertically along the stabilizer length. longer life. Blade stabilizers are designed to center the drill collars in the hole. Roller stabilizers have throwaway or replaceable rollers. blade and roller (Figure 2-10).1. The bit is steadied and hole maintained on grade by the use of stabilizers. Figure 2-9: Components in rotary drilling 2.1 Stabilizers The efficiency of rotary cutting is lost when the drill pipe is allowed to bounce around or when it fails to run smoothly in the center of the hole.2. They have the advantage over blade stabilizers by providing better stabilization. Dessureault 1/14/2004 8 . They are run between or above the drill collars.MGE 415 – Rock Excavation 2004 2. Blade stabilizers come in many forms.1.2. There are two primary types of stabilizers.1.1 Rotary drill string components The drill string consists of three main systems: The suspension and drive that connect the string to the rig and power system The drill pipe that transmits power. and lower torque requirements than blade types. some have spiraled blade arrangements. The smooth bore produced with adequate stabilization permits faster rock bit and drill steel retraction from the bore hole. ledges and other unconformities are obtained. The inner row cutting structure is more uniformly worn and inner and outer flank wear is reduced. Without stabilization. Stabilization assures that the bit will do this and thus cause the energies and forces exerted on it to be most efficiently utilized in an axial direction. Gage wear is uniform and less severe. the smooth bore sloughs less than a rough bore. Rock Bits are designed to rotate about their own center. Stabilizers also affect drill availability. More footage per bit and an increased penetration rate is achieved by proper utilization of the forces applied to the bit. Dessureault 1/14/2004 9 . These costs include reduced cost per ton of rock removed. The possibilities of crooked hole are enhanced. Drill steel rotating in these rough and crooked bores scrubs and scrapes against the bore wall and thereby abrades. Dull conditions of bits give evidence of this effective stabilization. Lateral movement or stumbling is restricted and hole is produced in the direction intended.MGE 415 – Rock Excavation 2004 Figure 2-10: Blade and roller stabilizer The advantages of using stabilizers should be considered when designing or selecting a drilling process. the more efficient use of expendable items such as rock bits and drill steel can be achieved. Shirt-tail wear problems are reduced. Re-drill time required is normally eliminated or drastically reduced. This means that less rock particles fall to the bottom of the hole. Stabilizers also affect drill stem life. rough spiral bores. The primary reason is costs. Furthermore. Theoretically.2. The stabilizer should therefore be held at the largest diameter practicable. They are also used at points of heavy wear to provide a readily replaceable thread. the guiding elements should have the same diameter as the bit. Scraping and gouging of the guiding elements are eliminated and co-centricities are maintained.as circles inside one another) of guiding elements with the axis of the bit and steel is also quite important to proper stabilization. Unfortunately this is not practical because of the normal attrition of rock bit gage wear surfaces. Concentricity (the quality of having the same center . Rib-type stabilizers. The conventional welded rib or cast rib stabilizer (also known as blade) does an adequate job when new and unworn. Due to the cost in maintaining close-to-hole wall contact these stabilizers are left in the drill string long after adequate stabilization has been lost. Unless the guiding elements of a stabilizer are very near the bore wall. provided the ribs are concentric. The stabilizer which maintains guiding elements close to hole wall is most efficient. Shock absorber subs are run above hammers or above the bit or rod in rough drilling situations. Roller-stabilizers provide adequate stabilization without imposing additional torque 2. necessarily drag and scrape against the bore wall and abrade rapidly. Eccentricities (different centers) of these elements tend to void any hope of reducing drilling costs with a stabilizer.MGE 415 – Rock Excavation 2004 In the selection of a stabilizer. A swivel mount shock sub is also used to dampen vibration effects on the rig structure thereby reducing maintenance on the rig and mechanisms for pulldown and rotation.1. adequate stabilization of the bit is not achieved. thus reducing drill string vibration and increasing bit life. Figure 2-11: Swivel mount shock sub Dessureault 1/14/2004 10 .1. The most efficient stabilizer is one with true-rolling centralizers that are in rolling contact with the bore wall. one must take into account the adequate stabilization. The problems with this type of stabilizer are wear rate and maintenance cost.2 Substitutes (Subs) Substitutes (known as subs) are used as adapters where threads of one size or type must be coupled with threads of another size or type. due to their construction. C. no modification of the drill is required and installation can be completed in a short time. Increases bit life by dampening cyclical shock loading normally transmitted to the bit bearings and cutting structure. Normally. The absorbed energy is released to the atmosphere in the form of heat. The pulldown or drilling weight to the bit is transferred through the segment rubber placed in shear.1.3 Rotary Drill Stems Factors involved in selecting drill stems (aside from bailing requirements) includes: Fabricated (welded) or integral (machined from single steel bar) drill steel. On multi-pass drills where length is not critical. B.2. The accelerated changes in axial and torsional loadings are minimized by the segment rubber.1. Increases drilling rates by keeping bit in more uniform contact with formation.MGE 415 – Rock Excavation 2004 Shock and vibratory energy transferred between drill and bit are absorbed through a series of rubber bonded segments. A. Figure 2-12: Vibration reducing effects of using shock subs The shock absorber is most beneficial when drilling in fractured formations. intermittent hard and soft layers or hard formations. 2. Decreases operator noise level by eliminating the metal to metal contact between rotary drive and drill pipe. The benefits include: Reduces drilling machine maintenance by dampening torsional and axial shock loads. Thread size and type Wall thickness Types of connections Dessureault 1/14/2004 11 . a saver sub is recommended for use between rotary shouldered connection and drill pipe. Allows use of more weight and higher rotary speeds in rough drilling areas. The rotational torque (or torsional loading) to the bit is transferred through the segment rubber placed in compressive loading. Degree of weathering.MGE 415 – Rock Excavation 2004 Figure 2-13: Blasthole components with respect to drill rig 2. Grain size.2 Rock drillability / Penetration rate Rock drillability is defined as the penetration rate of a drill bit into the rock.1. Texture. It is a function of several rock properties such as: Mineral composition.2. Dessureault 1/14/2004 12 . penetration rate is one of the most important factors in drilling. AQUILA’s drill performance monitoring Dessureault 1/14/2004 13 .g. The Bauer and Calder method states that penetration rate. in thousands of pounds Rpm = revolutions of drill per minute The adjustable factors in the above equations are the variables controlled by the operators. Rock compressive strength is not but can be estimated using the graph in Figure 2-14. can be calculated using the following equation: P (61 28 log10 S c ) W rpm 300 where p = penetration rate in ft/hr Sc = uni-axial compressive strength. in thousands of psi W/ = weight per inch of bit diameter. P. e. Figure 2-14: Penetration rate vs.MGE 415 – Rock Excavation 2004 As is obviously important to costs and productivity. Rock strength Several other indices exist for rock drillability which include: Drilling Rate Index. Some empirical equations have been developed from extensive rotary tests in iron ore. Classification of rock types based on the drillability of Barre granite Mohr’s test Indices (sensor-measured). The bit may be turned by: 1. Figure 2-15 illustrates recommended bit loadings for different bit sizes. Overloading the bit results in severe loss of bit life as illustrated in Figure 2-16. pulldown weight for 9 1/4 inch diameter rotary bits in hard formation There are several method of rotary drive. rotary head and drill stems to the drill bit.MGE 415 – Rock Excavation 2004 2. the bearing size increases thus allowing an increase in the tolerable load.2. Figure 2-15: Recommended pulldown weights per inch of bit diameter. As the bit diameter increases. Dessureault 1/14/2004 14 . rotating a rotary table which turns the pipe as it slides through.1.3 Rotary Drill Pulldown Weight A portion of the machine weight is applied by the pulldown motor via the pulldown chain or chains. Figure 2-16: Bit life vs. using a down hole turbine drive.MGE 415 – Rock Excavation 2004 2. Figure 2-17: Rotary drive types Figure 2-18: Pulldown mechanisms Figure 2-19: Rack and pinion pulldown Dessureault 1/14/2004 15 . 3. rotating the bit. rotating the pipe directly by a drive unit which moves down with the pipe. A bailing velocity of 1800 mpm/s (6000 fpm) is usually adequate to bail 13 mm (1/2 in. Table 2-3: Average Drill Bit Footages (from large Iron Ore mine in Canada) 2. the chips begin to move.MGE 415 – Rock Excavation 2004 General parameters for drill penetration can be seen in Table 2-3.4 Flushing medium Air is used to bail the drill cuttings from the hole as well as cool the bit bearings and. The velocity of the drill cuttings in this air is dependent on the chip size. Figure 2-20 illustrates a typical air requirements chart. The air volume is the primary requirement for bailing cuttings from the hole. Approximately 20% of the air is forced through the roller cones for cooling purposes by adjusting the air pressure across the bit using the bit nozzles. when used.1. Dessureault 1/14/2004 16 . the balancing air velocity in feet per minute is given by: Um = 264p½d½ Where: d = diameter of the chip in inches p = density of chip in Ib/ft3 At air velocities above this balancing value.) chips. Air velocity up the hole is dependent on the air volume per minute as well as the hole annulus (ring-shape where hole and stem meet). roller stabilizer bearings. and shape. density. their velocity being approximately one half the excess air velocity above the balancing value.2. Experimentally. will help cater for hole cavities. 5. reduce the volume of cuttings in the hole for a given penetration rate. will give increased stabilizer and pipe wear. and 7. etc. increased dust deflector and deck bushing wear 3. The drawbacks of increased bailing velocities include: 1. give higher bailing velocities. 6.. tend to give higher bit life. 2.MGE 415 – Rock Excavation 2004 Figure 2-20: Air Bailing Chart Factors involved with choosing the air velocity are that higher velocities: 1. 4. may give higher penetration rates and possibly lower cost per ft. 2. 3. will bail larger chips. will help cater for drill stem wear. may damage borehole walls in soft drilling Dessureault 1/14/2004 17 . After an idle period and before reusing a bit. ground-bearing limits. Nonuniform cone temperatures may indicate obstructed air courses and potential bearing failure. 4. 16. This can pinch the cones. other equipment. 2. Watch for people. reduce down pressure and rotation. care must be taken when blasthole drills are moved around the job site. 2. High speed impact is a common cause of damage. Break in a new bit by drilling at reduced down pressure and rotational speed for a short period. 1.MGE 415 – Rock Excavation 2004 2. Re-establish bottom hole pattern with reduced down pressure and rotation when drilling is interrupted. 6. width and weight of the machine. 10. Open air valves before starting to drill with a bit and keep the air on until the bit is out of the hole.Before an idle period. 15.2. 8. Use straight drill steel with properly maintained threaded connections.Rotary speed should be decreased as down pressure increases. Know the height. and steep hills or uneven terrain.2.Do not use more water than is necessary to control dust and maintain hole wall.6 Safety Tips The following safety tips should be taken into account around large surface blasthole drills: 2.In wet holes. To collar or start new hole. and gauge teeth.1 Moving the Drill As with all heavy equipment. Always secure on-board equipment before traveling. damaging the bearings. 13.1.1.6. low or narrow clearances. 14. 9. Never finish an old hole with a new bit. clean the bit by passing air or water through it while rotating the cutters by hand.2 High Voltage dangers 18 Dessureault 1/14/2004 .2. overhead electrical power lines. 5. 3. 11. Use a signal man when traveling and use an audible warning signal to alert people in the area before traveling.Guard against dropping the bit and drill steel to the bottom of the hole at high speed. 7.Maintain rotation while tripping into or out of a hole.2.1.Periodically inspect the bit for damage or impending failure. make sure all cones turn freely by hand. Use a good grade of thread grease and maintain connections properly.1. maintain as high an air pressure and volume as is possible.6. 12.5 Operational tips The following are some operational tips on bit and rig for top performance when using air circulation bits. Use care attaching and removing a bit from the drill pipe. Always maintain drilling air pressure at appropriate levels. Do not travel with people riding on the outside of the machine or inside the machinery house. 2. propel the machine with the operator's cab on the downhill side of the machine. After storing the mast.6.000 volts or less. When the mast is down.5 Crushing Danger from Lowering Mast Before lowering the mast. Some local regulations require greater distances than OSHA or MSHA. keep all people off the drilling platform and drill mast. always determine the slope limitations for your specific machine and never travel or operate on site slope conditions which exceed these limitations.2.2. and away from drill stems. Greater clearances are required for lines with higher voltages.1. Moving components or rotating drill stems can entangle clothing and can crush. Maintain cable slack while operating and traveling.2. always lower the machine slowly and in stages.6. lowering. Rotational Dangers During drill operation.6 Chips and Dust Drilling produces flying debris and dust which can cause serious respiratory disease.3 Trailing Cables Blasthole drills with electric tail cables can break the cable connection or crush the cable when traveling. 2. notify personnel to evacuate the roof and drill deck and inspect mast storage area for obstructions. 2.MGE 415 – Rock Excavation 2004 Hazardous voltage from the machine contacting power lines can cause electrocution and burns to anyone near the blasthole drill. MSHA and OSHA regulations require at least ten (10) feet of clearance form overhead lines carrying 50.2.1. maintaining a level condition until the crawler belts contact the ground and are supporting the machine. crawlers and machine during the process. Always lower dust curtains before drilling and inspect the curtains to Dessureault 1/14/2004 19 . or leveling the machine with jacks.1.4 Tipping Dangers Exceeding the slope or grade limitations given for your specific machine and machine configuration can cause machine tip-over. When propelling with the mast elevated. Sudden tip-over can occur when raising. Inspect the ground for lifting support and add cribbing or support mats before lowering jacks. Always level the machine at the lowest practical height that will unload the crawler belts.6. pinch or strangle personnel. Prior to propelling. 2. 2. When lowering the machine to the ground. Use a signal person to assist in watching the jacks. Use signalmen during traveling to prevent damage to the cable or cable connection. always position the machine with the mast on the uphill side of the machine.1.6. check for secure mast attachment on the front jack caps. For best chips. rapid clearing and transport up the hole to prevent regrinding of the chip. either by blast design or through purchasing equipment with taller masts. eating or drinking. 2. cavernous or other formations prone to lost circulation problems. large chips often make better samples. fewer chips are produced and the total energy used per metre of hole drilled is less (less surface area produced.MGE 415 – Rock Excavation 2004 make sure all dust curtains are installed and in the lowered position. Large chips need: bits with sharp cutting edges. exposure limitations.2. Keep personnel off drilling platform and away from the drill hole while drilling is in process. low solids fluid (must be balanced against erosion of hole wall). Use high thrust or feed. or structures which must be moved or avoided during movement of the drill. Single pass drilling is Dessureault 1/14/2004 20 . Use reverse circulation techniques for broken. Although larger amounts of energy are required for each large chip. Use high flushing flow of low viscosity. Keep hydrostatic head to a minimum. large amounts of energy per chip. The best chips are large chips. and protection methods for crystalline silica exposure.8 Single pass Most drills in surface mines use single pass drilling.1.2. Curtains should not be lifted to remove cuttings while the drill is drilling in the hole. fractured. Specific design/plan-related safety issues to consider Where and when the blasting is scheduled for the day The other equipment. Follow mine procedures for air monitoring. The ground surface strength in terms of having adequate weight bearing ability for the drill mass The grade of the slope which the drill will be working on (too slopedwill cause a tip-over) Steps necessary to keep unnecessary people and equipment at a safe distance from drill area. Many rigs produce good chips at the bottom of the hole but fail to clear them or lift them out of the hole before they are broken up.1. Usually penetration efficiency is improved when large chips are produced and cleared quickly.7 Chip Sampling Penetrating always involves" chip making". powerlines. Avoid dust contamination from work cloths. 2. the operator should use the drill as follows: Use sharp blade bits or long toothed roller bits with bottom clearing nozzles. 1. idle periods.9 Capacities. Key operating capacities for a drill rig includes: Maximum mast loading capacity Maximum hoist or hook load capacity Head torque capacity. Disadvantages of single pass are: High masts make the drill more instable Extra care is needed in moving drill long distances Pulldown chains become long and may require special attention 2.MGE 415 – Rock Excavation 2004 where the entire length of the hole is drilled without having to add more drill stem. The advantage to single pass drilling includes: Eliminates adding stems Reduces associated thread damage Reduces machine downtime for rod changing equipment and tool racks Facilitates the cleaning of boreholes Permits a continuous air flow through the bit at all times. bit and stem changes penetration rate Burden and spacing bit diameter (alters the penetration rate) bit life (changes the frequency of bit and stem changes) hole depth (changes the frequency of the movement time) number of holes (changes the frequency of set-up time) Dessureault 1/14/2004 21 . Productivity calculations would be site-specific and should take into account: set-up time. movement time. 2.2.2.10 Productivity Estimate Figure 2-21 shows a generalized productivity estimate for various materials.1. Drilling records or statistics are therefore important to record so that accurate budgeting can be undertaken both in terms of costs that are incurred along with the output from the drill. warehousing. The reason for this split is that the drill consumables cost. These consumables are dependant on penetration rate and wear rate and would include items such as stabilizers. and drill bits (at times one of the most significant cost). predominantly the drill bit cost. the drill consumables per unit length of drilled hole and the balance of the drill cost. direct and indirect labor. is independent from the drill penetration rate. and consumables. Determining the drilling costs must take into account. the following factors: Ownership costs: include amortization and depreciation. whereas all other costs are penetration rate dependent.MGE 415 – Rock Excavation 2004 Figure 2-21: Tons drilled per operating hour for rotary drills for various hole diameters. drill stems. Dessureault 1/14/2004 22 . and taxes and insurance. The balance of the drill cost is converted to the drill cost per unit length of hole using the drill penetration rate as the drill cost would be time dependent. assuming correct operating methods (poor operation can cause increased bit wear life). maintenance. The drilling costs are made up of two different parts. Operating costs: includes power. Figure 2-22 shows an estimate of the cost per meter. interest on borrowed money (if not purchased out of cash-flow. GPS based navigation systems are available from companies such as AQUILA (Caterpillar owned).2. Figure 2-23: Cost per ton of material excavated 2.1.11 Drilling technology Most advanced modern drilling technology focuses around the navigation and positioning of the drill. There are five main options for the drilling systems: DM-1: Production Monitoring System DM-2: Material Recognition System DM-3: Drill Control System DM-4: Guidance system for vertical drilling DM-6: Guidance system for inclined drilling Dessureault 1/14/2004 23 . or can be installed as retrofits.MGE 415 – Rock Excavation 2004 Figure 2-22: Rotary Drilling costs estimates by meter drilled and diameter of hole. AQUILA DM systems are available in two ways: as factory installs on new drills. The company has a both a drill and shovel product line that uses GPS and PLCs to enhance drilling productivity and provide susbstantical information from which better mine planning can be undertaken. so the operator doesn't have to zero the bit depth. The result is immediate. Figure 2-25: DM-2 Material Recognition System Dessureault 1/14/2004 24 . which means faster work with fewer errors. then delivers real-time information via a color LCD screen.11. accurate-to-the-centimeter information for both the operator and planning engineers in the mine office. Figure 2-24: DM-1 production Monitoring System 2.MGE 415 – Rock Excavation 2004 2.1.2. and determine hole geology while drilling. The DM-1 is designed to minimize the amount of operator effort. DM-2 pinpoints the location of ore and waste interfaces.1. eliminating errors in determining hole depth.2 DM-2 Material Recognition System DM-2 uses a vibration sensor and pattern recognition software to automatically process and analyze drill variables.1 DM-1 Production Monitoring System DM-1 uses AQUILA's Advanced Monitoring Platform (AMP) with Graphical User Interface (GUI) software to give the operator immediate feedback on drilling productivity and performance. Steel changes are also automatically sensed.2.11. For example. the start of drilling can be automatically detected. 5 DM-5 Guidance System The DM-5 Guidance System for Vertical Drilling combines high-resolution. real-time geological information on a location-by-location basis. More consistent hole depths and more even benches.1. calculates the required drilling depth and displays the information on an easy-to-read LCD screen.3 DM-3 Drill Control System DM-3 delivers consistent drill operation and performance in all kinds of conditions.11.MGE 415 – Rock Excavation 2004 DM-1 and DM-2 functioning in tandem would enable the following: Detailed. By interfacing the DM-1 or DM-2 to Programmable Logic Controller (PLC) modules and electrohydraulic actuators. Wireless connectivity to and from the mine office. And the ultimate benefits of longer drill bit life and more productivity. This allows the operator to place a blasthole within centimeters of his target. better fragmentation and reduced ore dilution.4 DM-3 provides: Reduced maintenance and consumables costs. 2. without the need for traditional surveying or staking.2. 2. Once the drill is leveled and the hole started. Figure 2-26: DM-3 Drill Control System 2. This means more consistent performance. Comprehensive production reporting. with optimized penetration rates for changing downhole conditions. for operators of all skill levels.2.1. It regulates pulldown pressure and rotary speed within the most productive limits for torque and vibration. the DM-5 automatically determines collar elevation.11.11. This technology allows the operator to navigate to each 25 Dessureault 1/14/2004 . Improved explosives usage.2.1. Optimized operation for overall productivity. DM-3 makes automatic hole drilling possible .from collar to design depth. And all the information on blasthole positions is stored by the DM-5 and transmitted to the mine office for use in blast design and updating the geological model. Real-Time Kinematic (RTK) GPS receivers and an AMP platform. 6 DM-6 Guidance System for Inclined Drilling 26 Dessureault 1/14/2004 . The positioning information is provided by differential GPS. Figure 2-27: Operator’s view when navigating The location where the drill finally punched the hole. from two receivers on the head of the mast of the drill.2.1.MGE 415 – Rock Excavation 2004 blasthole via a navigation screen.11. Figure 2-28: GPS receivers 2. along with sensor data (that can be interpreted as geological information) can be recorded and sent back to the engineering office. The figure below shows the view provided to the operator during the navigation phase of drill positioning and the second picture shows the operator’s view when within 1 meter of the designed hole location (the view automatically zooms-in). Dessureault 27 . And like the DM-5.7 Case Study: Highland valley Copper mine A post-audit was undertaken in 1998 of the implementation of AQUILA technology on three B&E 49Rs at the Highland Valley Copper mines.) The most important change incurred by the implementation of the technology was the increase in collaring accuracy (previously. 6. This has also tended to increase the productivity of the drill through a reduction in tramming and positioning time. By designing the blast pattern on an office computer (as opposed to on-site by the surveyors). engineering considerations can more readily be taken into account.1. Design time has been reduced (no surveying is required). The elimination of pre. 9. Adverse climatic conditions no longer impede the blast design and implementation. the position of each hole must be known so that bootlegs are avoided.11. 2. Using both DM-5 and DM-6 allow: Greater hole location and depth accuracy for both inclined and vertical drilling. 4. Any need to re-survey the drilled pattern was eliminated (according to mine regulation. Supervision was increased as the productivity and delays are recorded and delivered in the log file. 3. Much less skilled drill operators can operate the drill. HVC was able to place up to five surveyors (who formerly would be marking blast patterns) on other revenue generating projects within engineering. There is consistency in blast design (only one technician does all of the designs).and post-blast pickups. 7. the DM-6 uses a servo-controlled Automatic Leveling System (ALS) that automatically maintains the GPS antennas level with the horizon. but is specifically designed for use on machines that routinely drill inclined blastholes with angles between five and 30 degrees from the vertical. For inclined drilling. Time and money savings from no-stake marking and better blast results. errors of 1/14/2004 8. 5. It should be noted that further development or utilization of the the data provided by the drill were used in other initiatives.2. since positioning the drill is made easy through the ‘video game’ type navigation interface. this system also stores all blasthole positions. 10.MGE 415 – Rock Excavation 2004 DM-6 has the same basic features as the DM-5. which are transmitted to the mine office for updating the design file. 2. The DM-6 also uses software designed to facilitate drill setup and improve alignment accuracy when drilling inclined holes. The benefits achieved after only 1 year of utilization are listed below. 1. The creation of a pattern no longer requires a large amount of survey consumables. the current average error is 10 cm in X & Y).1 Detonation Theory Blasting theory is one of the most controversial topics in the rock excavation industry. Engineering staff consider that such oversize caused increased wear on shovels. The misalignment of the blastholes induced poor fragmentation. resulting in decreased recovery. It was also observed that poor fragmentation would adversely affect the mill. No single concept has been developed and accepted that fully explains the mechanisms of rock breakage in every situation. Rotterdam: A. exothermic reaction.3.MGE 415 – Rock Excavation 2004 up to a meter in an X & Y direction were common. under elevated temperature and very high pressures.2 Drilling equipment This sub-module will familiarize students with the equipment that. It is the sudden rise in temperature and pressure from Dessureault 1/14/2004 28 . yet a vast amount of research work has contributed valuable information and insight into blasting theories. Carlos Lopez. Throughput would fluctuate when processing such ore. ADDITIONAL RESOURCES V The above information is additionally supplemented by: Lecture module 3. Some of the theories and findings are discussed in this section. (optional) Assignment 4: Drill selection 2. Balkema. they will manage as part of a wider fleet to ensure system efficiency. The stable end products are gases that are compressed. and Francisco Javier Ayala Carcedo. as engineers. The second sub-module will cover the properties of explosives products.1 (mandatory) Readings. Available upon request 2. trucks. ADDITIONAL RESOURCES VI The primary component of this sub-module is provided by: Lecture module 3. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Emilio Lopez Jimeno. An explosion is a self-propagating.2 (mandatory) PowerPoint presentations: material developed for 2002’s 415 Rock Excavation course.3 Explosives Introduction The physics and chemistry behind explosives and their detonation are introduced in this first sub-module on explosives. Chapter 4: Rotary Drilling with Tricone Bits.A. 2. from: Jimeno. and crushers. are formed resulting in nonideal detonations and somewhat less than ideal pressures and energies. It is associated with heat only and carries no shock due to its much slower reaction rate. Deflagration is the chemical burning of explosive ingredients at a rate well below the sonic velocity.on one side of the wave is the un-reacted explosive and on the other. and (c) a steady state wave where pressure and temperature are maintained. (b) a rapid pressure rise at Y leading to a zone in which chemical reaction is generated by the shock. An oxygen-balanced mixture is one in which there is no-excess or deficiency in oxygen. The maximum energy release upon detonation occurs when the explosive mix is formulated for oxygen balance. Jouquet as a space. or a detonation traveling through the un-reacted explosive. CO (carbon monoxide). which is commonly referred to as the Chapman-Jouquet (C-J) plane. The velocity of detonations (VOD) lies in the approximate range of 5000 to 30. Deflagration occurs when less than ideal hole-loading conditions or explosive formulation are involved. such that the gaseous products formed are chiefly H2O (water vapor). Commercial explosive formulation attempts to achieve an oxygen-balanced mixture. Chapman and E. This condition for stability exists at hypothetical X. CH4 as NO (nitric oxide). the exploded gases.000 fps (1500 to 9000 m/s).L. The work done by chemical explosives in the fragmentation and displacement of rock depends on the shock energy as well as the energy of the expanding gases. The self-sustained shock-wave produced by a chemical reaction in a gaseous medium was described by D. In actual blasting practice. CO2 (carbon dioxide). NH4 (methane).MGE 415 – Rock Excavation 2004 ambient conditions that results in a shock wave. well above the speed of sound in the explosive material. as seen in Figure 2-29. small amounts of noxious gases such (ammonia). This space of negligible thickness is bounded by two infinite planes . Figure 2-29: Detonation Process of explosive cartridge There are three distinct zones: (a) the undisturbed medium ahead of the shock wave. and solid carbon. and N2 (nitrogen). and proceeds until complete at X. Between the Dessureault 1/14/2004 29 . A simplified and approximate velocity of detonation (VOD) can be obtained from the following empirical relation: Cd J (1 1. producing an increase in pressure. pa. At the end of this zone. The steady-state chemical reaction takes place behind the shock front within the reaction zone.3 ) Cd is the VOD in m/s J is heat of reaction in MJ/kg is the specific gravity Where: The detonation pressure Pd in N/m2 which exists at the C-J plane can be estimated with comparable accuracy using the equation below: Pd 2 Cd 4 Explosion pressure. a nonsteady-state region exists. Note Figure 2-30 as being the idealized detonation wave traveling through a cylindrical explosive shape. It is created by a flow of expanding gases in a direction opposite to that of the traveling wave front. momentum. 2. where the reactions are considered complete. These are: p – pressure V – velocity T – temperature E – internal energy or Q – heat of formation and – density Dessureault 1/14/2004 30 .MGE 415 – Rock Excavation 2004 two planes X and Y there is a conservation of mass. and energy. Once again. This is also the plane where all the thermodynamic properties are calculated. The explosion pressure is approximately half the detonation pressure.1.1 C-J Plane A deeper exploration of the individual zones in the C-J theory should be undertaken. is the pressure when the product gases have the same density as the unreacted explosive.3. the C-J plane is seen as the boundary between the steady and non-steady state. MGE 415 – Rock Excavation 2004 Figure 2-30: Detonation Process for cylindrical explosive The maximum pressure and duration of a wave pulse is directly proportional to the shock energy and gas pressure of the explosive. as seen in Figure 2-31. Figure 2-31: Pressure shape for A-high explosive and B-commercial explosive containing high gas volume Dessureault 1/14/2004 31 . In this case. The pressure pulse for less-sensitive commercial explosives shows a decreased pressure amplitude and a longer pulse length. They liberate gaseous products very quickly. respectively. High explosives such as military explosives or highly sensitive commercial explosives are characterized by an intense shattering effect upon detonation (known as brisance). The distance between the shock front and the C-J plane is very short and results in a pressure pulse of high amplitude and short duration. the reaction is slower and the gas volume is greater. When the acoustic impedance ratio is 1.3. then radial failure from the explosive source is evident. the material in which the energy wave originates is labeled A. When the impedance ratio is greater than 1. As seen in the figure. These possibilities depend on the material's acoustic impedance Ir which is defined as the product of the material's density and sonic velocity V. nearly all of the energy will be reflected back as a tension wave. most of the energy will be spent in peeling off layers of rock at the free face.1. there could be three possibilities at the interface as seen in Figure 2-32. some of the energy is transferred to material B as compression waves. and the material into which the wave travels. Figure 2-32: Shock wave propagation Dessureault 1/14/2004 32 .2 Quick note on Shock Wave Propagation As the stress wave front generated from a blast travels outwards.MGE 415 – Rock Excavation 2004 2. When the ratio of acoustic impedance of material A to material B is less than 1. B. all the energy is transmitted into material B as compression waves. If the distance between the free face and the explosive charge is relatively small. it has a tendency to compress the material at the wave front through a volume change. When the compression wave front travels from one material to another. When a compression wave traveling through a rock medium encounters an interface such as a free face. whereas the rest is reflected back as compression waves as well. If this tangential stress is strong. some of the energy travels through the interface as compression waves and the rest is reflected back as tension waves. There is a tangential or hoop stress at right angles to the compression wave front. MGE 415 – Rock Excavation 2004 2.3.3.1 Fume Class Class-Fumes are noxious gases that are produced from the detonation of explosives. borehole pressure. 