United States
Environmental Protection
Agency
Industrial Environmental Research
Laboratory
Cincinnati OH 45268
EPA-600/7-80-175
October 1980 "
Research and Development
vvEPA
User's Manual for
Premining Planning of
Eastern Surface Coal
Mining
Volume 2
Surface Mine
Engineering
Interagency
Energy/Environment
R&D Program
Report
-------
EPA-600/7-00-175
October lO'^
Region 111 Lil;
Environmental Pro^
USER'S MANUAL FOR PREMINING PLANNING OF
EASTERN SURFACE COAL MINING
VOLUME 2
SURFACE MINE ENGINEERING
by
R. V. Ramani, C. J. Bise,
C. Murray, L. W. Saperstein
The Pennsylvania State University
Department of Mineral Engineering
University Park, Pennsylvania 16802
EPA Grant Number R803882
Project Officer
John F. Martin
Energy Pollution Control Division
Industrial Environmental Research Laboratory
Cincinnati, Ohio 4526S
INDUSTRIAL ENVIRONMENTAL RESEARCH LABORATORY
OFFICE OF RESEARCH AND DEVELOPMENT
U.S. ENVIRONMENTAL PROTECTION AGENCY
CINCINNATI, OHIO 45268
-------
DISCLAIMER
This report has been reviewed by the Industrial Environmental Research
Laboratory, U.S. Environmental Protection Agency, and approved for publi-
cation. Approval does not signify that the contents necessarily reflect
the views and policies of the U.S. Environmental Protection Agency, nor
does mention of trade names or commercial products constitute endorsement
or recommendation for use.
11
-------
FOREWORD
When energy and material resources are extracted, processed, converted,
and used, the related pollutional impacts on our environment and even on
our health often require that new and increasingly more efficient pollution
control methods be used. The Industrial Environmental Research Laboratory-
Cincinnati (lERL-Ci) assists in developing and demonstrating new and
improved methodologies that will meet these needs both efficiently and
economically.
This document is the second in a series of six reports designed to
provide the surface coal mining industry and its regulators with a
comprehensive review of the best available methods for extracting coal
while protecting the environment. This volune summarizes the- engineering
and economic analyses that are necessary for prer.ining planning of surface
coal rining operations in the eastern coal fie las of the United States. For
further information on this series of reports, contact the Energy Pollution
Control Division.
David G. Stephan
Director
Industrial Environmental Rese.-irch Laooratory
Cincinnati
111
-------
ABSTRACT
The purpose of this research effort was to study the surface mining of
coal in the Eastern United States and to establish guidelines for developing,
evaluating, and selecting the least environmentally detrimental mining and
reclamation practices. The study was to consider the geological and hydro-
logical settings before mining as the basic inputs to premining planning, and
guidelines were to be developed for assessing alternatives in the areas of
surface mine engineering, water management, and land use planning.
The recommended methods, techniques, and alternatives for selecting and
designing mining systems are based on a review and critical evaluation of
the methods reported in the literature and applied in the field. They are
presented in a series of six reports, which together make up a user's manual
for premining planning of surface coal mining operations in the Eastern
United States.
This report, which is the second in a series of six, addresses the sur-
face mine engineering aspects. A comprehensive summary of the analyses
required to layout a surface mine, taking into consideration the limitations
imposed by geology, equipment, reclamation, economics, and environmental
control requirements is provided. Methods, techniques, and alternatives are
recommended for selecting and designing mining systems that include soil
handling and storage, overburden removal, and minimization of environmental
effects of drilling, blasting, off-highway transportation, coal loading,
and steep slopes.
This report was submitted in partial fulfillment of Grant No. R803882 by
the Department of Mineral Engineering of The Pennsylvania State University
under the sponsorship of the U.S. Environmental Protection Agency. This
report covers the period July 1, 1975, to May 30, 1978, and work was com-
pleted as of June 1, 1978-
IV
-------
CONTENTS
Foreword iii
Abstract iv
Figures . . . . . • ix
Tables • • xvi
Project Staff • • xx
Conversion Factors • • • xxii
Acknowledgements ... ... .... . . . . .... xxiii
1. Introduction 1
General 1
Energy Forms . . 1
Surface Mining 3
2. Exploration 7
Introduction 7
Types of Exploration . . . 7
Geological Analyses 9
Exploration Procedures .... ... 9
Summary 13
3. Surface Mining Equipment 23
Stripping Shovel . 23
Dragline 23
Bucket-Wheel Excavator 35
Loading Shovel .... 37
Scraper ... 37
Bulldozer ... 37
Front-end Loader 40
Hydraulic Excavator (or Hydraulic Shovel) . . 40
Trucks 40
Equipment Trends 44
Summary . . ... 44
4. Surface Mining Methods 49
Overview 49
Unit Operations 52
Area Mining .... . . 58
Dragline Overcasting . 69
Shovel Overcasting ... . 69
Bucket-Wheel Stripping . 69
BWE-Dragline Stripping . 69
BWE-Shovel Stripping 69
Smaller Scale Approaches . . 69
-------
CONTENTS (continued)
4. Surface Mining Methods (continued)
Contour Mining 79
The Head-of-Hollow Fill Technique 79
Block Cut 80
Mountaintop Mining 92
Open-Pit Mining 100
Special Considerations Multiseam Mining 100
Summary 108
5. Fragmentation Practices 110
Introduction 110
Drilling Ill
Percussion Drilling 112
Rotary Drilling 116
Vertical and Horizontal Drilling 119
Blasting 119
System Components 120
Design of Blasting Rounds 132
Blasting Pattern Problems 137
Summary 150
6. Surface Mine Haulroad Design 151
Introduction 151
Class I Roads 152
General 152
Design and Construction 152
Maintenance 152
Other Transportation Facilities 152
Surface Mine Haulroad 153
Haulroad Design 153
Alignment 154
Roadway Bearing Capacity 157
Plastic Deformation 157
Embankments 164
Drainage 170
Ditches 174
Culverts 174
Summary 176
7. Slope Stability 177
Introduction 177
Structures of Concern 178
Elements of Stability 179
References 180
Theory of Stability 181
Mechanics 181
Effect of Geology 184
Effect of Water 185
vi
-------
CONTENTS (continued)
7. Slope Stability (continued)
Design of Safe Slopes 187
Factor of Safety, First
Considerations . 187
Slip-Circle Analysis 188
Factor of Safety, Additional
Considerations 192
Constructing a Safe Slope 193
Monitoring 197
Summary 204
8. Operation Considerations 205
Introduction 205
Correction Factors 205
Cycle Time 209
Surface Mine Design . . • 219
Engineering Design and Pit Layout 219
Overburden Yardage to be Moved . . . 222
Calculation of Bucket Size . 222
Calculation of Reach 225
Machine Weight Using the Maximum
Usefulness Factor (MUF) 230
Dragline Digging Its Own Bench 230
Multiple-Seam Mining 233
Bucket-Wheel Excavator Selection 237
Soil Classification 237
BWE Output Considerations 242
Bucket-Wheel Design . 244
Other Considerations 245
Mode of Operations 245
Cut With a Crowding Machine 245
Cut With a Noncrowding Machine 247
Falling Cut Method 248
BWE Operations in USA 248
BWE Selection Parameters 250
Drilling and Blasting 251
Selection of Drilling Equipment 251
Number of Drills 252
Explosives Required Per Month 252
Coal Loading 254
Truck Fleet Selection 255
Secondary Equipment Selection 256
Bulldozers 260
Scrapers 265
Front-end Loaders 265
Summary 267
vii
-------
CONTENTS (continued)
9. Reclamation 270
Introduction 270
Land Effects 270
Reclamation Practices 271
Reclamation Planning 280
Land-Use Analysis 288
Economic Feasibility Studies 289
Technical Feasibility Studies 291
Monitoring and Maintenance • 299
Summary 309
10. Surface Mining Costs 310
Introduction 310
Stripping Ratio 310
Stripping Ratio in Multiple-Seam Mining 316
Stripping Ratio and Costs 318
Cost Analysis 319
Cost Estimation 321
Surface Mining Cost Models 322
Summary 335
11. References . 337
12. Bibliography 343
viii
-------
FIGURES
Number
1 Historical Trend of Domestic Coal Production
By Method of Mining
2 Average Thickness of Coal Seam Mined and
Overburden Excavated in the U.S ............ 5
3 Exploration Decisions Flow Diagram ........... 22
4 Stripping Shovel .................. 24
5 Stripping Shovel Working Dimensions .......... 31
6 Dragline ........................ 32
7 Dragline Cable Configuration .............. 33
8 Dragline Working Dimensions .............. 34
9 Bucket-Wheel Excavator ................. 36
10 Loading Shovel ..................... 38
11 Scraper ........................ 39
12 Bulldozer ....................... 41
13 Front-end Loader ................... 42
14 Hydraulic shovel loading into a truck ......... 43
15 Hydraulic shovel digging into a bench ......... 43
16 Equipment Application Zones .............. 45
17 Approximate Application Zones ............. 46
18 Optimum Shovel-Truck Combination ........... 47
19 Past Practice of Conventional Contour Mining ...... 50
ix
-------
FIGURES (continued)
Number
20 Auger Used to Drill Laterally into a Coal
Bed Exposed in Contour Mining 51
21 Typical Area Mining Method with Stripping Shovel .... 53
22 Typical Open-Pit Mining Method (Thick Seam) 54
23 Typical Open-Pit Mining Method (Dipping Seams) 55
24 Influence of Geological Factors in Surface
Mine Design 56
25 Equipment Application for Surface Mining Systems .... 66
26 A 45-Cubic-Yard Dragline at Work in a
Pennsylvania Coal Operation 70
27 Pit Layout at Bucyrus-Erie 4250-W Operation 71
28 A 90-Cubic-Meter Dragline With a 90-Meter Boom 72
29 Plan and Section Views of a Bucyrus-Eris 1950-B Pit • • • 73
30 Details of a Shovel Pit 74
31 Plan and Section Views of a Dragline and Bucket-Wheel
Excavator in Tandem Operation 75
32 Bucket-Wheel Excavator, Dragline, Coal-Loading
Shovel, Trucks, and a Dozer Ripper are Employed
in this Operation 76
33 Pit Layout at a Shovel and Bucket-Wheel Excavator
Tandem Operation in Illinois 77
34 An Operation in Illinois Using Bucket-Wheel
Excavator, Stripping Shovel, Loading Shovel
and Trucks 78
35 Head-of-Hollow Fill 81
36 A Check Dam, Downstream From a Hollow Fill 82
37 A Distant View of the French Drain 83
-------
FIGURES (continued)
Number
38 The Size and Gradation of the Rock that Form the
Core of the French Drain Is an Important
Design Consideration 84
39 The Valley Fill is Built From Bottom Up in Layers . 85
40 A Completed Head-of-Hollow Fill 86
41 Contour Mining Around a Hillside With the
Block System 87
42 Block-cut Method 88
43 Block-cut Method: Stripping Phase 89
44 Block-cut Method: Backfilling Phase 90
45 Mountaintop Removal Method: First Cut (Box Cut) .... 93
46 Mountaintop Removal Method: Second Cut 94
47 Mountaintop Removal Method: Fourth Cut 95
48 Mountaintop Removal Method: Mountaintop after
Final Grading and Topsoiling 96
49 Mountaintop Removal Method: Note the Head-of-Hollow
in the Front Being Prepared for Valley Fill 98
50 Large Area of Level Ground Created as a Result of
Contour Mining in West Virginia 99
51 Plan View of an Open-Pit Coal Mine 101
52 Multiseam Stripping Operation With a Dragline 102
53 Dragline Exposing Lowest Seam From Leveled Spoil .... 103
54 A Two-Seam Operation With a 30-Cubic-Yard
Dragline in Pennsylvania 104
55 Shovel-Dragline Tandem Operation for Multiseam Mining. . 106
56 Multiseam Stripping Operation With a Shovel 107
57 Multiple-Seam Mining Method: Two Seams Less Than
25 Feet Apart in the Same Highwall 109
xi
-------
FIGURES (continued)
Number
38 Single Row of Holes Connected by "Primacord" and
Detonated by Electric Blasting Caps or Caps and
Fuse Attached to "Primacord" at A 122
59 Double Row of Holes Connected by "Primacord" 122
60 Three Rows of Holes Connected by "Primacord" 122
61 Single Row Shot Fired in Sequence. "Primacord"
M.S. Connectors Inserted at Points "X" 123
62 Multiple Row Shot With Holes and Rows Fired in
Sequence From One End 124
63 Multiple Row Shot in Sequence Initiated at
Center Front 124
64 Multiple Row Shot Entire Row in Sequence 124
65 Single Row Horizontal Hole Shot in Bituminous
Stripping 125
66 A Cartridge of Explosive Undergoing the
Process of Detonation 127
67 Nomograph for Determining Loading Density 129
68 Nomograph for Finding Detonation Pressure 130
69 Bench Cross-Section View Showing D , B, H, T, and L . . 133
70 Generalized Blasting Patterns Showing B, S, b, and s . . 134
71 Representative Sedimentary Vertical Jointing System . . 139
72 Effect of Delay Time Between Shotholes on
Average Fragment Size 141
73 Comparison of Multiple Row Patterns Using
Echelon Firing on 1:1 and 2:1 Slopes 142
7~ Particle Velocity Versus Frequency With Recommended
Safe Blasting Criterion 143
75 Particle Velocity Versus Frequency for No
Damage Data . 144
xii
-------
FIGURES (continued)
Number
76 Proposed Blasting Plan (Example) ... . 147
77 Some Commonly Used Drilling and Blasting Patterns . . 148
78 Metalized Slurry is Used to Improve Blasting
Results With ANFO ... . . 149
79 Superelevation Runout Considerations . ... 156
81 Dual Wheels on an Axle .... 161
82 CBR Curves . . 162
83 Haulroad Materials and Layer Thicknesses ... . 163
84 Relationship Between Moisture Content and Density 169
85 Rainfall Intensity, Duration, and Frequency for the
Southern One-third of Ohio 172
86 Drainage Area Adjacent to a Haulroad . . 173
87 Pipe Culvert Capacity Graph . . . .... . . 175
88 Coulomb's Failure Surface 182
89 Critical Slip Circles For Purely Cohesive
Slopes, £=0 189
90 Slices on a Toe Circle for Use in Slip Circle Analyses
of Slope Stability . . 191
91 The Forces on Slice 5 191
92 Shovel Pit With Coal Fender . 223
93 Shovel Pit Without Coal Fender . ... 224
94 Dragline Pit .... . 228
95 Dragline Pit With Working Bench ... ... 229
96 Shovel MUF Graph ... ... ... 231
97 Dragline MUF Graph . . .
xiii
-------
FIGURES (continued)
Number
98 Multiseam Dragline Operation With Upper Seam
Exposed 234
99 Dragline Positioned on Leveled Spoil Removing
Parting 235
100 Block Digging Operation With BWE 246
101 Fall Cut Block Digging Operations with BWE 249
102 Front-end Loaders Variable Time 258
103 Scraper Distance Versus Travel Times 266
104 On Land Reclaimed with Grass and Trees after Strip
Mining, Responsible Coal Companies Have Created
Public Recreation Areas Like this Roadside Park
and Campground in Southeastern Ohio 274
105 Charolais Cattle and Giant Canadian Geese on
Mined Lands 275
106 Corn, as Well as Forage, can Grow on Former Coal
Land. The Ponds in the Background (Former Pits)
Provide Watering Places for Cattle 276
107 Charolais Calves Fatten in a Feedlot on Strip Mined
Land in Indiana 277
108 A Golf Course on Strip Mined Land 278
109 Various Phases of Surface Mining and Reclamation in a
Kentucky Coal Mine 279
110 Analysis of the East Pit Highwall, a Problem Wall
5 Years Old 284
111 Integrated and Interactive Surface Mining,
Reclamation, and Land-Use Planning 290
112 Movement of Spoil Within a Block to Insure
Segregation and Burial of Acid-Bearing Material . . . 296
113 Spoil Placement For Better Environmental Impact .... 297
114 A Pusher-Scraper Combination for Topsoil and
Overburden Removal 302
xiv
-------
FIGURES (continued)
Number
115 Large Crawler Tractor in the Process of Rough
Grading Coal Mine Spoil Banks in Hilly Terrain .... 303
116 A Giant Bulldozer With an Experimental 40-Foot Blade
Levels Mined Land in Southeastern Kansas 304
117 Completed Rough Grading of Coal Mine Spoil Banks
in Hilly Terrain 305
118 Bulldozer Spreading Topsoil on Graded Spoil 306
119 Coal Mine Spoil Banks Reclaimed to Farmland 307
120 Area Which was Mined and Reclaimed in 1975 308
3
121 Stripping Ratio in m /Tonne for Various Coal
Values and Stripping Costs 317
122 Production Versus Overburden Removal 323
123 Overburden Removal Versus Machine Capacity 325
124 Capital Requirements for d Surface Coal Mine 326
125 Graph for the Computation of Depreciation Charges . . . 327
126 Productivities for Surface Coal Mine Operations
in the United States, 1973 329
127 Graph for the Calculation of Labor Costs Per Ton .... 330
128 Relationship of Labor Costs and Direct
Operating Costs 331
129 Capital Structure for Start-up Mine Investment 333
130 Graph for the Calculation of Financing Charges
and Rates of Return 334
xv
-------
TABLES
Number
1 Selected Forecasts of Total U.S. Energy
Consumption in 1985 and 2000 ............. 2
2 Features to be Identified on an Exploration Map . . .10
3 Stages of Exploration for Coal ............ 15
Areas of Technical Expertise in Geotechnical
Investigation and Design . . .......... 17
:> The Importance of Geotechnical Factors to Specific
Mining Considerations .............. 18
6 Relationships of Investigation Techniques to
Mining Considerations ... ......... 19
Alternate Investigation Techniques .... ...... 20
S Qualitative Evaluation of Present Predictability
of Mining Considerations Behavior ........ 21
9 Stripping Shovel Specifications . . ....... 25
10 Dragline Specifications ............... 28
11 Equipment Rating - Topsoil Removal and Replacement ... 59
12 Equipment Rating - Overburden Preparation ....... 60
13 Equipment Rating Overburden Removal ......... 61
l-» Equipment Rating Coal Loading ............ 62
15 Equipment Rating Regrading and Backfilling . . . . 63
16 Equipment Rating - Revegetation ............ 64
1~ Equipment Rating Coal Hauling ......... 65
IS Single-Phase Stripping Possibilities .......... 67
xva
-------
TABLES (continued)
Number
19 Multiple-Seam Stripping Possibilities 68
20 Specifications of Some Rifle-Bar Rotation and
Independent Rotation Percussive Drills 113
21 Down-the-Hole Drill Specifications .... .... 115
22 Three-Cone Rolling Cutter Bit Dimensions and
Hole Volume Produced . ... 117
23 Bit Weight as a Function of Rock Strength and
Bit Size .... . . 118
24 Worksheet for Design of Blasting Rounds ... . 138
25 Maximum Explosive Charges Using Scaled
Distance Formula . . 1-6
26 Relationship Between Speed and Radius of a. Curve
and the Superelevation Recommended . . . 155
27 Bearing Capacities of Materials . . . 160
28 Classification of Highway Subgrade Materials 165
29 General Guide to Selection of Soils on Basis of
Anticipated Embankment Performance 167
30 Recommended Minimum Requirements for Compaction
of Embankments . . ..... 165
31 Runoff Coefficients . . . . 171
32 Steps in Computing Slope Safety Factors 190
33 Summary of Methods for Prevention and Correction
of Landslides ... . . . 198
34 Effect of Depth of Cut and Angle of Swing on
Shovel Cycle Time . . 206
35 Effect of Depth of Cut and Angle of Swing on
Dragline Cycle Time t .... 207
36 Minutes of Operating Time Per Hour and
Equivalent Operating Efficiencies .... .... 205
xvi i
-------
TABLES (continued)
Number
37 Fill Factor 210
38 Material Weights, Percentage, Swell and Load Factors . . 212
39 Operating Efficiency Factors 213
40 Walking Dragline Machine Output 215
41 Approximate Cycles Per Hour for Stripping
Shovels and Draglines 216
42 Walking Dragline Cycle Time Calculation 217
43 Walking Dragline Operating Delays 220
44 Specific Cutting Forces of Virgin Soils for BWE
Excavation 240
45 Specific Cutting Forces of Materials Suitable for
BWE Loading Operations 241
46 Application of Drilling and Penetration Methods
to Types of Rock 253
47 Fixed Times for Front-end Loaders 257
48 Miscellaneous Correction Factors 259
49 Bulldozer Correction Factors 261
50 Swell Factor - Bulldozed Material 262
51 Bulldozer Selection Chart for Ripping and
Dozing Application 263
52 Wheel-Loader Production Estimating Table 268
53 Potential Major Effects of Surface Mining Unit
Operations on Land, Water and Air Resources 272
54 Information Needs on Ecosystem Factors for Surface
Mine Planning and Land Use 281
55 Chemical Analyses of the East Pit Highwall 285
xviii
-------
TABLES (continued)
Number
56 Estimated Average Production Costs @ Average
Stripping Ratio =8.1 and 0.5 Acres
Disturbed Per Thousand Tons of Coal
Produced 292
57 Strip Mine Reclamation Project Variables Affecting
Backfilling and Grading Costs . 293
58 A Time Sequence For Reclamation Activities 300
59 Analysis of Stripping Ratios 313
60 Factors in Stripping Ratio Calculations 314
61 Calculation of Selling Price 336
xix
-------
PROJECT STAFF
The materials contained in the manual were prepared by an inter-
departmental and interdisciplinary group of the College of Earth and Mineral
Sciences of The Pennsylvania State University. Overall management for the
project was provided by the Department of Mineral Engineering. The project
staff was comprised of the following personnel:
Dr. R. V. Raziani
Professor of Mining Engineering
Dr. L. W. Saperstein
Professor of Mining Engineering
Dr. H. L. Love 11
Professor of Mining Engineering
Dr. R. R. Parizek
Professor of Geology
Dr. C. G. Knight
Associate Professor of Geography
Dr. R. Stefanko
Professor of Mining Engineering
Prof. R. L. Frantz
Professor of Mining Engineering
M. L. Glar
Research Assistant in Mining
Engineering
L. 3. Phelrs
Instructor in Mining Engineering
C. J. Bise
Instructor in Mining Engineering
?. J. Duhaine
Graduate Assistant in Geography
- Project Manager and Co-Principal Investi-
gator (Surface Mining Engineering, Land
Use Planning, Mine Planning)
- Co-Principal Investigator (Federal and
State Laws and Regulations, Land Use
Management)
- Co-Principal Investigator
(Water Quality Managenent)
Faculty Associate (Geology and
Hydrology)
Faculty Associate (Land Use
Managenent)
Faculty Associate (Project Consultant)
- Faculty Associate (Project Consultant)
In-House Project Co-ordinator
Surface Mine Engineering
Surface Mine Engineering
- Land Use Management
-------
D. Forsberg
Graduate Assistant in Mining
Engineering
K. W. Grubaugh
Graduate Assistant in Mining
Engineering
C . Murray
Graduate Assistant in Mining
Engineering
D. Richardson
Graduate Assistant in Mining
Engineering
J. Sgambat
Graduate Assistant in
A. Weiner
Graduate Assistant in
Engineering
Geology
- Water Quality Management
Laws and Regulations
Mining Engineering
- Water Quality Management
Geology
- Water Quality Manager>ent
Mining
-------
CONVERSION FACTORS
In this report, wherever practicable, the data are provided in metric
units. However, in many instances, particularly in figures and tables, the
units used are those that are commonly designated and understood by the per-
sonnel in the mining industry. Conversion to metric units, in these cases,
may not only be not useful but frustrates the reader with unfamiliar units.
Therefore, this conversion table is provided for use with this report.
Mass
1 short ton
1 long ton
1 pound
Length
1 inch
1 foot
1 yard
1 mile
Square Measure
1 square inch
1 square foot
1 square yard
1 acre
Cubic Measure
1 cubic foot
1 cubic yard
0.907 metric ton
1.016 metric ton
0.4536 kilogram
Power
1 horsepower = 745.7 watts
Pressure
1 pound (force) per square inch
6895 newtons per square meter
2.54 centimeters
30.48 centimeters
0.9144 meter
1.609 kilometers
6.452 square centimeters
0.0929 square meters
0.836 square meters
0.405 hectare
0.0283 cubic meters
0.0765 cubic meters
Liquid Measure
1 gallon = 3.785 litres
Density
1 pound (mass)/cubic foot = 0.01602 gram/cubic centimeter
Force
1 pound (force) = 0.4536 kilogram force
1 pound (force) = 4.448 newtons
Energy
1 British Thermal Unit = 1055 Joules
1 foot pound (force) = 1.356 Joules
xxii
-------
ACKNOWLEDGEMENTS
The assistance and cooperation of many mining companies and the various
State and Federal agencies are gratefully acknowledged.
Sincere thanks are extended to Mr. John F. Martin, Project Officer,
Industrial Environmental Research Laboratory, Cincinnati, Ohio, and
Mr. Elmore C. Grim, former Project Officer and presently Director, Division
of Forestry, the Kentucky Division of Natural Resources, Frankfort, Kentucky,
for their support and guidance.
Sincere appreciation is also extended to Mr. Richard D. Ellison,
Executive Vice President, D'Appolonia Consulting Engineers, Inc., Pittsburgh,
Pennsylvania, for permission to use information from a report "Geotechnical
Investigations and Design Guidelines."
Assistance provided by Mr. Sukumar Bandopadhyay, Graduate Assistant in
Mineral Engineering (Mining Section), and Dr. James M. Riddle, formerly
Research Associate in Mining Engineering, The Pennsylvania State University,
is gratefully acknowledged.
xxiii
-------
SECTION 1
INTRODUCTION
GENERAL
By the year 2000 A.D., the annual U.S. demand for energy is expected
to double and that for the world, to triple (Starr, 1971). Gordon (1976)
has summarized a sample of forecasts of U.S. energy consumption for the
years 1985 and 2000 from several sources and task forces dealing with various
aspects of the appraisal (Table 1). These projected increases will try man'^
ability to explore, discover, extract, and beneficiate the fuels in necessary
volumes, to transport them safely, and to dispose of effluent wastes with
minimum harmful environmental effects. Considering the difficulties experi-
enced at present in mining without environmental damage, in finding accept-
able locations for new power plants, and in controlling stack emissions
from existing plants, the energy projections for the year 2000 indicate the
need for a thorough evaluation of the available options and cautious
planning.
ENERGY FORMS
In order to meet the growing demand for energy, all fuels will be
required in increasing quantities. Economics and technology must also assure
that each fuel is put to the use for which it is best suited. Coal has
gradually overtaken natural gas and petroleum in the field of electricity
generation, and the use of nuclear energy has increased. Over a longer
period of time, nuclear energy may replace the fossil fuels (coal, oil
shales, tar sands, natural gas, and petroleum) as the leading fuel in
electricity generation, provided that technological advancements occur as
predicted. The use of geothermal, solar radiation, and tidal energies is
not yet widely practical and/or economical. However, coal, particularly
surface-mined coal, may compete with nuclear energy for some time. Forecasts
indicate that as technology in the field of coal gasification progresses,
demand for coal may shift from electricity generation to synthetic gas and
liquefied products. This would tend to remove some of the burden from
natural gas and oil and extend the projected peaks of each fuel. If coal
gasification becomes feasible by the early 1980's, the demand for coal may
accelerate more rapidly than predicted. Although there is no exact accuracy
in forecasting energy demand and fuel usage, coal should have an assured
position throughout the future because of its abundant widespread distribu-
tion and chemical versatility. These forecasts are made in a complex
ever-changing world where much depends upon international politics, tech-
nological innovation, and economic decisions. A projection of energy or
-------
TABLE 1. SELECTED FORECASTS OF TOTAL U.S. ENERGY CONSUMPTION
IN 1985 AND 2000 (QUADRILLION BTU) (GORDON, 1976)
1985
2000
Dupree and West
National Petroleum Council — Low
High
Intermediate
116.63 191.90
112.50
124.90
130.00
Project Independence — BAU* $ 7 Oil (no conservation) 109.06
AS* $11 Oil (with conser-
vation) 96.25
Energy Policy Project — Historic Growth 116.00
Technical Fix 92.00
Zero Growth 88.00
Organization for Economic Cooperation and Development —
Base 114.32
$ 6 Oil 105.83
$11 Oil 100.09
ERDA* — No new initiatives 107.30
Conservation 96.97
Synthetic fuels 107.28
Electrification 106.77
Limited nuclear 107.05
Conservation, Synthetic fuels and Nuclear 98.14
EEI* — High growth, Low energy prices 124.20
High growth, High energy prices 112.90
Medium growth, High energy prices 104.30
Low growth, High energy prices 98.90
Dupree and Corsentino
FEA* — 1976 Reference $ 8 Oil 103.29
1976 Conservation $16 Oil 92.51
n.a.
a .a.
n .a.
n .el.
n.a.
187.00
124.00
100.00
II . d .
11 .a .
11. a.
165.47
122.48
165.42
161.16
158.01
137.03
193.93
179.10
155.10
104.50
103.54 163.43
11 . OL .
u. a.
BAU: Business as Usual.
AS: Accelerated Growth Scenario.
ERDA: Energy Research and Development Administration.
EEI: Edison Electric Institute.
FEA: Federal Energy Administration.
1973 Energy consumption was about 76 Quadrillion Btus
1 Btu 0.2930 watts.
-------
fuel demands may be rapidly altered due to one or a combination of changes.
Recent factors such as ecological conservation can also effectively block
construction of new facilities and lead to a greater delay in supplying the
required energy.
SURFACE MINING
Surface mining is a broad term which refers to the removal of the soil
and strata over a mineral or fuel deposit and the removal of the deposit
itself. The various unit operations in surface mining include:
1. preparing the surface
2. drilling
3. blasting
4. overburden removal
5. loading of the deposit
6. haulage of the mined deposit
7. reclamation
Other operations include the building of drainage systems, waste
handling, and topsoil and subsoil storage. Surface mine planning and engi-
neering, therefore, transcends the fields of engineering and physical
sciences. Primary considerations include: geology, rock and soil mechanics,
excavation and materials handling systems, and environmental aspects.
Surface mining of coal is an answer to many of the underground coal
mining problems but in itself creates new problems. Recovery rates with
surface mining are much higher than that from underground mining, generally
80 to 90 percent, and coal losses are limited to spillage and losses in
transit. Safety problems are not severe in surface mining, but acid drainage
from pit and spoil areas must be considered. Land reclamation is an expen-
sive undertaking. A limiting factor with coal surface mining is the depth
to which coal can be profitably mined. Currently in most areas the maximum
depth of overburden for economic mining is slightly over 60.96 meter
(200 ft). Surface mining at greater depths may be feasible in the future,
depending upon technology, economics and environmental concerns. The rising
trend in coal production in the United States should continue through the
rest of the century. The contribution of surface mining to coal production
has progressively increased. Today, surface-mined coal accounts for more
than 50% of the total coal production (Figure 1). However, the average
thickness of coal removed has remained constant, as opposed to an increasing
average overburden thickness (Figure 2), a result of major advances in
surface mining technology.
Considering the problems which confront the mining and electricity
generation industries in their quest to meet the demands for energy, surface
mining will be challenged to provide coal at a low cost with minimal distur-
bance to the environment. With respect to the regulation of surface coal
mining operations and to the establishment of performance standards, the
passage of the Surface Mining Control and Reclamation Act of 1977 represents
a significant effort on the part of the federal government. The U.S.
-------
o
z
I
LL
O
o
O
h-
01
>
CO
a?
z
o
p
o
D
O
o
cc
0.
o
o
CO
D
1950
1955
1960
1965
1970
1975
Figure 1. Historical trend of domestic coal production by
method of mining (EPRI, 1977).
-------
70
60
ui 50
LJ
CO
o
5
Q
cc
:D
oo
40
20
10
0
1946
OVERBURDEN EXCAVATED
COAL SEAM MINED
I960
1955
I960
1965
1970
Figure 2. Average thickness of coal seam mined and overburden excavated in the U.S.
(EPRI, 1977).
-------
Congress, according to the Act, finds and declares that "surface mining and
reclamation technology are now developed so that effective and reasonable
regulation of surface coal mining operations . . . is an appropriate and
necessary means to minimize as far as practicable, the adverse social,
economic, and environmental effects..." Among the purposes of the Act,
important ones are to assure that (1) surface mining operations are not
conducted where reclamation required by the Act is not feasible, (2) surface
coal mining operations are conducted to protect the environment, and (3)
adequate procedures are undertaken to reclaim surface areas as contempora-
neously as possible with the surface coal mining operations. While the
increasing energy demand promises a good market for coal, increasing federal,
state, and local statutes with regard to surface mining make the exploitation
of coal deposits a difficult but challenging task. In order to meet this
challenge, a need exists to critically study and evaluate surface mining of
coal in detail and provide suggestions for solutions to these problems.
Attempts to provide solutions to problems require a thorough knowledge of
existing methods and criteria. It is the purpose of this report, therefore,
to briefly discuss the various aspects of surface mining of coal and lay
out important planning and design considerations.
-------
SECTION 2
EXPLORATION
INTRODUCTION
Exploration includes activities and evaluations necessary to gather
data for making decisions on such topics as the desirability for further
exploration, feasibility of mining, size, initial flowsheet, and annual
output of new extractive operations (Bailey, 1968). However, the main pur-
pose of exploration activity is to facilitate the discovery and acquisition
of coal deposits that may be extracted and processed economically, now or
in the future. The extent of exploration programs will vary according to
company practices and the area under investigation. The program should go
to the point of detail where additional study would not provide greater
reliability.
The bulk of exploration work is usually assigned to an exploration
department staffed with geologists and engineers. Other staff departments
that perform considerable work in exploration are legal, public relations,
and accounting. The highly simplified, primary role of the exploration
department is to determine the existence, location, quantity and quality of
a coal deposit; assess the possible favorable and unfavorable factors of
mining; and begin the acquisition program for mineral and surface rights and
ownership of the required lands. A legal staff should determine prevailing
federal and local laws and practices and provide backup for any required
negotiations.
TYPES OF EXPLORATION
There are two general types of exploration approaches, regional and
detailed. Regional exploration is a reconnaisance activity and is intended
to determine if and where a deposit may exist. Detailed exploration is
performed on a specific target area where a deposit is known to exist and
delineation and verification are desired.
Areas subjected to exploration procedures may be broadly classified
into lands where coal seams are known to exist and virgin areas. Coal
exploration programs may be defined as follows:
1. Geologic prospecting with minimum drilling. This type of explora-
tion is common in Appalachian regions where outcrops are abundant
and geologic formations are well known.
-------
2. Random drilling. This method is applied to previously explored
regions with high geologic predictability.
3. Pattern drilling and statistical approaches. A program of this
type is necessary in virgin areas or in areas of high geological
variability.
The importance of a thorough exploration program to a coal mining
company cannot be over-emphasized, particularly when viewed with regard to
surface mine engineering and design. In the eastern coal mining areas with
which this study is concerned, the presence or absence of coal is usually
known. However, coal mining companies still require much information about
overburden and coal characteristics, especially with respect to environmental
considerations and present laws and regulations.