2. and strength.33 0.3. sensitivity. Density is most useful in determining the loading density or the weight of explosives one can load per unit length of borehole (in pound per foot or kilogram per meter). in waterfilled holes. and often the reverse is true. Table 2-4: Standards for Fume Class Class 1 2 3 Volume of poisonous gas per 200g of explosive. Many factors affect the volume of poisonous gas produced including oxygen balance and adverse loading of explosives. The now-disbanded US Bureau of Mines (USBM) limits the volume of poisonous gases produced by permissible explosives (those used in underground coal and other gaseous mines) to 2. Explosives with a density less than 1 will float in water. for water-based explosives. detonation pressure. temperature effects.44 lb (200 g) of un-reacted explosive. Institute of Makers (IME) has developed a fume class classification scheme as seen in Table 2-4. The production of these gases is most critical in underground and other confined workings.2 Density The density of an explosive is defined as the weight per unit volume or the specific gravity.2. These properties include fume class.16 0. this is not the case.14 kg).5 lb (1. water resistance.2.16-0. Dessureault 1/14/2004 33 .2 Comparative Explosive Properties Explosives and blasting agents are characterized by various properties that indicate how they will perform under field conditions.33-0. Therefore. Commercial explosives range in density from 0. and is calculated in English units as: LD = 0. density correlates to the energy released in a given borehole volume.3405 Where: D2 is density D is explosive column diameter in inches. The fume class is a measure of the toxic gases in cubic feet per 0. in ft3 0.5 to 1. However. detonation velocity. Each of these properties will be covered.7. density. an explosive with a density greater than 1 is required. For certain granular explosives such as dynamite. Note that knowledge of loading density is required for blast-design calculations.67 2. and less than ideal energy and pressure evolve from the slower detonation rates.2. 2. Dessureault 1/14/2004 34 . For most explosives. the strength of an explosive increases with detonation rate. For high explosives such as dynamite. there is a minimum diameter Dmin below which detonation velocity increases nonlinearly with increasing borehole diameter as can be seen in Figure 2-33. contributing to the formation of noxious fumes and lower energy output.5 Detonation Velocity The detonation velocity is the speed at which the detonation front moves through a column of explosives.3 Water Resistance The ability of an explosive to withstand exposure to water for long periods of time without loss of strength or ability to detonate defines the water resistance. Such conditions include (not exclusive list): borehole diameter density confinement within the borehole the presence of water The speed of detonation is important when blasting in hard. For dry blasting agents and water-based explosives.2.3. 2. competent rock where a brisance effect is desired for good fragmentation. However.MGE 415 – Rock Excavation 2004 2. fair. The sensitivity and detonation velocity are hampered for certain water-based explosives at low temperatures while dynamites can become dangerously unstable below freezing temperatures. This represents a loss in terms of dollars spent on explosive energy. Gelled granular products have good water resistance.3. or poor rating. and certain waterbased mixtures have an excellent rating. The presence of moisture in amounts greater than 5% dissolves chemical components in dry blasting agents and alters the composition of gases produced. At this point. explosive manufacturers individually rate products based on a relative basis as good. At diameters less than Dmin. Above Dmin the explosive has reached its steady-state velocity.2. field loading conditions greatly affect detonation velocity. all thermodynamic properties are at a maximum as the reaction front approaches a plane shock front. Explosives manufacturers recommend the appropriate range of temperature for storage and use.3. A numerical rating is used based on the results of tests performed on the explosive.4 Temperature Effects Extreme low temperatures affect the stability as well as the performance of explosives. complete reactions do not take place. for example (in English units): P = 0. In general. is density XXX V is detonation velocity in fps.2.6 Detonation Pressure The detonation pressure is the maximum theoretical pressure achieved within the reaction zone and measured at the C-J plane in a column of explosives.3. 2. Such measurements cannot be made directly and are done during underwater tests performed for energy and strength determinations.2.7 Borehole Pressure Borehole pressure is the maximum pressure exerted within the borehole upon completion of the explosive reaction measured behind the C-J plane. There is little agreement in the literature regarding specific estimates of actual borehole pressures. a number of simplifying formulas are available for estimating detonation pressure for granular explosives based on detonation velocity and density.00337 where V2 P is detonation pressure in psi.MGE 415 – Rock Excavation 2004 Figure 2-33: Generalized relationship between VOD and Diameter 2. Most commercial explosives achieve pressures in the range of 0. The actual pressure achieved is somewhat less than this maximum due to non-ideal loading conditions always present in practice and due to certain explosive formulation.29 to 3.3. With the use of hydrodynamic computer models.48 x 106 psi (2 to 24 GPa). theoretical calculations of borehole pressures are made. Although detonation pressure is related to the temperature of the reaction. pressures after detonation within the borehole are estimated to be less than 30% of the theoretical detonation pressure. Dessureault 1/14/2004 35 . and the minimum critical diameter test are used.3. Standardized tests for high explosives have been adapted for commercial explosives that include the friction (pendulum). air-gap test. In recent years. Of these methods.MGE 415 – Rock Excavation 2004 2. These energy values have been useful in predicting the rock-breaking capabilities of explosives for comparative purposes. The cap sensitivity test measures the minimum energy required for initiation and is used to classify explosives (e. impact (fallhammer).8 cap is an industry standard cap of specific dimensions and charge characteristics. cap sensitive vs. The terms "weight strength" and "bulk strength" were useful many years ago when explosives were primarily comprised of nitroglycerin cartridges.2. Underwater tests were developed to measure both the shock energy and the gas (bubble) energy released during the detonation of standard test samples. and underwater tests.8 strength blasting cap test. friction. with the development of bulk blasting agents and less sensitive ingredients. Below this critical diameter. packaged in 50 lb (23-kg) boxes. In this respect. and heat determine its storage and handling characteristics. the sensitivity of an explosive to shock.8 Sensitivity The definition of explosive sensitivity is two-fold. Dessureault 1/14/2004 36 . noncap sensitive products) or the ability to initiate an explosive directly with a standard cap. as well as the volume of liberated gases and the heat of the reaction.9 Strength The strength of an explosive is a measure of its ability to break rock..g. The term properly used to define the propagating ability of an explosive is sensitiveness. explosives may deflagrate or "dead press. ballistic mortar. A number of methods are used to establish this energy including the use of theoretical computer models and tests such as crater. It includes sensitivity against accidental detonations in addition to the ease by which explosives can be intentionally detonated.2." Dead pressing occurs when an explosive is densified to a point that no free oxygen is available to ensure the start or progression of detonation. The No. The air-gap test measures the distance between the ends of adjacent cartridge explosives for which reliable initiation can be propagated from one cartridge to another. From the standpoint of safety and accidental detonations. The performance potential of an explosive is a function of the detonation velocity and density. impact. among others. The critical diameter of an explosive is the smallest diameter at which an explosive will maintain a steady-state detonation. tests such as the No. 2. and projectile tests.3. underwater tests give the best correlation to rockbreakage performance. new testing methods have been established to determine relative energies for all commercial products regardless of ingredients or packaging. 2% RBS B cm 3 cm 3 135.85) = 723 cal/cm3. Most manufacturers of explosives will include either value with technical product literature. Other common strength terms are the relative weight strength (RWS) and relative bulk strength (RBS) in which the relative measure of energy available per unit weight or volume of an explosive is compared to an equal weight or volume of the standard commercial explosive ANFO.25) Relative Bulk Strength RBS A ABS A ABS ANFO ABS B ABS ANFO 645 cal 723 cal 980 cal 723 cal cm 3 cm 3 89. explosive A is less powerful than ANFO while explosive B is more powerful.MGE 415 – Rock Excavation 2004 Other terminology widely used by manufacturers is based on the theoretical heat of reaction determined by explosive formulation.85 and an AWS of 850 cal/g. the ABS = (850 cal/g) (0. Note ABS and AWS can be computed from one another if density is known.2% Therefore on a volume basis. As a dry. The RWS and RBS are computed as a percentage of that available from ANFO. the heat of reaction is approximately 850 cal/g.5% Relative Weight Strength RWS A AWS A AWS ANFO RWS B AWS B AWS ANFO cm 3 (0. For a typical density of 0.9% 92. free-running blasting agent.8)850 cal 3 cm 980 cal 3 cm (1. Dessureault 1/14/2004 37 .8) and B (ABS 980 cal/cm3 and density 1. and it is the volumetric basis of reaction heat which correlates with energy. ANFO is capable of being loaded or packaged at varying densities. Absolute bulk strength (ABS) in calories per cubic centimeter and absolute weight strength (AWS) in calories per gram are computed from the heat liberated during the detonation and formation of gaseous end products. Depending on the proportions of the mix.25)850 cal 3 cm 645 cal 94. Example 1 Determine the relative strengths of explosives A (ABS 645 cal/cm3 and density of 0. A mixture of ammonium nitrate and fuel oil (ANFO) is by far the most widely used commercial blasting product. Although there are a number of methods available. HO.1 Oxygen Balance As previously explained. Furthermore. whereas water-based blasting agents react more efficiently. or free oxygen (if the mix has an excess of oxygen). and that these values are not affected by changes in the reactants (explosive mixture).MGE 415 – Rock Excavation 2004 2. do not completely react. such as particle size and borehole diameter. releasing optimum energy as predicted by formulation. With an oxygen-balanced formula. Oxygen balance for an explosive component or a mixture of components is usually reported in percentages or as moles of monatomic oxygen. assuming an oxygen balance. and N2 but not noxious fumes such as CO. an oxygen-balanced explosive formula is one in which the amount of oxygen O2 is sufficient to form the desired detonation gases CO2. temperature. many granular explosives. Example 2 Determine the oxygen balance for nitroglycerin. Dessureault 1/14/2004 38 .3 Thermochemistry of Explosives and the Detonation Reaction. It is possible. and pressure are achieved. Changes in the physical nature rather than the chemistry of the explosive mixture. Thermodynamic data used for these calculations are found in Table 2-5. the procedures selected herein are the least difficult to apply. In order to achieve maximum rock-breaking efficiency from an explosive. such as ANFO and nitroglycerin. Unfortunately. assuming the formation of ideal gaseous products. It is computed using the mass balance relationship for a reaction of 1 kg of explosive. C3H509N3. commercial explosives are non-ideal materials. 2. it is assumed that optimum values of thermodynamic properties such as energy. NO.3. an oxygen-balanced mixture is formulated to ensure the formation of nonnoxious gases upon detonation. to estimate the thermodynamic properties of an explosive reaction. however. The following calculations are made to illustrate the methods used to estimate explosive properties of interest. This is the case for ideal reactions. vary the rate of detonation and hence affect thermodynamic variables. and CH4 (if the mix is oxygen deficient). Table 2-5: Thermodynamic data for some explosive components and gases.3.3. Hence.0352 gO/g of C 3 H 5 0 9 N 3 .02O39.5228% or 0. 2.3: Introduction to Explosives. Balkema. Vol. 2 – Theoretical foundations.A.02/2 H2O + 13. for example grams O2/100 grams of explosives.62N13.” Jimeno.21 CO2 + 22. Emilio Lopez Jimeno.21H22.21/2 N2 + C13. and Francisco Javier Ayala Carcedo.53% Note that the OB for an oxygen deficient component is: OB = (O deficient) / (O deficient – O required) Note that some authors (Persson) consider that oxygen balance is given as the mass of oxygen which needs to be removed or added to the composition to achieve oxygen balance. therefore the reaction has an unbalance total of 2. oxygen.19 moles of oxygen.21 (39.2 Explosive Energy Other parameters of explosive thermodynamics are best described in the readings assigned with this lecture. Readings: “Chapter 9: Thermochemistry of explosives and the detonation process.3. (mandatory) Lecture Module 3.19 / 39. A simpler method of calculating oxygen balance for explosives that contain only carbon.21 – 5. 32 3 oxygen balance 5 9 100 4 2 227.81 – 13. the oxygen balance (OB) can be calculated by: OB = (O excess ) / (O available) = 2.09 3.01/2) O2 from the above equation the O2 is (19.62 = 0. and oxygen. Blasting Principles for Open Pit Mining. ADDITIONAL RESOURCES VII The material in the above submodule is further supplemented by: Readings: Hustrulid.095. is given by: oxygen balance Where CxHyNzOw y w 100 4 2 explosive molecular weight 32 x Therefore. Carlos Lopez. William.3. Rotterdam: A. it is expressed as a fraction or percentage of the explosive formula mass.0553 = 5.505) = 1. 1999. consider the previous example where nitroglycerin = C3H509N3.MGE 415 – Rock Excavation 2004 The mass-balance equation is: 13. In this case. nitrogen. “Drilling and Blasting of Rocks” Dessureault 1/14/2004 39 .62/2 – 13.21 – 11. 2.4. o Gunpowder also sped up the very slow process of digging up stones.4 Explosive Products Note that section 2. The key developments are summarized and described below: Black Powder Saltpeter or “Nitre” Probably originated with the Chinese around the 10th Century (mostly for fireworks.1 (below) is optional. and recorded. o After he found out the perfect ratio he wrote the ingredients and the amounts in code in his diary.1 History of Explosives The history of explosives is interesting from its profound influence on world development. o With gunpowder they could blow the stones out of the ground. 2. o He studied fireworks it and tested it over and over again. Blasting with powder replaced “fire setting” for loosening rock around the beginning of the 17th Century.) Roger Bacon published a formula for Black Powder around 1242 o In the 13th century Roger Bacon. o Bacon did not get credit for the making of black powder because he didn't use his invention. (Optional) Rotterdam: A. rockets. sulfur. o Roger Bacon had made.4.A. Berthold Schwarts saw this and exploited it. o Schwarts used the black powder to launch a pebble at high speed out of a metal tube. etc. Those interested in the history of explosives are welcome to become more informed. After many months he found the perfect ratio of saltpeter. o Now even some poor people cold have a house of stone.MGE 415 – Rock Excavation 2004 translated by Yvonne Visser De Ramiro. Berthold Schwartz invented the gun around 1300 which resulted in further refinement of black powder. was interested in the new knowledge from far east. This was a great technological step. the first black powder (the early form of gunpowder). Balkema. and a new ingredient. charcoal. Nitroglycerin Dessureault 1/14/2004 40 . The remainder of this sub-section is mandatory readings. a European. Mercury Fulminate Discovered by Howard in 1800 Later used as detonator for dynamite by Alfred Nobel. nitro fumes and nitro headaches Safety Fuse William Bickford of England devised the safety fuse. In 1865 Nobel invented the blasting cap. meaning power): o When ‘blasting oil’. politics. destroyed his notes and warned against its use Dynamite Alfred Nobel and his father built a NG small factory in 1861 in Sweden Loading and transporting nitroglycerin was dangerous (liquid poured in to holes and ignited with various types of black powder igniters. Ammonium Nitrate First synthesized in 1659 by J. Nobels saw that the kieselguhr abosrobed about 3 times its weight of NG. higher detonation velocity and more effective in breaking rock (not only moved. first used successfully in the late 19th century. o Nobel began to sell the 75/25 NG/kies. literature. Glauber by combining ammonium carbonate and nitric acid. but broke them apart) Cons: Would freeze. Eventually went to wood pulps which increased the energy output of the NG. o Patented in 1867 Nobel knew the amount of destruction his invention would cause and he did not want to be associated with thousands of deaths. accurately timed detonations. Two major uses: Fertilizer and Blasting Initially used to replace a portion of nitroglycerin in dynamite Dupont introduced NITRAMON in 1935 Dupont called his products “Blasting Agents” due to their safe handling. Pros: more powerful than black powder. providing the first safe and dependable means for detonating nitroglycerin and thereby considerably expanding its use for industrial purposes. o This development allowed the relatively safe transport transportation of NG. the first of the dynamites.. NG proved to be very dangerous and resulted in the death of many people including his brother Emil. originally a textilewrapped cord with a black powder core. Electrical firing.MGE 415 – Rock Excavation 2004 Nitroglycerin (NG) and Nitrocellulose (NC) discovered by Ascanio Sobrero in 1846 in Switzerland He was afraid of it. R. NG spilled into kieselguhr (NG was packed in it). so he left a large amount of money to the awarding of prizes in science. which for the first time enabled safe. Nobel discovers dynamites by accident (Dynamite is derived from the word Dynamis. low cost and non-headaches Dessureault 1/14/2004 41 . allows greater control over timing. etc… every year. 2. An explosive is a material that can undergo an exothermal chemical reaction resulting in a rapid expansion of the reaction products into a volume larger than the original. Brest France In 1955. cheaper and safer than dynamite Cons: desensitized by water What can be concluded about the history of explosives are the key aspects in the design of commercial explosives: o Explosive power o Safety in transportation and handling o Cost o Controllable and predictable detonation.2 Energetic material Energetic materials are all materials that can undergo exothermal chemical reaction releasing a considerable amount of thermal energy. Further sensitized by powered aluminum ANFO eventually replaced dynamite in dry holes Pros: easier to load. Ammonium nitrate/fuel oil (ANFO) was discovered to greatly increase the energy output of AN prills. An explosion is basically any rapid expansion of matter into a volume much larger than the original.4.MGE 415 – Rock Excavation 2004 Two disastrous shipload explosions resulted in development of ANFO (Ammonium Nitrate & Fuel Oil) Texas City. Watergel slurries invented simultaneously by Dupont USA and CIL Canada. Figure 2-34: Explosives Family Tree Dessureault 1/14/2004 42 . Figure 2-34 shows a simplified version of an explosives family tree. 4. The solid explosive trinitrotoluene. and H2O. TNT (C7H5(NO2)3).4 Single Molecule and Composite Explosives There are two fundamentally different kinds of explosive materials.4. and nitroglycerine. N2. The molecule decomposes into mainly gaseous reaction products. a mixture of a fuel and oxidizer. Propellants are intended to burn steadily at a rate determined by the design pressure of the rock or gun breech. Explosives are intended to function by detonation following shock initiation by a detonator or a booster charge. Single explosives are chemical substances that contain in one well-defined molecule all that is needed for an explosion.MGE 415 – Rock Excavation 2004 Figure 2-35: Explosive Classification Scheme 2. A composite explosive can be mixture of two single explosive substances. Dessureault 1/14/2004 43 . are examples of single explosive substances. or an intermediate mixture containing one or more single explosive substances together with a/or oxidizer ingredients.3 Explosives and Propellants The difference between an explosive and a propellant is often functional rather than fundamental. 2. such as CO2. and they are ignited to burning by a flame that provides a spray of hot burning particles. namely single explosive substances and composite explosive mixtures. MGE 415 – Rock Excavation 2004 Table 2-6: Single chemical explosive substances. Dessureault 1/14/2004 44 . MGE 415 – Rock Excavation 2004 Table 2-7: More Single Chemical Explosives Dessureault 1/14/2004 45 . to tertiary explosives. through the secondary explosives. and also because the reaction products have a high molecular weight. TNT.MGE 415 – Rock Excavation 2004 2. and additives to provide a moldable consistency. Others. if ignited with a flame (it is strongly advised against performing a demonstration of this without due precautions against the chance event of a detonation). 1947) Texas City explosion of ammonium nitrate. is an example. secondary. are not classified as HE. For example. and that the grain size exceeds a certain minimum size. B4.) They are nonetheless explosive. 2.4. However. Tetryl. AN (NH4NO3). This classification scheme shows how military explosives are often segregated from military applications. as demonstrated by some of the largest accidental explosions in history. A more detailed classification scheme for explosive material can be seen in Figure 2-35 which shows primary. Reaction products are generated at the surface faster than they can expand away from the surface. a stick of dynamite can burn as a candle. and Tertiary explosives. such as lead azide (PbN6). can burn to detonation. The secondary explosives. The primary explosives are able to transit from surface burning to detonation within very small distances.2 mm thick grain of lead azide when ignited will transit from burning to detonation within a distance less than the grain thickness. such as the (April 16. tertiary explosives are extremely difficult to explode and are in fact officially classed as non-explosives provided that certain conditions are observed (such as that an oxidizer not be mixed with fuels or sensitizers. PETN. and tertiary explosives and the differences between single and composite explosives. and require boosters or primers of HE for initiation. but only in relatively large quantities. which results in a quick build-up of pressure at the burning surface. Dessureault 1/14/2004 46 . whereas a truckload of dynamite may burn to detonation. This is because the lead azide molecule is very simple. RDX (cyclonite). Examples of military explosives include TNT. slowly. Under normal conditions. C4. which are mixtures of RDX.6 Commercial Explosives Figure 2-34 shows the breakdown of commercial explosive categories. such as dry blasting agents. decomposing in a very simple two-step reaction. NM. and TNT are examples. Single molecule explosives range with respect to the strength of the stimulus required for initiation or a self-supporting chemical decomposition reaction: from primary explosives. too. of which NG. of which ammonium nitrate. Secondary. small amounts of HE are added to commercial explosives to increase strength and sensitiveness. Many commercial or industrial explosives are classified as HE because they contain critical amounts of military explosives or nitroglycerin. which are used as igniting charges in detonators. B. These explosives are chiefly used in the weapons industry.4. and compositions such as A3.5 Primary. and usually they are cap sensitive. A 0. Dynamites are packaged in cylindrical cartridges from 'Is in. 2. which are all considered HE. a number of nitroglycerin (NG)-based products have been developed of three basic types: granular. sodium nitrate (SN). TNT. densifiers. that incorporate liquid explosive components such as nitroglycerin. Many of these components are also referred to as sensitizers and can also act as absorbents. including absorbants such as wood pulp and flour that also act as combustibles. and thickeners.4. (22 mm) in diameter and 8 to 24 in. The quality of the waxed paper wrapping is important for: water resistance. water. black powder. and include nitrated salts such as ammonium nitrate (AN). gelatins. fume production. permissibles. and AN. Often the difference among these products is formulation. Gelatins and semi gelatins contain nitrocotton that combines with NG to form a gel structure whose consistency is controlled by the percentage of nitrocellulose. are products. gum. however. It is comprised primarily of a stable yet powerful mix of nitroglycerin (nitrostarch).MGE 415 – Rock Excavation 2004 Industrial explosives are classified as one of the following: nitroglycerin-based. granular aluminum. Since its invention. cotton. and cellulose. product packaging and consistency can also change a classification Explosive components are referred to as: Oxidizers: Oxidizers contribute oxygen for oxygen balance. absorbents. and boosters. emulsifying agents.7 Nitroglycerin-based Explosives Dynamite is a trade name introduced by Alfred Nobel. such as wood pulp. SN. and calcium nitrate (CN). Straight dynamite derives its energy source from NG. gelatin. dry blasting agents. a common category. (for underground coal and will not be discussed) primers. (203 to 610 mm) in length. or any carbonaceous material that produces heat. actually contain mixtures or characteristics that fall in other classifications. sawdust. water gels. and semigelatin. Dessureault 1/14/2004 47 . Two-component explosives. carbon. ease and safety of loading. emulsions. Fuels: include fuel oil. Stabilizers: include flame retardants. 8 Dry Blasting Agents Dry blasting agents are one form of a general category of blasting agents. and size of primer used for initiation. ANFO is the most widely used blasting product.000 fpS (4500 m/s) and is achieved in borehole diameters greater than 15 in. percent fuel oil added to ammonium nitrate Typical values of specific gravity range from 0. prilled into porous pellets onto which a liquid fuel oil or propellant is absorbed. density. and cannot be detonated with a No. Each has similar mixtures as straight and ammonia (extra) dynamites with the addition of nitrocellulose for a gel consistency.75 to 0. The critical diameter of ANFO is between 2 and 4 in. The exact values of Dessureault 1/14/2004 48 . (381 mm).95. water conditions. Gelatin dynamites are either straight gelatin or ammonium (extra) gelatin. a mixture of fuel and oxidizer. those who blast underground using small-diameter holes. It is not classified as an explosive. They are also used as primers and boosters. confinement. No. free-running mix of a solid oxidizer (usually AN).5% industrial-grade ammonia nitrate and 5. in particular. 2.5% fuel oil.MGE 415 – Rock Excavation 2004 Ammonia dynamite (or "extra" dynamite) is a granular mix that contains a smaller quantity of NG mixed with AN and SN. Semigelatins are ammonia gelatin with a small amount of nitrocellulose and a 65% weight strength. A dry agent is a granular. with approximately 94. particle size. by definition. The properties of dry blasting agents vary significantly with borehole diameter. (51 and 102 mm) and is a topic of controversy among blasters and. The steady state detonation velocity of ANFO is over 15.4. Figure 2-36 shows the varying effect of the addition of fuel oil.2 grade diesel fuel oil is used for a nearly oxygen-balanced mix. Figure 2-36: Energy output vs. A blasting agent is.8 blasting cap. Figure 2-37 shows the varying effect of diameter confined on various explosives. led by slurries in the 1950s. ANFO does not detonate reliably within the range cited above. and heavy ANFO. a measurable increase in fragmentation energy is noted. came about in response to the disadvantages of ANFO in certain applications. For increasing percentages of aluminum by weight up to 6%. borehole diameter for selected industrial explosives Aluminum in granular form can be added to ANFO to increase the heat or energy output.4. is not cost effective. These were: lack of water resistance low bulk strength due to low density.9 Wet Blasting Agents Blasting agents that contain more than 5% water by weight are referred to as wet blasting agents. Within this category are: water gels or slurries. 2. Heavy ANFO is a combination of prilled ANFO and emulsion. emulsions. therefore. Dessureault 1/14/2004 49 .MGE 415 – Rock Excavation 2004 the critical diameter depends on the loading conditions. The development of wet blasting agents. The cost of additional aluminum beyond 6% does not result in proportionally increased work output and. however. Figure 2-37: VOD vs. high density and bulk strength. and coupling within the borehole. containing highexplosive sensitizers. In general. 20% water is used. Up to 18% aluminum by weight has been found to provide increased energy output. rendering the oxidizer water resistant while thickening the explosive mix. 2. Composition B. 2. Slurries are characterized by: excellent water resistance. or water gels. but rather as a slurry explosive. added as a sensitizer. The critical diameter of wet blasting agents is often less than 1 in. Certain mixtures. Granular aluminum. ethyl alcohol.9. Three varieties of wet blasting agents are in common use in the mining industry: slurries. (25 mm). and Heavy ANFO.1 Slurries Slurries. increases weight and bulk strength.MGE 415 – Rock Excavation Table 2-8: Typical Compositions of selected slurries and Emulsions 2004 Table 2-8 shows some typical compositions of water gels and emulsions. using cross-linking agents.9. confinement. The gels (guar gum) effectively surround the solid AN.4. good oxygen balance. nitrostarch.2 Emulsions Definition of emulsion: A suspension of small globules of one liquid in a second liquid with which the first will not mix. fuel oil.4. Emulsion explosives and 50 Dessureault 1/14/2004 . emulsions. are cap sensitive and hence should not be classified as a blasting agent. and glass bubbles (microspheres) are dissolved or added to the liquid phase. Fuels and sensitizers such as TNT. are a colloidal suspension of solid AN particles suspended in a liquid AN solution that is gelled. Densities range from 1.45 to 1.or Vaseline-like. that can act as hot spots to initiate the chemical reaction upon shock compression. emulsion explosives have excellent water resistance since each AN/water droplet is surrounded by a thin film of oil which repels water. emulsions can be cap sensitive 2. Emulsions provide high detonation pressures of 1. the AN solution stays as a super cooled liquid without crystallizing even upon cooling to sub-zero temperatures.15 to 1. a variety of emulsion explosives or blasting agents of different sensitivity levels can be produced. and droplets combine to form larger particles sizes whose bulk surface area is reduced. the oil phase migrates. Less fuel is in contact with the oxidizer. The cost of emulsion products is within the range for slurries. Microspheres.9. or paraffin using an emulsifying agent. all such emulsions have a high degree of inherent safety. and also because of the desensitizing effect of the water content. In contrast to ANFO which cannot be used in water-filled drillholes because of the high water solubility of AN. and less than ideal explosive properties are achieved. Over time. Depending on product diameter and sensitizers used. By distributing in it finely dispersed voids in the form of glass microballoons or gas bubbles.4. The final product has improved strength and provides good water resistance in comparison to ANFO. In the resulting margarine. smooth mixture. Emulsions are a two-liquid phase containing microscopic droplets of aqueous nitrates of salts (chiefly AN) dispersed in fuel oil. with a price 51 Dessureault 1/14/2004 . The extremely small droplet size. making the mixture a "repumpable" consistency. the intimacy of mixing of the fuel and oxidizer approaches that of a solution. microscopic glass. As they contain no ingredient that is an explosive in itself.3 Heavy ANFO Heavy ANFO is a product comprised of up to 45 to 50% ammonium nitrate emulsion mixed with prilled ANFO. or plastic air-filled bubbles and the AN droplets form the oxidizer. thereby eliminating the need for expensive explosive compounds. The waterin-oil structure depends on entrapped air or microspheres for sensitivity. Premixed emulsions are available in plastic tubes in a variety of diameters and lengths. while the emulsion contains no solid fuel. The only fuel component is in the ANFO (or a liquid fuel). . while the fuel oil exists as the oil phase. and the sub micron thickness of the oil film gives a very large contact area between the fuel and the oxidizer solution.45. Emulsions have excellent waterresistant properties regardless of packaging. It was developed in an attempt to increase the bulk density of ANFO. Emulsions can be mixed on site and pumped from bulk trucks.74 x 106 psi (10 to 12 GPa). wax.MGE 415 – Rock Excavation 2004 blasting agents are the latest development away from ingredients that are in themselves explosive substances. using TNT and PETN. Dessureault 1/14/2004 52 . and 1000 ms are available for short. is questionable. which. or time delays (in milliseconds) used in delayed blasting practices. 2. They are manufactured with an instantaneous (no delay train) time of initiation. Other primers are cast into specific shapes and weights. detonating cord. in turn. initiates a noncap-sensitive explosive or blasting agent. or emulsions are used with blasting caps or detonating cord. thereby reducing blasting efficiency. starts a chain reaction of explosives burning within the metal cap shell. Research has shown. A primer contains cap-sensitive high explosive ingredients. Initiators are used as components within a system of explosives and other devices to start the detonation of all other components.4.4. This process detonates a high-explosive base charge. Boosters are placed within the explosive column where additional breaking energy is required.10 Primers and Boosters- A primer charge is an explosive ignited by an initiator. Table 2-9 summarizes typical delay time intervals available for electric and non-electric initiating systems. Often-times. highly sensitized slurries. Time delays with intervals of 25. that the ability of emulsion to prevent ANFO from being dissolved in the presence of water.4. 2. in turn. 2. Boosters are often used near the bottom of the blasthole at the toe level as an additional charge for excessive toe burden distances. Boosters are highly sensitized explosives or blasting agents. 100. upon receiving an appropriate mechanical. safety fuse.(ms) or long-period (LP) delays. Initiation systems are either electric or nonelectric. produce a detonation or burning action. Electrical energy (ac or dc) is sent through copper or iron legwires to heat an internalconnecting bridgewire. This heat. Often cartridges of dynamites. or electrical impulse. Short delays are used in surface blasting operations.11 Initiators and Initiation Systems Initiators are devices containing high explosives that. 50. and include blasting caps. used either in bulk form or in packages of weights greater than those used for primers. through a powder delay train.MGE 415 – Rock Excavation 2004 range between that of ANFO and emulsions. however.12 Electric Caps Electric blasting caps are a commonly used method of initiation. igniting a cap-sensitive explosive. They are also placed within the explosive column adjacent to geological zones that are difficult to break or intermittently within the main explosive charge to ensure continuous detonation. designed with wells for initiator acceptance. or non-electric shock tubes. cartridge or plastic-bagged dynamites or sensitized wet blasting agents are used as primers as well as boosters. 500. while longer delays are used underground where blasting conditions are more confined. Surface delay elements. Safe blasting practices dictate that precautions are used to avoid blasting in the vicinity of extraneous electricity such as stray current.13 Non-electric caps Non-electric initiation systems include a cap similar to that of an electric cap. This is to ensure that the power source supplies the correct current to each cap in the circuit. Ac power lines and capacitor-discharge dc power sources approved for blasting are used to energize caps. The next generation of high-precision detonators will contain an electronic circuit instead of pyrotechnical delay elements. static electricity.MGE 415 – Rock Excavation 2004 The use of time delays in blasting enhances fragmentation and the control of ground vibrations. including all accessory connecting wires. Precise calculations are needed to determine the entire blasting circuit resistance. The plastic tube itself does not detonate. The energy source in the tubing is either a gas mixture or an internal coating of special explosive. In recent years. Table 2-9: Typical Delay times for donators 2. Non-electric tubing is not used in underground coal or gassy mines as it carries an open flame. therefore. electrical storms. and radio frequency energy when using electric caps. but they are connected to plastic tubing or a transmission line that carries an initiation (shock and heat) to initiate the cap. Caps and tubes of varying lengths are connected with special connectors between holes to configure unique blast pattern arrays. improvements have been made in the manufacturing of blasting caps that increase the accuracy in detonation time.4. when used in conjunction with in- Dessureault 1/14/2004 53 . The integrated circuits will permit microsecond rather than millisecond timing accuracy and allow programmability for onsite selection of each cap detonation timing. the only noise source is the cap itself. Advantages compared to regular electric or nonel detonators: Higher timing precision (10 s instead of 1-10 ms delay scatter) Same high timing precision at long delay times (10 s at 5 second delays) Increased control over time delay Greater safety against accidental ignition (coded firing signal) Current disadvantages include: Higher price because of chip and capacitor cost Back to electric wiring – risk of ground faults or poor contacts 2. provide nearly infinite numbers of delays in blasting patterns. Figure 2-38: Generic Design detonators 2. concerns about the effects of accidental detonations of electric caps due to stray currents are eliminated with the use of non-electric systems.MGE 415 – Rock Excavation 2004 hole delays.14 Electronic Detonators Electronic detonators have an electronic counter on a microchip in place of the pyrotechnic delay charge. Delays are available in short and long periods as well as in-hole and surface delays.4.15 Detonating cord Detonating cord is a flexible but strong continuous detonator that can be several hundred meters long. and a capacitor to supply the discharge energy for ignition. The advantage of non-electric systems over electric systems is the ability to design blasts with a greater number of holes than traditional electric blasting.4. In addition. A detonator is required to initiate a length of Dessureault 1/14/2004 54 . 5 g/m to 70 g/m. depending on the application. Detonating cord (sometimes known as Cordtex) consists of a core of PETN enclosed in a tape wrapping that is further bound by counter-laced textile yarns.MGE 415 – Rock Excavation 2004 detonating cord which cannot be normally initiated by fire. which generally varies from 1. Dessureault 1/14/2004 55 . ADDITIONAL RESOURCES VIII The material for this sub-module is additionally supplemented by: Lecture module 3. thus avoiding the need for multiple electric or plain detonators to provide continuous initiation of the full length of an explosive column in a blasthole. 10 g/m is the PETN weight of standard detonating cord whose VOD is about 7000 m/s.4: Explosive products. as distinct from point initiation with individual detonators. Their energy release depends on the amount of PETN in the core. However. tensile strengths and protective coatings. Detonating cord has two functions: to provide simultaneous detonation of several interconnected blasthole charges. The cord is either reinforced or completely enclosed by strong waterproof plastic. Detonating cords are available with a variety of charge weights. M. Paul Lever’s 415 course notes Hartman. Robert. Maintenance. Persson.ac. ed. Occupational Health and Safety.html Hustrulid. SME: Port City Press. Balkema. Engineering in Rock Masses. tools and techniques used for most blasting applications. Blasting Principles for Open Pit Mining. Holmberg. Jaimin. R. Ed. and desired products. 1992. More unique applications of blasting will be covered in the next module. Potash Technology: Mining. Dessureault 1/14/2004 56 . William. 1983 McKercher. Per-Anders. IRECO Chemicals: Salt Lake City. 570p. 1999. Edition 2002 Hartman. CRC Press: New York. 28th Ed. The History of Explosives. 1992 Kennedy. search: continuous mining Bell. 2 – Theoretical foundations. Pergamon Press: Toronto. Cook. From a historical perspective. and initiation. Rock Blasting and Explosvies Engineering. The Science of Industrial Explosives. Howard L. B. 1974. System Approach The first step in the process of design is to determine the goal or purpose of the intended design. http://sis. Transportation. 1983 Mining-Technology. and Jan M. SME Mining Engineering Handbook.com. Sydney: University of New South Wales Press. Coal Mining Technology Theory and Practice. 449p. 1990. ML. SME Mining Engineering Handbook. 2001 Class notes. Roger. Rotterdam: A. from Bob Cummings 1997 Class notes. 2nd ed. G. Blasting Technology for Mining and Civil Engineers. and Lee. Sen. Howard L. Processing.A. 2nd ed. Introductory Mining Engineering. Littleton CO. Howard L. 2002. explosives.. Utah. 534 p.uk/~dj9006/explosives/history. 1995. 2nd ed. explosives. Melvin A. F. Stefanko.bris. sequencing. geology. it has been seen that 7 These notes were assembled directly from the following references: From Dr. Editor. New York: John Wiley & Sons. Ed. Atlas Copco. along with underground blast design. 1992 Hartman. The Raise Boring Handbook 2nd ed. Mutmansky. Surface Mining Manual. Suface Mining 2nd Edition. Note that these are the key theory.Wirtgen America Inc.MGE 415 – Rock Excavation 2004 Module 3: Blast Design7 The section will cover the basic aspects of blast design in terms of geometry. Gour C. 2000. 146 p.:Society of Mining Engineers. Ed. from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department. ISBN: 0 7506 1063 8 Caterpillar Performance Handbook. Environment. Butterworth-Heinemann: London. Vol.A. Blast Design Basics Blast design is a semi-empirical systematic method that involves balancing numeric and qualitative assessments of rock properties. crushing. as can be seen. the variables upon which an engineer is to optimize based on lowest overall cost can be seen in Figure 3-2. noise. and grinding processes are to be taken into account. This often limited the blast design to enable the efficient loading and hauling of the material from the stope to the crushing circuit.MGE 415 – Rock Excavation 2004 drilling and blasting has been a process by which the mine has undertaken full responsibility. These aspects of the design are facilitated in that they are often specified by law. The internal environment is also less limiting as rock damage to walls that are not permanent are also not to dominate the design. Dessureault 1/14/2004 57 . bench sizes. Other constraints not shown are the limitations posed by pit design. Its key goal was to deliver to the mill the least expensive product. However. All these factors must be balanced with cost. limited operating room. equipment. Figure 3-1: Systems Approach In visualizing the system. an optimum design for simply the mine will most likely change when the mill is included. and other detrimental aspects of working with explosives. the more recent ‘systems design’ approach whereby the scope changes from simply the mine to the overall mine-mill system. flyrock. For example. the internal and external environments specify a minimization of wall damage. etc… The proper fragmentation to optimize the loading& hauling. MGE 415 – Rock Excavation 2004 Figure 3-2: Cascading constraints and goals Figure 3-3 shows the numerous controllable variables that can be manipulated to achieve the optimum design. The uncontrollable variables and output requirements should be taken into account in the design. Dessureault 1/14/2004 58 . however. This focus of design assumes that the optimization is to be taken at the blasting stage. it should be noted that changes in downstream processes may reduce some of the cost upstream where an expensive aspect of the design can be removed. They are presented in their original form in Error! Reference source not found. the base of the Dessureault 1/14/2004 59 .MGE 415 – Rock Excavation 2004 Figure 3-3: Controllable and uncontrollable input variables and output goals. As can be seen some of the costs decrease with increasing fragmentation while others increase. By adding the curves together one obtains the overall cost versus degree-offragmentation curve presented in Figure 3-8. It has the form of a saddle indicating that there is a certain degree of fragmentation for which the overall cost is a minimum. to Figure 3-8. Idealized fragmentation curves Some 30 years ago. MacKenzie presented his now classic conceptual curves showing the cost dependence of the different mining unit operations on the degree of fragmentation. In the particular case shown. At standard operating costs per hour (for all practical purposes independent of the production rate) this will result in lower costs per ton or cubic yard moved. and bridging time will decrease and the crushing rate per hour will increase. repair and maintenance. Figure 3-6: Crushing Cost Curve Crushing An increase in the degree of fragmentation gives lower crushing costs as more material passes through as undersize. Beginning with the loading. Liner costs. The effect of wear and tear will also decrease. giving lower operating cost per hour. size and type of truck and haul road conditions. hauling and crushing curves the logic. as presented by MacKenzie. lift. As indicated decreased bridging time also cuts down on truck delay time at the crusher which in turn gives higher truck and shovel Dessureault 1/14/2004 60 . At a standard operating cost per hour.MGE 415 – Rock Excavation 2004 saddle is quite broad suggesting that the overall costs change little over a wide fragmentation range. is as follows: Figure 3-4: Loading cost curve Loading An increase in the degree of fragmentation will give the shovel a higher rate of productivity. There will be a consequent decrease in cycle time. truck production per hour will increase with greater degree of fragmentation due to faster shovel loading rates and a decrease in bridging (and hence waiting time) at the crusher. Before discussing the development and application of these curves it is important to understand the logic behind them. this increase in truck speed or productivity will result in lower unit operating costs. Figure 3-5: Hauling Cost Curve Hauling Under similar conditions of haul. presuming that the same hole diameter and pattern is used.MGE 415 – Rock Excavation 2004 productivity.) the drilling cost could: remain constant. combinations of both. The % bridging time is one indicator. geologic structure. increase. The same is not true for the drilling and blasting costs. These have been the easiest to explain since the unit costs always decrease with increasing fragmentation. For blasting case (1. Thus there would be more drill holes required to blast a given volume. In his original presentation MacKenzie has explained the drilling dependence as follows: Dessureault 1/14/2004 61 . changing to an explosive having greater energy content per unit hole volume (higher energy content! density) 3. along with shovel loading rate of this degree of fragmentation.).) the associated drilling cost would increase if the explosive quantity were to be increased by simply drilling the same diameter drill holes but on a tighter pattern. or decrease depending upon the situation. If the same fragmentation is desired and a more energetic explosive is substituted for the one currently in use. an increase in the degree of fragmentation may be achieved by: 1. Figure 3-7: Drilling and Blasting Cost Curves Drilling & Blasting For a given rock type. the drilling costs would remain constant independent of the fragmentation. There are many possible combinations which can occur depending upon the particular design. Any increase in degree of fragmentation means less work for the crusher. increasing the consumed quantity of a given explosive 2. and firing sequence. If larger diameter drill holes were substituted and the increased hole volume (explosive quantity) achieved in this way then the rate of increase or decrease would depend upon the comparative drilling cost per foot of hole. then the unit drilling cost could decrease due to the possibility of increasing the hole spacing (spreading the pattern). For case (2. For case (3. The readings prescribed in the additional readings provide additional theory and equations. However.MGE 415 – Rock Excavation 2004 Generally speaking. as will be shown in future modules. The rate of increase or decrease or decrease will be dependent upon the drilling cost per foot. for a given type of drilling and of explosive. the cost per cubic yard or ton will remain constant or increase with the degree of fragmentation. the drilling cost per yard will decrease. These preliminary guidelines provide the reasoning (proofs) for the five key relationships for blast design. If higher energy explosives are substituted. Figure 3-9: Isometric view showing nomenclature The key variables for blast design include: Dessureault 1/14/2004 62 . Note that the basic guidelines shown here apply to surface mines. Preliminary guidelines for blast layout The preliminary guidelines for blast layout were taken directly from the Hustrulid (1999) which in turn were taken directly from Ash (1967). the same key variables are also applicable to underground mine blasts. Figure 3-9 shows the basic nomenclature for the key variables discussed here. Figure 3-8: Overall cost curve. the overall optimum fragmentation can be assumed to be the lower section of the saddle. Therefore summing all cost curves together. is equal to the bench height plus the sub-drill. The volume of rock to be broken by A UNIT of hole length is: Dessureault 1/14/2004 63 .is defined as the distance between the individual rows of holes. When the bench face is not vertical the burden on this front row of holes varies from crest to toe.is equal to the hole length minus the stemming. Burden to hole diameter relationship Note Figure 3-10.is the distance between holes in any given row. This distance is referred to as the subgrade drilling or simply the sub-drill Stemming (T) . The selection of one. Sometimes the explosive strength is varied along the hole. for example the hole diameter. length of the explosive column (Le) . Subgrade (J) . the different dimensions involved in a blast design are not arbitrary but closely related to one another. This column may be divided into sections (decks) containing explosives of various strengths separated by lengths of stemming materials. As will be seen in the next section. Figure 3-10: Plan view of bench showing first row. many of the others.e. the hole spacing (S) and burden (B) can be directly related through the following relationship: S KS B Where Ks is a constant relating spacing to the burden. Bench height (H) – is the vertical height from the toe to the crest. where it is shown that for a particular burden and spacing. a higher strength bottom charge with a lower strength column charge. drilled length (L) .A certain length of hole near the collar is left uncharged.MGE 415 – Rock Excavation 2004 drilled burden (B) . spacing (S) . This will be referred to as the stemming length (T) whether or not it is left unfilled or filled with drill cuttings/crushed rock.Generally the holes are drilled below the desired final grade. Spacing to burden relationship As can be seen in Figure 3-10. each hole diameter is expected to break a particular volume of rock. i. fixes within rather strict limits. It is also used to describe the distance from the front row of holes to the free face. However. Therefore the explosive available is determined using: EA 4 De2 E e Since Ee is a constant related to explosive type. ER B2 The amount of explosive energy available is determined by the explosive volume that can be loaded into that unit of length of borehole: Ve 4 De2 Where De is explosive diameter and Ee is the explosive bulk strength (strength by unit volume). the charge diameter (De) may be less than the diameter (D) of the hole. This would make the explosives available also directly proportional to hole diameter: EA D2 Dessureault 1/14/2004 64 . where bulk blasting agents are used. the entire cross-sectional area of the hole is filled with explosive. since S K S B 2 EV Since KS and EV are constants. The total energy for a unit of hole length is therefore ER ER VR EV B S EV KS B however.MGE 415 – Rock Excavation 2004 VR B S 1 A particular amount of energy (EV) is required to break the a unit volume of rock. the required amount of explosive energy is directly proportional to B2. as is sometimes the case at pit perimeters where perimeter blasting techniques are employed. Therefore hole diameter (D) is equal to explosive diameter (De). the De can be seen as directly proportional to Ea EA De 2 If using packaged explosives. There are several different rationales used for selecting the appropriate length. Here an explanation based upon explosive run-up distance will be presented. The results are essentially the same with all techniques. this would also require that the required explosives and available explosives to be equal since additional explosives cannot be added beyond the hole diameter. Dessureault 1/14/2004 65 . Figure 3-11: Effect of hole diameter on burden Subdrill to burden relationship The toe region is a highly confined volume. EA ER however. the diameter should also increase proportionally. relating to the hole diameter: B KBD Therefore as the burden increases. as seen in Figure 3-11. This extra explosive power is generally provided by extending the drill hole below the toe elevation and filling the so-called subdrill length J with explosive. Therefore extra explosive energy must be applied to assure adequate fragmentation.MGE 415 – Rock Excavation 2004 Furthermore. remember that: ER B 2 therefore D 2 B 2 therefore D B therefore the diameter is can be determined from a proportionality constant KB. In addition. Therefore. Figure 3-13: Run-up distance to achieve steady state VOD. To be conservative the minimum run-up distance will be assumed to be 6D. the distance from the drilled end of the hole to the toe elevation (the subdrill distance J) should be: J J 8D therefore since D B B resulting in a proportionality relationship J 1/14/2004 KJ B 66 Dessureault . the explosion (borehole wall) pressure (Pe) is proportional to the square of the detonation velocity: Pe VOD 2 The elevation in the hole at which steady state velocity is reached should not be higher than the bench toe elevation. As seen in previous classes.MGE 415 – Rock Excavation 2004 Figure 3-12: Toe confinement There is a certain distance (called the run-up distance) characteristic of the initiating system/explosive which the shock wave must travel away from the point of initiation before steady state conditions are reached in the explosive column. To break the confined toe. A normal offset is of the order of 2D. the primer is seldom placed directly at the bottom of the blasthole due to the presence of cuttings and water. the borehole pressure should be as high as possible. The other relationships are derived in a later section however. Therefore the “as or more difficult” relationship can be summarized as: V B a) b) Figure 3-14: Section view comparing the spherical charge (a) and cylindrical charge (b) minimum distance from the collar An initial assumption is made in that the degree of equivalence of the charges will depend on the proximity to the charge and that this relationship is linear and expressed as: TC K TC B From Figure 3-14.MGE 415 – Rock Excavation 2004 Stemming to burden relationship Near the hole collar. this relationship proof is explained here. the following relationship is evident: T TC 2 considering the " as difficult" relationship where V V K TC B 2 KT B B KT 1 B1 K TC 2 B and considering how TC K TC B : T T which can be simplified to where K TC 2 Determining KTC is difficult. Langefors & Kihlstrom (1963) have empirically derived the 67 Dessureault 1/14/2004 . In bench blasting. as desired. This can be seen conceptually. toward the vertical free face. the rise of the explosive should be controlled so that the possibility of breaking upward toward the horizontal free surface should be ‘as difficult’ or more difficult than breaking. by the placement of a spherical charge having the same distance from the collar as the burden. consider a bench containing two vertical side-by-side blastholes. The length of the elongated charge is expressed in multiples of the burden B. In the second blasthole a cylindrical charge with a linear charge concentration of 1 kg/m of hole is emplaced. The bottom of the charge is at toe elevation and then the column extends upward towards the collar. The burden is the same for both. the total charge would be B x 1 (remember that the explosive volume per length emplaced is 1 kg/m). Figure 3-16: Langefors & Kihlstrom’s Toe breaking equivalence of spherical and cylindrical charges Dessureault 1/14/2004 68 . The energy density by the time it reaches the toe is much less than that produced by energy which has traveled a shorter distance. this example will involve the toe region. Figure 3-15: View facing bench comparing equivalent cylindrical and spherical charges. The reason for this is that the explanation is easier and the principle is the same.MGE 415 – Rock Excavation 2004 spherical/cylindrical charge equivalence is as shown in Figure 3-16. For a cylindrical charge of length B. as seen in Figure 3-15. This is understandable since the energy contained in that part of the elongated charge near the collar must travel a much longer distance to reach the toe and in the process the energy is spread over a much larger volume of rock.6 x 1 x B. Consider a spherical charge of quantity Qo placed at the toe elevation in one of the holes. Rather than discussing the collar region which is the subject of this portion. From the Figure 3-16 one can see that at the toe this elongated charge has only the equivalent breaking power of a spherical charge of weight 0. To explain the significance of the curve. The limiting condition is when the center of charge reaches the toe elevation (Figure 3-19). If one continues to increase the scale (hole diameter) as shown in Figure 3-18. This can be seen in Figure 3-17. The value of KH is therefore: Dessureault 1/14/2004 69 . For charges shorter than 0.3 x 1 x B. the effect at the toe diminishes rapidly with increasing length.6 x 1 x B placed directly at the toe elevation. Thus an elongated charge extending from 0.6 2 1.e.3B above the toe elevation (for a total explosive weight of 0.6 B and noting that TC 0. according to the curve have the same breaking capacity as a spherical charge with a weight of 0.6 and where 1 0. This occurs for a hole diameter which yields a burden just equal to the bench height.MGE 415 – Rock Excavation 2004 For a linear charge of length 0. Figure 3-17: Equivalent spherical and cylindrical charges In transferring this concept to the collar region one finds that TC K TC KT 0.3 x 1 x B.4 2 0. i.3B the total charge has a mass of 0. the center of the charge progresses further and further down the hole. From the curve it is seen that this has the same effect at the toe as a spherical charge placed directly at the toe elevation with a mass 0.7 KT 1 K TC B K TC .3B. The fifth and last of the fundamental relationships is: H KH B where KH is a constant relating bench height to the burden. The same effect could be achieved by considering the elongated charge extending from the toe elevation downward.6 x 1 x B) would. 2 Bench height to Burden relationship To this point in the discussion there has been no specific mention of the bench height.3B this relationship holds as well.3B below the toe to 0. For elongated charges with lengths greater than 0. the elongated charge of a given weight has the same effect at the toe as a spherical charge of the same weight. The remainder provides the reasoning and proof behind these initial values. there is little difference between Ks=1 to Ks=1. Note that a staggered pattern provides more uniform energy coverage. Table 3-1 provides the summary of this section. the best energy coverage is with Ks = 1.5.MGE 415 – Rock Excavation 2004 KH 1 As a rule of thumb. the optimum burden and spacing ratio depends on the energy coverage of the bench (among other variables). five relationships were derived for preliminary blast design. Figure 3-19: Limiting the charge diameter and burden. Dessureault 1/14/2004 70 . This section will discuss the general values of the constants which is used as base designs from which to further optimize based on other design constraints. the best energy coverage is with Ks=1 however. note how: B K B D therefore H H KB K H K B D and since K H 1. empirically. Ratios for initial design In the previous section. to derive a method of limiting the choice of hole diameter. Ratio Ks As will be covered in a later lecture. we can derive H D K B D resulting in a limiting Figure 3-18: Burden to diameter relationship. When using a square pattern. For a staggered drilling pattern.15. Many also require knowledge of various qualities of the rock and explosives. Its actual value will depend on a combination of variables including the rock characteristics. nor is it understood. however. while dense explosives. As a rule. such as field-mixed ANFO mixtures. A convenient guide that can be used for estimating the burden. permit the use of KB near 40. But when the rock is completely fragmented and displaced little or not at all. such as tensile strengths and detonation pressures. etc. Light density explosives.MGE 415 – Rock Excavation 2004 Figure 3-20: Square pattern Figure 3-21: Staggered pattern Ratio KB The most critical and important dimension in blasting is that of the burden. is the KB ratio. the explosive used. Dessureault 1/14/2004 71 . the blaster can usually expect satisfactory results for average field conditions. etc. necessarily require the use of lower KB ratios (20 to 25). There are two requirements necessary to define it properly. the necessary information is not readily available. Experience shows that when KB = 30. There are many formulae that provide approximate burden values but most require calculations that are bothersome or complex to the average man in the field. an amount slightly less than the critical value is preferred by most blasters. Usually. To cover all conditions. such as slurries and gelatins. and subsequently finer sizing is also expected to result. The final value selected should be the result of adjustments made to suit not only the rock and explosive types and densities but also the degree of fragmentation and displacement desired. the KB value could be reduced below 30. one can assume the critical value has been approached. the burden should be considered as the distance from a charge measured perpendicular to the nearest free face and in the direction in which displacement will most likely occur. To provide greater throw. It must be remembered that the subdrill region generally forms the future crest/bench top for the bench below. densities for the latter exerting the greater influence. stemming refers to the filling of blast holes in the collar region with materials such as drill cuttings to confine the explosive gases. Collar and stemming are sometimes used to express the same thing. This increases the chances for cutoffs (misfires) and overbreak. for most rocks requiring blasting. As a rule of thumb consider: For light explosives in dense rock use KB = 20. For heavy explosives in average rock use KB = 35.7 far the most common value. the blaster can. Ratio KJ The most common value of KJ is 0. stressing will occur much earlier in the solid material than compaction of the stemming material could be accomplished. the latter being the unloaded portion of a blasthole.MGE 415 – Rock Excavation 2004 To estimate the desired KB value one should know that densities for explosives are rarely greater than 1. the subdrilling may be increased over that indicated by using KJ = 0. Unwanted damage done at this stage may have a long and costly life. The use of stemming material then assists in confining the gases by a delayed Dessureault 1/14/2004 72 . In addition excessive subdrill results in: A waste of drilling and blasting expenditures An increase in ground vibrations Undesirable shattering of the bench floor.0. In very hard toe situations. Since an energy wave will travel much faster in solid rock than in the less dense unconsolidated stemming material. For heavy explosives in light rock use KB = 40. the density in g/cm3 rarely exceeds 3. determines the degree of stress balance in the region.8 g/cm3.3. It accentuates vertical movement in the blast. Some specialists suggest the use of KT = 1.6 or less than 0. whether or not stemming is used. abandoned blastholes and deviations for the bench below.3. perform other functions in addition to confining gases. Thus.7. In certain sedimentary deposits with a parting plane at toe elevation subdrilling may not be required. This in turn creates drilling problems. But stemming and the amount of collar. Also. However it is probably better to consider using a more energetic explosive. by first approximating the burden at a KB of 30 make simple estimations toward 20 (or 40) to suit the rock and explosive characteristics. However.2 nor is less than 2. Thus the amount of collar that is left. Ratio KT The minimum recommended value for KT for large hole production blasting is KT = 0. For light explosives in average rock use KB = 25.2 with 2. with back break and possible violence. A key indicator of blast design is the specific charge. On the other hand. Too small a KH value will result in substantial cratering.Burden Burden – Diameter Subdrill – Burden Stemming – Burden Bench height . However. also known colloquially as the ‘powder factor’. Ratio KH Currently most open pit operations have KH values which are approximately 1. Summary of Ratios The table below simply summarizes some of the basic ratio values and equations. increasing the length of stemming may reduce the energy concentration in the collar region to the point where large boulders result. SANFO=1) Bench Height (H)=15m 73 Dessureault 1/14/2004 .3 0. reflection and refraction of the energy waves reduce the effects in the direction of the charge length.6 Powder Factor The following example is provided to illustrate the relationships developed in the last two sub-sections. the amount depending upon the degree of energy reduction at the density or structural interfaces. Placing the charge too close to the collar can result in backbreak. Field experience shows that a KT value of 0. Thus the KT value can be reduced under such circumstances.6 or more. For stress balance in bench-blasting of massive material. Design a blast considering these initial variables: Rock = syenite porphyry (SG=2. flyrock and early release of the explosive gases with resulting poor fragmentation. the value of T should equal the B dimension.7 1.6) Explosive = ANFO ( =0. Usually a KT value of less than 1 in solid rock will cause some cratering.7 is a reasonable approximation for the control of air blast and stress balance in the collar region.5 25 =0. if there are structural discontinuities in the collar region. Table 3-1: Ratio Summary Name of Relationship Spacing .Burden Equation S KS B B KBD J KJ B T KT B H KH B Value 1-1.MGE 415 – Rock Excavation 2004 action that should be long enough in time duration to permit their performing the necessary work before rock movement and stemming ejection can occur. particularly for collar priming of charges.8. MGE 415 – Rock Excavation 2004 Hole Diameter (D)=381mm (15”) Staggered drilling pattern, vertical holes 1 blast = 4 rows of holes each containing 6 holes from the relationship, we see: B K B D , where KB 25 (assumed), therefore B=(25)(0.381)=9.5 m S J T K S B , where KS=1.15 (staggered), therefore S=(1.15)(9.5)=11 m K J B , where KJ=0.3, therefore J=(0.3)(9.5)=3 m K T B , where KT=0.7, therefore T=(0.7)(9.5)=6.5 m H B 15 1.6 , considering the 9.5 L=H+J, hole length therefore equals to 15+3=18 m As an assurance, in verifying KH, we find K H value is usually equal to or greater than 1.6, we find these parameters to be acceptable. To calculate the powder factor, the volume and weight of the explosive used is calculated: Ve We 4 Ve D 2 (L T ) 4 1.31 m 800 kg / m 3 3 (3.81) 2 (18 6.5) 1.31 m 3 1049kg per hole 25,176 kg total weight of explosive TEXP 1049 kg 6 4 The weight of the rock that will be broken is calculated: TR VR number of holes n B S H 24(9.5)(11)(15)(2.6 t/m 3 ) 97,812 tons Hence, the powder factor using the explosive ANFO is found to be: PFANFO TEXP TR 25,716 97,812 0.26 kg ton Determination of KB As seen in the above discussion, KB is by far the most important constants in the design of blasts. Selecting the proper burden is therefore a key step in blast design, and the factor KB allows the selection of an appropriate burden and diameter as seen in the equation: B KBD As mentioned above, the best initial estimate for KB is 25, when using ANFO in rocks with an average density. However, consider the requirements for selecting a new KB when explosives or rock type varies. The approach described below can be used as an approximation until field results are available to guide the designed toward more optimal solutions. Note that the 74 Dessureault 1/14/2004 MGE 415 – Rock Excavation 2004 proof below is valid for the metric system. Consider this blast design has must be designed to other variables, including: SGE = specific gravity of explosive used SGR = specific gravity of the rock PFEXP = powder factor (kg/ton) TF = tonnage factor (m3/ton) As seen in a previous example, the total weight that one borehole is expected to break can be calculated using the following: TR VR B S H rock B 3 K S K H SGR Once the total amount of rock material to be removed is known, the amount of energy required is calculated using the powder factor: E Re qyured TR PFEXP B 3 K S K H SGR PFEXP The amount of explosive available is defined by the size of the borehole, which can be given as: E Avail 4 De ( H 2 J T ) SG E 4 De ( BK H 2 BK J BK T ) SG E B 4 De ( K H 2 KJ K T ) SG E Setting the amount of explosives to the amount required results in: B 3 K S K H SG R PFEXP B KB De B 4 De ( K H 2 KJ K T ) SG E K B D and assuming a bulk agent is being used, SG E 4 SG R PFEXP SG E 4 SG R PFEXP (K H (K H K J KT ) , and since B KS KH K J KT ) KS KH The powder factor based on the explosive used is replaced with the equivalent ANFO powder factor, denoted by the variable PFANFO: Dessureault 1/14/2004 75 MGE 415 – Rock Excavation 2004 PFEXP S ANFO KB PFANFO , where relative weight strength to ANFO of an explosive, S ANFO is equivalen to S ANFO QEXP , where Q is the energy per unit weight (usually cal/gm), which results in : Q ANFO SG E 4 SG R S ANFO PFANFO (K H K J KT ) KS KH (K H K J KT ) KS KH note that this equation in english units is equivalent to KB 2000 SG E 4 SG R S ANFO PFANFO where PFANFO ANFO equivalent powder factor in lbs/ton at 2000 lbs/ton The above formula can be used for several purposes, as will be seen in the following examples. Effects to pattern design in changing explosives One of the major ways the equation can be used is to study the effect of changes in the explosive on the blasting pattern while keeping other factors constant. Note that the bench height, KH, depends on the burden which also depends on KB. Therefore the burden can change. The approach can be seen as follows: Explosive 1: K B1 SGE1 4 SGR S ANFO PFANFO (K H1 1 K J KT ) K S K H1 Explosive 2: K B2 SG E 2 4 SG R S ANFO PFANFO (K H 2 2 K J KT ) KS KH2 In most blast design alternatives, a constant powder factor is maintained. A ratio between the two KB values, where the powder factor ratio is eliminated, can be expressed by: K B2 K B1 SGE 2 SGE1 S ANFO (2) S ANFO (1) K S K H1 (K H 2 K S K H 2 (K H1 KJ KJ KT ) KT ) Taking this ratio and for now, ignoring the changes in KH with changing burden, the equation is reduced to: Dessureault 1/14/2004 76 MGE 415 – Rock Excavation 2004 K B2 K B1 SGE S ANFO SGE S ANFO 2 1 The above is a first approximation of the KB2 and is equivalent to the square root of the bulk strength ratio for the explosives in question. To refine the KB2 value, an iterative process is used where: 1) Initial value of KB2 is substituted in the formula: B2 K B 2 De H B2 2) Then the new KH2 is derived from the equation KH2 3) This new value of KH2 is then inserted into the equation K B2 K B1 SGE S ANFO SGE S ANFO 2 1 K S K H1 (K H 2 K S K H 2 (K H1 KJ KJ KT ) KT ) 4) If the resulting KB2 is then compared to the original estimate. If they are the same, the process stops. If not, then the new KB2 is used in step 1 and the process continues, until the value of KB2 converges. Effects to pattern design in changing rock types The effect to the pattern design when rock types change is very similar to that discussed in the previous subsection. Here, the initial approximation can be made by the ratio: K B2 K B1 SG R1 SG R 2 Once again, the iteration process involving three equations is undertaken: 1) Initial value of KB2 is substituted in the formula: B2 K B 2 De H B2 2) Then the new KH2 is derived from the equation KH2 3) This new value of KH2 is then inserted into the equation K B2 K B1 SG R1 SG R 2 K S K H1 (K H 2 K S K H 2 (K H1 KJ KJ KT ) KT ) Dessureault 1/14/2004 77 6 KJ = 7/25 = 0.5 lbs/ton KB 2000 SGE 4 SGR S ANFO PFANFO (K H K J KT ) KS KH using imperial units Example 1: Changing diameter of holes What would the pattern be if changing to 15 inch diameter holes? The first step in this problem is to derive KB. the process stops.6 0.82 Q = 912 cal/gm Rock: SG = 2.3 KT = 17/25 = 0. then the new KB2 is used in step 1 and the process continues.15) 25.15 These values are plugged into the following equation: KB 2000 SG E 4 SG R S ANFO PFANFO (K H K J KT ) KS KH 2000 0.65 PFANFO = 0. First. Numerical Examples Consider the mine where the current design has the following parameters: Hole diameter = 12 ¼ inches Bench height = 40 ft Burden = 25 ft Spacing = 29 ft Subdrill = 7ft Stemming = 17ft ANFO: SANFO = 1 SGANFO = 0.82 4 2. If not. For the 15 inch diameter holes.5 (1.5 ft Dessureault 1/14/2004 78 . until the values of KB2 converge.65 1 0 . If they are the same.3 0.7) (1.2 15 inches 12 inches foot 31. the other ratios need to be derived: KH = 40/25 = 1.6)(1. the first approximation for the burden would be: B K B DE 25.7 KS = 29/25 = 1.MGE 415 – Rock Excavation 2004 4) If the resulting KB2 is then compared to the estimate at the beginning of the iteration.2 This is about what is expected from the recommended ranges discussed previously. 