The initial step in all exploration programs is a literature search
to obtain all public information and previous exploration data concerning
a property. Publications and maps may be available from the U.S. Geological
Survey, the U.S. Forest Service, state geological services, and other federal
and state agencies, and county offices of clerk, assessor, and surveyor.
Extensive drill hole information may be available in areas where oil and
gas exploration has occurred. Records of mining companies that have
previously operated in an area are helpful if they can be obtained. Volume 3
of this research effort discusses the geologic structure of the coal basins
and provides a useful summary of the reported literature.
Maps are important tools in all phases of a mining operation. Maps
should at least show geology, political boundaries, and surface contours.
Special maps may yield aerial views and paleotopographic and paleoenviron-
mental information. Good maps and mapping techniques will provide a means
for planning and accomplishing exploration, development, reclamation,
day-to-day operations, and equipment moves. Calculation of material volumes,
location of physical elements, and determination of mining conditions are
expedited by the use of maps. Maps provide a method for recording data so
that they can be organized and analyzed for ready reference.
Maps suitable for use as base maps are available from a variety of
sources. A wide range of map scales and types are available. Various
laws require certain scales for reclamation and mining plans but this may
vary from state to state. The most commonly used scales are 1" 500',
1" = 400', and 1" 100'- U.S. Geological Survey 7-1/2' or 15' topographical
maps can be easily enlarged to 1" 1000' and are the type most commonly
used for exploration base maps. Other map scales are available but are
suitable for more specific uses. These will be discussed in reference to
the appropriate topic.
Aerial photography and mapping methods (photogrammetry) are increasing
in usefulness in the surface coal mining industry. For some companies,
they are indispensable. They serve as a base for mining plans, vegetation,
and geologic and structural identification. Infrared aerial photography is
useful for historical/archaeological investigations. Photogrammetric methods
are relatively easy and cheap, can be adjusted to any scale, and are highly
accurate in any terrain.
8
-------
Aerial photography done at an altitude designed to produce maps of
the scale 1" 100' is quite popular. This method will provide for control
that is accurate to one meter in 5000 meters horizontally and one-half of
a contour interval vertically. The usual contour interval is 1.52 meter
(5 ft) in normal terrain and 3.04 meter (10 ft) in rough terrain. An advan-
tage of this type of mapping is that it may show drainage configuration,
roads, buildings, lakes, streams, timber, power lines, railroads, and fences
or other features that might be missed by a ground survey. Table 2 provides
a list of data that should be included or at least considered in a mapping
program.
GEOLOGICAL ANALYSES
Geological analyses are employed by most companies and yield crucial
information with regard to pre- and post-depositional forces and the effects
of migrating chemical constituents from the overburden on coal. Physical
features that can be identified include faults, folds, rolls, dikes, coal
water courses, and the presence of concretions and sulfur. The occurrence
of seam pinch-outs, local variations in properties and thickness of the coal,
and the presence of erosion and channel sands can also be determined by a
geologic analysis in conjunction with drilling information. The probability
of soil siltation or dust problems can be assessed. Physical properties of
coal and overburden that can be determined by physical inspection are
specific gravity, hardness, color, luster, size stability, weathering, grind-
ability, friability, fracture, cleat and cleavage. Geologic analysis will
also identify chemical and geochemical factors such as toxic or neutralizing
elements in overburden and coal.
EXPLORATION PROCEDURES
Good and rapid exploration programs will facilitate pit design and
equipment selection whereby costly lead times for start-up can be reduced.
Analyses for mining method, preparation processes, and market conditions
will be completed sooner so that more profitable decisions can be made.
General procedures for exploration are as follows:
1. Property for a new surface mine.
Step a. Acquiring and consolidating sufficient reserves to
justify capital expenditures for a viable mine plan.
Step b. Gathering reliable information about coal beds and
mining conditions. Exploration, especially drilling, must be performed with
proper engineering control to ensure accuracy of the information required
for evaluation. The correctness of reserve calculations, along with other
factors such as size of property, topography, depth to coal, highwall
stability, and thickness of the seam, will have a high influence on equip-
ment selection. Occasionally, the target area may have surface features
which preclude mining, thereby reducing available reserves. The scope of
the study will vary from property to property and according to the apriori
and actual information already available.
-------
TABLE 2. FEATURES TO BE IDENTIFIED ON AN EXPLORATION MAP
Topography
Natural features
Surface contours
Drainage configurations
Streams continuous and intermittent
Lakes
Rock croppings
Coal outcroppings
Natural cavern openings
Man-made features
Fences
Wells
Reservoirs
Railroads
Power lines
Oil and gas lines
Water and sewage lines
Highways, roads and streets
Buildings
Old mining areas surface and underground openings
Location of drill holes - collar elevation, coal elevation,
thickness of coal, depth of hole
Geographic features
Private and political property boundaries
Property ownership surface and mineral
Forests and vegetation
General land use
Annual precipitation
Centers of population labor supply and housing
Unusual concentrations of plant or animal life
Climate altitude
Geological features
Washouts
Burn lines
Faults and fractures
Major geological formation
Groundwater courses, water table classification, relation
to coal
(continued)
10
-------
TABLE 2. (continued)
Geological features
Topsoil thickness and condition
Overburden special characteristics, thickness, chemical
features
Coal - depth, thickness, physical and chemical properties
Dip of beds
Presence of methane concentrations
Partings
Hydrology
Properties distribution, and circulation of surface and
groundwaters
11
-------
2. Property with an existing surface mine. This type of exploration
is usually intended to expand reserves or further delineate known reserves.
The analysis required may be different from the one above and involve merely
the accumulation and organization of additional data.
Before actual field exploration can begin, several requirements must
be satisfied. The two important items are the landowner's permission for
entry and a prospecting permit.
Formal land acquisition procedures are usually initiated after explora-
tion begins. Acquisition procedures, however, should be a. continuing part
of exploration procedures. Drilling programs are conducted under options
to lease or purchase a particular property. Leases are negotiated with the
federal government, state governments, and private landowners. Purchase
agreements are negotiated with private landowners.
It is important that a mining company control the surface as well as
the minerals during the period of mining activity. Complex land ownership
patterns in eastern coal mining areas make this objective difficult to
attain. Another problem resulting from the complex ownership patterns is
the difficulty to put together a block of land containing reserves large
enough to be mined economically. This problem is most common east of the
Mississippi River.
After a thorough review of the available information from public and
private sources, additional information may be obtained only from additional
studies. Regional exploration may provide opportunities to use geophysical
methods in conjunction with a drilling program. Gravity surveys and seismic
reflection and refraction techniques may be helpful in determining the
location and geometry of marker beds, faults, and basement rock. Large area
exploration may also employ satellite or airborne photography, infrared
imagery, and side-looking radar in connection with a drilling program.
Side-looking radar is particularly helpful in the Appalachians where the
surface growth is relatively dense.
The most common type of field exploration is a drilling program. The
purpose of a drilling program is to provide a comprehensive data base cover-
ing the location of the coal and the overburden. Therefore, drilling must
be planned to yield maximum information. Drill hole patterns will vary
according to the terrain, geological formations, the presence of coal out-
croppings, and the type of exploration program.
When very large areas are being studied, hole spacings will vary greatly
and will probably not be in any set pattern. When the program is narrowed
to a regional study, a grid pattern is most common. Detailed exploration
is performed in areas where coal is known to exist, and therefore, closer
spaced drill patterns are required.
One of the main advantages of a drilling program is that it can provide
physical samples of the coal and the overburden for chemical and physical
analyses. This is accomplished by coring. In most programs, a percentage
12
-------
of the holes are cored for overburden and coal. Usually, this ranges from
10 to 25 percent. Holes should be cored from the surface, through the coal,
and then below it in order to maximize information about 'spoil' quality.
Air coring and wireline recovery allow for increased recovery up to 100
percent in favorable conditions. Photographing cores as they come out of
the hole can provide greater reliability of data. Gamma ray and conductivity
logs should be used. The coal core recovery should closely approach 100
percent. If it does not, the analytical information obtained should be
considered suspect.
Prospecting along outcrops may also be accomplished with dozer cuts at
regular intervals. This gives the attitude of the coal and will help to
determine how much outcrop coal (degraded or oxidized coal) must be left as
a barrier and the nature of the overburden with respect to machine operation.
Whatever method of exploration is used, the data from drill holes or
outcrop openings must be large enough to present an accurate picture of the
deposit and conditions. All data should be plotted on maps as soon as it
is obtained. Detailed exploration results should be plotted on large scale
maps whose scales may vary from 1" = 100' to 1" 2000' The scale of
1" = 100' is the most popular. These maps indicate the depth to the coal
and the thickness of the coal seam, isopachs (contour lines representing
uniform thickness) for coal and overburden, and stratigraphic sections
showing a representative profile of the overburden. There are too many
instances where not enough information was gathered and investment decisions
were made incorrectly.
SUMMARY
In summary, the details of exploration programs for locating and
defining surface mineral coal deposits vary widely, depending on the prior
availability of data and information about the area. The major steps that
are involved, and some of the techniques that can be used are identified in
Table 3. Basically, the major steps require a search of available material
regarding the nature and occurrence of the coal, a regional exploration
program using as many of the airborne geophysical methods as is practical
or possible, and then site specific ground work using drilling and bulk
sampling techniques.
While it may be possible to bypass some of the steps in some areas, it
is becoming increasingly clear that exploration programs today must provide
mine planners with information not only on the geology and hydrology of the
coal deposits but also on factors such as land use, historical and archae-
ological features, and vegetation and wildlife. Since geological informa-
tion is basic to many mine planning processes, and exploration is one of
the first steps in acquiring this information, it is important to choose
the right personnel for geotechnical investigations, identify the information
needs and importance and their relationships to mine planning, and use suit-
able data acquisition systems or techniques. The use or the limitations of
the collected data for mine planning purposes must also be known. Tables 4
13
-------
through 8 are taken from a paper by Ellison and Thurman (1976) and can
serve as a guide in this area. Finally, as shown in Figure 3, exploration
is a sequential process where after each stage a decision must be made to
either continue or abandon the exploration program.
14
-------
TABLE 3. STAGES OF EXPLORATION FOR COAL (BAILEY, 1968)
Exploration
Stage
Stage 1
Regional
Appraisal
Purpose
o-Availability for Mineral
Extraction
o-Identification of known
problem areas (i) water,
(ii) historical/cultural
features, (iii) endangered
fauna/flora, etc.
o-Geologic compilation
o-Market availability and needs
Stage 2 o-Infrastructure Development
Reconnaissance
of the Region/ F-Field check on geologic
or Target Area compilation
F-Identification of anomalies
F-Watershed identification
F-Regional land use plans
F-Detailed mapping of coal
outcrops/coal quality, quantity
F-Overburden quality and thick-
ness
F-Site location for plants and
facilities
F-Waste disposal areas
F-Wildlife and vegetation
Method/Source of Data
Federal, State, Local,
Zoning Laws, Land Use
Plans
Previous history,
records, government
and private studies
U.S. Geological Survey
State Geological
Survey
Intelligence data,
growth potential, plans
of utilities
Aerial photography
Geophysical survey
Infrared photography
Pilot-hole drilling
(continued)
15
-------
TABLE 3. (continued)
Exploration
Stage
Stage 3
Detailed
Target Area
Inves tigation
Purpose
F&L-Geochemical analysis of
streams/sediments
F-Historical and cultural
features
o-Analysis to define more
closely probable extraction
problems
o-Analysis to identify major
environmental problems
F-Detailed drilling, core
drilling of coal
F-Logging (down the hole,
geophysical)
L-Physical, Chemical and
Washability of coal
F-Overburden depth,
stratigraphic sequence
L-Physical, chemical and toxic
characteristics of overburden
o-Ore reserve compilation
F-Groundwater table/
permeability/charge potential
F-Sites for box-cut
o-Preliminary economic
evaluation
Method/Source of Data
Core drilling
Geophysical surveys
Aerial photo
o - Office study; F Field study; L Laboratory study.
16
-------
TABLE 4. AREAS OF TECHNICAL EXPERTISE IN GEOTECHNICAL INVESTIGATION AND DESIGN
(ELLISON AND THURMAN, 1976)
PLANNING PHASE
GEOTECHNICAL INPUT
TYPICAL PRINCIPAL^1>2^
GEOTECHNICAL INVESTIGATORS
INVESTIGATORS PROVIDING^
NECESSARY SUPPLEMENTAL DATA
Feasibility Studies
General Site
Evaluation
Mining Engineer-Geologist
Structural Geologist
Photogeologist
Geotechnical Engineer
Hydrogeologist
Geochemist
Petrologist
Seismologist
Conceptual Planning
Site-Specific
Investigations
Mining Engineer-Geologist
Geotechnical Engineer
Hydrogeologist
Geophysicist
Photogeologist
Structural Geologist
Geochemist
Petrologist
Preliminary Design
Final Design
Geotechnical
Designs
Geotechnical Engineer
Hydrogeologist
Mining Engineer-Geologist
Geophysicist
Structural Geologist
Engineering Geologist
(1)
(2)
Principal Investigators may vary depending upon site conditions.
An Investigator with proper experience and training may provide necessary expertise in more
than one area.
-------
TABLE 5. THE IMPORTANCE OF GEOTECHNICAL FACTORS TO SPECIFIC
MINING CONSIDERATIONS (ELLISON AND THURMAN, 1976)
18
-------
TABLE 6. RELATIONSHIPS OF INVESTIGATION TECHNIQUES TO MINING
CONSIDERATIONS (ELLISON AND THURMAN, 1976)
19
-------
TABLE 7. ALTERNATE INVESTIGATION TECHNIQUES (ELLISON AND THURMAN, 1976)
LEGEND
- IMAT TM£ M.TtlWATE TtCMWlOUE CA» PMCTICALLT
*SYS k USCD IN PIKX or TMC usic TECHNHUC
b-l»OIC*TtS THAT THt ALTEMlAre TtCMNBiK CA11 SOMCTUltS
K US£0 • n-UX. Of THE BASIC rtCHMIOUt
I - IXCHCATtS THAT TMC U.TEJOIATE TECHNIQUE SHOULD COST
ICSS THAN TME BASK TECHNIQUE
2 - INDICATES THAT THE ALTERNATE TECHNWOE USUALLY COSTJ
MODE THAN THE BASIC TECHNIQUE
NOTE
TME USER OF THIS TABLE MUST RCC06NIZE LIMITATIONS
IN COMPARING TECHNIQUES FOR OENCRALItEO SITUATIONS
THIS TABL£ IS A PLANNING GUIDE ONLY
20
-------
TABLE 8. QUALITATIVE EVALUATION OF PRESENT PREDICTABILITY OF MINING
CONSIDERATIONS BEHAVIOR (ELLISON AND THRUMAN, 1976)
\v Mining
\Considerations
Basis \^
For \
Predicting N.
Behavior \^
Transfer of ^ '
Technology
Field (1>
Inve s t iga t ions
Analysis
Model Tests
Empirical
Relations
Overall (2)
Predictability
Requirement
For Monitoring
SURFACE MINING
Cut Slope
Stability
2
3
2
5
2
2
2
Floor Heave
3
3
2
5
4
2
3
Spoils Stability
1
4
1
5
3
1
5
c
w
•H
co
cu
O
cfl
s
3
cO
2
3
1
1
5
Handling of Water
2
3
2
4
2
2
Rippability
2
2
2
1
1
Rehabilitation
2
2
2
2
2
1
Monitoring System
Design
2
3
2
4
3
2
Spontaneous
Combustion
2
3
3
5
2
3
1
SURFACE FACILITIES
60
C
•r-l
4-J
•H
co
CO
0)
•H
4-1
•H
I—I
•H
CJ
CO
fn
1
2
1
5
5
1
5
Structures
1
2
1
5
5
1
5
CO
-------
NJ
Stage I
Regional
Appraisal
Region not
Attractive
at the Present
Time
Stage 2
Regional
Reconnaissance
of Target Area
Reject: Region
Unfavorable
T
I
I
I
Stage 3
Detailed Target
Area Investigations
Oetai led
Mine
Planning
Target Area
Region Unattractive
at the Present Time
v
"I
Uneconomic
Mineral
Deposit
Normal Exploration Steps
Conceivable Recycling Sequence
Exploration Decision
Figure 3. Exploration decision flow diagram (Bailey, 1968).
-------
SECTION 3
SURFACE MINING EQUIPMENT
Several types of equipment are available for surface mining, and because
of this, a degree of selectivity can be exercised in choosing equipment for
a particular job. A brief description of the equipment commonly used in
surface mining and their performance characteristics is provided here.
Extended descriptions of their physical construction and operational modes
are available in many references (Peurifoy, 1970; Pfleider, 1968; Caterpillar
Tractor Co., 1975, 1976).
STRIPPING SHOVEL
The stripping shovel (Figure 4) is a highly productive machine capable
of handling all types of material, including large blocky material. However
it is restricted to comparatively rigid operating conditions and has a
limited ability to spoil. The shovel digs material from its operating base
upwards. Factors affecting shovel productivity include low flexibility,
low maneuverability, slow tramming rates, bank and bucket slides, and water
seepage into the pit. Table 9 gives manufacturer specifications for various
stripping shovels, and Figure 5 diagrammatically explains these relationships
to the stripping shovel.
DRAGLINE
The dragline is an adaptation of the power crane for earthmoving
purposes (Figures 6 and 7). These machines have the ability to dig above
and below grade; can function under less rigid operating conditions than
shovels; can perform optimally when digging an unconsolidated nonmuddy
material; can handle blasted rock; and usually have greater reach than the
shovel. Since the dragline operates on the ground surface, the problems of
bank slides, water runoffs, and seepage are virtually eliminated. The drag-
line^ ability to "chop down" some fraction of the overburden to dig its own
bench (i.e., dig above its operating base) is an advantage for extending its
operating range, though this method of operation should be avoided as far as
practicable because of significant drops in the machine's production capacity.
In any case, due to less precise motions, draglines are only 75% to 80 per-
cent as productive as shovels of equal capacity. Table 10 gives manufacturer
specifications for draglines, and Figure 8 shows the relationship between
the dimensions.
23
-------
Figure 4. Stripping shovel,
24
-------
TABLE 9. STRIPPING SHOVEL SPECIFICATIONS (STEFANKO, ET AL, 1973)
1-1
01
J-l
3
4J
U
to
14-1
3
C
BE
BE
BE
BE
BE
BE
M
BE
M
M
BE
BE
M
BE
BE
BE
M
M
,— i
0)
t>0 '•""•'v
IH ^
•<* s_x
B •
O 00
O 0)
CQ Q
45
45
45
45
40
47 1/2
45
47 1/2
47
46
49
47 1/2
47
50
52
50
45
45
4-J
f]
00
*n /"™N
QJ PQ
i-M ^"""^
§•
C
•H e H
a -H i
1 CO 4J
P a PH
104-0
106-6
111-9
96-0
122-6
133-9
134-0
101-0
143-6
150-0
117-6
135-0
125-0
153-0
ca
CO 4-J
3 rC
•H bO
tl -H ,—v
CO (D O
£*! 33 ^
oo a •
C 3 C3
•H B M
CX -H 1
§x •
CO 4J
Q S PM
230-0
140-0
156-3
127-6
162-6
180-0
185-0
150-6
195-6
210-0
151-6
192-0
181-0
211-6
4-1
£1
00
•H * > — *
CU H O
33 *rH ^""^
o
00 (X •
C co c
•H M
0, M 1
3 > 4J
Q O Pn
68-0
64-0
76-0
71-0
106-3
110-3
123-6
127-6
118-0
133-6
(0
3
•H
""O ** •'""V
CO ^H W
« -H ^-^
0
00 O. •
C CO C
•H M
Du M I
6 0) •
3 ? W
Q O PM
100-3
96-0
114-6
109-6
158-0
167-0
184-0
200-9
184-0
217-9
/ — s
GJ v_/
4-1 C
J3 M
bO 00 I
•H C •
0) -H 4J
BC CX P4
B
00 3 «
C Q co
•H 3
CX • -H
B X T3
43-0
43-0
48-0
47-0
83-0
69-6
81-9
64-0
88-3
92-6
104-0
76-4
104-6
' 108-0
83-6
102-0
95-0
113-6
co
3
•H
""O
£6. O
* V '
oo B •
C 3 C
•H B M
a -H i
& ^(:
3 CO 4-1
« S Pn
106-6
101-6
121-0
115-0
225-0
147-6
162-0
135-0
169-6
189-0
193-9
157-0
203-6
218-0
163-0
200-0
186-6
219-9
£
(T!
00
•H '•""v
a) 33
oo B •
d 3 c
•H B M
4J -rl 1
4J X •
92-0
88-0
104-0
99-6
141-3
152-9
163-6
176-9
162-0
189-6
« 33 ^
oo B •
C 3 C
•H B M
4J -H 1
•MX •
3 CO 4J
^j ^r^ r^j
158-6
165-0
182-0
199-0
185-6
218-6
y— N
•t-1 C
i-C ^^
00 00 1
•H C •
(1J -H 4J
4J
00 3 ~
G 0 CO
•H 3
4J • -H
4-1 X T3
3 CO cfl
0 S K
88-9
95-6
99-6
112-6
104-0
122-0
^
^ V^rf'
CO
3 •
•H C
T3 M
cO 1
4-1
00 PM
C
•H -
4J •
4-1 X
51
114-0
110-0
131-6
126-6
245-0
158-0
173-0
145-0
181-3
200-9
207-6
170-0
216-3
232-0
172-0
215-0
201-6
236-9
S3
Ul
(continued)
-------
TABLE 9. (continued)
t-J
CU
)^
3
4-1
O
CO
IH
3
C
3
BE
BE
BE
BE
BE
BE
M
BE
M
M
BE
BE
M
BE
BE
BE
M
M
, — i
CU
3
550B
550B
1050B
1050B
1850B
1650B
5761
1650B
5860
5861
1950B
1850B
5900
3850B
1950B
3850B
5960
6360
lr-*i
S
•4-1 ~^
0 CL.
3 •
co 1 C
3 C M
•H co 1
CO iH 4-1
PS o fe
66-0
66-0
75-6
76-0
125-0
97-0
95-3
84-0
99-3
1-5-6
113-6
109-0
126-3
135-0
96-6
129-0
124-0
156-9
^^
25
U-l ..^
O CO
M .
42 CU C
4-1 ,-H H
&0 S 1
c to •
CU M 4-1
hJ 0 fe
17-3
17-3
21-2 1/8
21-2 1/8
34-0
22-7 1/2
23-4
22-7 1/2
30-0
30-0
35-0
34-0
34-0
41-0
35-0
41-0
42-0
45-0
,~^
0
ff* N — '
U-l (0
O V-i
CU C
43 <-H H
T3 cfl •
•H (-1 4-1
& C_> fe
9-0
9-0
10-2
10-2
16-0
13-5
14-0
13-5
16-2
18-0
19-6
16-0
21-0
25-8
19-6
25-8
26-4
30-0
„
4-1
i-H
CU
« -^
PU
O
CO
rf (U
-1 i f^
-0 o
jH £
30
30
36
36
66
54
60
54
66
72
78
66
82
100
78
100
104
120
„
CO
)_l 0) x-v
CU H 0-
O to
M •
X <_> C
t> H
00 .*"^ I
C w •
Q) 0 4-1
t-J CQ fe
47-0
47-0
54-11 1/8
54-11 1/8
75-6
58-7 1/2
59-4
58-7 1/2
72-0
72-0
80-0
75-0
78-0
89-0
79-0
89-0
92-0
103-0
„
CO
J_J
CO
0 P •
0 C
43 "— ^
4-1 43 I
"O 4-1 •
£ ° r^
13 pq fe
38-9
38-9
43-11
43-11
57-0
49-5
50-0
49-5
58-2
60-0
68-6
57-0
65-0
73-8
63-6
73-8
76-4
88-0
CU /— v
4-1 C/3
-0 cfl ^
c3 e
CO " v) •
f* i^ j«^
r^-l rS M
43 4J O M
4J 00 M 1
T3 C! &< •
•H CU O. 4J
3 iJ <; fe
22-9x48-1 1/2
22-9x48-1 1/2
26-9x52-0
26-9x52-0
58-0x79-6
36-0x61-8
45-0x65-0
36-0x61-8
50-0x72-0
52-0x80-0
64-0x85-2
58-0x79-6
60-0x86-8
70-0x87-0
64-0x85-2
70-0x87-0
74-0x95-0
72-0x104-0
co
3
•H
'O
cti ^**
p2 H
**-/
13 •*
C3 B •
W 3 C
B ^
)-< -H |
tO ^ •
-------
TABLE 9. (continued)
co
n
01
3
4-1
0
CO
<4-l
3
C
eg
s
BE
BE
BE
BE
BE
BE
M
BE
M
M
BE
BE
M
BE
BE
BE
M
M
rH
0)
-o
Q
S
550B
550B
1050B
1050B
1850B
1650B
5761
1650B
5860
5861
1950B
1850B
5900
3850B
1950B
3850B
5960
6360
ti,
Q) x—.
s >
CD CO v-'
U M
C r^f *
cfl C
t i L| J— -|
cfl CU 1
QJ t> •
H C 4J
19-0
19-0
21-9
21-9
29-0
24-6
23-6 3/4
24-6
27-9
27-9
33-1 1/2
29-0
33-2
38-6
33-1 1/2
38-6
35-6
39-0
O /-N
H -13
l|_l 4-1 ^
0 C 0
0 O •
M *H P^i C
A
4J
G W)
0 C
O <1)
PQ i-4
95
95
113
llw
235
145
170
135
180
200
200
150
200
210
170
200
190
215
01 T3
H 3
TJ iH *
C 0 H
cfl C o)
SB M (X
M X -H C
0) U Q M
& 00 1
p, p] hn .
•H CU C 4-1
Q hJ -H fe
61-6
56-6
70-0
64-0
146-0
86-0
102-6
80-6
109-0
118-0
122-0
90-6
126-0
137-0
102-6
125-0
117-0
133-0
4-1
•H
U
Cfl W
& M
U cfl
^~*
}_i
01 O
a -H
P , r>
•rl 3
O 0
25
28
40
45
65
70
75
75
80
90
100
100
105
125
130
140
150
180
-------
TABLE 10. DRAGLINE SPECIFICATIONS (STEFANKO, ET AL, 1973)
VJ
0)
1-1
3
4J
0
tfl
U-J
3
C
01
S
M
M
M
P
M
BE
BE
M
P
P
BE
BE
P
M
BE
M
BE
BE
M
P
P
P
M
BE
M
BE
M
M
BE
BE
M
M
BE
.-i
cu
'O
o
s
15 1M*
183M*
D/E
7400
732
7500
4 SOW
Diesel
4 SOW
19 5M*
736
740
800W
1260W
752
7820
1300W
7920
1350W
1370W
8000
757
762
862
8050
1500W
8200
2450W
8400
8750
2560W
2570W
8850
8950
4250W
• T— »
CU <
i-H •>— '
00
C CO
< d)
CU
e ^
O 00
o cu
PQ O
46
42
30
30
30
40
40
43
30
30
35
30
30
30
34
30
38
38
32 1/2
30
38
33
38
37
38
30 1/2
37 1/2
30
30
30 1/2
34
38
*.
Si
00
c
cu
>J]
C3 4J
o cu
o cu
pa fa
80
120
235
175
220
175
175
120
170
195
195
225
215
275
235
275
267
267
325
255
245
275
315
267
320
275
305
275
275
285
360
325
310
CO
3
•H
TJ
3
K
/— ^
00 CQ
C N>— '
•H
p, 4J
6 cu
3 cu
Q fa
65
100
220
173
209
154
154
98
171
197
179
215
221
262
215
263
241
241
299
259
230
275
290
241
282
249
290
246
274
283
340
300
302
4-1
JT
00
•H X~N
cu o
DC N— ^
00 •
c c
•H M
0. 1
s •
3 JJ
(~} pt^
38-0
60-6
104-0
64-0
84-0
63-0
59-0
70-0
75-0
112-0
104-0
136-0
95-0
113-0
115-0
135-0
137-0
120-0
132-4
146-0
137-0
*
(-;
J-l
0.,-N
CU Q
a ^
00 •
c c
•H M
00 1
00 •
•H 4J
O f-^*
40-0
40-6
117-6
115-0
130-0
80-0
80-0
75-0
115-0
120-0
135-0
130-0
135-0
150-0
150-0
135-0
150-0
130-0
180-0
155-0
135-0
165-0
200-0
115-0
150-0
130-0
200-0
126-6
175-0
165-0
180-0
158-0
185-0
cu
u
cu
e
nj x-~*i
•r-l W
Q v
CU •
*o c
•H M
CO 1
4J
3 4-»
O fa
31-0
37-0
37-0
36-0
36-0
41-0
45-0
45-0
55-0
56-0
50-0
50-0
52-0
52-0
58-0
65-0
60-6
70-0
66-0
65-0
63-6
58-0
63-6
65-0
75-0
65-0
74-0
80-0
80-0
105-0
f,
cu
o
to fa
N^/
14-1
0 •
c
4_| |
T3 •
•rH 4-1
s fa
3-6
5-6
6-0
6-0
6-0
6-0
4-7
7-6
8-0
7-6
10-0
10-6
9-0
9-0
10-0
10-0
10-0
13-0
11-6
12-6
12-6
13-0
12-0
11-0
12-0
12-0
15-0
12-0
14-0
15-0
16-6
20-0
cu
o
CO /— ^
o
o
& c
4J M
00 1
c
a; 4J
hJ fa
20-9
30-0
36-6
36-6
42-6
42-6
31-3
35-0
40-0
50-0
56-0
49-0
48-0
54-0
55-0
55-0
62-0
65-0
54-6
59-0
59-0
65-0
70-0
55-0
70-0
55-2
70-0
65-0
72-0
70-0
70-0
130-0
(continued)
28
-------
TABLE 10. (continued)
[Manufacturer
M
M
M
P
M
BE
BE
M
P
P
BE
BE
P
M
BE
M
BE
BE
M
P
P
P
M
BE
M
BE
M
M
BE
BE
M
M
BE
CD
TJ
151M*
183M*
D/E
7400
732
7500
4 SOW
Diesel
4 SOW
19 5M*
736
740
80 OW
1260W
752
7820
1300W
7920
1350W
1370W
8000
757
762
862
8050
1500W
8200
2450W
8400
8750
2560W
2570W
8850
8950
4250W
•* .• — \
>-l CO ffi
CD CD ^-^
> 0
0 f. •
to C
^J 4J .
•HOW
££ pO pu
17-6
22-6
44-6
51-2
49-6
49-6
22-7
62-6
77-0
71-0
70-0
75-0
75-3
81-3
94-0
94-0
89-6
83-0
90-10
92-0
109-0
93-0
104-6
114-0
116-6
151-0
Clearance
Radius,
Ft. -In. (J)
18-0
23-0
38-0
30-10
39-0
38-0
38-0
26-6
41-0
48-8
50-0
52-0
57-0
50-0
57-0
66-0
66-0
66-0
66-0
63-0
75-0
63-0
66-0
68-0
68-0
69-0
66-0
77-0
80-0
80-0
84-0
78-0
105-0
4-1 •> —
0
o - •
fe co C
P M
g -H I
O 'o •
0 Cd 4J
CQ M fe
7-5
8-3
16-2
21-0
16-2
16-8 1/2
16-8 1/2
9-5
22-6
26-0
15-3
17-2 1/2
32-6
20-3
17-2 1/2
21-6
25-0
25-0
21-6
36-6
35-0
39-0
21-6
25-0
21-6
25-0
23-3
24-0
30-0
30-0
24-0
26-0
50-0
Clearance
Height,
Ft. -In. (L)
6-3
5-0
4-4
5-4
4-4
4-8 1/2
4-8 1/2
5-9
6-9
7-7
8-3
9-0
7-8
7-4
9-0
8-7 1/2
8-6
8-6
8-8
9-2 1/2
9-7
10-3
8-8
8-6
11-0
8-0
13-0
15-10
14-0
14-0
13-0
16-0
16-7
£
4J -~s
O
O - •
(n 4J C
g W) 1
O "f~i •
O CD 4-1
8-1
10-1 1/2
7-3
10-0
7-3
8-6
8-6
10-2
10-6
11-8
14-9
17-2
12-0
13-6
17-2
12-3 1/2
15-2
15-2
13-9
15-0
16-0
17-5 1/2
13-9
15-2
14-10
16-0
19-2
21-4
16-0
16-0
21-4
22-0
24-0
Dumping
Clearance,
Feet (N)
33
33
35
42
46
48
50
52
55
65
71
129
4-1
C
0 -CD
6 00 4->
O i-l CD
O CD (D
^f\ *T^ r** i
120
120
127
129
148
180
180
180
185
153
158
214
(continued)
29
-------
TABLE 10. (continued)
Manufacturer
M
M
M
P
M
BE
BE
M
P
P
BE
BE
P
M
BE
M
BE
BE
M
P
P
P
M
BE
M
BE
M
M
BE
BE
M
M
BE
cu
151M*
183M*
D/E
7400
732
7500
4 SOW
Diesel
480W
19 5 M*
736
740
800W
1260W
752
7820
1300W
7920
13 SOW
1370W
8000
757
762
862
8050
1500W
8200
2450W
8400
8750
2560W
2570W
8850
8950
4250W
Point Sheave
Pitch Diameter
Inches (P)
38
38
54
80
54
66
66
50
80
80
90
85 1/4
80
82
85 1/4
92
120
120
96
90
110
102
96
120
120
144
126
110
144
144
132
132
144
Bucket
Capacity
Range, Cubic
Yards
4-8 1/2
5 1/2-10
9-14
13-19
10-20
12-18
12-18
16-18
16-24
21-32
16-26
25-40
36-48
35 --4 5
42
40-60
40-50
52
45-65
45-64
62-74
55-68
50-70
65
60-75
75
60-80
80-115
90
115
110-140
140-160
220
* Crawler Mounted; BE Bucyrus-Erie Co.; M Marion Power Shovel Co.;
P - Page Engineering Co.; D/E Diesel Electric. All other electric unless
specified.
NOTE: All dimensions based on largest bucket capacity available. For
other bucket sizes consult manufacturers' specifications.
1 ft 0.3048 m
1 cu yd 0.7645 cu m
30
-------
Figure 5. Stripping shovel working dimensions.
-------
OJ
jo
Figure 6. Dragline.
-------
Hoist chain
Bucket
y/A^w/AWy/jsW/^Xw^w/^
Figure 7. Dragline cable configuration.
-------
Figure 8. Dragline working dimensions
-------
BUCKET-WHEEL EXCAVATOR
The bucket-wheel excavator has its greatest application in unconsol-
idated overburden where little or no preparation is required. The excavator
is adaptable to soft, thin or thick layers of overburden or broken ore, pre-
ferably consisting of earth or clay, or sand or soft shales with no hard
rock formations present. The bucket-wheel is a continuous mining machine,
independent of cyclic operation methods. Most wheel excavators are designed
to allow the wheel to excavate below, above, or on the working level, and
discharge into belt conveyors which are normally low-cost transportation
devices. The long stacker allows a much greater discharge radius while con-
suming less power per cubic yard of material removed than other equipment
(Price, Manula, and Ramani, 1973).