3 This results in a pattern with the following dimensions: B = 30 ft S = 34.10 Q = 815 cal/gm We will need to use the weight strength of this explosive product with respect to ANFO.15) 24.65 1 0.3 0.89 We use the initial estimate. and using the initial KB value from the previous question: K B2 K B1 SGE S ANFO SGE S ANFO 2 1 (25.5 Dessureault 1/14/2004 79 .5 (1.82 1.82 4 2. as derived in a previous section.89 0. Example 2: Changing Explosives What would be the change in the pattern in changing the explosive from ANFO to heavy ANFO with the following properties (original hole diameter of 12 ¼ inches): SG = 1. which is: KB = 24.MGE 415 – Rock Excavation 2004 However.6)(1.5 ft T = 21 ft J = 9ft The resulting powder factor may be slightly different from the original on account of the rounding off.1 this process is repeated through several iterations until a stable KB is found.7) (1. To maintain fragmentation.10 0.2) 1.6 0.5 This new value is then input back into the original KB equation: KB 2000 SG E 4 SG R S ANFO PFANFO (K H K J KT ) KS KH 2000 0.27 31. this new value for the burden would result in a change in KH. the powder factor would have to be increased.00 27. Therefore the new KH is found to be: KH H B 40 1. however note that this pattern would result in increased coarseness in fragmentation. which is calculated: S ANFO 815 912 0. 89 0.2) ( K H 2 0.95 12.3 0. which for this example.7 J = 8.25 inches 12 inches foot 27.10 0.42 28.5 12.82 1 1.1 ft The new KH is then calculated to be: KH2 H B2 40 1. initiated on the same or different delays. separated by inert stemming material.6 0.51 ft this process is repeated until a stable value of KB2 results.7) 1.2) (25.1 This value is then substituted into the equation: K B2 K B1 K B2 K B2 K B2 K B1 SG E S ANFO SG E S ANFO 2 1 K S K H1 (K H 2 K S K H 2 (K H1 2 1 KJ KJ KT ) KT ) KT ) (K H 2 KJ KH2 KT ) SG E S ANFO SG E S ANFO 1.3 0.25 inches 12 inches foot 28. faults or clay seams Dessureault 1/14/2004 80 .6(1.MGE 415 – Rock Excavation 2004 The new burden would therefore be: B2 K B 2 DE 27.6 (1. Decking is employed to: Reduce explosive use adjacent to weak rock zones.7) 26.82 1 (K H1 K H1 KJ (25.6 S = 31.3 0.3 T = 19.3 Stemming & Decking Many blast designs use decked charges formed by dividing the explosive column into two or more individual charges. the final KB2 = 27.95 The new burden is calculated to be: B2 K B 2 DE 26.3 0.42 0.7) 1.89 0.10 0.42(1.6 0.7) KH2 1.0 The blast pattern would therefore be (in feet): B = 27. lowering ground vibrations Bring the powder column up higher in the hole to assure good breakage near the collar Decked charges should be separated by stemming materials at a length beyond which two adjacent decks do not affect one another.8. therefore B=(25)(0. and subgrade drilling remain unchanged. Therefore assuming if three decks are used. where KB 25 (assumed). Solution: The amount of explosives per unit length is calculated as: Wexp lo L 4 D2 exp lo 4 (0. The basic data is: Rock = syenite porphyry (SG=2. SANFO=1) Bench Height (H)=15m Hole Diameter (D)=381mm (15”) Staggered drilling pattern. Note that the collar stem. above.6) Explosive = ANFO ( =0. keeping the same powder factor. where KJ=0. we see: B K B D .381) 2 (0. therefore J=(0. we see that the powder factor is 0. Example of Deck design Determine the change in layout if decking is needed in the layout by limiting the weight of explosive to 275 kg/delay. bench height.5)=11 m K J B . where KT=0.381)=9.15 (staggered).5)=3 m K T B . A rule of thumb for the design of interdeck stem length is to employ the hole radius dimension in feet.MGE 415 – Rock Excavation 2004 Reduce charge quantity detonated at one time delay. This situation is referred to as sympathetic detonation and may lead to excessively high ground vibrations or flyrock and a loss of fragmentation due to confinement from improper timing.26 kg/ton and that the amount of explosives per hole is: 1049 kg.7. where KS=1. vertical holes from the relationship.3. the deck designed to initiate on the earlier time delay may prematurely initiate the second deck.15)(9. hole length therefore equals to 15+3=18 m A following calculations from section 0. therefore S=(1.7)(9.5)=6. the overall explosive column Dessureault 1/14/2004 81 . therefore T=(0.8) 91 kg/per unit length Charge decks per hole is found to be: 275 kg / 91 kg = 3 meters per deck.5 m L=H+J. The following example gives the design procedures for a blast design in which the explosive charge is limited to control ground vibrations.5 m S J T K S B .3)(9. If interdeck stemming is too small. (3)(275 kg) B (1. B. be aware that blasting is an never-ending process of fine-tuning and modifications. and powder factor are the most important variables in blast design. Higher powder factors result in coarser fragmentation and are typically used for rock removal using draglines and large shovels. This approach is necessary due to the many factors that cannot be controlled such as geology and explosive loading conditions.15)(8. It is desirable to select a size that will provide an adequate powder factor (the ratio of explosive distributing the explosive evenly throughout the hole depth). The capacity of the excavation equipment or requirements of downstream processing dictates the required fragmentation. Table 3-2: Powder Factor Dessureault 1/14/2004 82 .MGE 415 – Rock Excavation 2004 length is 9 meters leaving 1. Burden should be selected based on geology and explosive energy output. the following is determined : PF 11. Fragmentation and particle size distribution are a function of hole diameter and burden.15) B 15 (2. the drillhole diameters and burden are typically lower for quarries or mines using smaller equipment in weaker rock.4 m therefore the new S (1.75 meters of stemming between each deck.7 B 2 (kg / m 2 ) 825 (kg) therefore B 825(kg) 11.7m Concluding Notes on D.26kg / ton 8 . burden. and PF Drillhole diameter. Although this lecture provided the preliminary blast design layout equations.6ton / m 3 ) solving for B.7(kg / m 2 ) (n decks)(W per deck) B KS B H rock 0.4) 9.5 meters of stemming length available meaning 0. Note that since the powder factors are lower. Higher powder factors result in fine fragmentation and are therefore required for small capacity removal equipment such as a front-end loader. Hole diameter is usually set by the drill rig capacity which is matched to the range of hole depths anticipated for the job. from: Jimeno.also considered readings for Module 4. It has been found that the explosion pressure should not exceed 5% of the dynamic Young's modulus in order to obtain optimum results from blasting. This crushed zone is detrimental to the transmission of strain waves in the surrounding rock. Assignment 5 (you will learn much when doing this assignment. or adjacent to the blast site. and cracks formed as a result of previous blasting. Chapter 19: Controllable parameters of blasting and Chapter 20: Bench Blasting. Density The density of rock is closely correlated with its strength. care should be taken to review existing geology and geologic structure. around. (Mandatory for assignment 5) . The following properties of rock may have a significant influence on blasting results. The post-detonation gas pressure exerted in the cracks between the blasthole and the free face pushes the burden forward and produces heave. Dynamic Compressive Strength If the explosive's outgoing strain wave exceeds the dynamic compressive breaking strength of rock an annulus of crushed rock is formed around the charge. and Francisco Javier Ayala Carcedo. An increase in rock density often results in a decrease in the displacement of a rockmass Dessureault 1/14/2004 83 .A. Rotterdam: A.1 – Basics of Blast Design Geological Impacts on Blast Design Prior to developing each blast design. natural fissures. Balkema. Lecture Module 4. Elastic Moduli The elastic moduli give the behavior of rocks under stress and should be determined by sonic techniques (dynamic moduli) rather than by the use of mechanical tests (static moduli). This can be accomplished by simply review existing outcrops in . The bulk modulus of a rock has to be known in order to calculate heave.3 – Patterns and sequencing. may also be used.MGE 415 – Rock Excavation ADDITIONAL RESOURCES IX 2004 Additional learning resources include: Readings. Thus there are planes of preferential fracture orientation in anyone or more of an infinite number of directions relative to the blasthole's axis. Carlos Lopez. Emilio Lopez Jimeno. A detailed drill log indicating discontinuities at various depths. and an explosive of low density and velocity should be used in rocks with low compressive strength values. Hence it is important that the dynamic compressive strength of the rocks in situ be determined. Rock Properties Blasting performance is usually influenced more by rock properties than by the properties of the explosive. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rocks show numerous planes of weakness. owing to the filling of pores with water. There is also a strong possibility of the closing of microcracks by the static stress in the rock mass when the confining pressure is above 100-300 MPa. This can be achieved by bottom priming and by having adequate stemming to prevent premature venting of gases. Porosity Porosity tends to reduce the efficiency of blasting operations. The lengths of strain-wave-induced cracks in a highly porous rock are calculated to be only about 25% of those in a non-porous rock of identical mineralogy. permeability. owing to the lower friction characteristic between grain surfaces. A typical example of such a phenomenon is where some of the radial cracks from the blasthole tend to curve off into the direction of the static stress field. In Situ Static Stress High in situ static stresses often exist well within the rock body and blasting results can be affected by these stresses. or changing to an explosive which has stronger heave energy. and hence slope instability. It increases with a high degree of porosity. fluids in a porous rock reduce both the compressive and tensile strengths. when a stress field exists in a direction normal to pre-existing radial cracks around a Dessureault 1/14/2004 84 . It is therefore advisable in open-pit mining to dewater a rock mass where a 'permanent' slope. If water is present in discontinuities adjoining a block of rock which is being blasted. internal friction values for igneous or metamorphic rocks are lower than for sedimentary rocks. has to be formed. This has a wedging action which will have a considerable influence on overbreak. reducing the blasthole pattern. Water Content Water saturation considerably increases the velocity of propagation of strain waves. Generally. strain waves may have a greater ability to weaken that rock mass by means of water being jetted considerable distances through interconnected fissures. Internal Friction Internal friction is a relative measure of a rock's ability to attenuate strain waves by the conversion of some of the mechanical energy into heat. Hence. However. post-detonated gases have to be kept trapped at high pressure until they have performed their task. Again. This implies that highly porous rocks are fragmented mainly by heave energy. if the rock pores are filled with water. which require high energy explosives for satisfactory blasting. which is a good medium for elastic wave transmission. Adequate displacement of higher-density rock can be achieved by following one of three courses: increasing the blasthole diameter. the internal friction factor reduces considerably. However. which is intended to remain for several years with minimum maintenance.MGE 415 – Rock Excavation 2004 fragmented by blasting. and jointing of the rock mass. giving easier passage of the strain wave and improved fragmentation. MGE 415 – Rock Excavation 2004 blasthole, it can be sufficiently strong to prevent extension of these cracks. Moreover, it may induce the formation of new cracks in the direction of the stress field. These types of unexpected blast results are often due to in situ stresses in the rock mass. Structure The bedding planes and joints in a rock mass tend to dominate the nature of the blast-induced fracture pattern. Maximum fragmentation is generally achieved where the principal joint planes are parallel to the free face. Where the angle between joint planes and the face is within a region between 30° and 60°, the blastholes may produce an irregular new face, owing to the formation of wide cracks behind the blastholes. When the joint planes are at right angles to the face, each block requires at least one blasthole in order to obtain satisfactory fragmentation (Figure 3-22). Figure 3-22: Blasthole placement in relation to joint planes. If there is a high density of joint planes normal to the face, it is worth while to consider adopting smaller diameter blastholes at closer spacings. Pronounced bedding planes can also inhibit the explosive's strain energy transmission from one stratum to the next. This has particular importance where the stratum near the collar of the blasthole is hard and is filled with a prescribed length of stemming material to avoid fly-rock or early escape of the explosive's gas energy. This situation can be dealt with in two ways. The first is to place a small 'pocket' charge centrally within the stemming column (Figure 3-23). The second method is to drill 'stab holes' half-way between the drilled burden and spacing, and lightly charge them although this may not be practical from an operational standpoint. Dessureault 1/14/2004 85 MGE 415 – Rock Excavation 2004 Figure 3-23: Use of Pocket Charges Figure 3-24: Stab hole The physical distance between the structures and how that would affect blasting should be considered. Pre existing cracks in this area will direct or even dictate the fragmentation size. The three most negative geological effects (structurally related) on blast performance include: Rock that has open structures Zones of incompetence within the rock, in which structures are unpredictable Rock with alternate zone of competent and incompetent rock. Close and tight rock structure are preferable as the explosive energy is not lost or vented. The problems arise when the energy is not confined or when the transmission of stress waves within the rock mass are interrupted. Open or widely separated structures can result in poor fragmentation due to: Interruption of the explosive generated stress waves. Causes an inconsistent formation of cracks Disruption of confinement resulting in oversize. Venting and airblast can also occur in weak seams or open layers of rock. This can be corrected through stemming and decking. These areas can be identified in the drilling process when the drill experiences slower penetration rates related to poor hole flushing. Dessureault 1/14/2004 86 MGE 415 – Rock Excavation 2004 Some solutions to addressing these problems include: Closer initiation intervals (achieves desired rock breakage before allowing existing cracks to open further) Altering design (burden, spacing, and hole diameter) Selecting more dense explosives or blasting agents (detonate at higher velocity) Use of multiple decks or cartridged explosives smaller in diameter into the stemming zone. Structurally induced Interruption of stress waves Figure 3-25 shows the type of radial cracking which one might expect when blasting a single hole in a brittle, massive rock formation. There will be a relatively few long cracks (6-8) spaced uniformly around the hole. As one approaches the hole the cracks will be shorter and more numerous. Figure 3-25: Idealized radial cracking surrounding a single hole. The maximum length (Re) of the radial cracks for a given explosive and rock type can be shown to be directly dependent on the hole radius. Thus as the hole diameter is increased from 150 mm to 310 mm the length of the longest cracks would be expected to about double. This is consistent with the design relationship B = KBD presented earlier since the burden should be related to the lengths of the cracks generated B Re If the strength of the explosive used in a hole of given diameter is increased or decreased, the outer crack radius should change accordingly. This is reflected in the value of KB chosen. Since in general, a larger diameter hole is less expensive to drill than one of smaller diameter (on a cost/volume basis) the natural conclusion would be to drill as large diameter holes as possible. Unfortunately fragmentation considerations would suggest just the opposite, Dessureault 1/14/2004 87 MGE 415 – Rock Excavation 2004 i.e. the holes should be smaller to better distribute the explosive throughout the rock mass. To illustrate this some simple geometric reasons will be given. Figure 3-26 shows two possible blast patterns using different size holes but the same explosive. The specific energy (powder factor) is the same for both. A simplified representation of the radial cracks after blasting is shown in Figure 3-27 for each pattern. As the hole diameter is increased and the pattern expands, the distance between adjacent crack tips becomes greater. For the case shown L>L' Thus even though the energy density is the same, the fragmentation is more coarse. Generally as the pattern is spread, the powder factor (energy factor) must be increased to maintain acceptable fragmentation. Figure 3-26: Extent of cracking for two patterns with different hole diameters and same PF Dessureault 1/14/2004 88 MGE 415 – Rock Excavation 2004 Figure 3-27: Maximum block dimensions for hole patterns. One way of maintaining proper fragmentation is to increase the powder factor by limiting the pattern spread to some proportion of the theoretical value. It is well known that an actual rock mass generally contains many discontinuities of different types. the story becomes even more complex. Another way of accomplishing this would be to increase the energy of the explosive being used. As shown in Figure 3-28. Dessureault 1/14/2004 89 . there are limiting hole diameters/burdens/spacings which yield acceptable fragmentation. If such structures (joints in particular) are now introduced. such as is shown in Figure 3-29. there is now an overlap of the longest fractures. because of the point introduction of energy into the rock and the fracture geometry. even in massive rocks. Figure 3-28: Fragmentation enhancement achieved by reducing the spacing Therefore. MGE 415 – Rock Excavation 2004 Figure 3-29: Effect of jointing on fragmentation The radius of influence for any given hole is significantly reduced since The radial cracks will not cross the gaps formed by the joints. The high pressure gasses can be short-circuited by the less resistant joints compared with the fresh cracks. Therefore the primary fracturing effectiveness is reduced as well as that produced by a sustained heave of the fractured material. Although these pre-existing cracks limit the formation of new cracks and provide avenues of escape for the explosive gases, mobilization of these is a major reason why the specific breakage energy in blasting is much lower than other processes which must attack the intact rock. Figure 3-30 shows two potential drilling patterns in the jointed rock. The smaller diameter, closely spaced holes yield almost one hole per block and the fragmentation would be expected to be good. On the other hand, the larger holes on wide spacings could yield a large number of substantial blocks largely isolated from the effect of the explosive by the joints. Pattern (a) would have higher associated drilling and blasting costs than Pattern (b). By assigning costs to the degree of fragmentation and knowing the overall ideal fragmentation requirements, an 'optimum' pattern can be determined. Figure 3-30: Possible blast patterns in Jointed rock Structure Orientation The orientation of the major structures can have a significant effect on blasting results. There are three cases to be considered: Shooting with the dip Shooting against the dip Shooting along the strike Dessureault 1/14/2004 90 MGE 415 – Rock Excavation 2004 In shooting with the dip (Figure 3-31) one finds a tendency to get more back break less toe problems a smoother pit floor more movement away from the face and therefore a lower muckpile profile. Note: less sub-drilling may be required in this case owing to the fact that the explosive energy may follow the strata downward, eliminating toe problems. Furthermore, inclination of the drillholes in the direction of dip may reduce backbreak but will also tend to ‘cast’ (throw) and spread the muckpile. A slight addition of delay time in the back row may provide relief to the back of the shot resulting in a more stable highwall. Figure 3-31: Shooting with the dip When shooting against the dip (Figure 3-32) one finds less backbreak since the strata is dipping into the wall. the toe would be more difficult to pull. a rougher floor condition. the muckpile may be higher with less movement from the face. In this situation, the rock tends to move upward, parallel to the plane of the joints. As a result of explosive energies migrating into the strata, a rock unit may cause back break. This situation may result in the creation of an unstable highwall. Also the muckpile would tend to be poorly displaced (creating more work during excavation). If a rock unit is massive, a sawtoothed effect along the floor and overhand at the crest of the slope could result. To eliminate toe problems, the blaster may consider: a blast using angled drill holes, high energy explosives in the toe area and/or additional subdrilling. Decking, satellite holes or small diameter explosive charges in he crest area might help to alleviate an overhang. Pre-splitting may also be a highwall stabilization option. Dessureault 1/14/2004 91 MGE 415 – Rock Excavation 2004 Figure 3-32: Shooting against the dip Finally, when shooting along the strike (Figure 3-33) one finds that: the floor can be highly sawtoothed due to the different rock types intersecting the floor. for the same reasons the back break is irregular. these are some of the worst conditions for those involved in drilling and blasting. To overcome this, the working face may be reoriented to a more favorable conditions. Figure 3-33: Shooting along strike ADDITIONAL RESOURCES X Additional learning resources include: Readings, Chapter 17: Rock and rock mass properties and their influence on the results of blasting, from: Jimeno, Carlos Lopez, Emilio Lopez Jimeno, and Francisco Javier Ayala Carcedo. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Rotterdam: A.A. Balkema. (Mandatory for assignment 5) - also considered readings for Module 4.3 – Patterns and sequencing. Lecture Module 4.2 – Geological impacts on Blast Design. Dessureault 1/14/2004 92 MGE 415 – Rock Excavation 2004 Patterns and Sequencing The formations should be examined to identify the strike and dip direction of the most prominent joints. In igneous and metamorphic rock formations, one should consider aligning the rows of holes parallel to the alignment of the dominant joint system. In sedimentary rocks, the drill holes should be placed in rows drilled parallel to the formation strike line. When considering the option of altering the blasthole pattern by increasing the spacing (parallel to joints) and reducing burdens (perpendicular to joints) or using a staggered pattern, the direction of movement or angle of movement controlled by delay intervals should be considered. A diagram of blast hole location and firing time intervals will assist in predicting the true burden and spacing firing angles relative to existing geologic structure. Millisecond (MS) delay blasting was introduced in open pit quarry blasting many years ago. Even when blasting to a free face, the rock movement time can be an important factor. This is particularly true in multiple row blasts. For a typical quarry with 15-foot (4.6 m) spacings, the initial movement at the free face may occur in 10 to 12 milliseconds, but the burden only moves about 0.5 foot (15 cm) in 10 milliseconds. With one or two rows of holes, the prime movement is directly out from the face. As the number of rows increases, the rock movement will tend toward the vertical. This is caused by the low velocity of the broken rock successively reducing the relief toward the free face. This can contribute to "tight" bottom as well as flyrock Figure 3-34: Increasing numbers of rows increases upward movement due to lower velocity of previously broken rock. Delay Blasting techniques are employed to improve fragmentation, control of rock movement, overbreak, and to reduce ground vibrations. Delays are incorporated into the blast design using electric or nonelectric caps or delay connectors with detonating cord. The delay patterns used in design will determine the sequence of hole or deck initiations, thereby, dictate the Dessureault 1/14/2004 93 Figure 3-35 illustrates the difference between two adjacent blasts. It may also be important to provide additional nomenclature for the various types of blasts in terms of degree of fixation. Depending on the S/B ratio. Figure 3-36: Corner and face blasts. The effective spacing is the distance between holes in a row defined by adjacent time Dessureault 1/14/2004 94 . The figure shows a variation of timing used for surface blasting called echelon or half chevron. Figure 3-35: Muckpiles illustrating the difference between millisecond timed and instantaneous. the actual timing (in milliseconds) between detonating charges will determine muck pile displacement height and distance from the bench. an effective burden Be and effective spacing Se result as shown in Figure 3-37. The blast on the left used millisecond timing and the blast on the right instantaneous firing. Effective Burden and Spacing Depending on initiation sequence.MGE 415 – Rock Excavation 2004 overall direction of blasted rock movement and resulting fragmentation. Figure 3-36 shows a corner blast and face blast. (Figure 3-38) is applicable to most types of formations.07-foot (2. It can be readily adapted to the square or rectangular pattern. Figure 3-37: Echelon or half chevron. the angle of movement is 45 degrees to the open face.6) = 15. When a V/MS delay pattern is used in conjunction with a square drill pattern.14-foot (4.6) = 6. Effective burden is the distance in the direction of resultant rock mass movement. Therefore.MGE 415 – Rock Excavation 2004 delays (e.26 ft Be = 13* sin(31.8 ft Dessureault 1/14/2004 95 . also known as chevron. The “V” (Vee) pattern..3 m) spacing or the burden is only one-half the spacing. delays by rows). the angle of movement will vary in relation to the relative (also known as ‘effective) burden and spacing dimensions. If a rectangular drill pattern is used.6˚ Se = 8 / sin(31.1 m) burden and a 14. The formula to determine the angle of movement in relation to the open face for a rectangular pattern is: tan A B S where B = burden S = spacing A = angle of movement Therefore effective spacing Se and effective burden Be would be calculated by: Se B and Be sin A S sin A Example: Determine the effective burden and spacing for a V initiation pattern with an 8 ft drilled burden and 13 foot spacing on a square drilled pattern.g. a 10 by 100 foot (3 by 3 m) square pattern becomes a rectangular pattern with a 7. Arctan (8/13) = 31. If the interhole delay is too short. If interhole delays are too long. This detachment forms an internal free face (or relief) to which successive detonations will interact with the reflection of stress waves. the movement of row burdens is restricted and fragmentation is poor.MGE 415 – Rock Excavation 2004 Figure 3-38: “V” (Vee) or Chevron initiation pattern on a square drill pattern. Research has shown that stress wave travel time is a fraction of the time required to develop radial cracks. High ground vibrations result. and backbreak along the new high wall may persist. Furthermore. contributing to the detachment of the rock mass in the vicinity of the hole. cutoffs of surface delays may occur. The minimum timing is. studies using high-speed photography indicate that the burden moves within a timeframe which is between 2 to 10 times the wave travel Dessureault 1/14/2004 96 . therefore: t 2 Be 10 3 Cp where t = stress wave travel time in milliseconds (ms) Be = effective burden or distance form the hole to the free face in feet Cp is velocity of sound for the rock in fps The maximum timing is that at which the burden is fully detached and accelerating as gas pressures build. jeopardizing the stability of the slope. Designing the Timing The design of initiation timing for multiple-hole blasting is critical to the blasting effectiveness. The minimum time for design is controlled by the stress wave travel distance (= 2 Be) in order for radial cracking to begin to develop. 5 to 2. Reduced-scale research using a variation in delay ratios suggests improved fragmentation for timing between 11 to 17 ms/ft of Be. it has been recommended that a 1. some suggest delays of 1 to 5 ms/ft within rows and 2 to 15 ms/ft of Be between rows (or on the echelon).2 to 1. For single-row production shooting and S/B of 1. provides the effect of drilling patterns and S/B ratios on the area covered by fracture circles.2 ms/ft of Se and 10 ms/ft of Be while forward throw is minimized.2 to 4. Generalized Timing When taking both pattern shape and timing into account. in a general way. with ratios of 1. the existence of fractured ground where existing Dessureault 1/14/2004 97 . Some researchers have recognized the need to qualify delay ratios. Timing studies have been performed to investigate resulting fragmentation and muck pile shapes. based on existing fracture density. it was established that 3. However.MGE 415 – Rock Excavation 2004 time to the face. and explosive type.6 that timing ratios greater than 1. multiple-row blasting resulting in recommended timing to improve various aspects of the resulting muckpile.4 ms/ ft of Be.3 ms/ft of Se and 1. from measured flyrock velocity. Even more research has demonstrated improved fragmentation for S/B ratios of two at timing ratios of 1 ms/ft of Be or greater. The timing ratios cited are found to vary over a wide range. and gas venting.5 ms/ft of Be. Other research has shown that the time to burden movement ranges from 5 to 50 ms. while weak fractured rock fragments best with higher delay ratios. and suggests an optimum range of timing for design between 1. the most desirable overall drilling/initiation pattern would the drilling pattern with the best energy distribution.: For optimum fragmentation. A great deal of research on the effects of initiation timing cannot be compared due to the lack of similar variables such as geology. Competent dense rock requires lower delay ratios to achieve fine fragmentation. For optimum breakage and forward movement. Subtantial research has been undertaken in production-scale. resulting in high muck piles.2 ms/ft of Be for multiplerow production blasting in hard rock.2 ms/ft of Be are ideal.3 ms/ft of Be is to be used.4 ms/ft of Be was recommended for soft rock with long stem lengths and low powder factors.5 to 2 ms/ft of Se and 5 to 6 ms/ft of Be. A 2. Table 3-3Error! Reference source not found.4 ms/ft of hole spacing and 8. Rock Types: One researcher recommended 1. scale. Ground vibrations: to control ground vibrations. while using high powder factors and short stem lengths. Forward throw and muckpile shape: similar work in which muck pile profiles were mapped indicates that optimum forward throw and muck pile height reduction occur for delay ratios of 4. through the collar stemming. Alternatively. as discussed in previous lectures. simultaneously initiated blastholes should be far enough apart to prevent mutual interaction between their stress fields. Table 3-3: Effect of drilling patterns and S/B ratios on the area covered by fracture circles (energy distribution) Figure 3-39 provides a summary of events and their timing for a single blasthole allowing the reader to visualize the chain of events. Figure 3-39: The events and timing in a blasthole Dessureault 1/14/2004 98 . Overall initiation patterns are best where each blasthole is initiated separately and in sequence.MGE 415 – Rock Excavation 2004 fracture planes limited the development of new radial cracks would reduce the efficiencies involved in maximizing the energy distribution. This is usually best achieved through the “chevron” or V patterns. Check with the suppliers to avoid overlap or crowding. No.MGE 415 – Rock Excavation 2004 Skipping a Period It is a common practice of many blasters to double the delay time on the last row. and height of face all have a pronounced effect on the number of rows that can be fired successfully without excessive stacking or without encountering high bottom. 15 through No. This practice. will also reduce the upward ripping action and materially reduce the backbreak on the face. No. it will occupy on the average 30 percent more volume (swell factor) than it did in the solid. When the blast consists of as many as eight or nine rows.15 (200 MS through 500 MS) provided a nominal 50 milliseconds between each period.8 periods (25 through 200 MS) provided a nominal 25 milliseconds between each period. Remember that swell factor will vary with the type of rock. The hole diameter.000 MS) provided a nominal 100 milliseconds between each period. the No.1 through No. called "skipping a period". 