Bucket-wheel excavators for strip mining may be divided into two main
types, German and American-Kolbe. The wheel, or digging component, of both
types consists of a relatively large diameter wheel with buckets arranged
around its circumference and an arrangement to transfer the mined material
from these buckets to a conveyor system (Figure 9).
The German type was developed for use in their lignite fields. It
is usually crawler mounted, with mountings arranged to give three-point
support, and has a relatively low ground-bearing pressure, approximately
103.42 x 103 Pascals (15 psi). The ladder (loading) and stacker (discharge)
booms swing independently of each other and therefore require counterbalances.
Each boom is equipped with a conveyor belt and the transfer of material from
the ladder boom to the stacker boom is accomplished through a somewhat com-
plicated chute arrangement. These machines may excavate overburden, lignite,
or both overburden and lignite.
Early model German-type bucket-wheel excavators were equipped with both
independent swing of the ladder and stacker booms and with "crowd", a feature
whereby the ladder assembly is mounted on rollers and a track so that it may
be advanced or retracted. As the size of excavators increased, it became
uneconomical to retain both independent swing and crowd, so most late model
German-type excavators do not have the crowd feature.
Most Araerican-Kolbe-type machines are crawler-mounted with four-point
support and have hydraulic cylinders for leveling. Groundbearing pressures
are high, approximately 314.76 x 103 Pascals (45 psi) because the machines
are designed to operate from the coal surface. The ladder and stacker booms
do not have independent swing and are not as heavily counterweighted as the
German-type machines; since their horizontal relationship does not change,
each tends to counterbalance the other. Transfer of materials from the
loading to the stacker boom conveyor is simple, since the direction of
material flow does not change. These machines usually excavate overburden
only and may be used in conjunction with a stripping shovel. Overburden is
usually spoiled but may be transferred to an external haulage system.
The Kolbe-type bucket-wheel excavators have crowd. Due to mechanical
design restrictions, these excavators do not have independent swing of the
35
-------
CO
Figure 9. Bucket-wheel excavator,
-------
stacker boom. Several other types of bucket-wheel excavators have been
developed in the United States for special applications. These may be either
crawler, rubber tire, or rail mounted and have a wide range of capacities.
In contrast to a shovel or dragline, the wheel excavator has a lower
instantaneous power demand, no shock loadings, and less weight for greater
outputs. Bucket wheels are also an asset for land reclamation, because spoil
piles are more regular in size and contour and loose soil can be placed over
the hard rock removed from above the coal seam in tandem operations.
Limiting factors for the wheel excavator include less flexibility due
to the great bulk of the machine, high maintenance costs, and lower operating
efficiency. Other limiting features in bucket-wheel utilization are the
considerable working surface preparation required, and the difficulty encoun-
tered in handling hard materials or rocks sometimes present in unconsolidated
overburden. The loading problems combined with a low break-out force also
limit the application of wheel loaders. Safe operation on high benches can
also become a problem.
LOADING SHOVEL
The loading shovel in most cases is a scaled-down version of the strip-
ping shovel. The shorter boom of the loading shovel is equipped with either
a bottom-dump or a rigidly attached side-dump bucket. The loading shovel,
widely used for coal loading in surface mines, is an efficient machine when
great mobility is not required (Figure 10). However, it is being replaced
at many mines with front-end loaders which can perform better than shovels,
especially in contour mining. The loading shovel can load above grade and
over the end of a track; can handle coarser materials than the front-end
loader; has a longer productive life and better availability; and can load
selectively under various mine conditions (Wendertz, 1970).
SCRAPER
Scrapers are of two distinct types, the wheel-tractor scraper (Figure 11)
and the track-tractor scraper (Drevdahl, 1961). The main distinction is that
the wheel-tractor scraper or rubber-tired scraper, as it is sometimes called,
has a speed advantage over the track type, making it more suitable for in-
creased haul distances. The track type, on the other hand, has superior
traction in mud; can make its own roads in rough terrain; and can handle
larger amounts of material. In surface coal mines the wheel-tractor scraper
is more commonly used, although it often requires pushers to assist in load-
ing. When haul distances exceed one mile, scrapers are uneconomical when
compared to shovel-truck combinations.
BULLDOZER
A bulldozer is a tractor mounted with a blade which is secured by pivot
arms. The blade comes in different types and capacities: U-shaped, "S" or
37
-------
00
Figure 10. Loading shovel.
-------
I
Figure 11. Scraper,
-------
straight-blade, and "A" or angling-blade. Most surface mine dozer applica-
tions utilize special "U" shaped blades for light noncohesive material
(Figure 12). The bulldozer is a versatile machine, normally limited to a
short operating radius of about 152.4 meter (500 ft). It is applicable for
site clearing and preparation; construction and upkeep of access routes and
haulroads; cleaning up behind shovels and draglines; boosting trucks from
pit; towing disabled equipment; maintaining spoil piles, moving cable skids
and stumps; and ripping and disposal. It is also used extensively in strip-
ping and leveling for reclamation.
FRONT-END LOADER
A front-end loader is a wheel tractor with a loading bucket mounted on
the front (Figure 13). These machines are efficient for loading coal from
seams of small thicknesses (0.304 to 3.048 meter thick (1 ft to 10 ft) where
it becomes necessary to gather coal before loading. They have high mobility
and can operate under favorable conditions with high productivity. They
also have found increased application in contour stripping for load and carry
functions over short distances.
HYDRAULIC EXCAVATOR (OR HYDRAULIC SHOVEL)
For mining and other earthmoving operations, hydraulic excavators are
adaptations of the traditional backhoe for excavation and loading purposes.
Among the advantages claimed for this equipment are the greater break-out
force and the variable pitch dipper. Figures 14 and 15 show the use of a
hydraulic excavator to remove overburden and load it into a truck.
TRUCKS
In surface mining, off-highway trucks are used primarily for haulage
because of their inbuilt ability for continuous uninterrupted operation,
while regular highway trucks are used only as service vehicles. There are
mainly two types of off-highway trucks, the bottom dump and the rear dump
truck. The bottom dump will handle aggregate and unconsolidated material
and dumps while in motion. They operate efficiently on grades of less than
four percent, and are primarily used for coal haulage in surface coal mines.
The rear dump truck comes in various forms: 1) small to medium two-axle
trucks, ideal for small pits and development haulage; 2) large two-axle
trucks which can handle large loads and negotiate steep grades; and 3) large
size three-axle trucks which have identical attributes as the earlier two,
but are uneconomical on haul roads due to exhorbitant tire costs (Burton,
1975, 1976).
Off-highway trucks are characterized by engine or drive train type,
usually diesel mechanical, diesel electrical, and gasoline internal combus-
tion engines. There are trucks available with either mechanical transmission
drive or electric wheel drive. Electrical drive has the advantage of down-
hill braking the vehicle to eight kmh (five mph) before mechanical or
40
-------
Figure 12. Bulldozer.
41
-------
r-o
Figure 13. Front-end loader.
-------
Figure 14. Hydraulic shovel loading into a truck.
Figure 15. Hydraulic shovel digging into a bench.
43
-------
hydraulic brakes are applied for final stop. Off-highway trucks are
available in ranges from 26.41 x 103kg to 355.61 x 103kg (26 tons to
350 tons). Figures 16, 17, and 18 show truck and scraper application
zones and the match requirements for trucks and shovels.
EQUIPMENT TRENDS
It is difficult to generalize the equipment preference of operators
and manufacturers. Depending on the type of equipment, the trend is toward
the selection of medium and large capacities. For instance, the trend in
draglines is towards medium capacity equipment with long booms. Long boom
draglines have greater operating range, reduce the amount of rehandle, and
have improved operating costs.
While no large stripping shovel has been selected in the recent past,
a vastly increased use of quarry mine shovels is seen in surface coal
operations for both overburden and coal handling. The trend in the loading
shovels is towards large capacity machines, i.e., 15.29 cu m to 19.11 cu m
(20- to 25-cu yd) shovels. This is a result of the shift to large truck
capacities. While several truck manufacturers have 152.4 tonne (150-ton)
units in service, in 1976, Wabco introduced its 3200B Haulpak, a 238.76 tonne
(235-ton) truck (Felix, 1976). It is evident that the normally small
secondary equipment are increasing in size for application with the already
stabilizing medium to large size primary stripping equipment.
For secondary equipment such as dozers, front-end loaders, and scrapers,
the trend is also toward larger sizes. In contour mining, there is an in-
crease in the use of this equipment for methods which are referred to as
"haulback" methods. In these methods, the overburden is transported by
scrapers, and front-end loaders over short distances; when the haul is long,
usually front-end loaders and trucks are employed. Haulback methods have
become more common because of their flexibility in selective spoil placement
and in reclamation. While small front-end loaders have not been displaced,
large loaders are being employed in coal surface mines. The Michigan 6750,
a 18.34 cu m (24-cu yd) loader, has proven useful in eastern surface mines.
Large diameter drill holes with longer drill rods are possible with
modern rotary drills equipped with air filtering and ventilating devices.
Additional efforts have been directed toward improving the productivity of
the current models by increasing the rotary speed and adding automatic
controls, thereby allowing operators to fully utilize the drill capabilities.
SUMMARY
This chapter has briefly reviewed the major equipment currently used
in surface mining of coal. The need for overburden segregation, topsoil
removal, storage and replacement, and selective placement and burial of
toxic materials has dictated the use of several pieces of equipment that
have loading and carrying capabilities. From a reliability point of view,
the largest available machines have not been widely accepted. At
44
-------
C.C.
20
18
•o
o 16
>-
u
5 14
3
O
c
o
5 10
/ Xr^1 «s.r»*VO^ .••'*'''*' ^r\^
.^/ S& ..• 'Ix' * *fet*fc^
/ • O .••**"** ^^^^y^G^
4? « ^ v ^ •*** « ^JkttC^
~ o' •tt** ^"**{O*X
C? X ^* fcr _^^^^^^0>-
«, y .Xo^ ^"^^
~ ^ 1 ^t>
j'c\*
/f ^
/I
1
1 Conditions: 4% Rolling Resistance
- / 0% Grade Resistance
Haul-12 M.P.H. Max.
Return-ISM.RH. Max.
Note : Total Resistance Less
Than 10% Will Hove Little
Effect On Economic Point
i i I i 1 I I
1
-^^^»<
***•*-
—
—
—
_
-
-
—
i
200 400 600 800 1000 1200
Haul Distance-One Way (Ft.)
1400 1600
Figure 16. Equipment application zones (Haley, 1974).
-------
c/2
CO
UJ
o
z
10
8
6
4
2
APPROXIMATE APPLICATION ZONES
UJ
o
DC
UJ
0.
END DUMP HAULERS
ELE-
VATING
PUSH
LOADED
^SCRAPERS
SCRAPERS
BOTTOM DUMP
HAULERS
i
_i_
j_
_L
I 23456789 10 II !2xioooFT.
610 1219 1829 2438 3048 3667 METERS
HAUL-ONE WAY
Figure 17. Approximate application zones
(International Harvester, 1975).
46
-------
t
0
(A
o
4- O
CO C
o —
o *-
0)
o
o
Best Shovel-Truck Combinations
_L
I
I
I
I
40 60 80 100 120 140 160
Truck Capacity (Tons)
180 200
LEGEND
II Yard Shovels
22Yard Shovels
Figure 18. Optimum shovel-truck combinations (Pfleider, 1968).
47
-------
the present time, there are several research projects funded by the
Department of the Interior, Bureau of Mines, for equipment and methods
development in surface coal mining. For the near future, draglines, trucks.
and shovels are seen as the main overburden removal equipment.
48
-------
SECTION 4
SURFACE MINING METHODS
OVERVIEW
Surface mining of coal is conducted in a relatively simple sequence of
operations which includes: 1) preparing the surface, 2) drilling, 3) blast-
ing, 4) overburden removal, 5) loading the deposit, 6) haulage of the
mined deposit, and 7) reclamation. Mining techniques for a particular
region are largely dictated by geologic and topographic conditions. Even
where the techniques are generally comparable, economics of alternative
equipment choices and utilization are not easy to generalize. The surface
mining techniques can be broadly classified into: 1) contour mining, 2)
area mining, 3) open-pit mining, 4) quarry mining, and 5) auger mining.
Other methods include dredging and hydraulic mining. Auger mining is usually
associated with contour mining where coal near the final highwall is recov-
ered by boring into the seam to a depth of 60.96 meter (200 ft), with
cutting heads up to 2.13 meter (seven ft) in diameter. Since quarrying,
dredging, and hydraulic mining are not employed for surface mining of coal
in the east, they will not be reviewed.
Contour mining is commonly practiced where the coal seams occur in
rolling or mountainous terrain. Basically, this method consists of re-
moving the overburden above the coal seam by starting at the outcrop and
proceeding along the hillside. After the exposed seam is loaded out,
additional cuts into the hillside are made until the economic limit for
surface mining is reached.
Contour mining, as practiced in the past, created a bench on the side of
the hill (Figure 19). On the inside, it was bordered by the highwall while
a high ridge of spoil, the slope of which may cause severe erosion and land-
slides, was on the outside of the bench. It was not uncommon to punch into
the coal seam from the highwall and extract the seam by underground methods
(punch mining). It was also common, where there was not enough coal for
punch mining, to recover the coal by auger mining (Figure 20). This method
of contour mining had come under great criticism. Demands for grading the
mined—land to original contour and for the burial of the exposed highwall
has made it necessary that the spoil be hauled back into the mined-out
pit. Therefore, coal companies are experimenting with several extraction
and reclamation techniques that aim at controlling adverse environmental
impacts and increasing land values after mining. Several new methods with
innovative "spoiling" sequence have been proposed and are being practiced
in the steep slopes of eastern United States. In many instances, haulback
49
-------
Ul
o
NO DIVERSION
DITCH
TOXIC MATERIAL,
BRUSH & TREES IN FILL SECTION
O BARRIER
Note: Downslope is
not scalped
TOE
OF FILL
Figure 19. Past practice of conventional contour-mining which is
no longer an allowable method.
-------
Ui
Figure 20. Auger used to drill laterally into a coal bed exposed in contour mining.
-------
methods have replaced conventional overcasting methods. These methods
provide the flexibility to separate top soil, segregate toxic materials,
and generally facilitate placement of spoil in a desired sequence. Two
such methods are the block-cut and mountain top removal methods.
Area mining, applied where the terrain is flat, commences with a trench
or "box cut" made through the overburden, which exposes a portion of the
coal seam (Figure 21). This first cut may be extended to the limits of the
property in the strike direction. As each succeeding parallel cut is made,
the overburden is deposited in the cut previously excavated.
Simple overcasting, explained in Figure 21, is still the most common
form of stripping. Shovels and draglines continue to be popular, with
draglines increasingly favored over shovels. However, the requirements of
topsoil removal, storage, and respreading on the top of graded spoil, and
the requirements of textured material in the graded spoil for safe operation
of farm machinery have dictated the choice of scrapers or other mobile
machinery to handle the upper layers of the soil in a haulback mode of
operation.
In open-pit mining of coal, the amount of overburden removed is small
when compared with the amount of coal recovered (Figure 22). Thick western
coals under shallow cover are often removed by open-pit mining. Open-pit
mining may also be used for steeply dipping seams (Figure 23).
Geology of the deposit is a big factor in the selection of the method.
All non-stratified deposits and steeply pitching and thick stratified
deposits require an outside dump or around the pit transportation of waste.
The general method can be classified as open-pit mining. Area mining is
applicable to gently pitching, thin deposits. This influence of the geology
and the topography are illustrated in Figure 24.
UNIT OPERATIONS
As previously mentioned, surface mining consists of a sequence of
seven operations. A discussion of these unit operations follows. This
section will also mention the types of equipment that can be employed.
Preparing the surface includes the removal of all vegetative cover in
preparation for other mining operations. While this operation is not always
necessary, it may be necessary to clear the land of trees or other obstruc-
tions. Sometimes, this operation may produce commercial quality timber or
may be required for relocation purposes.
It is not easy to define what is topsoil though regulations in several
states and federal statutes require that topsoil be removed and stored for
later use. Topsoil is generally understood to be the soft layers of soil
over which the current vegetation has established its roots. Since this
material is soft, elevating scrapers, bulldozers, or front-end loaders
52
-------
:Vr ^" ••- •"*•"'.,;;'"> .».v-""A-«M
RECLAIMED AREA ',.">.
^. y JNt. r. •. > /n jj* * t^«-'r •* •'Tt'i*"
£^!>>;^S
W^w^^^wa
ORIGINAL SURFAC
COAL BlD
STRIPPING BENCH
Figure 21. Typical area mining method with stripping shovel (Grim and Hill, 197A).
-------
Ui
JS
OVERBURDEN
Figure 22. Typical
open-pit ndning method (thick seam) (Skelly and Loy, 1975)
-------
Ln
Figure 23. Typical open-pit mining method (dipping seams) (Skelly and Loy, 1975)
-------
mineral deposit accessible for surface mining*
1
stratified
1
th
overb
back
by tr<
aroun<
1
Horizontal
I
1
ck thin
jrden overburden
1
1 1
thick thi
seam sea
1
filling backfill!
ansport by dire<
d the pit castinc
n
m
r>9
:t
}
inclined <$tf
1 1
gentle
inclination; steep narrow
greater than
nnnle nf r»pos* ._. ,.,
outside
dump
outside
dump
1
non-stratified
1
1
massive
(stock or pipe)
1
' >^- wide circular
or irregular
1
outside
dump
gentle inclination
thin seam
thick burden/or thin burden
^Surface topography should also
be considered for each deposit.
backfilling by direct casting
Figure 24. Influence of geological factors in surface mine design (Thomas, 1973).
-------
are used for its removal. According to the Permanent Regulatory Program under
the Surface Mining Control and Reclamation Act of 1977 (Title 30, Chapter VII,
Subchapter A) :
Soil horizons are contrasting layers of soil lying one below
the other, parallel or nearly parallel to the land surface. Soil
horizons are differentiated on the basis of field characteristics
and laboratory data. The three major soil horizons are:
1. A horizon. The uppermost layer in the soil profile often
called the surface soil. It is the part of the soil in
which organic matter is most abundant, and where leaching
of soluble or suspended particles is the greatest.
2. B horizon. The layer immediately beneath the A horizon
and often called the subsoil. This middle layer commonly
contains more clay, iron, or aluminum than the A or C
horizons.
3. C horizon. The deepest layer of the soil profile. It
consists of loose material or weathered rock that is
relatively unaffected by biologic activity.
To allow easier handling by stripping equipment later, the unit opera-
tions of drilling and blasting to fracture rock masses may or may not be
required. Where the ground cover is hard, it is usually drilled and blasted.
Soft strata may often be directly excavated. Some strata can also be suf-
ficiently prepared through use of ripper-bulldozers.
Overburden removal is the most important aspect of the mining system.
The equipment and methods used to remove overburden must be carefully chosen
to provide the required production at the minimum cost. The stability of the
highwall and the spoil is an important consideration in pit design. The
equipment used varies from fleets of mobile equipment such as end-loaders to
giant draglines, shovels, and bucket-wheel excavators. A factor that has
become important is the ability to segregate the soil into stratigraphic
layers at desired locations for reclamation and environmental control
purposes.
Coal loading is usually done by a loading shovel. Other methods in-
clude use of end-loaders, hydraulic excavators, or fine graders. Coal
transport is typically by bottom dump trucks.
Reclamation includes backfilling,regrading, surface stabilization,
revegetation and restoration operations.
jackfilling is achieved through use of virtually any of the equipment
used in overburden handling. However, approaches to segregation of material,
specific burial, and layering and compacting can be distinguished as distinct
backfilling steps.
Regrading is typically achieved by wheeled or tracked dozers. The amount
of regrading depends upon the care with which backfill is done, and on the de-
gree of topsoil restoration and grading, which is either desired or practicable.
57
-------
Considerations of surface stabilization include provision for water
quality maintenance, compaction and layering, and sealing the spoil layers.
It also encompasses amendments and therapeutic actions which revitalize the
surface or prepare it for revegetation, such as mulching, fertilizing, liming,
stabilizing, disking, harrowing, etc. One phase is usually the spreading of
topsoil on the regraded soil. Scrapers are often used to reclaim soil from
the stockpile, transport it, and spread it on the spoil. End-loaders may be
used with bottom-dump trucks which spread the topsoil "on the run." Seeding
is actually the final stage of rehabilitation.
Revegetation and restoration takes place over a longer time period.
The growth of plants and establishment of cover is an involved process.
Careful attention over several growing seasons may be required. Of course,
this step is not necessary where alternative land use, such as a shopping
center or a school, is envisioned.
Tables 11 through 17 summarize the unit operations, types of applicable
equipment, and ranges of operation, and Figure 25 provides some combinations
of equipment that can be employed in surface mining systems for these unit
operations.
AREA MINING
Area mining is practiced in the eastern and interior provinces as local
conditions permit. It is commonly practiced in Ohio, western Kentucky,
Illinois, Kansas, and Oklahoma. In this method, a trench, usually called
box cut, is made through the overburden to expose the coal seam. Overburden
material (or spoil) thus removed is placed on adjacent unmined land, stabi-
lized and contoured. As the coal is removed from the first cut, a second cut
is started immediately adjacent and parallel to the first cut. Overburden
removed from this cut is placed in that portion of the first cut from which
the coal has been removed. Cut by cut, the mining process proceeds across
the property. The final cut leaves an open trench, bordered on one side by
the spoil and on the other side by the unbroken column of overburden or high-
wall. If restoration were not affected, the spoil banks from stripping would
resemble a giant washboard or the ridges of a plowed field. The preferable
practice in most cases, includes grading the spoil to a level terrain and
planting grasses and/or trees.
The practice, depicted in Figure 21, is known as simple overcasting.
Simple overcasting, so named because the overburden is cast over the pit to
the spoil bank, is the most common form of overburden stripping. The suit-
ability for reclamation, efficiency, and capability of equipment operation,
and other considerations, have resulted in the development of alternatives
to simple overcasting—passover, crossover, or haul-around techniques. Area
mining incorporates a wide range of earth- and rock-moving equipment. The
range of stripping applications is presented in Table 18. Table 19 presents
some of the combinations which can be practiced in a two-part (multiple-seam)
stripping operation. If a three-part stripping operation were to be under-
taken, then the third lift would be taken by any method feasible under the
second lift.
58
-------
TABLE 11. EQUIPMENT RATING - TOPSOIL REMOVAL AND REPLACEMENT
(SKELLY AND LOY 1975)
LEGEND
1. Should be considered
2. May be considered
3. May be considered under
certain conditions
4. May be considered special
situation
A. High
B. Moderate
C. Low
Topsoil Thickness
Haul Distance
Flexibility Under
Varied Field Conditions
0 -2'
2'-5'
0 -300'
300'-500'
500 '-1000'
1000 '-1500'
1500'-5000'
good
fair
poor
CO
CD
N
O
a
1
1
1
2
-
A
A
B
co
p
0)
CO
0
T3
C
1
C
0
1
1
1
2
-
-
A
A
B
en
cu
a.
CO
o
oc
C
•H
J-l
C3
;>
CU
td
1
1
2
1
1
1
2
A
A
B
^
1
o
i
iH
3-
1
1
2
1
1
1
2
A
A
B
5_,
0
4J
0
S-J
H
j.
CO
3
c-
•fj
-H
1
1
1
1
1
1
2
A
A
B
cfl
O
X
DO
i— 1
0)
CU
r"
5
4J
CD
O
3
-
3
-
B
B
C
o
3
u
H
M3
-------
TABLE 12. EQUIPMENT RATING OVERBURDEN PREPARATION
(SKELLY AND LOY - 1975)
LEGEND
1. Should be considered
2. May be considered
3. May be considered under
certain conditions
4. May be considered special
situation
A. High
B. Moderate
C. Low
Composition of Overburden
Depth of Overburden
Mobility
Flexibility Under Varied
Field Conditions
Hardness of
Rock
Very Consolidated
Moderately
Consolidated
Slightly Consolidated
Unconsolidated
0 -30'
30'-60'
60'-90'
>90f
Good
Fair
Poor
Very Hard
Moderately Hard
Soft
Tractor-Ripper
3
1
1
1
1
2
3
3
A
A
A
A
3
2
1
Drills
Percussion
1
1
2
4
1
1
1
1
A
A
A
A
1
1
1
x
}-i
cfl
4-J
O
Pi
1
1
2
4
1
1
1
1
A
A
A
A
1
1
1
Jet Piercing
2
1
2
4
1
1
1
2
B
A
A
B
2
1
1
1 ft. = 0.3048 meter
60
-------
TABLE 13. EQUIPMENT RATING OVERBURDEN
REMOVAL (SKELLY AND LOY - 1975)
LEGEND
1. Should be considered
2. May be considered
3. May be considered under
certain conditions
4. May be considered special
situation
A. High
B. Moderate
C. Low
Thickness
Characteristics
Transport Distance
Coal Seam Support
Characteristics
Segregation
Capability
Production
Capability
Flexibility Under
Varied Field
Conditions
F Mobility
0-30'
30'-60'
60'-100'
>100'
Poor Fragmentation
(Blocky-Large Breakout Force)
Moderately Blocky
Good Fragmentation
(Low Breakout Force)
Unconsolidated
SO'-ISO1
150'-300'
300'-500'
500 '-1000'
> 1000'
Good (Hard)
Moderate
Poor (Soft)
-
Good
Fair
Poor
-
[Dragline
2
1
1
1
3
2
1
1
1
1
2
-
-
1
1
1
A
A
A
A
A
B
(Shovel
1
1
1
2
1
1
1
1
1
-
-
-
-
1
2
4
C
A
A
B
C
B
|Shovel & Truck Comb.
1
1
2
3
1
1
1
1
-
2
1
1
1
1
2
4
A
B
A
B
C
B
Front-End Loaders
1
2
3
4
3
2
1
1
1
1
2
-
-
1
1
2
A
A
A
B
C
A
[Dozers
1
2
3
4
1
1
1
1
1
1
2
-
-
1
1
2
B
A
A
A
B
A
Front-End Loader &
Truck Comb.
1
2
3
4
3
2
1
1
-
-
3
1
1
1
1
2
A
B
A
A
B
A
[Bucket Wheel
[Excavator
1
1
2
2
-
3
1
2
1
1
-
-
1
2
4
A
A
B
B
C
C
Scrapers
[Elevating
1
1
2
3
2
1
1
-
-
3
1
1
1
1
1
A
A
A
A
B
A
i->
0)
0
PL,
i—l
rH
ft,
1
1
2
3
2
1
1
-
-
3
1
1
1
1
1
A
A
A
A
B
A
With Push
Tractor
1
1
2
3
2
1
1
-
-
3
1
1
1
1
1
A
A
A
A
B
A
1 ft = 0.3048 meter
61
-------
TABLE 14. EQUIPMENT RATING COAL LOADING
(SKELLY AND LOY - 1975)
LEGEND
1. Should be considered
2. May be considered
3. May be considered under
certain conditions
4. May be considered special
situation
A. High
B. Moderate
C. Low
Coal Seam
Thickness
Fragmentation
or Ease of
Digging
Floor
Conditions
Mobility
Flexibility
Under Varied
Field Conditions
Production
Requirements
l'-3'
3'-5T
5'-10'
10 '-25'
>25'
Soft or Highly Frag.
Moderately Frag.
Hard or Low Frag.
Very Soft
Moderate
Hard
-
Good
Fair
Poor
High
Medium
Low
Shovels
2
1
1
1
1
1
1
1
4
1
1
B
A
B
C
1
1
1
(Rubber Tired)
Front-End Loaders
1
1
1
2
3
1
1
3
1
1
1
A
A
B
B
1
1
1
High Lift (Tracks)
1
1
1
2
3
1
1
2
1
1
1
A
A
B
B
3
1
1
Backhoe
2
3
-
-
-
2
3
-
2
2
2
B
A
A
B
3
2
1
w
M
-------
TABLE 15. EQUIPMENT RATING - REGRADING AND BACKFILLING
(SKELLY AND LOY 1975)
LEGEND
1. Should be considered
2. May be considered
3. May be considered under
certain conditions
4. May be considered special
situation
A. High
B. Moderate
C. Low
Spoil Configuration
% Spoil Rehandled
Transport Distance
Final Surface Contour
High Peaks
Moderate Peaks
Low Peaks
>75%
75% 50%
50% 25%
<25%
50'-150'
150'-300'
300'-500'
500'-1000'
>1000'
Flat & Smooth
Flat & Rough
Steep & Smooth
Steep & Rough
co
CU
N
O
Q
1
1
1
1
1
1
1
1
1
2
1
1
1
1
CO
CU
TD
tv)
0
3
1
-
_
2
2
-
-
-
1
2
-
co
0)
a,
o
C/3
2
1
1
1
1
1
1
1
1
1
1
1
1
1
-
CO
C
•H
00
cfl
O
2
2
2
2
2
3
-
2
2
-
-
2
2
CU
-o
o
-o
C
1
4-)
C
o
M
3
1
1
2
1
1
1
1
1
2
-
3
2
3
2
to
o
3
H
"°
0)
'O
O
T-,
C
la
1
C
o
fa
3
2
1
1
1
1
1
-
4
1
1
-
63
-------
TABLE 16. EQUIPMENT RATING - REVEGETATION
(SKELLY AND LOY 1975)
LEGEND
1. Should be considered
2. May be considered
3. May be considered under
certain conditions
4. May be considered special
situation
A. High
B. Moderate
C. Low
Acreage
Seed Mixture
Mulching
Fertilizer & Lime
<3
3-10
10-25
25-50
>50
Grasses
Crops
Seedlings
^
pr;
1
1
1
1
1
1
3
-
1
1
w
CO
U
ID
cfl
O
CQ
1
1
2
2
3
1
2
3
00
C
•H
-------
TABLE 17. EQUIPMENT RATING - COAL HAULING
(SKELLY AND LOY 1975)
LEGEND
1. Should be considered
2. May be considered
S.May be considered A. High
under certain B. Moderate
conditions C.Low
A.May be considered
special situation
Material
Length of Haul
Condition of
Road
Type of
Haulage Roads
Daily
Production Rate
Flexibility
Under Varied
Conditions
Maximum
Adverse
Grade
Rough-Blocky
35 inch size
24 inch size
Fines
300'-500'
500'-1000'
1000 '-1500'
1500 '-5000'
5000 '-10, 000'
10,000'-15,000?
>15.000'
Good
Wet. Soft
Mine Roads Only
Combination:
Mine & Public
High
Medium
Low
Good
Fair
Poor
+3%
+5%
+10%
+15%
+20%
>+20%
Trucks
Off Road
D.
e
Q
T5
W
1
1
1
1
1
1
1
1
1
1
1
1
1
1
2
1
1
1
A
A
A
1
1
2
3
-
-
o o,
4J 6
4J 3
0 Q
PS
-
-
2
1
2
1
1
1
1
1
1
1
1
1
2
1
1
1
A
A
A
1
2
3
-
-
-
Trailer
End Dump
1
1
1
1
2
1
1
1
1
1
1
1
1
1
2
1
1
1
A
A
A
1
2
3
-
-
-
Highway
a
D
O
-o
c
w
1
1
1
1
2
1
1
1
1
1
1
1
2
2
1
1
1
1
A
A
R
1
1
2
3
-
-
^ a
Q) g
•H D
-H Q
cd
t-i TD
H C
W
1
1
1
1
2
1
1
1
1
1
1
1
2
2
1
1
1
1
A
A
R
1
2
3
-
-
-
Scrapers
>j
0)
o
(X
fH
r-l
3
fn
-
-
1
1
2
1
1
1
3
-
—
1
3
1
4
1
1
1
A
A
R
1
1
2
2
3
4
Elevating
-
-
1
1
2
1
1
1
3
-
—
1
3
1
4
1
1
1
A
A
R
1
2
2
3
4
-
Under Powered
w/Tractor
-
-
1
1
2
1
1
1
3
-
_
1
3
1
4
1
1
1
A
A
R
1
2
2
3
4
-
Conveyor
-
2
1
1
4
4
4
3
2
1
1
—
—
-
-
1
2
—
B
B
r,
1
1
1
1
2
3
Train
1
1
1
1
-
-
-
-
•>
L.
1
1
-
-
-
-
1
2
—
C
C
r
i
2
-
-
-
-
Pipeline
-
-
-
4
4
4
4
4
3
3
2
-
-
4
4
_
—
-
_
1
1
2
2
3
3
65
-------
surface mining systems
unit operations
t
ground
preparation
excavation
transport
dumping
reclamation
continous
\
none
I
multi-bucket machines
bucket-wheel
bucket-chain
non-continous
\
blasting, ripping or none
I
single-bucket machines
I
I
crowd shovel
shovel loader
surge-hopper .
I crusher feeder
**X ^__
1 I
dragline |
drag scraper bowl scraper
I I
casting scraping bowl
belt conveyor
I
stacker
bull dozer
railway truck
dump bins dump bins
or areas or areas
I
rail ploughs
bull dozers
scrapers
Figure 25. Equipment application for surface mining systems (Atkinson, 1971).
-------
TABLE 18. SINGLE-PHASE STRIPPING POSSIBILITIES
Number
(1)
(2)
(3)
(4)
(5)
(6)
(7)
(8)
(9)
(10)
(11)
(12)
Approach General
Name Method
Dragline Overcasting Overcasting
Shovel Overcasting
Bucketwheel Stripping Passover
Loader-Conveyor
Stripping*
Shovel-Conveyor
Stripping*
Bulldozer Stripping Crossover
Scraper Stripping
Loader Stripping
Dragline-truck Haul-Around
Stripping
Shovel- truck
Stripping
Loader- truck
Stripping
Dredging*
Equipment Used
Excavation Transport
Dragline
Stripping
Shovel
Bucket-wheel
Excavator
Front-end
Loader
Loading
Shovel
Bulldozer
Scraper
Front-end
Loader
Dragline
Loading
Shovel
Front -end
Loader
Dredge
Dragline
Stripping
Shovel
Belt Conveyor
(Stacker)
Belt Conveyor
Belt Conveyor
Bulldozer
Scraper
Truck
Truck
Truck
Dredge
Pipeline
These methods are not commonly practiced in coal in the Eastern
United States.
67
-------
TABLE 19. MULTIPLE-SEAM STRIPPING POSSIBILITIES*
(RAMANI, 1968)
Upper Seam Method Feasible Lower Seam Method
1 1, 2, 6, 7, 8, 9, 10, 11
2 1, 2, 6, 7, 8, 9, 10, 11
3 1, 2, 6, 7, 8, 9, 10, 11
4 1, 2, 6, 7, 8, 9, 10, 11
5 1, 2, 6, 7, 8, 9, 10, 11
6 6, 7, 8, 9, 10, 11
7 6, 7, 8, 9, 10, 11
8 6, 7, 8, 9, 10, 11
9 6, 7, 8, 9, 10, 11
10 6, 7, 8, 9, 10, 11
11 6, 7, 8, 9, 10, 11
12 1, 2, 6, 7, 8, 9, 10, 11
* Refer to Table 18 for the description of the
numbered method.