19 (500 MS through 1. the timing on MS delays should provide the additional time without skipping a period. This provides additional time for the rock ahead of the last row to move forward so that the relief on the last row will be increased. Dessureault 1/14/2004 99 . Even with additional time between rows. As shown in Figure 3-40.8 through No. When the rock is broken. burden and spacing. This sequence is provided only as an example for discussion: the actual sequences and intervals of detonator timing vary with manufacturer. the tendency still exists for the rock to stack if the number of rows is excessive. Figure 3-40: Blast timing sequence skipping periods Always base timing designs on the limitations of detonator accuracy since delays of a given period have a range of actual firing times. as seen in Figure 3-41. If a rectangular drill pattern is used. Depending on the formation and the number of holes per row. The forward movement is controlled within reasonable limits and the broken muck is deposited in a windrow 90 degrees to the open face. Another advantage may be that the impacts between rock in different rows may additionally fragment the muck. but is not as practical for ease of loading under field conditions. Dessureault 1/14/2004 100 . and rectangular. Note that various initiation systems can be used to achieve the various timing options. Remember that there are three basic drill pattern types into which these patterns can be designed: square. Note that his type of drill pattern also results in an effectively staggered initiation sequence. Pattern Types The next few sections will provide general guidelines for a few blast designs which can be altered by mine planners to meet needs and conditions. direction of movement and relative timing are what the figures are intended to represent. forward movement is limited. It may be used with a staggered pattern. “V” (Vee) Pattern The V pattern (Figure 3-38) is applicable to most types of formations. the angle of movement will vary in relation to the relative burden and spacing dimensions. to open the blast using two holes on the first delay. Note that this pattern results in a high concentrated and centralized rockpile. Obviously. it is sometimes desirable. staggered.MGE 415 – Rock Excavation 2004 In most cases the material has only two directions to move. This will result in slightly more forward movement in the muckpile. It can be readily adapted to the square or rectangular pattern. The specific numerical dimensions and timining of the figures reflects only suggestions or possible options. The shape. If the number of rows is excessive. excessive movement in either direction will result in dangerous flyrock. to the front and vertically. thus additional space for forward expansion cannot be provided. flatter muckpile. Note that these patterns would result in a lower. Figure 3-42: Echelon Delay Pattern Dessureault 1/14/2004 101 . Figure 3-42 is an example of an echelon delay pattern and is also known as a half-chevron. Other advantages include a minimization of tight toe problems and an easier connect-up (pattern tying) since the paths can be easily seen. Echelon Delay Pattern When the blast area is open on two adjacent sides in an external corner.MGE 415 – Rock Excavation 2004 Figure 3-41: Chevron or V pattern with double hole initiation for slightly more forward movement. the blast should be designed to take advantage of the reduced degree of fixation. Figure 3-43: Channel pattern Flat-face pattern The flat face pattern will move the broken material farther from the face and usually will result in rock of larger size.MGE 415 – Rock Excavation 2004 Channel Delay Pattern Operations are frequently opened on side hill terrain where there is not sufficient area opened in front of the blast for forward movement associated with the V or echelon patterns. Figure 3-44: Flat face pattern Dessureault 1/14/2004 102 . The design of the blast must be relatively narrow. In these cases a channel delay pattern can be used to confine the broken rock to the blast area. The channel pattern will not hold all the broken rock within the blast area if the terrain is steep. This pattern should only be used for special conditions. It is almost always used with a staggered pattern unless large blocky material is desired. in certain formations it may be desirable to skip a delay period to allow additional time for rock movement because only the direction of movement for the initial holes is vertical. Figure 3-46: Sinking shots Dessureault 1/14/2004 103 . In most formations it is necessary to decrease the burden and spacing of the initial holes in the delay pattern in order to open an area of relief to which the remaining holes may break. A sinking blast varies from most blasts because there is no open face or relief. Since the entire blast will be ‘in the tight’. also known as a sinking shot or sink. it may be necessary to make a sinking blast. which levels of vibrations and the generation of flyrock are more likely.MGE 415 – Rock Excavation 2004 Alternating hole delay pattern This pattern has been used with limited success in thinly laminated formations on very wide spacings. It is not recommended for most formations. and the direction of blasted rock movement must be vertical. Figure 3-45: Alternating delay pattern Sinking Blast Pattern When opening a new pit or starting a new lift in an existing pit. MGE 415 – Rock Excavation 2004 Sequential Pattern Variant Note that considering that various initiation products (det. electric detonators.A. Rotterdam: A. Emilio Lopez Jimeno. and Francisco Javier Ayala Carcedo. Balkema. (optional) Lecture Module 4. Figure 3-47: Sequential Firing ADDITIONAL RESOURCES XI Additional learning resources include: Readings. Chapter 27: Initiation Sequence from: Jimeno. etc…) can provide any timing sequence possible. cord.3 – Patterns and Sequencing Dessureault 1/14/2004 104 . variants of the timing can be created allowing movement planes not directly aligned along drill holes. programmable detonators. det. Note each hole in Figure 3-47 is initiated at different times yet is similar to the movement plane of an echelon blast. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Carlos Lopez. Blasting each hole individually may also aid in reducing vibrations from the blast as less explosive material is initiated simultaneously. shock tubes. cord delays. Sydney: University of New South Wales Press. lack of an the tunnel drilled in a in multiple 8 These notes were assembled directly from the following references: From Dr. Cook.html Hustrulid. can be developed stages as seen in Figure 4-1. The Raise Boring Handbook 2nd ed. Note that for tunnel rounds that are too large to be single pass (as seen in many civil applications). 570p. 1995. 1983 Mining-Technology. 1999. Balkema.ac. http://sis. Howard L. 1992 Kennedy. William.A. Stefanko. Surface Mining Manual. ML. 1974. 1992.Wirtgen America Inc. ed. 28th Ed. 1990. 2nd ed.:Society of Mining Engineers. Maintenance.bris. Environment. Butterworth-Heinemann: London. Paul Lever’s 415 course notes Hartman. Atlas Copco. Introductory Mining Engineering. Roger. 534 p. 449p. The Science of Industrial Explosives. Editor. Potash Technology: Mining. Persson. Mutmansky. SME Mining Engineering Handbook. Sen. Occupational Health and Safety. G. Ed. Utah. 2nd ed. Coal Mining Technology Theory and Practice. CRC Press: New York. Ed. Suface Mining 2nd Edition. Engineering in Rock Masses. Drift Design The blasts in tunnels and drifts are characterized by the initial available free surface towards which breakage can occur. Blasting Principles for Open Pit Mining. Melvin A. IRECO Chemicals: Salt Lake City. Holmberg. Blasting Technology for Mining and Civil Engineers. Ed. B. Pergamon Press: Toronto. and Jan M. and Lee. 2000. 2001 Class notes.com. The History of Explosives. Per-Anders. Littleton CO. F.M. search: continuous mining Bell. 2002. Howard L. ISBN: 0 7506 1063 8 Caterpillar Performance Handbook. Rock Blasting and Explosvies Engineering. Howard L. only heading itself. 1992 Hartman. Dessureault 1/14/2004 105 . Rotterdam: A. 2 – Theoretical foundations. Gour C. Transportation. SME Mining Engineering Handbook. Vol. Jaimin. R.MGE 415 – Rock Excavation 2004 Module 4: Blasting Applications8 The section will cover the more advanced aspects of blast design in terms of its various applications. 146 p.uk/~dj9006/explosives/history.. Robert. New York: John Wiley & Sons. from Bob Cummings 1997 Class notes. Edition 2002 Hartman.A. Processing. SME: Port City Press. 1983 McKercher. 2nd ed. from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department. Figure 4-2: Different zones in a drift round Contour holes are those which establish the final shape of the tunnel and are placed with little spacing and directed towards the interior of the mass to make room for the drills in collaring and advance. as seen in Figure 4-3. the lack of co-operation between adjacent charges and. The opening usually has a surface of 1 to 2 m2. This is due to drilling errors. the demand made by swelling. In fan cuts. Stoping can be geometrically compared to bench blasting although it requires powder factors that are 4 to 10 times higher. there is a negative action of gravity as happens in lifter holes. the cut and cut spreader blastholes usually occupy most of the section. the absence of hole inclination. in some areas. Figure 4-3: Orientation of the contour blastholes to maintain tunnel profitelookouts Dessureault 1/14/2004 106 . although with large drilling diameters it can reach up to 4 m2.MGE 415 – Rock Excavation 2004 Figure 4-1: Systems for advancing large faces The principle behind tunnel blasting is to create an opening by means of a cut and then stoping is carried towards the opening. which requires sufficiently insensitive explosives to avoid sympathetic detonation and at the same have a high enough detonation velocity. This phenomena consists of the explosion gases pushing the air that exists between the column charge and the wall of the blasthole. They are blasted at or near the center of the face to break a roughly cylindrical opening to the intended depth of the round. The boreholes that surround this cut area are sequenced to fire later and break to this newly created opening. To be successful. as well as calculation of the patterns and charges in the rest of the sections which are. whereas those of the second have fallen in disuse due to the difficulty in drilling. to prevent channel effects in the cartridge explosives placed in large diameter blastholes. the latter is usually chosen as it avoids the free fall of the material. Burn Cut (Hopler) This is a closely spaced group of boreholes drilled parallel to the direction of advance and perpendicular to the existing face. The blasts in tunnels and drifts are much more complex than bench blastings owing to the fact that the only free surface is the tunnel heading. the different types of cuts are explained in their order of importance. it must break the rock and move it forward. lower center and upper center. independent from the type of cut applied. Cuts. It is important that the bum cut holes be drilled accurately and parallel to each other. Of the three positions. destroying the hot spots or excessively increasing the density of the explosive. generally speaking. fragmentation and also on the number of blastholes. The first group is most used in operations with mechanized drilling. On the other hand. They are only applied in small excavations. compressing the cartridges in front of the shock wave. burdens are small." (lengths of borehole left after the blast that may contain un-exploded explosives. Cuts can be classified in two large groups: Parallel hole cuts Angled hole cuts. the profile of the broken rock is more extended. less compact and better fragmented. Improper location of the burn cut holes may result in "bootlegs. In the following. This creates a void which provides additional relief for the remaining holes to be fired later in a predetermined sequence. Bootlegs may be caused by: Dessureault 1/14/2004 107 . The primary function of the cut remains the same regardless of the type of cut or its variations.MGE 415 – Rock Excavation 2004 The position of the cut has influence on rock projection. comer. The powder factors are elevated and the charges are highly confined. above 3000 ms. the pyramid. Types of angle cuts include the Vee. Dessureault 1/14/2004 108 . Usually the angle cut will move the rock further down the heading and produce coarser fragmentation out of the cut area. Figure 4-4: Various types of burn cuts (solid dots are loaded holes) The varied rock types and structures determine the drill spacings. Angle Cut An angle cut is a group of boreholes drilled at various angles inclined to the free face to provide as much freedom of movement for the rock as possible. and the drnw or hammer. However. thereby destroying the delay sequence excessive burdens between the bottom of the holes. and types of explosive that will successfully fragment and remove the rock to the full depth of these boreholes. Longer drill steel is required in the cut holes to achieve the same depth as the surrounding boreholes in the round. In narrow headings it is difficult to drill a sufficiently wide angle to insure "pulling" the cut. Angle cuts generally require fewer holes and less powder-per-round than burn cuts. detonator delays. they are generally more difficult to drill and require more experienced miners.MGE 415 – Rock Excavation 2004 sympathetic detonation from propagation between holes that are too close. bottom left. All the blastholes in the cut are placed with little spacing. three-hole pyramid. double Vee or baby cut.MGE 415 – Rock Excavation 2004 Figure 4-5: Angle cuts (Vee or wedge. top left. and is the most frequently used in tunnelling and drifting. The type of cylindrical cut most used is the four section. uses the Swedish theories and empirical updates. regardless of their dimensions. in line and parallel. an experienced miner would modify the pattern as ground conditions Dessureault 1/14/2004 109 . as it is the easiest one to mark out and execute. bottom right) Cylindrical cut The cylindrical cut can be considered a parallel hole cut. Beginning with an initial design. The large diameter blastholes (65 to 175 mm) are drilled with reamer bits which are adapted to the same drill steel which is used to drill the rest of the holes. and a draw cut. This type of cut consists of one or two uncharged or relief blastholes towards which the charged holes break at intervals. The calculation method for patterns and charges of this cut and for the rest of the tunnel zones. top right. which explains the frequent use of jumbos which come with automatic parallelism. It is considered to be an evolution or perfection of the bum cuts which will be discussed later on. Figure 4-6: Cylintrical cut Blast Layout Designing the actual blast is considerably difficult. and Francisco Javier Ayala Carcedo. Carlos Lopez. ground conditions. drilling into bootlegs is not permitted. Emilio Lopez Jimeno. to be modified depending on the cut location. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. ADDITIONAL RESOURCES XII Additional learning resources include: Readings. Chapter 22: Blasting for tunnels and drifts from: Jimeno. There are two mechanisms used to sequence the holes. and lengths of the blast. some blast the back last. Note that in most countries. Therefore cuts usually alternate from side to side. densities.1 – Drift Development (mandatory) Assignment 6 – in class (part of lecture) Dessureault 1/14/2004 110 . Balkema. or other factors that may improve the results. to ‘fluff-up’ the muck that has compressed from downward blasting). Rotterdam: A. (mandatory) Lecture Module 5. other the floor (ostensibly. hole spacing. Figure 4-7: Sequencing drifts Follow the mandatory readings to determine the charge lengths. Design specifications are typically made for a ‘standard’ for each rock type in a mine.A.MGE 415 – Rock Excavation 2004 change or as assessing blast results point to necessary changes. Figure 4-7a shows the general areas that are initiated in order based on blasting the floor last while Figure 4-7b shows a layout when blasting the back last. Development involves tunnelling. cross cutting.5 to 0. The most common types of explosive used include a cartridged nitroglycerine-based semi-gelatine or emulsion explosive with a composition density of 1.0 to 1. development and production. The length of the blastholes is about 1.. Short-hole Production In short-hole blasting the diameter and length of shotholes are usually limited to 43 mm and 4 m respectively. However.9 g cm-3. Capped safety fuse and igniter cord (burning speed of about 18 s m-1) make up the carrier of the initiating system. to prevent the hanging wall and footwall being damaged. In an average stope.MGE 415 – Rock Excavation 2004 Production Blasting Metal-bearing orebodies are extracted in underground mines by various methods. Short-hole blasting can be used in both stope and room-and-pillar mining.2 m.8 and 0. orientation. The usual blast pattern in a South African narrow reef of about 1.25 g cm-3 and a column of pneumatically loaded ANFO with a density of between 0. the present practice is not to detonate two or more shotholes simultaneously. Short-hole 'breast stope' blasting is most commonly applied in narrow. tubular orebodies such as gold or platinum reefs. there are about 120 shotholes per panel and. shaft sinking. The ends of the safety fuses protruding from the charged blastholes are connected sequentially with a trunkline of igniter cord.6 m. The production work can be subdivided into two categories: short-hole and long-hole blasting. raising.2 m stope width (or height) comprises staggered rows of 35 mm diameter blastholes. the burden of each hole being 0. Both rows are drilled at 70° to a line parallel to the stope face and close to the hanging wall and footwall. so that the ore bodies are easily accessible and transportable after excavation. depth and geological characteristics of the deposit. the excavation work is usually divided into two broad categories. depending on the size. Many mines use a pyrotechnic system of initiating explosives in the shotholes. as seen in Figure 4-8. etc. Dessureault 1/14/2004 111 . The method requires three distinct operations The formation of a tunnel.MGE 415 – Rock Excavation 2004 Figure 4-8: Short-hole production blast. The excavation of an empty space. Experience has established that the likelihood of igniter cord trunkline cut offs from rock projectiles decreases as the burning front increases. The drilling of sets of radial holes. The distance separating a detonating shothole and the flame front in the igniter cord is known as the burning front. parallel to the slot at appropriate spacing and burden. from a sublevel along the axis of the proposed excavation. called the ring drive. to the full width of the excavation. If a correct combination of burning speeds of safety fuse and igniter cord is achieved the shotholes are detonated in sequence at 4 s to 8 s intervals. The igniter cord itself is usually fired by an electric starter for igniter cord (ESIC). bench mining and vertical crater retreat. Longhole Production Blasting Basically there are three long-hole blasting systems: ring blasting. and is usually between 3 m and 5 m. at the end of the ring drive. Dessureault 1/14/2004 112 . The ESIC consists of a plastic capsule into which a standard electric fusehead (fitted with short lead wires to energise the system with an exploder) and a pigtail of fast igniter cord which has been crimped. called Rings. Ring Blasting Ring blasting has wide application in massive are bodies with their high rate of extraction at low unit costs. The powder factor varies between 1 to 2 kg/m3. called the slot. 3. measured at right angles to one. the spacing/burden ratio is about 1. the distance between two consecutive rings is called the burden. using construction lines (Figure 4-10). Normally. and straddling the outline of the are block. Dessureault 1/14/2004 113 . whereas the term spacing refers to the ends of the adjacent holes in the same ring.5. but it can be as high as 1.MGE 415 – Rock Excavation 2004 Figure 4-9: Ring Blasting In ring drilling. in order to achieve good fragmentation. A guide to the powder factor for some types of rocks is given in Table 4-1. Table 4-1: Powder factors for Ring blasting Since the blastholes in a ring radiate from a centre point of the ring drive. The blasthole Dessureault 1/14/2004 114 . an allowance of at least 30% expansion from the solid is recommended. Hence the blastholes need to have a variable stemming length (Figure 4-11) in order to avoid serious overcharging in the ore body close to the ring drive.MGE 415 – Rock Excavation 2004 Figure 4-10: Spacing construction for ring blast design. When blasting into a confined slot. Also the stemming length should not be more than two-thirds of blasthole length. It should be noted that the degree of success in ring blasting depends on the degree of accuracy in designing and drilling the blasting holes. the collars of these blastholes will be fairly close to each other. it is preferable to use one delay per ring.MGE 415 – Rock Excavation 2004 in each ring could be drilled upward as well as downward. Dessureault 1/14/2004 115 . Then. Figure 4-11: Stemming length for ring blasting Bench Blasting Bench blasting is essentially similar to surface excavation. If conditions permit. depending on the amount of pull (which could be 6. depending on the thickness of the ore body and/or the availability of drilling machinery. either vertical or horizontal blastholes are drilled to increase the height of the excavation (Figure 4-12). First a development heading is excavated at the top sublevel to provide drilling space. but this may generate a shattering effect in the adjacent rock. The blastholes can be from 32 mm up to 250 mm in diameter. This phenomenon can be avoided by using two or even three different delays per ring. These delays are not alternated between holes but are apportioned to whole sections of the ring. fragmentation requirement.0 m or more) and other factors such as quality of rock. etc. Dessureault 1/14/2004 116 .MGE 415 – Rock Excavation 2004 Figure 4-12: Bench Blasting Vertical crater retreat (VCR) Essentially. A cuboid of ore body can be excavated from the lower level upward by a number of horizontal slices using the same blastholes. vertical or subvertical blastholes are drilled downward from the top level to the bottom level. The average depth of slice achieved was 3. The optimum distance for positioning the explosive charge should be determined through small-scale crater tests using the same explosive-rock combination. The blasthole charging details have been well described in a case study where the blastholes were of 159 mm diameter and average length 35 m. in the middle of which was a primer attached to a 10 g/m detonating cord down line. followed by 1 m of crushed rock.4 m. Theoretically. In practice. to obtain maximum cratering effect. using a 6 mm diameter polypropylene cord. The size of this cavity can exceed the optimum distance of the charge from the back many times and its extent depends on rock properties and the local structural geology. two detonators of the same delay were used to initiate the down line. Finally. In the vertical crater retreat method.MGE 415 – Rock Excavation 2004 Figure 4-13: VCR It is imperative that the first set of charges in the blastholes breaks through into the undercut. spherical charges should be placed. Then a small amount of 16 mm crushed rock was poured into the hole to obtain a seal. The explosive charge was first stemmed by 1 m of sandfill. to the desired depth of blockage. however. First a square section wooden plug (100 X 100 X 200 mm) was lowered down the hole.8 m above the free face (Figure 4-14). The average explosive charge length was 0.8 m. gravity enlarges the crater dimensions by excavating the whole rupture zone. Dessureault 1/14/2004 117 . about 1. this is achieved when the deviation from the true spherical charge is not greater than a 1:6 diameter to length of explosive column ratio. A. It is important to use an explosive of high energy and high-detonation velocity in the VCR method to achieve maximum advance with minimum drilling costs. Balkema. Chapter 24: Underground production blasting in mining and underground tunneling.MGE 415 – Rock Excavation 2004 Figure 4-14: VCR loaded explosive column The main advantage of the VCR technique is the decreased possibility of damaging the surrounding rock which in turn reduces the risk of dilution. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Carlos Lopez. (mandatory) Dessureault 1/14/2004 118 . from: Jimeno. ADDITIONAL RESOURCES XIII Additional learning resources include: Readings. Rotterdam: A. This technique has been applied successfully in blasting out pillars between cemented cut and fill stopes. Emilio Lopez Jimeno. and Francisco Javier Ayala Carcedo. consider an area in the blast called the blast damage transition zone (BDT). The extent of each zone is characterized by a radius from the center of the production change. their extent (expressed as hole diameter) and the corresponding peak particle velocity (PPV values resulting when ANFO in medium strength rock are assumed can be similar to the value in Table 4-2.2 – Production blasting (mandatory) Controlled Blasting The following discusses more of the theoretical aspects of blasting design in avoiding damage control. As seen. The zones. it is a square pattern with 5 holes in each row. as seen in Figure 4-15. Table 4-2: BDT Characterization These zones have been drawn on Figure 4-16 for a blast containing 2 rows of production blast holes (9-7/8" diameter). a burden of 20' and spacing of 20'. Figure 4-15: Diagrammatic representation of the BDT of a fully charged hole.MGE 415 – Rock Excavation 2004 Lecture Module 5. To facilitate the discussion. In this example it has been assumed for simplicity that Rc=5D=4ft Rf= 12 D = 10 ft Ri = 55 D = 45 ft Dessureault 1/14/2004 119 . The object is to create an artificial plane of weakness which serves to limit the extent of the fracture and influence zones from both the production holes and any buffer (helper) holes placed between the final production row and the perimeter. Generally. these line-drilled holes are not charged with explosive but. This is shown with respect to the example case in Figure 4-18.MGE 415 – Rock Excavation 2004 Figure 4-16: DBT for a two row production blast After blasting. In the BDT zone there exists a crushed and a fractured zone surrounding each production hole. as the name implies. it is with detonating cord or a highly decoupled charge. one should take into account the BDT limit into account. Line Drilling Line drilling. The purpose for lightly charging the holes is to destroy the Dessureault 1/14/2004 120 . here the rock will be assumed to still be in place. The zone lying between the fracture zone and the boundary of the BDT consists of an inner portion which has been influenced by both rows of blastholes (to a distance of about 25') and an outer portion (20' in extent) influenced by just one row of production holes. Therefore in the design and implementation of any blast damage control techniques. if charged. Figure 4-17: BDT after blasting two production rows. the situation is shown in Figure 4-17. involves the drilling of closely spaced holes along the limit of the excavation. Although the crushed zone as well as a small portion of the fractured zone may have been loaded out along with the rock from production rows 1 and 2. The objective is to generate a line of cracks connecting the holes. However because of its high drilling cost. the hole spacing (c-c) should be 12". the method has not been commonly used in open pit work. Some recommendations for hole spacing as provided in Table 4-3. In this way. Dessureault 1/14/2004 121 .MGE 415 – Rock Excavation 2004 integrity of the rock web. close drilling control is essential for the method to succeed. Pre-splitting The pre-splitting technique also involves the careful drilling of relatively closely spaced parallel holes along the final perimeter.a smooth. it is intended to achieve nearly the same effects: Terminate the growth of the radial cracks Act as a barrier to the shock wave Provide an escape route for the explosive gases. Line drilling produces one of the best final surface . As can be appreciated. Figure 4-18: Line drilling positioned along the planned final perimeter Table 4-3: Factors for determining hole spacing In the example there is 45 ft wide zone between hole row 2 and the perimeter. Now however the holes are lightly charged and shot instantaneously. clean face with no backbreak or crest fracture. When line drilling 6" diameter holes in copper ore. To get hole spacing one multiplies the values in the table by the hole diameter expressed in the same units. The holes must be drilled so that they all lie in one plane corresponding with the dip of the final pit wall. For line drilling to be most effective: It must be used in conjunction with a buffer row The main excavation charges should be 1 to 3 rows from the pit limit. To excavate this rock one might consider using another row of production holes at normal burden and spacing and then one or two rows of smaller diameter buffer holes. For this type of blast. the presplit line is formed ahead of the main blast and allows the gas being driven back from the buffer row through the radial cracks to terminate at the presplit line.MGE 415 – Rock Excavation 2004 It attempts to produce the same effect as with line drilling but at a significantly lower cost. The choice can then be made between loading and firing the presplit line or infilling the main blast. Figure 4-19 illustrates typical presplit blast layout using 102-mm (4-in) diameter presplit holes for 250-mm (9y'-in) diameter production holes. As shown in the figure. The figure illustrates the upper bench where two benches will finally run together to form the final face between berms.) diameter hole and is inclined at 15° to the vertical. presplit holes would normally be drilled first. ahead of main production holes. The presplit angle is somewhat dictated by rock structure although a slight angle is preferred regardless of structure for long-term stability as well as for best initial results with large production holes. In the latter case. Figure 4-19: Presplit blast coupled to a 250mm production blast The presplit row in Figure 4-19 has a spacing of 2 m for a 102-mm (4-in. Obviously to be of any use the presplit line must be created prior to the blasting of any holes lying closer than I-BOT distance away. Dessureault 1/14/2004 122 . the presplit line would be fired instantaneously 100 to 150 millisecond before the main blast. Figure 4-20 is a graph of recommended hold spacings used in presplitting as a function of hole diameter. more accurately. termed the buffer row. The drill must be capable of drilling close to the previously produced bench face at an angle of 15° beneath itself so the face can be continued to depth. This is to prevent damage to the bench below or to the wall at that point. burden. Subsequently. Figure 4-20: Relationships between presplitting from various researchers.MGE 415 – Rock Excavation 2004 Presplit drill requirements become clear when presplit holes needed for the next bench are considered. hole diameter and spacing for Dessureault 1/14/2004 123 .) diameter. One further point to note from Figure 4-19 is the subgrade or. These data cover various types of material. Currently. The back row of the main production blast. lack of subgrade used on the presplit and buffer row holes. main blastholes after the buffer row are designed at regular spacing. must also be carefully designed with respect to standoff distance from the presplit row and spacing as well as explosives load. The inset sketch on the right side of Figure 4-19 shows how the top portion of the buffer row hole charge acts as a spherical crater charge breaking to the bench surface. but are not fully specified in the open literature. this means some form of drifter drill is required limiting the hole size to 102 to 127 mm (4 to 5 in. and loading for the type of material blasted. some minor crest fracturing or backbreak may result but the amount of damage is much less than would be produced by the main production blast if no control blasting was used. Often the relationship S/B= 0. the final pit perimeter lies in the zone of influence from the final row of production holes. Figure 4-21: Smoothwall damage zones There are five general followed in the design of the smoothwall row: Rule 1 The burden. as opposed to line drilling and pre-splitting. The hole size for the smoothwall and buffer row holes may be the same as in the production round with the required reduction in influence zone dimension occurring through pattern adjustments and decoupling or smaller diameter holes may be used with or without decoupling. This is shown diagrammatically in Figure 4-21. spacing and charge concentration of the smoothwall line of holes are selected so that the extent of the associated influence region does not exceed that of the production holes. Rule 4 The delay time between the helper row (that adjacent to the smoothwall row) and the smoothwall row should be chosen so that the smoothwall holes can shoot to a free face. Dessureault 1/14/2004 124 . Since the final row of holes lies in the influenced zone.8 is used Rule 3 The holes in the smoothwall row are shot on the same delay with detonating cord downlines to assure as simultaneous detonation as possible.MGE 415 – Rock Excavation 2004 Smoothwall blasting In smoothwall blasting. Rule 2 The hole spacing is less than the burden. bulk-loaded explosives in larger diameter holes is one way of improving the economics of the method. The boreholes are drilled in a line along the planned excavation limits. and blasted to remove the undesired material. Dessureault 1/14/2004 125 . Figure 4-22: Trim hole row trimming the fractured and influenced rock. Trim Blasting Trim blasting. as the name implies. Some rules for selecting the burden. Today. the holes should be drilled at the final pit slope angle. To achieve the best results. As noted earlier. The design process differs from the smoothwall technique in that the layout begins at the actual pit perimeter and works outward toward the desired final pit limit rather than vice versa. The trimming may be accomplished using one or several rows of blast holes depending on where in the BDT zone it is desired that the final crest should fall. involves trimming away some of the fractured and influenced rock from the pit perimeter after the production blast has been shot and cleaned up. smoothwall blasting is much more common to use than most other methods in mining operations because the method involves less drilling and less complicated blasting. loaded lightly. accurate drilling is important.MGE 415 – Rock Excavation 2004 Rule 5 All of the smoothwall and buffer row holes are shot together with the main production round. As in all other types of perimeter control. This is shown diagrammatically in Figure 4-22. spacing and charge concentration as a function of hole diameter are presented in Table 4-4. The use of low density. a reduced explosive load can be obtained in various ways. rows of drilled holes are loaded with explosives. a channel is created around large sections in the quarry. The explosives are detonated to free the sections of granite on all sides and on the bottom by explosives. This high-velocity flame. Likewise. lift these blocks to the quarry's rim (Figure 4-23). it is separated from the bottom by explosives. etc.MGE 415 – Rock Excavation 2004 Table 4-4: Recommended charge loads and blast geometries for trim blasting Blasting Ornamental Rock Ornamental rock is all stone that is used. The remainder is consigned to commercial applications such as street curbing and gravel. marbles and marmoreal limestones.000 degrees Fahrenheit. In this process. facility of elaboration. The most common types of rock can be generically classified in three large groups: granites. or derricks. Dessureault 1/14/2004 126 . when high-speed drills are used. diamond wire saws are used. etc. After a section has been completely wiresawed or channeled by the burner. In some quarries. for its aesthetic characteristics such as color. As the flame nozzle is moved up and down. or is sent to "grout piles" as waste products. is directed at the granite to be removed. A long loop of small steel cable. cuts the sections free from the bed of the quarry. causing a continuous flaking action. steel wedges are driven manually into holes previously drilled along the desired line of cleavage. shine. impregnated with industrial diamond segments. texture. Requirements for monumental granite are exacting. in blocks or slabs. Large cranes. The sections are readily forced apart and cross-wedged into rectangular blocks. The large sections are then broken into workable sizes by wedging. created by burning oxygen and fuel oil. polish. and technical such as strength. grain. Granite is cut from the "bed" of the quarry with a jet piercing machine that produces a flame burning at approximately 3. and only about 50 percent of the granite removed from the quarries finds its way into finished monuments. the following criteria should be of use: Drilling diameters. with mounted rock cutters.50 m and heights of 0. but generally around 25 to 45 mm. d = Diameter of the explosive charge (m). schistocity. fissurization. as there are many different rock types and exploitation conditions. Although it is difficult to give general recommendations for design in this type of blasting. but with slight variations as it is of maximum importance not to damage the rock and at the same time take into account the properties: strength. in parallel piped form. The usual interval is between 4 and 8 D.2 m. the formula suggested by Berta can be applied: S 2 PE S RT e d2 D D where: PEs = Specific pressure (MPa). generally lengths of 1. etc.8 to 3. Case studies show that the cutting methods consist in primary separation from the rock mass of a large block (100 to 4000 m3). Explosives. RT = Tensile strength (MPa). widths of 1 to 1. while for the horizontal planes explosives of low detonation velocity are also used.MGE 415 – Rock Excavation 2004 Figure 4-23: Ornamental stone materials handling – cranes. because cutting systems with helicoidial and diamond wire. It is established as a function of the rock properties and explosive charge characteristics. D = Diameter of the blasthole (m). The cutting technique is usually with explosives.9 and 1. The blasting techniques are a special type of presplitting.5 m. They depend upon the phase of excavation and the type of rig used. as they generate a large volume of gases. although not exclusively. which is subdivided afterwards to achieve sizes that are easily handled and within the ranges that the transformation industries require. To be able to make an analytic calculation. with flame torching and with water jet kerfing are often applied. homogeneity. Spacing. e = Density of the explosive (g/cm3). Dessureault 1/14/2004 127 . In the vertical benching planes detonating cords with a core of pentrite are usually used. 3g/cm3. The explosive columns are generally designed to be continuous and decoupled with an air chamber although. As the rock characteristics become poorer. They are necessary to use the maximum pushing power of the gases. Initiation. As in contour blasting. to increase the energy transmitted to the rock by the detonating cords. These vary greatly depending upon the type of rock. Example A block of granite is to be extracted by drilling blastholes and blasting with detonating cord. Drilling diameter D = 0. it is suggested that empty holes be used at the end of the line or next to the free surfaces. Apart from this. instantaneous initiation of all the blastholes with detonating cord is recommended. Specific pressure PEs = 1200 MPa. in some cases such as in hard rock. Powder factors. the structural properties of the rock mass should be used to advantage. made up of nitroglycerine.032 m. in vertical blastholes there is no subdrilling and they are usually drilled to a few centimeters above the horizontal plane. explosive and extraction phase. Charge configuration. with the detonating cords the stemmings are small.14m Dessureault 1/14/2004 128 . S 2 1200 1. there is extensive use of charges prepared in connecting plastic tubes that contain powdery explosives with low density and detonation velocity.MGE 415 – Rock Excavation 2004 In these last planes of the cut. Distribution of the charge in the borehole. In general. Diameter of the detonating cord core of pentrite d = 0.032 0. In order to eliminate breakings or fracturation in the corners of the blocks. the blastholes are filled with water. sodium nitrate and other ingredients.032 0. if a blackening by explosion smoke of the cut surfaces is to be avoided. Density of the pentrite charge e = 1.3 0. What should the spacing between boreholes be when the initial data is: Tensile rock strength RT = 10 MPa. whereas with powder a larger confinement is necessary.0034 m. In some countries. Also. Stemming.0034 2 10 0. the holes can be filled with sand or drilling waste. the heights are usually shortened to assure that the pressure of the gases do not act upon the rock for a long period and therefore produce damage. which is within the practical interval of 4 to 8 D. Most underwater blasting is done by method 3. Using divers to drill and to charge the blastholes. and the corresponding charge ratio varies between 3. the spacing should be increased to S = 0. The procedure is then the same as for land excavation. that is.26 m. Using a barge or platform to drill and charge the shots.MGE 415 – Rock Excavation 2004 The ratio S/D is equal to 4. This is accomplished in a number of ways: 1. from a platform. in deeper water. Individual charges are normally 10 to 25 kg. 3. depending on rock type. In deep water (say >10 m) this is the only method that can be applied economically in the present state of technical knowledge. Using shaped explosive charges fastened in a predetermined pattern onto a metal frame or high explosives packed in boxes and connected to a line of detonating cord and then lowered onto the sea bed. Method 1 is only economical if the water is shallow and the area of excavation is small. Dessureault 1/14/2004 129 . a charge becomes effectively tamped by the pressure of the water. or excavate trenches for pipelines. which may be of either the floating or the jack-up type. Its success depends on the intimacy of the contact between the explosives and the rock. 2. The most important factor is the placement of the charges in the designated area.37. 