68
-------
Dragline Overcasting
Figure 26 is a photograph of a 34 cu m (45 cu yd) dragline in western
Pennsylvania. Up to 26 m (85 ft) of overburden are cast across an 18 m
(60 ft) pit to exploit an 0.5 m (1.67 ft) coal seam. Figure 27 is a line
diagram of the largest dragline operation. A 168 cu m (220 cu yd) machine
removes 36.5 to 40 m (120 to 130 ft) of overburden to exploit a 1.2 m (4 ft)
coal seam in a pit 76 m (250 ft) wide. Small dragline operations are common
in the hilly terrain of Pennsylvania. Figure 28 shows a large dragline pit
with other surface mining equipment.
Shovel Overcasting
Shovel overcasting is similar to dragline overcasting but incorporates
a stripping shovel instead of a dragline. Figure 29 shows the second largest
stripping shovel working in 45 m (150 ft) of overburden in Ohio. Figure 30
shows details of a shovel pit.
Bucket-Wheel Stripping
Stripping with a bucket-wheel excavator (BWE) alone is not practiced in
the Eastern United States. The BWE can effectively handle soft material.
Its performance is poor in hard material or strata that has to be fragmented.
The applications common in the study area of this report incorporate a two-
phase stripping system, in which the upper, unconsolidated material is
removed by the BWE, and the lower, hard strata which must be fragmented is
handled by either a stripping shovel or a dragline.
BWE~Dragline Stripping
Figure 31 is a line diagram, and Figure 32 is a photograph of an opera-
tion of this type in Illinois. Both machines work at the same bench eleva-
tion, 0 to 20 m (0 to 65 ft) below the surface. Up to 37.0 m (120 ft) of
overburden are removed, 20.0 m (65 ft) by the BWE and 17.0 m (55 ft) by the
dragline. The pit width is 28.0 m (90 ft), and the coal seam is 0.91 m
(3 ft) thick.
BWE-Shovel Stripping
Figure 33 is a line diagram and Figure 34 is a photograph of an opera-
tion of this type in southern Illinois. The BWE cuts its own bench from 0 to
16.8 m (0 to 54 ft) below the surface. The shovel operates in the pit at a
depth of 14.0 m (46 ft) below the BWE bench. Pit width is 33.5 m (110 ft)
and the coal seam is 1.37 m (4 to 5 ft) thick.
Smaller Scale Approaches
The preceeding discussions have centered on the largest scale mines,
typical of Illinois, Indiana, and Ohio. In the east, area mines are usually
of such limited dimensions that large-scale and fixed equipment are not used.
A variety of approaches are taken in smaller-scale operations, approaches
which incorporate virtually every combination of equipment and methods.
69
-------
Figure 26. A 45-cubic-yard dragline at work in a Pennsylvania coal operation.
-------
Plan View
Surface
120-130'
u42-47" h
175'
*•+••-
Section View
175'
Figure 27. Pit layout at Bucyrus-Erie 4250-W operation
(Stefanko, Ramani, and Ferko, 1973).
71
-------
Figure 28. A 90-cubic-meter dragline with a 90-meter boom. Note the
coal-loading shovel and truck in the pit (Ramani and Grim, 1978)
-------
Surface
n n
] Shovel
Exposed
Sewickly No. 9
Coal
I
I
I
u u
Coal
Loading
Cut
Plan View
45" U- 60'-»|
120'
Section View
Figure 29. Plan and section views of a Bucyrus-Erie 1950-B pit
(Stefanko, Ramani and Ferko, 1973).
73
-------
Figure 30. Details of a shovel pit.
74
-------
\
\
Bucket Wheel
Excavator
Cut
Dragline
Spoil
Section View
Figure 31. Plan and section views of a dragline and bucket-wheel
excavator in tandem operation (Stefanko, Ramani, and
Ferko, 1973).
75
-------
CTX
Figure 32. Bucket-wheel excavator, dragline, coal-loading shovel, trucks, and a
dozer ripper are employed in this operation (Ramani and Grim, 1978).
-------
Stripping Shovel
Wheel Excavator
Coal Shovel
130'
Figure 33. Pit layout at a shovel and bucket-wheel excavator tandem
operation in Illinois (Ramani, 1968).
77
-------
. t.
00
Figure 34. An operation in Illinois using bucket-wheel excavator, stripping
shovel, loading shovel and trucks (Ramani and Grim, 1976).
-------
CONTOUR MINING
Contour mining is practiced where an outcrop of coal occurs on a hill-
side. The mining may progress into the hillside only to a certain point,
after which there is too much overburden on top of the coal for economic
removal. The mining may proceed around the hillside, however, at the given
elevation. Successive cuts are removed along the contour line, hence the
name contour mining. There are several methods of contour mining such as
outslope disposal, outslope reduction, box cut, block cut, haulback or step
blockcut, modified area, mountaintop removal, and head-of-hollow fill methods.
For all practical purposes outslope disposal of spoil is no longer prevalent
or allowed. The outslope disposal method is sometimes referred to as "shoot
'n shove" mining. As shown in Figure 19, the approach is to fragment the
overburden by blasting from the outcrop into the stripping limit, and simply
to shove the material over the edge. Outslope disposal contour mining
(Figure 19) was the traditional approach but is legally limited today, due
to its degrading effect on the environment of the mining area. Soil sta-
bility, erosion, sedimentation, water pollution, etc., are some of the
problems that are enhanced-by this method.
Outslope reduction methods are based on the theory that the stability
of the outslope spoil bench could be enhanced if the slope of that bench
were reduced and the spoil spread over a larger area. Grim and Hill (1974)
discuss the various outslope reduction methods.
Outslope reduction was accepted in the recent past as an approach to
mining in mountainous terrain and was often practiced as a means of re-
claiming abandoned contour strip mines. It can still be used to reduce the
slope of existing spoil pile or refuse bank areas. However, federal legis-
lation and some state regulations have made clear the undesirability of any
spoil disposal downslope from the outcrop or beyond the solid edge of the
bench on slopes over 20 degrees. Therefore, these methods will not be
discussed here. The interested reader is referred to Grim and Hill (1974)
for detailed discussion of many of these methods. On the other hand,
extended presentation of methods that have been developed to decrease or
eliminate the outslope disposal of overburden are presented.
The Head-of-Hollow Fill Technique
Head-of-hollow and valley fills basically consist of storing the over-
burden material in narrow, steep-sided valleys according to a well engineered
plan. Due to the swelling of material when broken, careful analysis of the
cut-and-fill problem associated with the method is required. This method
represents a controllable alternative to outslope disposal. Although the
disposal is technically downslope and outside the solid bench, a specific,
controlled situation is developed for such disposal.
While a wide range of valleys or hollows is acceptable for head-of-
hollow fill purposes, v-shaped, narrow, steep-sided valleys located near the
tops of ridges and free from connections to subsurface drainage (e.g., old
underground mine openings, seeps, and springs) are preferred. The size of
the valley selected for filling is a function of the amount of spoil which
79
-------
will be generated in the mining operation. The objective is to develop a
valley fill which completely fills the prepared valley, all the way to the
head.
While there are many references regarding the construction of a valley
fill, they are either general in nature or specific in certain aspects re-
lating explicitly to requirements of the state of operation. Therefore,
dimensions and specifications for valley fill construction that are provided
here should be interpreted only as generally indicative of current practices.
Several important procedures are followed in adapting the head-of-hollow fill
and mountaintop removal methods (to be discussed later in this section) to a
particular area. Figure 35 presents a drawing of the head-of-hollow fill
method. The head-of-hollow selected for filling is scalped of vegetative
cover. Topsoil is removed and stored for later use. Check dams or silt
control structures are built downstream from the hollow fill (Figure 36).
The natural drainway in the hollow is deepened. The minimum width recom-
mended for the drain is 4.56 meters. A French drain is built in this
drainway with rocks no less than 0.36 meters in size in any one dimension
(Figures 37 and 38). The fill is built in layers, 1.2 meters thick, be-
ginning at the toe of the valley (Figure 39). The overburden material is
deposited in uniform horizontal layers and then compacted with haulage
equipment. Each layer of fill is built so that the face of the fill has an
outer slope no steeper than two horizontal to one vertical, and usually has
crowned terraces (Figure 40). Some operators have graded the face to ap-
proximately 22 degrees from the horizontal, eliminating crowned terraces.
However, long slopes without interruptions, such as diversion ditches, are
more prone to excessive erosion from surface runoff; therefore, where the
face of the fill is not stepped, diversion ditches at a minimum of every 15
meters vertical height of the fill is recommended. Revegetation of the
hollow-fill face, usually with hydroseeding, is done as soon as the fill
height increases.
Block Cut
In the block-cut method, the general progress of mining is along a
contour. However, mining is accomplished by taking successive blocks along
the contour which may be from one to several cuts deep. Figure 41 shows
how these successive block cuts are carried along the contour. The ad-
vantages of this method are that it allows overburden spoiling in the pit and
on the bench and that each cut is worked to the ultimate limit of the
stripping ratio once it is opened. In this second consideration, the small
operator often finds the previously mentioned, against-the-contour advance
of each block cut easier to handle on an economic basis.
Several extensions of block-cut mining are practiced. They have a
number of different names, such as modified block cut, haulback, pit storage,
and put-and-take. Figure 42 is a line drawing of the sequencing of cuts, and
Figures 43 and 44 explain the stripping and backfilling phases.
Basic block-cut mining is frequently confused with these extensions.
In this method, the width of each block is the same. The depth each block
is carried into the hillside may vary, however, as each block is carried to
80
-------
oo
PROCEDURE:
l.SCALP ENTIRE AREA THAT WILL BE COVERED WITH FILL. REMOVE AND STORE TOPSOIL
2.CONSTRUCT FRENCH DRAINS IN THE HOLLOW WATER COURSES.
3.BUILD THE FILL IN COMPACTED LAYERS.
FACE OF FILL NO STEEPER THAN 2:1.
4.CONSTRUCT CROWNED TERRACES EVERY 20 FEET,
APPROXIMATELY 20 FEET WIDE.
5.CENTER OF COMPLETED FILL BENCH IS CROWNED
TOWARD THE HIGHWALL. SO THAT WATER
WILL FLOW ONTO EXCAVATED BENCHES.
6.BUILD SILT CONTROL STRUCTURES BELOW HOLLOW FILL.
MOUNTAIN TOP
BEING REMOVED
TOPSOIL
CROWNED FILL BENCH
CROWNED
TERRACES
LATERAL
DRAIN
ROCK FILLED ,(FRENCH DRAIN),
NATURAL DRAIN WAY
Figure 35. Head-of-hollow fill (Grim and Hill, 1974).
-------
• i&'^'M " tw
5 p,~*
Figure 36. A check dam, downstream from a hollow fill.
82
-------
Figure 37. A distant view of the French drain.
83
-------
00
-p-
Figure 38. The size and gradation of the rock that form the core
of the French drain is an important design consideration.
-------
Figure 39. The valley fill is built from the bottom up in layers
85
-------
oo
ON
Figure 40. A completed head-of-hollow fill,
-------
PLAN VIEW
Hill
ill\
Valley
Cut
.Valley
Cut 2
Cut
Cut
D
Spoil used
to backfill
Outcrop
barrier
B
Hill
ut4
Cut ',
Cut 3
Valley
Cut
DETAIL
Figure 41. Contour mining around a hillside with the block system
(Saperstein and Secor, 1973; Maneval, 1972).
87
-------
TOP OF RIDGE
HIGHWAU-
CUT 7
CUT 5
•*- —
CUT 3
-^ —
CUT 1
-^- -^-
CUT 2
— -^E-
CUT 4
— -^
CUT 6
-^^M
OUTCROP BARRIER-
HOLLOW
PROCEDURE:
1 SCALP FROM TOP OF HIGHWALL TO OUTCROP BARRIER,
REMOVE AND STORE TOPSOIL.
2.REMOVE AND DISPOSE OF OVERBURDEN FROM CUT 1
3 PICK UP COAL, LEAVING AT LEAST A 15 FOOT UNDISTURBED
OUTCROP BARRIER
4 MAKE SUCCESIVE CUTS AS NUMBERED.
5 OVERBURDEN IS MOVED IN THE DIRECTION, AS SHOWN BY
ARROWS, AND PLACED IN THE ADJACENT PIT.
6 COMPLETE BACKFILL AND GRADING TO THE APPROXIMATE
ORIGINAL CONTOUR
Figure 42. Block-cut method (Grim and Hill, 1974).
88
-------
RIDGE TOP
Figure A3. Block-cut method: stripping phase (Grim and Hill, 1974)
89
-------
RIDGE TOP
JOMPACTED
CLAY
BARRIER
Figure 44. Block-cut method: backfilling phase (Grim and Hill, 1974)
90
-------
its maximum allowable stripping ratio. One advantage is that the dragline
can work, for example, cuts 3 and 4 from one location. Another is that the
dragline can work uniformly along the highwall with short move time. Aside
from the initial cut, the spoil is placed in a previous excavation, which
limits rehandling and is environmentally preferable.
The modified block cut and haulback extensions capitalize on the pit-
spoiling aspects of block cut. In these, the first cut is sized larger than
the others to assure that the volumes of broken and swollen overburden from
the remaining cuts will be exactly equal to the mined-out volume. Therefore,
once the spoil of the first cut is disposed of in a valley fill or other
suitable manner, the remaining spoil can be stored in the pit. The dif-
ference in modified block cut and haulback is that modified block cut pro-
ceeds in one direction along the contour and may utilize any of several
techniques for overburden handling, while haulback proceeds in both direc-
tions from the initial cut and utilizes load and carry or truck transport of
overburden. The advantage of the haulback refinement of pivoting about a
central location seems to be proven through actual practice in both methods.
Haulback is unique, as it allows continuous overburden removal, coal removal,
augering and reclamation in one pit.
Block-cut methods have been accepted by industry and government alike.
Saperstein and Secor (1973, 1977) found that the block-cut methods can be
cost-competitive with any other contour mining system.
One key to success in the block-cut methods is the determination of the
ultimate stripping ratio. Given this ultimate ratio, the block depths must
then be established. For modified block cut to haulback, extensive pre-
mining planning is necessary as additional considerations are brought in to
assure that the proper allowances for swell have been developed and the first
cut and valley fill have been properly sized. A rigorous analysis of the
overburden handling unit operation is required. Many combinations of exca-
vation and transport equipment are possible. The sizing and logistical
configuration of many alternatives must be considered if an optimum is to be
found. Grim and Hill (1974) list the following advantages and disadvantages
of the block-cut method:
Advantages
1. Since the method is not limited by the slope of the area, currently
available equipment can be used. Further, the method can be
used in multiple-seam applications where unit operations must be
closely concentrated.
2. Less coverage is disturbed than in other mining methods. The
mined-out areas are completely backfilled, eliminating downslope
spoil disposal and the final highwall.
3, Reclamation costs are lower, and revegetation costs are reduced.
Also, overburden segregation is enhanced.
4. The problems associated with water quality control (acid mine
drainage, siltation, and sedimentation), are reduced, since toxic
or material which is difficult to handle can be selectively placed
91
-------
in isolation as the block cuts are filled. Fewer sediment control
structures are required and the size of the disturbed-area drainage
system is smaller.
Disadvantages
1. Extensive premining planning and scheduling is required. The
location and size of the initial box cut must be precisely deter-
mined. Further, investment costs for overburden transport equipment
are high, discouraging application in small operations.
2. The drilling and blasting processes are complicated and time-
consuming even though they are standard approaches. Overburden
must be controlled and fragmentation must be matched to the equip-
ment used in excavation and transport.
3. The actual long-term environmental consequences are not known or
documented.
4. Additional recovery techniques must be carried out concurrent
with initial mining. Grim and Hill specifically mention augering.
Any future recovery of coal from the seam by a surface approach
requires complete removal of the spoil that has been replaced.
The most attractive aspect of the block-cut technique is that the
recovery of the mineable resource, and the restoration and rehabilitation
of the surface are an integral part of the total mining process.
Mountaintop Mining
Mountaintop mining is an adaption of area mining methods to contour-
mining situations. The two approaches used are modified area mining and
mountaintop-removal mining. Where coal lies under ridges, mountaintops, and
knobs, there is usually a stripping-ratio limit to the depth into the hill-
side to which mining may proceed. Mountaintop methods are, however, the
exception in which the overall stripping ratio is taken as the guide rather
than local and ultimate maximum stripping ratios. It is based on the
analysis that area mining methods, even at a higher average stripping ratio,
allow the recovery of more coal (by removal of all the coal in a given tract)
at a lower total cost per ton, due to economies of scale in operation, than
is possible in contour mining. Figures 45 through 48 explain the sequence
of operations.
These methods are applied where a coal seam is present on opposite sides
of a hill or ridge to a great extent, and the overburden is sufficiently
thin (less than 50 m [165 ft]) across the coal, to allow area mining. The
method is applicable to previously contour-mined ridges and hills as well
as to coal seams in virgin areas. When restoration is to approximate
original contour, the best methods to follow are the dragline or shovel-
casting approaches of area mining, and the technique is modified area. When
the mountaintop is restored as essentially flat land, the best approaches
incorporate load-and-carry, bulldozer, or loader-truck techniques, and the
method is designated as mountaintop removal.
92
-------
MOUNTAIN TOP
FIRST CUT
(BOX CUT)
I I
HIGHWALL
BARRIER
BLOSSOM
Figure 45. Mountaintop removal method: first cut (box cut)
(Grim and Hill, 1974).