0D rigs are fitted with chucks so that the equipment can drive a hollow outer casing of steel tubing through soft material. Method 2 is fairly labor intensive but economical for excavating selected small areas. Overburden drilling (0D) rigs are used by both types and the blastholes are normally between 51 and 102 mm in diameter. whose thickness to be removed by blasting in one pass is generally limited to one meter.5 and 7 kg/m3. The technique can also be extended to winning underwater mineral deposits. Moreover. and consequently the efficiency of the explosive in breaking rock is increased. If the rock were of worse quality with a tensile strength of 5 MPa. Strict control on positioning each blasthole at its designed place is extremely important for the success of this method. Underwater Rock Excavation Underwater blasting of rock is usually done to deepen the sea bed in and around a harbor to allow the passage of larger vessels. as well as drill rock through the casing. Then the casing is uncoupled from the drill chuck and the normal drill string is inserted through the hollow casing for blasthole drilling (b). the drill string is withdrawn (c) and the charging procedure takes place through the hollow casing as follows: the primer cartridge with detonating cord or signal tube is pushed through the hole first by using a string of stemming rods with a flat-ended wooden coupling for easy handling of explosives (d). The sequence of operations is shown in Figure 4-24 and can be described as: The outer casing. The gantry is then moved to the next line of holes and the operation repeated.MGE 415 – Rock Excavation 2004 Figure 4-24: Sequence of operations in underwater rock excavation. the tube or cord is retrieved from inside the tube (e) and brought up to deck level. The required cartridges are then added. The explosive used in this type of operation should have high velocity of detonation and high density. a slipring attached to a hemp rope is lowered to the bottom. grips firmly into the top part of the bedrock (a). is driven through the water until the ring bit. and the latter is placed on a roller away from the drilling activity. by rotating. When the required depth is reached (Note: normally. It should have a good degree of water resistance so that 24-hour immersion does not affect its performance. A weight is attached to the end of the cord or tube. the subgrade length is the same as the burden). When charging is complete. The length of the cord or tube should be adequate to allow for rise and fall of the tide when it is brought up to deck level inside the casing. fitted with a hollow ring bit with serrated edges. Then as the outer casing is withdrawn. yet its properties should be such that it would be made inert by a long period (say Dessureault 1/14/2004 130 . homogeneous rock. weaker rock. or mass decked loads. face coverage 0. the higher ratios being necessary in rock of higher tensile strength.4 lb/ft in-hole. the higher values associated with stronger rock and/or larger hole diameters.MGE 415 – Rock Excavation 2004 one month) of immersion.1-0. Typical charge densities 0. 5. Decoupling avoids shattering and reduces gas penetration. If the ratio is too high a more rugged.2-0.0 kg/m3 depending on rock conditions.18 lb/sq ft. with the lower values necessary under less ideal conditions. 6.0. Trim blasting may be more cost-effective in such cases. in meters. decoupling ratio (dpowder/dh). or open fractures spaced 0.5 and 2.3. presplit powder (skinny segmental or linkable cartridges).6S or less. 2. 16 (to 18) for trim blasting. even closely-spaced presplit will tend to break to such fractures.3S. air decked loads. and R is the distance from the charge to the point affected by pressure. Typical decoupling ratios = 0. The powder factor for this type of operation varies between 0. 7.10 . fracture density. cartridges taped to cord at intervals. 3. Typical S/dh = 10-14 for pre splitting. 9. and there is essentially no way to compensate through loading or timing adjustments. and fracture orientation. Use linear. in order of increasing potential for rock damage. but may require decreasing spacing by 25-50%. It is estimated that a peak waterborne overpressure not exceeding 40 kPa is safe for humans and animals. charge density (charge/unit length) and drill accuracy. Rock with closed fractures spaced less than 0. Critical geologic parameters are rock brittleness. more damaged slope will develop. Dessureault 1/14/2004 131 . 1. 8. Fracturing striking less than 15 degrees to a presplit line is very difficult to overcome. distributed charges with cord (50 to 400 grain most common). will dissipate presplit elastic stresses. trim blasting in fractured. Controlled Blasting Rules of Thumb . Presplit works best in brittle. The waterborne blast-induced overpressure P (kPa). 4. Critical geometric parameters are hole diameter.Supplemental The following points are controlled blasting – presplitting and trim blasting – additions that blasting consultants consider important. S/dh is more likely to be too large than too small. can be calculated approximately from P m 3 55 10 R 3 1 where m is mass of explosives in kg. spacing/diameter ratio (S/dh. Unless the fractures are very tight and the rock is very weak. Compensate by closer hole spacing. Fracturing striking 15-45 degrees is less difficult. rock strength. flyrock. and then add time as necessary. hydraulic bursters. Refine standoff according to the buffer/production arrangement. there is a potential danger of environmental disturbance. and ground vibration. delaying 9 ms between holes or groups of holes. 11. More time is better. 13. etc. fumes. dust.MGE 415 – Rock Excavation 2004 10. and then to keep meticulous records of all blasts. to permit release of gases. It is now appropriate to analyze the energy distribution after the detonation of a quantity of explosives.Standoff between presplit and buffer (if used) or production holes should be 0.3-0. 0. Stem trim holes 0.1. Techniques involving explosives differ from other available systems in that the energy applied is released in a matter of milliseconds. Rock can also be excavated by methods other than blasting such as the use of rippers. changes in the natural profile of the ground. it is good practice to carry out a detailed survey of all properties that might conceivably be considered at risk of damage. 14. The blasting engineer should be aware of the need for defense against allegations of damage caused by the above factors.75 if trim blasting). ADDITIONAL RESOURCES XIV Additional learning resources include: Lecture Module 5.0S. add an extra delay period or two for trim hole relief. start by grouping presplit in as many simultaneously-fired holes as possible.Presplit shot together with production must be timed at least 50 ms ahead of small-diameter (<4") production holes.Angled controlled holes perform best. Before any explosives are used. To meet vibration concerns.Presplit holes ideally are not stemmed.3 – Controlled Blasting (mandatory) Guest Lecture – Bob Cummings Environmental Issues The use of explosives is probably the most widely used means of fracturing rock. The shot width should be at least 3 S (spacing lengths) for presplit confinement. 12. at least 100 ms ahead of medium-diameter (4"–8”) production holes. but with less effectiveness. and at least 200 ms ahead of large diameter (>8") production holes.5 x BaProduction (0. presplit holes can be delayed as much as 25 ms hole-to-hole. plugs and feathers. To control noise. Environmental disturbance includes the effects of airblast. The back break from production or buffer holes is needed for additional fragmentation in front of controlled holes. they can be stemmed. if the resulting charge weight exceeds vibration/scaled distance criteria. Trim blasting: hole-tohole timing is best.5S < T < S.5-0. The explosive's potential energy can be Dessureault 1/14/2004 132 .Presplitting: simultaneous firing is ideal. but seldom are drilled flatter than 50 degrees or so. If the energy release process is not adequately controlled.7S . However. On the other hand. and ground vibration. with the consequent wastage of explosive energy in the air. the majority of the explosive's energy will be manifested as ground vibration. Severe airblast is caused either by inaccurate charging. For example. whereas if the same quantity of explosive is buried very deep in the rock. The resulting airblast overpressure may be estimated from the following formula: 1.3 R m 1 3 When assessing the effect of airblast it is usual to correlate this with the possibility of cracking panes of glass. or by the firing of exposed detonating cord in a sensitive area. inducing the waves to bend upward away from the ground. an increase in sound wave velocity results. if the air becomes warmer with increasing altitude. the percentage distribution of the above factors may vary. The pressure waves generated by 133 Dessureault 1/14/2004 . Moreover. For confined borehole charges. except for heat emission. if there is a strong wind in a particular direction the refracted sound waves are channeled into the wind path. causing the waves to be returned to ground by refraction. if a quantity of explosive is detonated on the ground surface. airblast. which will be constant in all situations. m = explosive charge (kg) R = distance from charge (m).2 p 185 R m 1 3 where p = pressure (kPa). Depending on the placement of the explosive charge. fly-rock. Airblast Air overpressure due to an explosive charge (for a given amount) increases rapidly as the confinement decreases.MGE 415 – Rock Excavation 2004 manifested as heat. airblast overpressure may be estimated from 1. The ideal blasting condition for minimum airblast effect is where temperature decreases with increasing altitude. Each aspect which can cause environmental problems is now discussed. the major part of the explosive's energy will be converted into airblast. The most damaging cases of airblast are caused by unconfined surface charges. causing a decrease in the sound wave velocity. rock fragmentation. This phenomenon greatly increases the airblast pressure at focal points.2 p 3 . 69 6. In sensitive areas. exposed surface lines of detonating cord. causing the sound waves to bend downward towards the surface. Normally. While the most frequently mentioned complaint resulting from airblast is cracked plaster. A combination of wind and temperature inversion can take place several kilometres from the blast and the overpressure due to airblast may increase by a factor of 100.9 >6.03 0. The technology for reducing noise and vibration from pneumatic and hydraulic drill operations without losing performance does not yet exist. The above conditions apply on the assumption that there is no appreciable wind. The sources of objectionable blast noise are: the use of lay-on or plaster charges often used in secondary blasting. However. a covering of 20 cm of stemming material over detonating cord will bring about a significant reduction in noise. Atmospheric conditions such as temperature inversion and wind can lead to the phenomenon of focusing. temperature decreases with height. but can damage structures. poorly stemmed holes. and sometimes even painful. On the other hand. Dessureault 1/14/2004 134 . but most of which occur below 20 Hz frequency in the concussion range. Certain well-known symptoms caused by the peak overpressure levels are given in Table 4-5. Usually ideal atmospheric conditions for blasting exist in the early afternoon. in a temperature inversion (air temperature increasing with height). close attention to the design of the canopy for compressors can bring about almost silent machines. The engine should be fitted with twin silencers and the inside of the canopy should be coated with sound-absorbing plastic foam to eliminate natural panel resonances. causing the sound waves to curve away from the surface. which is not detectable by the human ear. the velocity of sound in air also decreases with height. some of which are audible. These airblast wave effects can be seen from Figure 4-25. blown-out shots resulting from poorly designed blast patterns. sound velocity increases with height. research has shown that window panes fail before any structural damage results.9 Overpressure (dB) 140 150 170 >170 Noise from drilling and blasting can be annoying. Table 4-5: Airblast Effect on Structures Symptoms Rattle of loose windows Failure of poorly-mounted window pane Damage to strongly-mounted window pane Cracking plaster Peak (kPa) 0.MGE 415 – Rock Excavation 2004 explosives consist of energy over a wide range of frequencies. correspondingly. Moreover. drilling pattern. The major causes of fly-rock in mining are shown in Figure 4-26. Dessureault 1/14/2004 135 . Proper stemming has an important role in trapping explosive gases in the blasthole to do useful work such as rock fragmentation and throw. Crushed angular rock is recommended as the ideal stemming medium and the stemming length should not be less than the burden distance. the explosive's gas energy is vented violently into the atmosphere and propels rocks in front of it. deviation of blastholes in surface mining operations can effectively reduce the burden. When this is compromised. causing flyrock. and charge ratio. In a particular type of rock there is a compatible relation between the height of the explosives column in the holes.MGE 415 – Rock Excavation 2004 Figure 4-25: Airblast waves in various conditions Fly-Rock Fly-rock is the term for undesirable projectiles of blasted material. MGE 415 – Rock Excavation 2004 Figure 4-26: Causes of fly rock. There are a number of different situations in which the actual conditions depart markedly from the ideal: The explosive extends too high in the hole so that cratering to the upper surface occurs. Moreover. an out-of-sequence shot has the same effect as an overburdened hole and is liable to cause fly-rock. The lack of confinement offered by both of these situations provides a weak link for the gas to exploit. When the specific charge (q) is q 0.2) where d = hole diameter (inches).12. In this case the maximum throw expression becomes: L=57d For a 10 inch hole diameter the maximum expected throw would be L = 570 m. q = specific charge (kg/m3). A typical specific charge in bench blasting is 0. For other values of q the maximum throw is expressed by L=143 d (q-0.2 kg/m3 there is no throw. The rock plug involved is pushed out in an early Dessureault 1/14/2004 136 . An irregular face brings the explosive column too close to the free face resulting in cratering.6 kg/m3. L = maximum throw (m). Some very limited field studies reported by various researchers that suggest that for granite the maximum throw (L) as a function of the hole diameter (d) and specific charge is as shown in Figure 9. In excavation work close to buildings. special precautions may be taken such as placing a thick rubber mat (such as heavy conveyor belting. Of all various types of flyrock. Dust Dust is mainly produced while drilling and could be a considerable nuisance in high winds. In a properly designed shot for rock blasting. in demolition work. If the weakest link in the system is the column of stemming and not the collar rock.MGE 415 – Rock Excavation 2004 stage of the gas expansion process and the expansion energy is expended in propelling a relatively small volume of rock at high velocity. the most dangerous situation occurs when cratering occurs on the top part of the bench near the collar. This is the application of a carefully designed blast pattern in the berm (bench). or using wet drilling. However. or can be filled with water to form a recreation area. very little dust is scattered. It was found that for granite with a specific gravity of 2. If the mine working is below the natural profile of the surrounding country the void can be infilled with waste material. dispersal of fine dust is inevitable. and has proved to be successful in providing an aesthetically pleasing landscape. a turmoil of the natural surroundings is often experienced with the end result of residual benches of bare rock. and this will be the medium for selective establishment of vegetation. = 0. This can resemble a deep scar in the immediate environment. Hence the throw distance can be very great. Apart from incorporating a dust collecting unit in the drill rig itself. The screen blast piles will mask the earlier scorch marks. Thus for a 10 inch diameter borehole the boulder size would be 2/3 = 0.1 (10) Disturbance of the Natural Ground Profile In an open-cut mining operation by blasting. = boulder diameter (m).47 m. Clearly the driller should wear a protective anti-dust mask. this can be ejected much like a projectile from a cannon barrel. the relationships for the maximum throw (L) involving rocks of diameter Lmax = 260 d2/3 = 0. if the workings are in elevated ground where older workings are actively eroding with rock fall and spalling onto the mine floor or benches beneath. This could be partly controlled by sprinkling a jet of water.1 d2/3 where d = hole diameter (inches). the problem requires different treatment.6. A method of restoration blasting has recently been adopted in various areas. very little can be done to suppress dust from this operation. particularly of a tall structure. and thereby creating an artificial curtain of fine mist to contain the falling Dessureault 1/14/2004 137 . However. Some empirical and computer simulation have examined the maximum throw and boulder size as a function of hole diameter. initiated to reproduce a predicted sloping profile with infilled materials. The longitudinal. Jets of water are also sprayed in some surface mining operations where dust has caused problems in nearby habitats. Since structural damage. The presence of water in the blasthole may also adversely affect the fumes produced by a blast. This aspect is adequately safeguarded by legal requirements laid down by the relevant authorities. in general. the use of peak particle velocity is accepted as the principal parameter for vibration measurement. etc. This excursion or oscillation of the individual particles is measured to determine the magnitude of the blasting vibration. As the pressure wave travels away from the borehole it forms a seismic or vibration wave by displacing the particles around it. Calculation of the resultant of these three components is sometimes preferred and countries have different standards as regards the suitability of measuring planes. in addition a small amount of undesirable poisonous gases. a pressure wave will be generated in the earth surrounding the hole. Ground Vibrations When an explosive charge is detonated in a borehole. known as fume or toxic gases. and transverse components of the seismic waves caused by blasting are generally measured by instruments. are hazardous in confined spaces such as underground mines. Careful thought must be given to mechanical ventilation of such areas with fans so that fumes are diluted to a harmless level. is strain related. However. carbon dioxide and nitrogen. such as carbon monoxide and oxides of nitrogen. vertical. The vibration is largely influenced by the prevailing constriction and by certain physical site factors. Fumes The detonation of a commercial explosive in ideal conditions produces water vapor. or the resultant. The same amount of charge at a fixed distance does not necessarily produce the same magnitude of vibration.MGE 415 – Rock Excavation 2004 dust in a limited area. The fume characteristics of a cartridge explosive relate to the conditions where the explosive is fired in its cartridge. Removal of explosive from its cartridge will upset the oxygen balance and unfavorably affect the explosive's fume qualities. tunnels. The peak particle velocity (ppv) of ground motion can be related to distance and instantaneous explosive charge by the following equation: Dessureault 1/14/2004 138 . For elastic wave transmission the strain is directly proportional to particle velocity. that is. The safest maximum charge per delay becomes 36 kg from: Dessureault 1/14/2004 139 . a test shot (using a small amount of explosives so that the SD value is well over 12) has to be monitored at various distances. respectively. The constants K and P are obtained. Allowing for a certain amount of scatter due to the variation in blasting constriction. several (at least 6) vibration recordings are taken. Such a graph is known as a regression curve. vertical (ordinate) for ppv and horizontal (abscissa) for SD. The simplest method of establishing the site constants is to use log-log graph and put the measured values straight onto the axes. three values of SD can be obtained: 30. from the intercept of the regression line on the ppv axis. Obviously. K = site factor constant. the magnitude of ppv decreases. When blasting has to be done close to a sensitive structure. In a given situation when the same blasting is further away (the SD is more than 12). From this. and by calculating the slope of the regression line. as the value of SD increases.MGE 415 – Rock Excavation 2004 P ppv K R m 1 2 where R = distance in meters. m = explosive charge in kg. P The ratio R m 1 2 is commonly known as scaled distance (SD). the following procedure determines the optimum charge. it can be assumed determined that SD = R m 1 2 = 50 will be the best solution. two other lines parallel to the first one are also drawn to encompass all points in the graph. the graph is a straight line. 40. Figure 4-27 is such a graph obtained by plotting ppv against SD for 10 monitored results. if there is any possibility of causing damage to the structure by normal production blasts. 50. and P = site exponent constant. SD is a useful parameter for comparing one set of vibration results to others. Example Consider that there is a sensitive structure where maximum ppv is stipulated as 5 mm s-1. On the other hand. and this structure is 300 m from the blast area. what will be the maximum instantaneous charge? By drawing a horizontal line from ppv = 5. After drawing the best fitting line. MGE 415 – Rock Excavation 2004 m R 50 2 300 50 2 62 36 Figure 4-27: Obtaining site factors for ground vibrations. The factors listed in Table 4-6 identify the phenomena important in blasting. Dessureault 1/14/2004 140 . It is essential for the blasting operator to know and to be able to control what will happen when an explosive charge is detonated. A collection of scaled distance data for determination of peak particle velocity of some 2500 blasts from 40 mines. The line labeled 50 is the best-fit line.MGE 415 – Rock Excavation Table 4-6: Factors that influence ground motion 2004 Figure 4-28: Ground vibrations results from 2500 blasts in 40 different operations. The factors responsible for this scatter are the changes of geological conditions. difference in the geometry of the blasts. the graph being drawn from a considerable amount of scatter. and experimental error. whereas the other two are bounds below which 84% and 95% of Dessureault 1/14/2004 141 . quarries and construction sites in the USA is shown in Figure 4-28. differences between types of explosives. Table 4-7: Recommended maximum peak particle velocities. a working recommendation is a maximum ppv of 5 mm/s. the size of the vibrations can be controlled. The maximum ppv recommended by Standards Australia is given in Table 4-7. particularly if they are in an indifferent state of repair. drilling pattern. overlay soil types Practical methods to reduce the ground vibrations by limiting the cooperating charging weight per interval are: Adapt the ignition pattern so that the charging level is spread over more intervals and the scattering in the delay elements of the detonators is utilized Reduce the number of holes and the hole diameter Use decked charges by dividing up the necessary charge level in a drillhole into more ignition intervals by means of sand plugs Use decoupled charges. charge diameter smaller than hole diameter Divide the bench into more benches Do not blast to the final depth at once. the size of the vibrations depends on: Cooperating charge Confinement conditions The character of the rock The distance from the blasting site The geology. so that all the holes will break the burden in the easiest way Increased hole inclination (of the drillhole) A voiding too large burdens and choke blasting For blasting at a shorter distance than 100 m. commercial 10 buildings not included in description below Commercial and industrial buildings or structures of 25 reinforced concrete or steel construction This table does not cover historic buildings. or some sensitive structures such as reservoirs and buildings with long-span or suspended floors.g. there should be as little confinement as possible.MGE 415 – Rock Excavation 2004 the data fall. and ignition pattern. e. Type of building or structure Particle velocity mm s-1 Houses and low-rise residential buildings. In the absence of particular site-specific data. Among other things. Vibration reduction By adapting the blasting method. charging scheme.. At the moment of detonation. the risk for interaction between the different intervals is small. This can be obtained by: A carefully adapted ignition pattern. The risk for cooperation between the Dessureault 1/14/2004 142 . 8 ms in order that the two charges be considered separate charges. or greater than. cooperate or not depends on the following factors: Time interval between initiations Dessureault 1/14/2004 143 . while at a long distance (B). Using delay detonators. The expression cooperating charge is somewhat inappropriate as it is only applicable at certain distances.smaller or greater number of drillholes. the charges do not cooperate. where any structures sensitive to vibrations are situated a large distance away. The increase in costs primarily depends on the following factors: Drilling . it was shown how the maximum detonating weight of charge can be estimated when the vibration level and the distance are known.more detonators and higher cost of labor Blasting . initiated one after the other. the USBM states that the delay interval should be equal to. interference and reinforcement of the two vibrations may occur. the larger a total charge weight per interval can be used because the scatter of the delay time increases with the nominal delay time. it is possible to blast rounds with considerably higher total charge weights per delay interval. The cooperating charge is defined as the total charge per interval. The higher the interval number. The weight of the charge in question is the maximum total weight of charges that can be initiated at the same time. In the USA. multiplied by the reduction factor appropriate for the interval used. Whether two charges. Charging .MGE 415 – Rock Excavation 2004 intervals increases with large blast in quarries: for instance. The size of the vibrations is then influenced by: Charging level Interval times The resonance frequency of the ground (which depends on the depth and the character of the ground) The local geology The cost for careful blasting near built-up areas increases very rapidly with decreasing permissible vibration level.more rounds and longer stand-up time The costs of planning and control work will also increase in: Blasting Visual inspection Vibration measurement Blasting record Insurance administration Cooperating Charges – Ground vibration In the preceding section. if the vibrations are observed at a short distance (A). When blasting two separate charges with detonators having the same interval number (Figure 4-29). If two seismic waves resulting from two blastholes meet at a point. The effect of Delayed Explosion Reduction of explosive weight per delay is perhaps the greatest factor in reducing the probability of structural damage due to ground vibration. the frequency will be reduced by half. However. the period and frequency are those of the single wave. and the expected ground vibration is given in three different situations. Table 4-8: Effect of delayed charges on ppv for a total charge of 2900 kg at a distance of 220m.MGE 415 – Rock Excavation 2004 Velocity of propagation of the vibrations in the bedrock The decay time of the vibrations Distance from the observation point to the charge The geometry of the charge Velocity of detonation Confinement. The cardinal point for successful delay blasting is that the seismic waves from any detonating blasthole shall pass all other blastholes before any of them are initiated. the ground vibration monitoring distance is 220 m. Table 4-8 highlights this point for a situation where the total charge is 2900 kg. Figure 4-29: Cooperating charge related to distance. Moreover. the resultant motion will be the sum of the two motions and the vibration level will be significantly increased. Mode of Detonation Instantaneous Two delays with equal charges Four delays with equal charges ppv in mm s-1 75 43 25 Each delayed charge generates its own seismic wave which is separated from the subsequent delayed wave. Since the period may approach twice that of a single wave. precise timing for the detonation of each hole is imperative for effective delay blasting. the seismic wavelength of the composite motion varies from a single wavelength to nearly twice the length of a single wave. This condition may Dessureault 1/14/2004 144 . At the point of maximum overlap. Where small-diameter holes are used. Risk Assessment Near most sites where blasting is necessary. To optimize the blasting work.MGE 415 – Rock Excavation 2004 produce a region of high seismic risk due to increased motion and reduced frequency of vibration. what size vibrations the environment will accept and second. which may have an influence in focusing seismic waves. the absorption characteristics of the medium. Has a geological examination been made regarding working site as a risk area? 2. The probability that a correct decision will be made increases as more information is made available. Too large a cooperating charge may result in damages to neighboring buildings. this maximum charge may be the charge having the same vibration effect as that from several holes detonated with detonators having the same nominal time delay. the moisture content of the medium and the ground-water level. how large a charge can be blasted at a certain distance without exceeding that vibration limit. there are structures whose sensitivity will limit the maximum permissible vibrations. Checklist for risk analysis: 1. the topography and the geological structure. a risk analysis should be made involving a careful examination of the factors that can affect the blasting operations. It is called the maximum cooperating charge. Client or contractor must then decide on the maximum charge that can be detonated without causing damage in the neighborhood. What is the nature of foundations and underground parts of buildings in the area? 4. which determine the propagation velocity of the seismic waves (predominant range and type of waves). the economy of the blasting job is greatly influenced by that decision. What is the type of construction and the condition of buildings within the risk zone? Dessureault 1/14/2004 145 . and even court disputes. Because the cost increases considerably if the maximum cooperating charge has to be reduced. first. Before the blasting operations can begin. Is there a potential risk for lowering the groundwater level? 3. damage claims. it is necessary to carry out risk analysis in order to determine. Too conservative a decision on the size of the maximum cooperating charge will result in excessively increased costs and project time. The decision data should be based on as many points as possible in the list below. Effect of Geological Factors The intensity of ground vibration is often influenced by the following properties of rock: the elastic properties of the medium. A. Rotterdam: A. and distances? 8.4 – Vibrations (mandatory) Assignment 6 – in class (part of lecture) Evaluation of Blast Results Once the blast has been carried out. fumes. Balkema. deal primarily with communities around the explosives consumers and manufacturers/transporters. To achieve a global evaluation. aside from esthetics of the excavation and implications of the operation itself. “Drilling and Blasting of Rocks” translated by Yvonne Visser De Ramiro. Presence of boulders in the pile. Has information about the blasting job been distributed to the neighbors? Sustainable Development Issues in Rock Excavation Sustainability issues in rock excavation. cable trenches. laser equipment. Geometry. ground vibration. telegraph cables. transportation or storage accidents. and dust. State of the remaining rock and bench floor. cooperating charges. oil cisterns. the issue of explosive theft is important as criminal elements can obtain explosives for illicit purposes. engineers should be more sensitive and knowledgeable of the needs of the community. Emilio Lopez Jimeno. Chapter 33: Land Vibration. and Francisco Javier Ayala Carcedo. relays…) which are sensitive to vibrations in the neighborhood? 6. As the importance of the community’s concerns continue to rise. Adequate storage and compliance with all licensing and transportation safety protocols should be taken. How are the inhabitants in the neighborhood influenced? 9. ADDITIONAL RESOURCES XV Additional learning resources include: Readings. district heating culverts. fly-rock. it is necessary to analyze the obtained results.. What connections are there between vibration values. Is there any equipment (such as computers. height and displacement of the muck pile. Are there any underground objects (tunnels. Carlos Lopez. (optional) Lecture Module 5... the following aspects must be analyzed: Fragmentation and swelling of the muckpile. fly rock and airblast produced by the blast. air blasts and their control from: Jimeno. Communities can be exposed to noise. Adequate information is key. This is the basis for the optimization process. Vibrations. Dessureault 1/14/2004 146 . In some regions. electron microscopes.) that might be damaged by blasting? 7.MGE 415 – Rock Excavation 2004 5. as its interpretation will give the successive modifications of the design parameters for the following rounds. In general. it does not allow an exact distribution of the sizes and very frequently there is no written report on the results. Due to high costs and necessary time to achieve the complete size distribution curve. there is no method which enables a quantitative evaluation of fragmentation in conditions that would be trustworthy. Size distribution is the basic tool within the optimization process of blasting. as seen in Figure 30. Photographic method This technique has been used in various ways and methods of analysis. placing scaled targets in the field of vision for dimensional purposes. even if the person in charge has vast experience.MGE 415 – Rock Excavation 2004 Fragmentation and Swelling of the Muckpile Apart from the classification of size distribution or screening of the muckpile in treatment plants. Dessureault 1/14/2004 147 . Some researchers used photographs of the muckpile and chose at random around 15% of the total to analyze sizes and number of the fragments with a superimposed grid. The muckpile and the general aspect of the blast is observed immediately afterwards and the engineer responsible for the evaluation makes a subjective assessment. Reid used a series of photographs of the muckpile at different moments of digging. changes in fragmentation can only be distinguished when the differences are great. After studying five blasts under similar conditions. it is only good for an initial contact with the blast results so that the specialists can later make a complete study.6% in the mean fragment size. the results showed a deviation of more or less 9. However. The application of this technique is not very precise. as it is the only means of comparing the fragmentation obtained when a study is to be done on the sensitivity of the design parameters. The most important source of error was in assuming that the surface fragment distribution was representative of the total mass. in mining operations the following approximate methods are used: Qualitative visual analysis Photographic methods Photogrammetry methods High-speed photography Study of loading equipment productivity Volume of material that requires secondary blasting Bridging delays at the crusher Partial screening Image analysis by computer. Qualitative visual analysis This is the most widely used and usually the only analysis applied. etc. Photogrammetric method The photogrammetric methods give more precision than in conventional photographs. The only inconveniences are: Time consuming in preparation and study. The advantage of the photogrammetry is that it allows a tri-dimensional study of the muckpile. Dessureault 1/14/2004 148 . thereby aiding in the calculation of each fragment size and the volume and swelling of the pile.MGE 415 – Rock Excavation 2004 Figure 30: Gird method of evaluating fragmentation and field environment This system is one of the most useful 'and also gives a graphic documentation for analysis and comparison of different rounds. design and. above all. Some have considered it as a technique for the evaluation of fragmentation. and It is difficult to quantify the small sizes. High-speed photography The use of high-speed photography in the evaluation. are the need for qualified personnel to use and interpret the method. apart from cost. The main problem is that the gases and dust in the environment obscure the vision of crack formation and muckpile displacement. The drawbacks. geological studies. it is only used as a complement to the main applications such as topographic control of the cuts and muckpiles. control of the blasting has been widely undertaken. As the investment in equipment and accessories is relatively high. Quantitative: o Exit time of the blast accessories o Time and efficiency of the gas confinement o Acceleration. along with detailed information of the design parameters of the blast and of the whole of the operation.MGE 415 – Rock Excavation 2004 The information taken from the analysis of the high peed photographs can be classified as: Qualitative: o First rock movements o Confinement of the stemming o Trajectory of the muckpile movement. Dessureault 1/14/2004 149 . of the presence of water and of the stemming Determination of the best initiation sequence Yield of the chosen initiation system Global movement of the muckpile Source of oversize at the face Muckpile displacement and Profile geometry of the muckpile. is very useful for the detection and definition of: The existence of misfires and their causes Incorrect explosive charges The effect of sub-drilling. direction and velocity of the fragments o Velocity of stemming ejection o Projection and displacement of the muckpile. The study of these data. MGE 415 – Rock Excavation 2004 Figure 31: High speed camera bench blast Dessureault 1/14/2004 150 . perimeter. Stereometric interpretation. A grey level is defined. such as waiting for the trucks. then the pixels with values above this level are turned to white (fragments). such as the area. erosion and line thinning is used. The scale of the image is defined. The presence of oversize. 6. 5. 3. The studies should be Dessureault 1/14/2004 151 . For example. a precise evaluation can be obtained. etc.MGE 415 – Rock Excavation 2004 Digital processing of images Computer technology has opened the door to image analysis of the evaluation of muckpile fragmentation. 2. Measurement. At this stage. The procedures take in the following stages: 1. shape. a shading filter which corrects illumination defects. If the technique is applied correctly. 4. number. Image segmentation. and subjected to an automatic digitization process. and they will be taken into account whereas the pixels below this level will be darker (back-ground) and are turned to black. an iterative process of dilation. size and orientation. mechanical breakdowns. This stage uses digital filters which permit an enhanced viewing of the fragments. In this step the distribution of two dimensional size is converted into volumetric sizes or tri-dimensional. should all be taken into account. low pass or Gaussian filters which eliminate noise. shifting the shovel or clean-up operations. The modem methods of image analysis quantify geometric aspects with images in two dimensions. The segmentation process is never perfect. 7. as the contours of some segments will bridge together and others will be confused with the background. This system. measures the diameter of an equivalent area circle and classifies them. The image is captured by a camera. the fragments are separated from the rest of the background to produce a binary (black and white) image. Image input. Binary image manipulation. Studies of loading equipment productivity This technique of fragmentation evaluation is based upon the assumption that the digging rates are an inverse function of the coarseness of the muckpile and a direct function of the swelling of the same. or grey level value from 0 (black) to 255 (white). Scaling. The lost time that is not directly related to the condition of the muckpile. and giving each one a certain brightness. To correct this. usually video. after identifying each object in the binary image as an independent fragment. This means the conversion of the optic image to a digital format of picture points pixels. Image enhancement. normally using a marker placed on the muckpile as a reference. This conversion demands the use of stereometric principles along with some empirical relationships. reduced swelling and poor toe condition will be immediately reflected in productivity. not only because of the high fragmentation cost. it would be impractical because of the high cost and time involved. burden. Initiation sequences and delay timing. In small operations this technique can be used with representative samples.MGE 415 – Rock Excavation 2004 made with the same machines and operators to eliminate the experience factor or erroneous estimates. Dessureault 1/14/2004 152 . Screening This is the only precise method of quantitative fragmentation evaluation. and Coarse size distribution. Bridging delays at the crusher The production of any crusher depends basically upon the coarseness of the material that enters. recording the signals on magnetic tape and processing the data on the computer. Its Height and Displacement The configuration of the muckpile is governed by: The geometric design parameters: height of bench. Boulder count and secondary breakage Any fragments of rock produced by blasting that cannot be handled by the mining equipment is referred to as boulders or oversize. but in large open pit mines. The sizes of these blocks depend upon each operation and during the same should be set aside for fragmentation. Insufficient swelling of the muckpile. Excessive displacement of the pile. blasthole inclination. with the procedures described in the following chapter. The interruptions at the crusher due to oversize should be controlled. Studies have been undertaken to monitor the continuous current motors of the rope shovels. The call for extra power can be caused by the following: Defective floor breakage with toe appearance. Geometry of Muckpile. Powder factors. as well as the wear of the steel plates. The relative volume of the oversize should be maintained at a minimum. and indirectly this can be an index of fragmentation taken from the yield and the energy consumed per treated ton. but because they affect the whole of the operation by giving low digging yield due to lost time dedicated to taking them aside and also delays at the crusher. spacing and stemming. the digging and haulage system used. Profile 3 reflects optimum conditions for rope shovels. dangerous. o Low clean up area. and the quantity of damage to the remaining mass. good productivity. low productivity for rope shovel. Profile I represents the ideal situation for front end loaders. aided by transversal photographs. it is possible to observe if there has been overbreak. high safety conditions. safe. the system in table below is one of the most used for its simplicity and pragmatism Dessureault 1/14/2004 153 . The evaluation of the damage caused by the blasts to the remaining rock can be carried out by any of the methods of geomechanic characterization of the rock mass. Profile 2 requires minimum cleaning labor and the productivity is high. high productivity.MGE 415 – Rock Excavation 2004 The optimum geometry depends upon. but there can be safety problems involved for the operators due to the fall of rock from great height. o Excessive clean up area. in each case. the yield will be low and time consuming in collecting the material and cleaning up around the shovel. but if the available equipment are rope shovels. This control procedure can be done with topographic equipment. Figure 32: Different geometries of the muckpile Condition of the Remaining Mass Once the muck pile has been loaded. o Low clean up area. but for what is wanted. the subdrilling and the between-row delay. The appearance of a high floor systematically could be due to a plane of weakness or to insufficient subdrilling and bottom charge. 154 Dessureault 1/14/2004 . This is corrected by increasing the explosive column or by placing a small pocket charge in the stemming.MGE 415 – Rock Excavation Table 9: Levels of damage to pit walls 2004 Analysis of the Bench Floor Once the blast has been evacuated. Within the pile. Low floor. the bottom charge and subdrilling should be diminished and the loading operation observed. Boulders Large sized blocks can appear in the muckpile in the following areas: On the top or crest. and In front of the pile. the bench floor can show the following: Toe in front of holes. They are eliminated by the same method as for toes: decreasing the burden and increasing the bottom charge. The boulders at floor level are due to a weakness plane in the lower part of the bench. On the floor (usually occurs with toe problems). The boulders on top could be caused by a hard level or poor fragmentation of the top part of the bench. When the floor level is lower than was designed. High floor. The toe between blastholes is due to overspacing and this should be decreased. Toe between holes. Dilution. and oversize is also a common method of characterizing the quality of a blast. In certain methods. bootlegs. but one of the most important is a premature escaping of the gases to the atmosphere after stemming ejection. characterizing the quality of the resulting excavation is a key method of evaluating the blast results. As to airblast. a poor blast yield because of defective priming or an alteration of the explosive such as local dampening of ANFO. Blast Evaluation in Underground Mining In underground mining. such as VCR. the following results will come about: Poor fragmentation and insufficient swelling of the muckpile Uncontrolled fly rock High vibration levels. giving a better size distribution of the muckpile and lower vibration intensity and frequency. By analyzing the vibrations registered with a seismograph and later modifying the indicated parameters. Overbreak is commonly calculated by comparing the planned excavation with the resulting excavation. especially for development.MGE 415 – Rock Excavation 2004 The boulders in the interior of the pile are due to incorrect drilling. which are potentially less dangerous. Damage to ventpipe or other utilities can also be a sign of inadequate stemming or improper blast preparation procedures. Hard toes. and Low frequency vibrations. For this reason. cratering at the top of the blasthole is a common indicator of a lack of stemming. the studies with high-speed cameras can help define the optimum stemming height and the ideal type of material so as to achieve a larger confinement of the explosive without negatively affecting the fragmentation in the zone where the inert material is placed. there are different mechanisms that cause it. The boulders in front of the pile could originate in an excessive breakage of the previous blast in the same zone. a factor of both the design and execution of a blast may also be used as a sign of blast evaluation. Vibrations and Airblast If the delay timing and the initiation sequence of a blasting are not adequate. and to a pernicious effect of the inner joints with respect to the opening of the grid. a better use of the explosive energy can be obtained. ADDITIONAL RESOURCES XVI Additional learning resources include: Lecture Module 5.5 – Blast Results (mandatory) Dessureault 1/14/2004 155 . Comparatively. Intenders are used more widely than drag bits primarily due to the shear loading of the rock by the drag bit. results from lab-based rock cutting experiments worldwide have reported variances in cutting conditions and variables such as pick shape and rock type. Furthermore. a brittle material weak in tension. generalities do emerge from these experiments and are regarded as fundamental features of rock pick cutting. The cutting edge of a drag bit is subjected to bending resulting in tensile stresses along the bit edge. Dessureault 1/14/2004 156 . this makes the possibility of catastrophic failure likely. 5. This will allow the two modes of the primary breakage process to be compared. a sharp drag bit applies the main force in a direction predominantly parallel to the rock surface. However. longwalls. Several will be presented with an emphasis on comparison between drag bits and indenters. Various theoretical models have been developed as presented in the SME Mining Handbook (Chapter 9.1) yet few have admittedly been proven empirically.1 Mechanical Excavation Basics The two main types of mechanical breakage for production purposes include indenters and drag bits (also known as picks). roadheadders. However the tools can be mounted in a fashion so that the bit contacts with the rock in a mainly compressive fashion. Figure 5-1: Longwall shearing machine whose cutting bits attack the rock in compression Various theoretical models have been developed that can compare drag bits with indenters. Both are considered to break the rock in the primary breaking process. and rippers. as seen in the Figure 5-1. Several groups of equipment will be considered in this section including continuous mining machines. The main difference between indenters and drag bits is that an indenter breaks rock by applying a force that is predominantly in a direction normal to the rock surface.MGE 415 – Rock Excavation 2004 Module 5: Mechanical Excavation This module on mechanical excavation considers the mechanical excavation basics and theoretical aspects of bits for excavation. Mechanical excavation can be considered the removal of rock from in-situ by mechanical means. The breaking mechanism for both is actually a tensile fracture. As most cutting tools are made of tungsten carbine. and pointed tool. it can be seen that this relationship is approximately linear. Simple chisel. = rake angle) Both FC and FN increase with d (depth of cut) for all pick shapes. vbottom.) Dessureault 1/14/2004 157 . However. Figure 5-2: Pick Shapes . simple chisel. such as. Round bottomed tool. chisel. round bottom. Pointed tool. and pointed.Pointed tool. ( = Back clearance angle. (not linear when using wide picks at depths much less than their width but machines such as these do not exist. From Figure 5-3. v-front. these many picks can be categorized into three main groups (as seen in Figure 5-2: Pick Shapes).MGE 415 – Rock Excavation 2004 Important variables in understanding these generalities include: FC = Mean force FN = Peak force ES = Specific Energy (relates to the cutting force to amount of rock produced Various picks have been tested. d) From Figure 5-4. depth of cut (FC vs. therefore: ES constant d Figure 5-4: Specific energy vs. d) Dessureault 1/14/2004 158 . it can be seen that Specific energy varies inversely with depth of cut for all pick shapes. depth of cut (ES vs.MGE 415 – Rock Excavation 2004 Figure 5-3: Mean force vs. beyond which further marginal improvement is at an increasing penalty to pick strength and its potential to survive. Figure 5-6: Mean force vs. each member of a monotone decreasing sequence is less than or equal to the preceding member. rake angle (FC vs. Each member of a monotone increasing sequence is greater than or equal to the preceding member. Dessureault 1/14/2004 9 159 . back clearance angle (FC vs.MGE 415 – Rock Excavation 2004 Cutting and normal forces decrease monotonically9 with increasing rake angle as seen in Figure 5-5. ) Figure 5-6 shows that increasing back clearance angle reduces pick forces up to about 5 . Figure 5-5: Mean force vs. Most of the benefit to pick forces has been achieved at a rake angle of 20 . beyond which forces are independent of this angle. ) monotonic defined for mathematics: Designating sequences. the successive members of which either consistently increase or decrease but do not oscillate in relative value. FC and FN increase linearly with pick width as seen in Figure 5-7. A further issue related to cutting forces is the degree of rock saturation as increased saturation results in a decreasing mechanical strength of the rock (the reason for this is not clearly understood). Dessureault 1/14/2004 160 . pick width (FC vs. and it has yet to be determined what characteristic of rock strength is the predominant influence. Hence speed has no direct effect on pick forces Cutting efficiency improves with cutting depth For a chisel shaped pick.MGE 415 – Rock Excavation 2004 Higher productivity will be achieved by cutting slowly and increasing depth than when cutting at a higher speed yet more shallow cut since: Cutting speeds of up to 5 m/s and beyond have no discernible affect on pick forces or specific energy. Figure 5-8 compares the mean cutting force with both the tensile and compressive strength of various rocks. w) Pick forces increase with rock strength. Figure 5-7: Mean force vs. whereas chisel bit cut more material for a given depth of penetration. it cuts with the lowest specific energy and therefore is the most efficient shape (pointed pick least efficient). Dessureault 1/14/2004 161 . t) In terms of pick shape. pointed picks operate more efficiently than the chisel bit. (Hence. The chisel pick requires the greatest forces. considering that the chisel pick cuts a considerably larger volume of rock than the other two shapes.) Figure 5-9: Chisel vs. c and FC vs.MGE 415 – Rock Excavation 2004 Figure 5-8: Compressive and tensile strengths upon varying mean cutting force (FC vs. the pointed pick requires the least cutting and normal force. Due to the increased penetrating capability of the pointed pick. picks are better. pointed picks can but deeper for a given level of force. for a given available normal force. However. when operating at the same rake and clearance angles and depth of cut. Operationally. pick bits. MGE 415 – Rock Excavation 2004 5. pp. The immediate entries on both sides of the panel are called the head entry and tail entry. such as conveyor design. The tail entry is used as the air exhaust. fewer residual pillars remain hence coal recovery is higher and surface subsidence is relatively uniform and complete.1 to 4 meters. Although most are unable to survive in the hard-rock environment.2 Longwall These lecture notes review the design and operational aspect of the longwall mining method. Figure 5-10 shows a typical plan view of a longwall layout and the nomenclature of the key design variables. E. Longwalls are used to mine flat horizontal coal seams of thicknesses ranging from 1. typically excavated by a continuous mining machine. especially as bit wear proceeds.. (mandatory) Readings – Speight. Panel sizes and length are generally determined by: Experience size and shape of the coal seam geologic conditions Dessureault 1/14/2004 162 . 1997. personnel.1. and supplies. The depth of overburden can range from anywhere between 60 to 820 meters. ventilation and roof support. provides continuous production and can provide a safer work environment. (Mandatory. Longwall panels are flanked by panel entries on both sides of the main entries. Longwall mining has one of the simplest layouts. 5. ADDITIONAL RESOURCES XVII The above information is additionally supplemented by: Readings – Chapter 9. Two main problems remain to be solved: Drag bit materials that have far improved wear resistance and strength Machines to provide sufficient thrust force to keep drag bits at an effective depth of cut.17-37 (mandatory). but this is for modules 1 and 2) Lecture module 2. Important issues related to longwall. The panel layout is conducive to good ventilation. Therefore this method is considered to be better than room and pillar methods used to mine similar coal deposits. are left to materials handling and rock mechanics courses. Since this system uses full caving.1. The head entry is used for air intake and the transportation of coal on conveyors. 1. they require substantially less energy than other mechanical cutters. Particular focus will be on the cutting mechanism and design. No. H.1 Synopsis Drag bits are an efficient means of breaking rock. “Obersvations on drag tool excavation and the consequent performance of roadheaders in strong rock” AusIMM Proceedings.1 – Mining Engineering Handbook. The economic factors pushing for longer and wider longwalls include: Reduction of the development cost as fewer panel entries would be required Increases in recovery and recovery rate (few longwall assemblies.MGE 415 – Rock Excavation 2004 location of surface structures (buildings can subside horizontally when in the middle of a panel but would subside differentially if it straddles a pillar). capacities of the transportation system ventilation power capacity (of equipment) Panel widths vary between 120 to 293 meters (centre to centre) and have lengths between 1220 to 1830 meters. downtime between panels) Other issues that begin to surface in longwalls include: Downtime (maintenance & repair) in large panels Capital cost of equipment in a single panel Figure 5-10: Typical Longwall retreat method Dessureault 1/14/2004 163 . Coal at the face is cut by a shearer or plough and loaded onto an armoured flexible conveyor (AFC) and conveyed to the head entry T.junction.1 Mining System Description Figure 5-11 shows a diagram of a coal operation.2. The area between the rear edge of the supports and faceline is called the face area while behind the supports is called the gob. The AFC and powered supports are advanced hydraulically after each cutting cycle of the shear and the roof behind the supports are allowed to cave. an area of roof is opened. Now that the basics of longwall mining have been established. at the tail entry where ventilation circuits are set up. Dessureault 1/14/2004 164 . However. Increasing panel width increases roof exposure time (when the shearer or plow moves across the face. cribs are usually required to strengthen the support. Panel entries are where the regular conveyors are maintained by roof bolting or other methods. a focus on the actual rock excavation aspects will be explored in more detail.MGE 415 – Rock Excavation 2004 5. Figure 5-11: Longwall shearing system Technical factors of longwall mining technology include: The longwall width is limited to the power and structural strength of the face conveyor. The coal is loaded onto a stage loader which empties onto the entry belt conveyor. Powered sectional supports are used to support the roof along the whole face. This small area may cave prior to the supports moving forward to secure the back). Single-ended ranging-drum (SERD) for thicknesses of 1. Major parameters in shear design are: Type of shearer. Ploughs are simply a large blade that is thrust into the rock and dragged across the face resulting in the removal of a slice of coal.3 Dimensions of Shearer Equipment suppliers provide various models with varying dimensions and power available. only shearers are discussed.mining height D . Dimensions. Is by far the most employed shearer.26 to 1. articulated so that coal seams of varying height can be mined.2 Selection of Cutting Machine The two main types of longwall machine are the shear loader and the plough.52 meters. Key variables in the selection of the type of cutting machine are: Mining height Seam structure Roof bonding strength of coal Cutting/ploughing resistance of coal. The key variables in selecting the shearer can be seen in Figure 5-12: Figure 5-12: Shearer dimensions Where: Hc Hb And where: Hc . Can mine thicknesses of 1.47 – 3. Haulage speed.MGE 415 – Rock Excavation 2004 5. Single-ended fixed drum (SEFD) for thicknesses of 1. There are several types of shearers: Double-ended ranging-drum (DERD): two shearing drums mounted on cutter. Since shearers are the most common machine.96 meters. Power capacity. Shearers are used more widely in the US.2.37 meters 5.2.diameter of the drum B 2 La sin D 2 Dessureault 1/14/2004 165 . body depth – angle of the ranging arm The maximum cutting height should not exceed 2 D. A nomograph for determining shearer dimensions and mining height is shown in Figure 5-13.MGE 415 – Rock Excavation 2004 Hb – height of ranging arm La – length of ranging arm B .4 mm. Note that the conversion factor is 1 inch = 25. Figure 5-13: Nomograph for shearer dimensioning Dessureault 1/14/2004 166 . R. Howard L. Maintenance. Stefanko.. Maintenance. 570p. Environment. Surface Mining Manual. Littleton CO. 2nd ed. Occupational Health and Safety. SME Mining Engineering Handbook.M. 1992 Hartman. 2nd ed. Transportation.com. Ed. (508mm) and =30˚ The dotted line on the nomograph can be followed from point a through to f in Figure 5-13. Introductory Mining Engineering. Robert. This results in a value of D being 30 inches (762 mm). 1992 Hartman. 2nd ed. in this example. Ed. The Raise Boring Handbook 2nd ed. Pergamon Press: Toronto. D would therefore be equal to 60 inches.:Society of Mining Engineers.1 – Longwall Mining. La =70in (1778mm). Surface Mining Manual. ADDITIONAL RESOURCES XVIII The above information is additionally supplemented by: Lecture for module 2. 1983 McKercher. Edition 2002 10 These notes were assembled directly from the following references: Hartman. (2. 1983 McKercher. Potash Technology: Mining. SME Mining Engineering Handbook. Mutmansky. Ed. Transportation. B = 20 in.com. and Jan M. Hb =50 in (1270mm). continuous surface mining equipment. Howard L. and Jan M. Atlas Copco.:Society of Mining Engineers. In practice. 2nd ed. 1983 Mining-Technology. New York: John Wiley & Sons. Mutmansky. Potash Technology: Mining. These notes were assembled directly from the following references: Hartman. Howard L.29m). The Raise Boring Handbook 2nd ed. 570p. given the mining height and major dimension of a shearer: where: Hc =90 in. Howard L. New York: John Wiley & Sons. in order to reduce the loading resistance of the drum and to increase loading efficiency. Atlas Copco. Introductory Mining Engineering. 2002. Littleton CO.3 will cover underground continuous mechanical mining technology and.2 – Longwall (mandatory) Readings – Chapter 20. Stefanko. Processing. R. 1983 Mining-Technology.MGE 415 – Rock Excavation 2004 Example 1: Shearer dimensioning Use the nomograph in Figure 5-13 to select the drum diameter D. search: continuous mining Wirtgen America Inc. Robert. Occupational Health and Safety. search: continuous mining Wirtgen America Inc. Pergamon Press: Toronto. 2002. Coal Mining Technology Theory and Practice. from SME handbook (optional) 5. Environment. Processing. typically 75-80% of the mining height. Edition 2002 Dessureault 1/14/2004 167 .M.3 Continuous Mining Machines10 Module 5.. the drum diameter is selected larger than the minimum value. Ed. Coal Mining Technology Theory and Practice. 1 Boring Machines Figure 5-14 shows a boring machine. produce a much coarser product than that of a ripper miner.1. Only those machines still in use today are reviewed. sometimes using very different processes.MGE 415 – Rock Excavation 2004 5. Maintenance issues therefore become more critical. These machines have rotating arms equipped with bits that bore out the coal. Underground blasting however. and loading into one operation by a single machine. which when working together. Trona. it cannot be used to bolt simultaneously as advancing. The continuous nature of this mining method eliminates the problems involved in the cycling of equipment. Continuous mining equipment has evolved significantly over the last 100 years. Since the machine uses its tracks to provide the thrust. The machine can have drills mounted on a slide carriages and have met with success at bolting and mining simultaneously.1 Underground continuous mining machines Underground continuous mining machines typically operate in room and pillar mines in coal. ripping out and loading the coal in one step. or potash (essentially soft material in horizontal orebodies). However. The machines are an alternative to drilling and blasting operations. these advantages are offset in some part by concentrating the reliance of production on a few headings. drilling. Dessureault 1/14/2004 168 . While most of this equipment attack the solid face. Other advantages of continuous mining is combining the operations of cutting. salt. The boring machine has a cutting chain on the bottom that creates a flat working bottom and an upper trimming chain. Figure 5-14: Four-Rotor Boring machine 5. minimizing gas and dust problems. so are manpower requirements.3. produce a roof configuration that is arched or ovaloidal opening.3. As the number of equipment is reduced. The arms revolve relatively slowly. is undesirable in coal operations due to the risk of secondary explosions caused by suspended coal dust or methane gas. blasting. however.2 Milling head miner Milling head miners are the most popular continuous miner in operation in US coal mines. usually the width of the head is less so that two passes are required to mine the width of a room or entry. Further discussions on potash mining will be discussed later. The disks rotate in a vertical plane.3. Dessureault 1/14/2004 169 . the arms to which the disks are attached oscillate or are provided with a side splay or horizontal movement.1.MGE 415 – Rock Excavation 2004 The main disadvantage of this type of machine is its size. Figure 5-15: Milling head continuous mining machine 5. Ventilation is also an issue when dealing with its size.6 m (15 ft) wide. the machine must move considerably slowly. The machine is a ripper with disks or wheels located on arms. not coal mines. therefore these machines are typically used in potash mines. It typically occupies the entire stope. While the machine can be equipped with a full-face head about 4. To mine the coal between the disks. The tips of the arms can be extended or contracted hydraulically over certain limits to provide some machine clearance during tramming. The design of the machine permits a concentration of power at a single cutting bit at a time. Loading arms located under the head load the coal. therefore. In this type. with the majority outside the US. 5. most miners use the same bit attack with a rigid head and the machine is usually referred to as the Hardhead miner. While the basic machine had an oscillating feature when it was introduced. enabling it to cut harder rock than drum-type continuous miners can excavate. Roadheaders were introduced in the post World War II era. it is. negating its widespread use for concurrent bolting. quite popular today.3.3 Boom-Type Miner A boom-type miner or road-header consists of a rotary cutter-head mounted on an articulated boom. although the optimum shape with regard to ground control should be heeded. a crawler assembly. some equipment allows both bolting and mining simultaneously. its primary use is for driving mine entries and tunnels in civil engineering projects. The milling machine sumps (pushes) into the top and shears downward. and at present there are over 2800 employed worldwide. 170 Dessureault 1/14/2004 . The usual mode of operation is to push the machine into the face on its crawler while the head is in an elevated position and the disks are rotating. drills for roof-bolting would have to be attached to the chassis in a flexible manner.1. In addition. accomplishing the same purpose without oscillating and with subsequently fewer maintenance problems. and a gathering arm system. Since this machine cuts by advancing the crawler to the face. While it is used for mining coal and other minerals. The usual procedure is to mine an opening cut the width of the machine [usually 3 m (10 ft)] then mine the other side of the face. its head can be tapered to provide an arched opening where desired. This machine is cheaper as well as requiring less maintenance. bits are spirally wound on a fixed shaft so that each bit cuts an infinite number of vertical planes during a single revolution. However.MGE 415 – Rock Excavation 2004 Figure 5-16: Miller-head continuous miner. and rectangular. Unlike conventional drilling and blasting. Thus lighter support may be used or less concrete. The wide flexibility in cutting at various angles also increases the utility of the roadheader when compared to the continuous miner. about 140 road-headers are in use in North America with approximately 75% of the machines working in mines. arched. aiding cost. horseshoe. Figure 5-17: Mining of Steep seams with roadheadder and miller head (drum-type) miners Dessureault 1/14/2004 171 . Roadheaders can excavate any size and shape of cross section: circular. as seen in Figure 5-17. The machine is being used to enlarge mine entries in a bituminous mine in central Pennsylvania by removing bottom rock.MGE 415 – Rock Excavation 2004 However. the smooth cutting action of the cutterhead does not weaken the rock due to overbreak. MGE 415 – Rock Excavation 2004 Figure 5-18: Mining in variable height seams Figure 5-19: Double head Roadheadder Dessureault 1/14/2004 172 . including rock-tool interaction models. TBM models work better. and are not good.MGE 415 – Rock Excavation 2004 Figure 5-20: MD1100 with traverse boom 5.30(ICR) when trimming Utilization is affected by: Support installation Surveying Pick replacement. the OCR is 0. both in terms of the rock and rock mass.60 x ICR. Maintenance/repairs Haulage delays Shift changes In the calculation of utilization.3.1. as before: SE=specific energy HP=headpower Dessureault 1/14/2004 173 . TBM performance remains to be calculated empirically.45–0. while 0.4 Performance Analysis Predicting the cutting rate of these machines requires information on: rock properties. Advance rate is calculated by: Advance rate OCR utilizatio n Penetration rate face area Be warned: theoretical cutting models for roadheaders vary in applicability for machine types. machine properties. A few key measures include: ICR: instantaneous cutting rate OCR: operational cutting rate (includes utilization) Typically. boom repositioning and final profiling are not included. however. can be used in conjunction Figure 5-21: Rock Mass Cuttability Index vs. Dessureault 1/14/2004 174 . in kg/cm = with Figure 5-21. RQD and UCS. The rock cuttability index.MGE 415 – Rock Excavation 2004 Where: SE = HP / ICR Most predictions use RMR. 2 c RQD 100 2 3 . cutting rate Note that Figure 5-21 can also be used for predicting road-header productivity as a function of rock mass properties. The thrust reaction force is reacted through gripper pads that are pressed hydraulically against the tunnel walls. Because these cutters break the rock by indentation. whereupon it is tipped onto a belt conveyor that runs through the center of the machine. rock class. The rock broken from the face by the cutters falls to the floor where it is scooped into buckets mounted around the gage of the cutterhead. these machines are characterized by very high thrust requirements.7 ft (1. In almost all cases breakage is effectuated by roller cutters mounted on the cutting head. This thrust is provided by hydraulic rams that press the cutterhead into the rock face.5 Tunnel Boring Machine This type of machine can be used to drive circular tunnels from 5. Figure 5-23: Tunnel boring machine Dessureault 1/14/2004 175 . as seen in Figure 5-23.75 m) to more than 36 ft (11 m) in diameter in rock types that range from weak.1.MGE 415 – Rock Excavation 2004 Figure 5-22: Roadheader performance vs. This debris is lifted in the buckets to the tunnel crown.3. 5. loosely consolidated to very strong and abrasive. MGE 415 – Rock Excavation 2004 Figure 5-24: Components of a TBM Figure 5-25: Disc cutters on TBM The most common type of cutting tool employed on these machines is the disk cutter. The most common cross section for a Dessureault 1/14/2004 176 . In some cases the cutting edge of this tool is a hardened steel surface and in other cases it is a row of cemented tungsten carbide buttons that are press-fitted into the disk rim. groove deepening is the cutting procedure most commonly employed with tunneling and with other boring machines. Third. Second. the disk-button cutter next. Consequently. whereas with drag bits an optimum s/d value of 2 to 3 is typical (see Figure 5-28). and the pineapple cutter. This is fortunate because in practice. it is necessary to increase both the machine thrust force and torque.MGE 415 – Rock Excavation 2004 hardened steel disk cutter is seen in Figure 5-26. First. less abrasive rocks. and therefore the capability of cutting strong abrasive formations. is the reverse of this efficiency ranking. and the pineapple cutter as least efficient. The value of this optimum spacing depends on the depth of cut taken and on the rock type. which is a frustrum with cemented tungsten carbide buttons press-fitted onto the surface. The advantage of this design is that the tool area presented to the face is maintained resulting in relatively constant penetration rate. Several other comments can be made about the cutting behavior of disk cutters. places the steel disk cutter as the most efficient. However. the efficiency of the rock breakage process does not decrease when disk cutters are used in a groove deepening (when multiple passes are made by a tool taking a series of shallow cuts in a kerf before producing major rock chips) mode. Figure 5-26: cross section of disc cutter Figure 5-27: Kerf and Pinapple cutters A ranking of cutting efficiency of these tool types. steel disks tend to be used for cutting weaker. an optimum spacing exists between an array of disk cutters working a rock face. which is simply multiple disks mounted on the same hub. the efficiency of the rock Dessureault 1/14/2004 177 . the wear resistance. in terms of specific energy. similar to the findings for drag bits. As the cutters become blunted. and pineapple cutters are used for machining the most abrasive and toughest formations. However. in contrast with drag bits. with disk cutters this value is more typically in the range 5 to 10. Other cutter types include the kerf cutter. or sequentially. TBMs are widely used in civil applications. A key environmental aspect is the lack of ground vibrations when compared to blasting. the advance rate would be about 33 feet per shift. The primary advantage is that this machine is quasi-continuous. and cost of the equipment has made TBM not fully accepted in mining. however. Some attempts have been made to develop hardrock continuous mining machines that are more flexible. compared with the drill-blast-muck cyclical method. The lack of flexibility.MGE 415 – Rock Excavation 2004 breakage process is independent of whether the grooves are cut simultaneously. the bit wear costs remain prohibitive. Dessureault 1/14/2004 178 . size. Roughly. Considering these disadvantages. Figure 5-28: spacing versus depth relationship Skidding has a major influence on cutter longevity (more will reduce longevity). A prototype oscillating disc cutter appears to show promise but has yet to be fully developed into a working production unit. the speed of advance for a drill-blast-muck crew would be 10 feet per shift (ideal). Therefore tunnel rates are somewhat more constant and higher where TBMs are used. A further considerable advantage is the lack of secondary damage caused by the machine when compared to the damage cause by blasting. This dragging or ploughing action is particularly acute for rollers as the center of the cutting head. a rule of thumb is that a tunnel of at least 2 km in length to justify the cost of a machine setup. The disadvantages are: High capital costs Tunnel cross section is circular Large turning radius (100 meters) Cumbersome machine and system installation. with independent disks. with multiple disks on a single hub. whereas on a TBM. Tunneling through rocks up to 30. and volcanic rocks have TBM specific excavation rates of 8 to 360 X 10-6 ft3 of rock extracted per lb of force per cutter per revolution (0. (5) specific excavation rate.00005 to 0. Note that the metamorphic. Several of these parameters deserve to be discussed further here. First.00005 to 0.0023 m3 of rock extracted per kN per cutter per revolution) and the rates in sedimentary rocks are 8 to 720 X 10-6 ft3 of rock per lb of force per cutter per revolution (0. Figure 5-29: Empirically derived excavation rates of TBMs by diameter for igneous and metamorphic rock types Dessureault 1/14/2004 179 .000 psi (207 MPa) has become fairly common.MGE 415 – Rock Excavation 2004 The TBM is considered to be the primary equipment in the field of rapid excavation. hardness of rock is an area in which some progress is being made.000 psi (350 MPa) in compressive strength TBM specific excavation rates have also gone up over the period of rapid excavation usage.0046 m3 of rock per kN per cutter per revolution). (3) diversity of application. granitic. however. some tunnels have been driven through rock approaching 50. (2) time percentage of machine availability. (4) rate of advance. Progress in rapid excavation can be measured by several performance parameters: (1) hardness of the rock. and (6) cost of advance. Case studies dating back to 1980 have resulted in empirical tables of rock types with corresponding excavation rate. Dessureault 1/14/2004 180 . Maintenance. Another important limitation of the TBM in mining projects is the economics of conventional versus rapid excavation development. This is evident in the application of a TBM to development in an underground mine (Stillwater). cutter replacement.MGE 415 – Rock Excavation 2004 Figure 5-30: Empirically deriven excavation rates of TBMs by diameter for sedimentary rock types Although improvements have been made in these parameters. see Figure 5-31). The original turning radius of the machine (350 ft or 106 m) was simply too great to be able to effectively maneuver the machine underground. One of the major deterrents to applying rapid excavation in underground mining is the rather massive dimensions of the typical TBM with its trailing gear. and other delays all contribute to the problem. This eliminates its use for many mine development projects. ground control. machine availability is still a problem area. There are many good reasons that the TBM is not engaged in the excavation process a greater percentage of the time. the TBM normally cannot provide a cost that is lower than the costs of conventional development practices.000 ft (6 km). Studies have shown that TBM availability is 35 to 50%. In this project the mining company had to work with a TBM manufacturer to reduce the turning radius of its TBM to about 200 ft (60 m) in order to effectively use the device in mine development. For tunnels of less than about 22. backup equipment. 3.MGE 415 – Rock Excavation 2004 Figure 5-31Cost comparison between tunnel boring and conventional drill and blast. particularly in the area of raise boring.1. In the advance of vertical openings. rapid excavation has greatly improved mine development and civil works. but they drive vertical or steeply inclined holes rather than tunnels. Although TBMs have been used for development at the Stillwater Complex in Montana and at the Ray Mine in Arizona. raise borers for development of stoping operations are routinely used and have replaced conventional development in all but unusual circumstances. 5. the rapid excavation revolution has still not established this process as commonplace in providing horizontal development openings in mines.6 Raise / Blindhole / Shaft Boring Machines These machines use the same button roller cutter technology as TBMs. Dessureault 1/14/2004 181 . In many mines. When this hole is completed. are reduced by using these machines. such as: safety--conventional drill-and-blast raising is notoriously dangerous. however. and replaced by a reamer head having a diameter with the same dimension as the desired excavation. The process of raise boring is shown in Figure 5-32. this raising system offers other significant advantages. and they have become very popular. particularly in hard-rock underground mines. First. the boring machine is set up on the upper level. The rock debris falls by gravity into the lower excavation where it is removed. Figure 5-33: Raiseboring in operation top & bottom Raise Borers are machines used to produce a circular excavation either between two existing levels in an underground mine or between the surface and an existing level. at the lower level. Frequently the direct costs of driving a raise. These machines are very effective in driving raises. and a small-diameter (of the order of 9 in.MGE 415 – Rock Excavation 2004 Figure 5-32: Steps in raisebore mining. Some type of roller cutters are mounted on the reamer head. down to the lower level. In addition. the drillbit is removed. improved excavation rates improved productivity. or 230 mm) pilot hole is drilled. usually with a tricone bit. in terms of dollars per foot (dollars per meter). This head then is rotated and pulled back up towards the machine. Dessureault 1/14/2004 182 . Storm water storage and drainage 6. Redirection and retrieval of water in hydroelectric projects 3. Reduced friction for ventilation raises. generally several feet (meters) in diameter. Penstocks and surcharge chambers in hydroelectric projects 2. Road and rail tunnel ventilation 5. pressurized gas. Dessureault 1/14/2004 183 . Material Transfers (ore or rock passes) 2. Equipment access (pipes. is excavated upward from this level. Storage of petroleum. and nuclear waste 4. The machine is set up on a level in an underground mine. Ore production 5. Slots for longhole Civil 1. and a steeply inclined hole. specific applications are summarized in Table 5-1. Water inlets and outlets for fish farms. Personnel access 4. Ventilation 3. hoses. Figure 5-34: Raise boring diagram – Up and downward boring.MGE 415 – Rock Excavation 2004 circular shape combined with the lack of blasting damage results in an excavation of greater strength and integrity than a hand-driven raise. cables) 7. Blindhole Borers are machines that will produce a circular excavation without the need for a pilot hole. Table 5-1: Applications for boring Mining 1. Applications of raise boring are in both civil and mining environments. Boring Site Preparation. Site planning considerations must include: Derrick mounting systems Selection of bailing fluid Bailing fluid and cuttings discharge Storage and positioning of drill string components Overhead clearances Floor space and equipment positioning Compressed Air Water Electric power Lighting Dessureault 1/14/2004 184 . Raise boring site preparation begins with a comprehensive plan of the site layout.MGE 415 – Rock Excavation 2004 Figure 5-35: Blind boring operation Figure 5-36: Blind boring bit head. Correct site preparation eliminates major delays and adds noticeable efficiencies to the raise boring operation. MGE 415 – Rock Excavation 2004 Communications Ventilation.7 Underground mining case studies The first case study of an underground continuous mining system is exploration of the Boulby potash mine in the UK. A special impermeable lining was then constructed Dessureault 1/14/2004 185 . Figure 5-37: Directional boring in urban environment.1.3. Combined with directional drilling (undertaken when drilling the pilot hole). Horizontal boring Use of horizontal raise boring in civil engineering projects is becoming increasingly popular as compared to alternatives since: no trench is to be dug no vibration or environmental side-effects of drilling and blasting less set-up time and cheaper than TBMs. Figure 5-38: horizontal borehead 5. Note that the shaft had to be excavated by freezing the surrounding rock mass so that a minimum of water inflow would occur. a directional tunnel can be built. Jeffrey and Joy continuous miners discharge to Joy electric shuttle cars.MGE 415 – Rock Excavation 2004 long the shaft in the section under the aquifer. which run to feeder breakers on the main conveyors to the hoisting shaft. This can be seen in Figure 5-39. Dessureault 1/14/2004 186 . CPL converted its continuous miners to remotecontrolled operation (so they can be safely used in shaly sections without prior blasting). Figure 5-39: Cross section of the Boubly potash mine strata To cope with greater stresses in the ore below 1. The mine is currently establishing a mine-wide digital microseismic monitoring system and has devised a two-road with stubs (rather than four-road) stress-relieving room-and-pillar technique which achieves more mining per shift.200m and to raise output above 3Mt/y by 2001. separated by pillars supporting the overlying strata. In addition to automating its Marietta miners to measure ore grades and mine selectively. These connect to the main haulage conveyors. The central control system. Five automated Marietta two. each capable of extracting 650t/h of ore. where it is hoisted to surface. form the production fleet. surface and underground. Dessureault 1/14/2004 187 . oversees all aspects of the mine and mill operation. The run-of-mine ore is loaded on to extensible conveyors attached to the continuous miners. which move the ore to skip-loading pockets at the shafts. is controlled. and has maximized ore extraction and recovery since it was commissioned. the company installed a central control room at Rocanville from which the entire mine. supplied by Allen-Bradley. Case Study: Rocanville Saskatchewan Rocanville uses a long room-and-pillar method of mining.and four-rotor continuous miners.MGE 415 – Rock Excavation 2004 Figure 5-40: remote operation allows distance mining and the operator to view from different angles. Ore is mined from rooms in three passes. which shows various boring type cuts. These can be seen in Dessureault 1/14/2004 188 . Figure 5-43.MGE 415 – Rock Excavation 2004 Figure 5-41: Marrieta miner in Rocanville Figure 5-42: Marrieta digging a trench. being maintained Several types of cuts are made in mining potash. In particular.250 bm³/h when engaged in lignite mining.3.MGE 415 – Rock Excavation 2004 Figure 5-43: Cross sectional view of four-rotor boring in potash 5. Eliminating dilution increases the efficiency of transportation and further processing. crushing and loading. minerals etc. Surface Miners have reached a rated capacity of 1. Selective mining of thin or sloping layers results in a clear separation of material (overburden.).3.1 Surface Miners Surface Miner combines the individual processes normally associated with traditional technology. dust. This means that layers of up to approximately 1. gasses. such as drilling. noise and flying rocks. transportation.2. Dessureault 1/14/2004 189 .2 Continuous Surface Mining Technology 5. extraction. the fact that blasting is no longer required leads to a considerable reduction in environmental pollution from. blasting or tearing.2 m thickness can be mined in one step. irrespective of discontinuities and without blasting. This includes: lignite coal anthracite bauxite phosphorite marl limestone hard sediment stone (sandstone. certain conglomerates. breccien clays.MGE 415 – Rock Excavation 2004 Figure 5-44: Various aspects of a Surface miner Surface Miners are capable of continuously mining stone with 50 MPa or 100 MPa. loam Dessureault 1/14/2004 190 . MGE 415 – Rock Excavation 2004 Figure 5-45: Surface miner ideal for thin strata Figure 5-46: Cutting capacities for various Wirtgenamerica products Figure 5-47: Surface miner ideal for defined thin vertical or flat seams Surface Miners are suitable for: Mining for the extraction of minerals and overburden Road construction for preparing roads. motorways and waterways Earthworks for preparing construction sites etc. Dessureault 1/14/2004 191 . sand. Dessureault 1/14/2004 192 . loam.3. and lignites). clays. One of the most successful applications of BWEs has been in German and Australian lignite mines. The teeth on the individual buckets are the primary ground engaging tools that break out the material from the ground.MGE 415 – Rock Excavation 2004 Figure 5-48: components of Surface miner 5. black coal. The machine digs out the material using a large wheel with buckets that revolve as the wheel turns. some limestones. and tar sands.2 Bucket wheel excavators Originally designed for relatively easy digging materials (gravel. bucket wheel excavators (BWEs) can now dig in relatively hard material. marl. BWEs are most often attached to a conveyor network where waste material is sent to a spreader or to an ore stockpile. These have included compact sediments such as shales.2. with modern systems. Dessureault 1/14/2004 193 . large boulders or blocky material cannot be handled Sticky material build up in buckets and can gum up the conveyor (although.MGE 415 – Rock Excavation 2004 Figure 5-49: Bucketwheel nomenclature Figure 5-50: Bucket cutting and discharge BWE considerations: Hard consolidated materials. sticky material can be handled). Therefore the mine must have strict discipline in terms of the mine plan. power consumption Reduced slewing (swing) speed.000 m3/hr. reduction in ground bearing pressure BWE can mine both thin overburden and deeper overburden Conditions possible where single bucket (dragline) is impossible Theoretical output of machine is: Qth 60 Fs Swell Factor where Q is theoretical output in yd3/hr (or m3/hr) bank. F is capacity of single bucket S is number of bucket discharges per minute Swell factor is that of the material being excavated. large variations in overburden thickness Specific cutting force The digging resistance of the material being excavated is limited by: The bucket wheel drive power The mechanical strength of the bucket Machine mass (service weight) Dessureault 1/14/2004 194 .MGE 415 – Rock Excavation 2004 Abrasive material produces excessive wear on the teeth (some modern tooth design can significantly increase tooth life). machine mass. maintenance costs. Largest disadvantage is their size. side-slewage conveyor belt system and spreaders Advantages of Continuous Excavators Have lower impact loading than comparable single-bucket machines Tends to reduce dynamic stresses. such as: Sufficient reserves to justify the capital cost Relatively horizontal stratification over a wide area of the deposit Uniform geologic conditions (i. severe undulations. The theoretical output of BWEs range between 200 to 20. Some advantages in direct reclamation (environmental benefit). which limits flexibility. reduced digging impacts.e. absence of major faulting. The digging face should be stable Actual output is usually only 45-110% of theoretical Very limited flexibility (can be flexible in certain geological and equipment situations) High capital costs but may be the most economical method of mining weak flat tabular deposits. Relatively little manpower Must closely match downstream equipment (as in all mining operations) Low total costs Long life There must be a matched system linking BWE. ..........26 feet Wheel Speed........* Wheel Diameter.8 feet (width) 29 feet (length) Dessureault 1/14/2004 195 ..*** Number of Buckets..............125 R....M............41 feet Width........P............ Texas General The Big Brown Mine......49 feet (receiving) 82 (discharge) Crawler Size.. About 12.......... is composed of two active mining areas that supply fuel to the Big Brown Steam Electric Station..S.......885 F......... Big Brown Mine fuel is used to generate approximately 7% of TU Electric's total electricity generation...33 feet Length....... Case Study: Big Brown Mine. TXU Mining receives about 17% of its total lignite production from the Big Brown Mine..... located near Fairfield in Freestone County............MGE 415 – Rock Excavation 2004 Despite empirical research..60 inches Belt Speed................................ and 1 discharge) Belt Width...... sandstones........2 (1 receiving.. no direct relationship mathematical model for prediciting specific cutting force between the key variables has been established................ The total amount of overburden have been excavated since mining operations began is approximately 1..08 yd³ Belts.................... BWE Specifications: Height.... can be excavated economically with BWEs.. and frozen Tar Sands.000 acres have been mined at since mining began in 1971... phosphates...........5... modern BWEs that have high wheel rotation speeds with slower slewing speeds.........144 feet Weight...4 billion cubic yards........ and that the ground’s strength ranges between 15-18 MPa............374 tons Bearing Pressure ..14...... These key variables include: Digging resistance Intact rock strength jointing Bedding Tooth shape and sharpness Angle of attack Wedge angle of the bucket cutting lip Hard Ground operation is possible with large.12 Bucket Size.......................M...** Boom Length......... Material that cannot be dug out with a cable shovel such as hard clays........ The Mine areas have a combined total pit length of approximately 7 miles....5 P..........I..P...... Provided that rock does not contain large hard boulders........1.................. Approximately 72 million cubic yards of overburden are moved per year.. .** Personnel Required. In this operation.. ed. Engineering in Rock Masses. as seen in Figure 2 through Figure 4 Radial. joints and fractures.. Students are welcome to view these sources in my office..MGE 415 – Rock Excavation 2004 Design Digging Height.. the ripper shank is pulled through the soil or rock to loosen or fragment the material. • Brittleness. or handled by the dozer blade. which can then be loaded by a scraper or front-end loader. from Sean Dessuresult’s Surface Mining Course Notes for UBC’s Mining and Mineral Processing Department. SME Handbook.. F. Caterpillar Ripping Handbook.P...... Front-end loaders then load the material into haul trucks.49feet Design Digging Width. from Bob Cummings 1997 Class notes..60 feet BWE will dig 3 feet below level of tracks Ground Speed.. This lecture and its notes were complied using sources from: Bell.. Rippers are used in quarries to excavate sand and gravel deposits.. (Optional) 5.. ISBN: 0 7506 1063 8 Caterpillar Performance Handbook.M.... 12th Ed.. 1 operator and 1 oiler ADDITIONAL RESOURCES XIX The above information is additionally supplemented by: Assignment 2 Lecture for module 2. Figure 5-51 shows a ripper in action (back of dozer). 1/14/2004 Dessureault 196 . Rippability is determined by: • Compressive strength • Bedding planes.. which is mounted to the rear of the tractor. • Softness (from weathering). G. 32. There • • • 11 are three ripper types.. 2001 Class notes.... Ripping is an inexpensive method of breaking discontinuous ground or soft rock masses. Rippers are also used to some extent in mining for surface preparation and in coal production. Butterworth-Heinemann: London. In the ripping process... Rippers are used widely in civil engineering projects to excavate slopes and wide swaths of surface area. Parallelogram.3 Reading – Chapter 13.....8 F..5 – Continuous Excavators. the dozer excavates the exposed coal and pushes the loosened material into windrows.4 Dozers and Ripping11 Dozers can be utilized for ground preparation when using a ripper.. 1992... 28th Ed.. Adjustable parallelogram.. MGE 415 – Rock Excavation 2004 Figure 5-51: Ripper in Action (a) (b) (c) Figure 5-52: (a) Radial ripper. The Adjustable Parallelogram ripper combines the features of both the hinge. Hinge-type rippers offer the advantage of an aggressive entry angle.type and parallelogram rippers. (c) Adjustable Parallelogram ripper In a Hinge-type ripper. but does not provide the aggressive tooth angle necessary for hardto-penetrate materials. As the shank enters the ground and penetrates to maximum depth. the linkage carrying the beam and shank pivots about a fixed point at the rear of the tractor. This type of ripper has advantages over the hinge-type when ripping above maximum depth. It can vary the tip angle beyond vertical for improved penetration and can be hydraulically adjusted while ripping to provide the optimum ripping angle in most Dessureault 1/14/2004 197 . the tooth angle is constantly changing. (b) Parallelogram ripper. A Parallelogramtype ripper allows the linkage carrying the beam and shank to maintain an essentially constant tip-ground angle regardless of tooth depth. but cannot be adjusted to compensate for varying conditions. also known as breakout – the maximum sustained force upwards generated by the lift cylinders Penetration force – the maximum sustained downward force.MGE 415 – Rock Excavation 2004 materials. generated by the ripper lift cylinders measured at the ripper tip. The ripper may have a multishank arrangement where multiple shanks are pulled by a single dozer (typically up to three). Figure 5-53: Ripper Nomenclature Other ripper nomenclature includes: Pryout. Dessureault 1/14/2004 198 . Down pressure. Dessureault 1/14/2004 199 . Good penetration is essential for high production and depends on: Material.MGE 415 – Rock Excavation Figure 5-54: Nomenclature for rippers (see Table 5-2) 2004 Table 5-2: Ripper Selection for Large CAT Dozers. in proper application. Ripper tip replacement should be included in hourly costs. Figure 5-55: Rock quality classification in relation to excavation processes. Ripping is an inexpensive method of removing discontinuous ground or soft rock masses. Ripping direction (material bedding planes. Some weaker fragmented sedimentary rocks (less than 15 MPs compressive strength.MGE 415 – Rock Excavation 2004 Point or tip angle with ground. Machine life average 8000 hrs (continuous ripping) May be 1/3 to ½ of blasting. ripping: Higher repair costs.1 Determining Rippability. 5. Figure 5-55 shows a rough chart of various excavation methods considering several ground types. Dessureault 1/14/2004 200 . slopes) Cost comparison: Blasting vs.4. example: mudstone) are not easily removed by blasting as they are pulverized in their immediate vicinity of the blasthole or may lift along bedding planes then fall back when gas pressure has been dissipated. Table 5-3 provides excavation characteristics for a wide variety of material. Figure 5-56 shows the seismic velocity of Dessureault 1/14/2004 201 .MGE 415 – Rock Excavation 2004 Table 5-3: Excavation Characteristics in relation to rock hardness and strength Table 5-4: Excavation Characteristics in relation to joint spacing Figure 5-56: Seismic velocities in relation to ripping The most common method for determining if a rock mass is economically rippable is seismic refraction. An adapted Q system of rock classification can be used to define an excavability index. However. The upper limit of ripper operations is ground with a seismic velocity approximately 2 km/s. N. and supplemented by Table 5-5. homogeneous material with no discontinuities RQD Rock quality designation Jn Number of joint sets Js Reducing effect which the block shape and orientation has on the force needed to break out the material Jr Roughness of the most unfavorable joint sets Ja Degree of alteration Table 5-6: Mass Strength Number for Rocks (Ms) Dessureault 1/14/2004 202 . N MS ( RQD) Js (J n ) Jr Ja Table 5-5: Variables in excavability index Ms mass strength number (amount of effort needed to excavate dry. Seismic velocity can vary as much as 1 km/s in identical materials. Rippability assessments based on rock classification is far more accurate. Table 5-6 through Table 5-10 provide comparative tables from which an excavability index can be calculated. which ca be calculated by the equation below.MGE 415 – Rock Excavation 2004 various rock types and their rippability. this generalized measurement only provides a very limited view of the issues in ripping operation. MGE 415 – Rock Excavation Table 5-7: Joint count number (Jc) and Joint set number (Jn) 2004 Table 5-8: Relative ground structure number (Js) Dessureault 1/14/2004 203 . MGE 415 – Rock Excavation Table 5-9: Joint roughness number (Jr) 2004 Table 5-10: Joint Alteration number (Ja) Dessureault 1/14/2004 204 . and lower shank again). Avoid ripping on or creating slopes the tractor cannot climb.3 Estimating Ripping Production Ripper production rates: can be determined empirically by: Determining the average cycle time. There are several key issues to be aware of when ripping downgrade.4. then calculate production rate. the machine must turn (takes 0. Avoid sideslopes.1 Ripping technique: Using hydraulic force to help fracture rock is an effective technique for ripping hard rock seams or pockets is to combine tractor drawbar power and hydraulic force.4. Engine speed is maintained high enough to allow the tractor to continue moving forward as the ripper shank angles back. or Record time spent ripping. Therefore 91m 26. The ripper shank is controlled to adjust shank angle back slightly. the ripper shank is returned to its original position (forward).MGE 415 – Rock Excavation 2004 5. the downgrade approach can be helpful when working a hard spot or seam. an estimate of ripper production in these terms is needed. The decelerator is used to control trackslip. 4. While the tractor moves forward. Consider that the ripper shank has a penetration depth of 61 cm. remove and weigh (through surveying & volume calculations). then remove the weight of the material. After every 91 meters.2 Ripping Downgrade Ripping downgrade can increase production.7m / min 3. Traction on rock is less than on dirt.25 minutes to raise shank. pivot dozer. rip spacing. For example.2.4.6 km/h = 26. If the job layout permits. consider an example where a CAT D10R dozer equipped with a single shank ripper excavates a path every 90 cm (width of tear) at an average speed of 1. When the ripper has contacted the hard rock. 5.25 (turn time) = 3.2 Operational issues: 5. the following procedure is undertaken: 1. 2. Since ripping is primarily compared to drilling and blasting. calculate production rate.66 min/pass Dessureault 1/14/2004 205 .7 m/min. 3. This is the equivalent of 1 pass.41 + .4. The simplest method is to record the time spent ripping. 5.6 km/h. combining ripper hydraulic force with tractor drawbar pull. Solution: 1.41 min 3. and depth penetration. on a per ton or volume basis.2. measure average rip distance. 90m = 49.3 passes per hour. is would be possible to make 45/3. However.1 BCM per pass Therefore the volume per pass is equivalent to 12.1 = 604 BCM/h Note that these estimates are approximately 10-20% higher than what would be expected on-site.3 x 49. Figure 5-58 also provides estimates for ripper productivities for various material type.61m x 0.MGE 415 – Rock Excavation 2004 if operator works 45 min/h. it should be recognized that in these charts: Rip full time (does not doze simultaneously) Power shift tractors with single shank rippers 100% efficiency (60 min/hr) For all classes of material. Figure 5-57: Ripper Production Chart Dessureault 1/14/2004 206 .66=12. such as those seen in Figure 5-57. The volume ripped can be estimated by: 91m x 0. Productivity can also be estimated using the ripper production charts. MGE 415 – Rock Excavation 2004 Figure 5-58: Performance chart for D11 Figure 5-59: D11 with ripper Dessureault 1/14/2004 207 . very definite limitations are usually placed upon the size of the rocks to be accepted. where there is equal opportunity for the use of either method. As mentioned before.MGE 415 – Rock Excavation 2004 5. Blasting is at times unpredictable.effective method to achieve these requirements. dozed material presents few problems. blasting can be made in terms of how the material is to be moved. loaders and hauling units . If this equipment can be used to finish the job . A cost analysis will indicate the economics of ripping over drilling blasting. in an urban area there may be restrictions prohibiting the use of explosives. It's soon apparent that considerable effort can be expended to rip the material in order to keep scrapers on the job. Ripper tips are the most expensive variable in the operating costs of ripping tractors. however. and transporting and loading methods. spacing. Generally. Scrapers can inexpensively haul materials which are well broken up and loosened. Variations in ripping depth.can be achieved only if the material is in small pieces and easy to handle." If the material is used to form an embankment. As we stated. reblasting). This is because many earthmoving jobs already involve track-type tractors and scrapers for a sizable portion of the total yardage. Their greatest advantage . as the desired rock fragmentation may be difficult to obtain and even require expensive secondary blasting (in effect. the end use of the material.high production . But in most situations. Elevating loaders and conveyors are high capacity systems. blasting decision. and direction of passes usually can produce the desired material size. making ripping a necessity.4. The final comparison of ripping vs. explosives.rather than bring in a rock crew with drills.4 Ripping vs. the first consideration is probably one of cost . This comparison indicates how ripper tip life is an important factor in deciphering the production needed for ripping to be cost effective. ripping is the most cost. Blasting Improvements in ripping tractor capabilities have made ripping a viable alternative to blasting. There are few size limitations when the rock is simply moved by a bulldozer and "wasted. End-use of the material also influences the ripping vs. Political factors or the threat of terrorism severely restrict the use of explosives in some countries. Appreciable increases in crusher production have been realized by cement plants and aggregate quarries after switching from a blasting to a ripping operation. accounting for approximately 30 to 40 percent of total operating costs on the Dessureault 1/14/2004 208 .it's not difficult to appreciate the savings involved.will it be cheaper to rip or blast? This initial-cost consideration must then be weighed with other influencing factors: the economics of fully utilizing equipment. Full utilization of the equipment available or already on the job can help determine the best method of loosening the material. environmental factors will undoubtedly play an important role in reaching this decision. Optimum compaction cannot be obtained if there are large rocks in the fill. For example. Material top loaded into hauling units cannot be larger than the loading bucket. blasting decision will to be loosened and moved. and on the blasting techniques. depend on the total volume of material production capabilities and costs of the size and relative efficiency of drilling and ADDITIONAL RESOURCES XX a ripping vs. In the final analysis.MGE 415 – Rock Excavation 2004 largest tractors.4 (Mandatory) Assignment XX? Dessureault 1/14/2004 209 . The above information is additionally supplemented by: Lecture for module 2. on the ripping tractor(s) used.