-------
SECOND CUT
ORIGINAL
GROUND
SLOPE
MOUNTAIN TOP
SECOND
CUT
HIGHWALL
SPOIL
BARRIER
BLOSSOM
Figure 46. Mountaintop removal method: second cut
(Grim and Hill, 1974).
94
-------
FOURTH CUT
MOUNTAIN TOP
BRUSH DAM
HOLLOW
BARRIER
DIVERSION
DITCH
Figure 47. Mountaintop removal method: fourth cut
(Grim and Hill, 1974).
95
-------
TOP SOIL
FLAT TO ROLLING LAND
BARRIER
BLOSSOM
DIVERSION DITCH
BARRIER
BLOSSOM
X
DIVERSION DITCH
Figure 48. Mountaintop removal method: mountaintop after final grading and
topsoiling (Grim and Hill, 1974).
-------
Modified area mining is practiced in areas such as western Pennsylvania
and western Kentucky. The topography and the requirements of restoration to
the original contour are two reasons for its practice. The spoil in modified
areas is placed in the mine excavation. The value of the land is taken as
better in original contour than in revised contour.
Mountaintop removal is practiced most extensively in West Virginia.
Mountaintop removal incorporates the use of valley fills to which a signif-
icant portion of the spoil is committed. For example, by this method, it is
possible to create from 50 hectares of coal land 75 to 100 hectares of "flat"
restored land surrounded by stable and controlled outslopes. In most cases,
such gently rolling land is worth much more in its state of revised contour
than in its previous sloping state. Near Welch, West Virginia, 30.A hectares
(75 acres) of mountaintop removal land were donated as a high school building
site, and the remaining land can be used for subdivision. In this area,
residential land is worth as much as $98,684 per hectare ($40,000 per acre).
The steps in the mountaintop removal method are essentially the same as
in valley fill and are as follows:
1. Determine the postmining volume of overburden. Given this volume,
determine the amount that will be spoiled in the mined-out area
and the amount that will be spoiled in valley fills.
2. Prepare the required valley fill areas (Figure 49).
3. Scalp (clear and grub) areas for temporary storage of topsoil.
4. Scalp the hilltop. Remove topsoil to temporary storage areas.
5. Initiate mining with a typical area-mining trench cut along one
edge of the property. Usually a 5 m (16.5 ft) barrier of blossom
coal is left along the outcrop edge.
6. Continue to mine relatively parallel to the first cut across the
mountaintop. In the modified area, the dragline or shovel spoils
most of the overburden in the previous pit to the approximate
original contour. In mountaintop removal, much of the spoil in
initial cuts is placed in valley fills by the mobile equipment
fleets; when these valleys are completely filled, spoiling is
carried out on the mountaintop.
7. Grade the mountaintop and fills to approximate original contour
(modified area) or to essentially flat, revised contour (mountain-
top removal).
8 Restore topsoil, amend and revegetate (Figure 50).
Mountaintop mining methods offer the same general advantages as the
block cut and other practices that do not include outslope-disposal. The
determination of the size and location of valley fills and the sequencing
of cutting and filling operations are crucial for success. The increased
value of the flat land created due to this method, may enable surface mining
of coal which heretofore were recoverable by either underground or auger-
mining techniques.
97
-------
Figure 49. Mountaintop removal method. Note the head-of-hollow in the
front being prepared for valley fill.
98
-------
Figure 50. Large area of level ground created as a result of
mountaintop removal mining in West Virginia.
-------
Open-Pit Mining
True application of open-pit mining techniques to the exploitation of
coal resources is limited to two specific occurrences. In one case, the coal
outcrops, but dips so steeply that simple overcasting cannot be easily
practiced, due to limitations on equipment reach. The value of the coal is
such that the overburden can be stripped in prismatic sections and the pit
can be advanced along the strike of the vein. In the other case, the coal
is very thick and is overlain by shallow or thick cover. Therefore, benches
are required in the coal and the overburden (Figure 51).
While open-pit mining of coal is practiced in the United States, these
are in places outside the geographic area of this report.
SPECIAL CONSIDERATIONS - MULTISEAM MINING
Multiple-seam surface mining is done with many types and combinations
of equipment. The mining of a second seam of coal often reduces the strip-
ping ratio and therefore provides an overall reduced stripping cost. More
selective machines are constantly being designed for increased production
and reduced operating costs in special applications, such as removal of both
thin partings and coal seams. Flexibility and mobility of this new equip-
ment, combined with lower capital investment, offer great inducement in
multiple seam operations or multimine situations.
Mining a second seam of coal and handling the parting in accordance with
economic and environmental controls presents a number of problems and poses
a challenge to planners. In the context of total resource extraction, all
available resources such as valuable claybeds, limestone strata, etc. are
best extracted with the extraction of the coal seam under consideration. On
the technical side, it is necessary to use a carefully planned sequence of
mining and correct combination of available equipment. Loading and hauling
units and overburden stripping must be matched.
Porter (1971) conducted a survey of multiple-seam strip mining opera-
tions and developed an economic feasibility model. All of Porter's case
studies involve stripping shovels, draglines, or tandem operations from the
East or Midwest. It is a comprehensive review of conventional multiple-
seam (usually two-seam) operations in these regions. Another examination of
multiple-seam operations is found in a report by Stefanko et al. (1973),
(Figures 52 and 53). Although intended as a preview of all strip mining
methods and equipment, the case studies presented cover some multiple-seam
mines, particularly in the Midwest and Northern High Plains. Figure 54
shows a photograph of a multiple-seam operation with a 22.93 cubic meter
(30 cu-yd) dragline in Pennsylvania.
The ratio of coal thickness to parting thickness is important in the
consideration of haulback methods. When the thickness of the parting layer
is large compared to the thickness of the coal, space for stoppage of spill
is limited and the parting must be hauled upgrade for spoiling. Short,
100
-------
Haulage Road
V V
V
Mined Out Area
Exposed Lower Bench
Exposed Upper Bench
Surface
Figure 51. Plan view of an open-pit coal mine.
-------
100-150'
Long
Keycut
Exposed
4' Thick
Seam
Plan View
7.5'
Figure 52.
Section View
Multiseam stripping operation with a dragline
(Stefanko, Ramani, and Ferko, 1973).
102
-------
Surface
19.5' Thick
Parting
Highwall
Exposed 7.5'
Coal Seam
Leveled
Spoil
Spoil
Plan View
Surface
Figure 53,
100-200'—H
Section View
Dragline exposing lowest seam from leveled spoil
(Stefanko, Ramani, and Ferko, 1973).
103
-------
o
4>
Figure 54. A two-seam operation with a 30-cubic-yard dragline in Pennsylvania,
-------
steep grades encountered with loaded haulage vehicles quickly nullify trans-
fer and stowage benefits.
Porter (1971) details a tandem operation using a stripping shovel and
dragline to mine three seams of coal. The operational configuration is
shown in Figure 55. A 1050B shovel exposes the first two seams, and a
Marion 8300 dragline exposes the bottom seam while simultaneously rehandling
36 percent of the shovel spoil. At intervals, the imbalance between the
primary stripping machines necessitates the use of the dragline on the middle
parting in order for the 1050B to maintain a proper lead. The rehandle is
nonproductive work and the dragline output is about 30 percent less in the
shown operating mode than would be expected in a conventional (highwall
position) stripping situation. Some of this loss can be attributed to the
180 degree swing required while stripping from the spoil side. Another
problem area noted, less related to overburden and parting handling but
relevant to multiple-seam operations, is haulage from three levels. Haulage
roads are brought into the pit about midway between the upper two seams.
Access to the lower seam is gained by an incline dug in the face of the
spoil parallel with the pit. The entire operation requires a longer than
normal pit to allow maneuvering room. Although the pit is a good operation
from a conventional view, the thicker seams of the West offer possibilities
for stowing. Stowing would eliminate rehandle, ineffective dragline ap-
plication, and complex timing sequences as machines pass one another.
Effective stowing could shorten the required open-pit length. Although the
thinner coal seams will not allow efficient backstowing, the problems
encountered in conventional multiple-seam mining are evident.
One multiple-seam operation in Illinois utilizes one shovel to uncover
two coal seams. A Marion 5761 shovel strips both seams as two active,
parallel cuts. An average of 12.19 meters (40 feet) of top overburden is
removed with a 180 degree swing and the seven-foot parting is stripped from
directly in front of the shovel with a 90 degree swing. The 1.67 meter
(5.5 ft) thick top seam and associated thin parting create this unique
situation. The shovel can selectively remove overburden as the top of the
parting is only 3.81 meters (12.5 ft) above the top of the second seam on
which the shovel sits. The volume of material in the innerburden is minimal
and leaves sufficient spoiling room for the top overburden. Although the
seams are exposed simultaneously, they are effectively separate slices and
the overburdens are shot separately. Utilization of the single unit is
good in this particular case. The reach available permits easy handling
in a pit 45.72 meter (150 ft) wide. One desirable result is burial of the
parting material. Exposure of both seams is another benefit as this allows
blending of the coals from the two seams.
Another example of the use of one shovel to mine two seams of coal is
near Central City, Kentucky (Figure 56). Although the mining method is not
identical to the plan proposed here, it is similar in that one shovel is used
to excavate two separate strata of overburden material simultaneously.
The stripping shovel, a Marion 5761 with a 49.69 cu m (65 cu yd) dipper,
operates on top of the lower coal seam to remove the overburden and the
parting layer. Overburden thicknesses average 12.19 meters (40 ft) and the
105
-------
NOTE: SEAMS IN DESCENDING ORDER ARE
No. 13, No. 11 AND No. 9
120'--J
19" 1ft"
30'-35' '8
| »66"-72"
55'-60'
'48"-54"
66"-72"
48"-54"
No. 9 SPOIL
66"-72"
48"-54"
Figure 55. Shovel-dragline tandem operation for multiseam mining
(Porter, 1971).
106
-------
TOP
BOTTOM.
PREVIOUS
SPOIL
SEAM ( • SEAM f PR|VIOUS
OF COAL'S OF COAL — SPOIL (
REMOVED! REMOVE'D -
-ADVANCED
irSPOIL
LINE —
A c~c -
^^===
POSITION „„ .«, •
SPOIL_LI-NE -
BOOM-DUMPING-^:
POSITION
* .."..SURFACE ..„, .....
......... •-. ...... -7- .,, ..... .,:
:.r>,, :,-
. i. «
..."•. ••>"•-. ..»,.' ;.„_ ',,; ..... i,.
'
•" '"'.». " j]' ••"• .
•" - .„. •.;,. - «
PLAN VIEW
5' COAL SEAM SHOVEL
10 20
CROSS SECTION
PREVIOUS
SPOIL
Figure 56. Multiseam stripping operation with a. shovel.
107
-------
parting layer averages 2.13 meters (7 ft) in thickness. Two working faces in
a double-wide cut are advanced at the same time. The parting layer is
excavated from in front of the shovel and dumped into the empty cut. The
5761 with a 57.91 meter (190 ft) boom then digs to the highwall side to
remove the overburden which is also spoiled into the empty cut and on top
of the parting. This type of operation requires a large shovel with good
reach, but the stowing and selective mining concept indicates that such a
shovel application is feasible. A variable pitch dipper would increase the
efficiency of such a mining method.
Grim and Hill (1974) outline three methods for multiple-seam mining
which are applicable in steep terrains:
Method No. 1—If the overburden from the upper seam will not reach the
bench of the lower seam, treat each seam as a separate mining operation,
mining the lower seam first. This bench may be used to store spoil produced
during stripping of the upper seam.
Method No. 2—If the overburden from the upper seam will reach the
bench of the lower seam, mine the lower seam in advance of the seam above.
Grading should be delayed on the lower bench in order to catch big rocks
from the upper seam and bury them in the pit. In no instance can spoil from
the upper seam extend more than one-half the distance from the highwall to
the edge of the solid bench of the lower seam.
Method No. 3—If both seams appear in the same highwall, separated by
more than 7.6 meters (25 feet), and two or more cuts are planned, the coal
should be recovered from the bottom seam first. If the seams are separated
by less than 7.6 meters (25 feet), mine from the upper seam down, recovering
both seams in one systematic operation (Figure 57, Steps 1 and 2). Lateral
movement of the spoil is recommended.
SUMMARY
This chapter has briefly reviewed the more commonly used surface coal
mining methods. Several examples were presented on specific equipment
applications, including multiseam mining operations. While haulback methods
are becoming popular in the hilly terrains of the east, area mining with
simple overcasting is most popular in the interior region. Tandem operation
of large strip equipment (bucket-wheel excavator, shovels, draglines) is
quite common in Illinois. The concern for concurrent land reclamation, and
topsoil replacement has seen the introduction of a fleet of wheeled vehicles
for topsoil and overburden removal.
108
-------
MINE UPPER COAL SEAM FIRST
ORIGINAL
GROUND.
SLOPED
HIGHWALL NO.
1st CUT
MOUNTAIN
\\\\\\\\\\ . .
PARTING LESS THAN 25 FEET
\\X\\\\Nlk\\\\\\\\\\\\\\\\
MINE LOWER COAL SEAM
HIGHWALL
COMPLETED
FOR 1st CUT
ORIGINAL
GROUND
SLOPE
MOUNTAIN
COAL SEA M
PARTING LESS THAN 25 FT
\m\\\
COAL SEAM
Figure 57. Multiple-seam mining method:
apart in the same highwall.
two seams less than 25 feet
109
-------
SECTION 5
FRAGMENTATION PRACTICES
INTRODUCTION
Fragmentation encompasses several different unit operations in the
surface mining of coal. These operations combine to provide for the fractur-
ing of both the overburden and coal, thereby simplifying the subsequent
materials handling procedures. In some favorable conditions, neither the
overburden nor the coal may exhibit such strength as to require fragmentation
prior to actual excavation. In the most restrictive case, the overburden
and the coal may both be so strong as to require noncurrent blasting. As is
typical in the mining industry, composite situations exist whereby portions
of the overburden or coal can be directly excavated while other portions
must be prepared prior to the digging operation.
As far as overburden preparation is concerned, a range of possibilities
is presented in the fragmentation unit operations. Overburden fragmentation
and preparation may be accomplished in some situations by ripping with bull-
dozers. As an intermediate case, widely spaced drilling and blasting may be
sufficient. In the limiting case, a regular and intensive drilling and
blasting approach may be required. Where blasting is required, various con-
siderations can affect the drilling layout. The size and nature of the
drillholes may vary, thus affecting spacing. Further, the proper type of
explosive, either high or low strength, must be carefully examined.
Although the stated purpose for rock fragmentation is to fracture the
overburden and coal for loading, there are other secondary, but vital,
purposes of a properly designed fragmentation process. These purposes
include product size, and rotation and translation.
The size of the broken materials should be small enough to fit into
the bucket of the excavating machine and to eliminate secondary blasting,
yet small particle generation should be kept to a minimum for dust control.
While the largest size is clearly dictated by the dimensions of the machinery,
the lower limit is rather difficult to determine. Suffice it to say that
fragmenting to a size smaller than that required is wasteful in explosives
and cost.
Rotation and translation (throw) refers to the manner in which the
blasted material comes to rest. If the explosive force achieved is too great,
then the material may move beyond the affected area, thus causing difficulty
in loading. Further, a properly designed blasting layout can rotate the
fragments to improve their loadability characteristics.
110
-------
Three additional factors in fragmentation relate to the precautions
needed to minimize the environmental effects of blasting. The blast design
must minimize the environmental effects of blasting. The blast design must
minimize fines, noise and vibration. Critical to the environmental effects
is fines production (dust). Over-designing the blasting rounds can, in
addition to fines production, cause conditions of excessive noise and
vibrations. If not properly controlled, the above-mentioned three problems
can have as great an impact on fragmentation practices as the production
requirements.
The parameters on which the design of fragmentation processes are based
relate to two general areas. One is the previously mentioned total opera-
tion design, i.e., the dimensions of the area to be prepared, the desired
fragmentation, and the economic considerations. The second area concerns
the properties of the material to be broken, i.e., primarily the strength
of the material and its response when subjected to high energy impulses.
A solid mass exhibits different responses to the application of energy.
Basically, the response differs with the absolute quantity of energy intro-
duced and the time over which the energy is applied. Rocks can absorb
small impulses of energy, or even large quantities of energy applied evenly
over a long period of time, and can deform without collapse. Therefore,
the goal in breaking rock is to apply a large enough force in a short
enough time to create fractures in the mass and to rotate and translate the
resultant fractured mass.
The simplest case of fragmentation is that in which no explicit fragmen-
tation process is required, or the case where the excavator (bucketwheel,
dragline, shovel, loader, scraper, or bulldozer) can dig the material with-
out prior treatment. In this case, the digging machine is capable of
delivering an excess of the total amount of energy that the mass can with-
stand in the static case. The next extension of the simple case is that in
which the excavator is just barely incapable of digging the material. In
this case, it is possible to concentrate energy through use of a ripper.
A ripper concentrates the tractive effort of the dozer (which can be con-
siderable) at one point, the ripper tooth. The applicability of direct
ripping or direct excavation without fragmentation is determined in the
specifics of equipment selection for either coal or overburden handling. In
many cases, small operators may choose to use such techniques at marginal
conditions in an effort to save cost in drilling equipment and personnel.
Large operators may choose to use direct excavation where mobility is a key
to successful operation. In the Eastern United States, drilling and blast-
ing are necessary prior to excavation in most operations.
DRILLING
Where the strength of overburden material is such that blasting is
required prior to excavation, drilling precedes the use of explosives.
Quite simply, drilling is the unit operation which produces the hole for the
111
-------
placement of explosives. Drilling, for the raining of coal by surface
techniques, is categorized as rock penetration by mechanical attack.
Mechanical attack is further subdivided into two methods - percussive and
rotary action.
Beyond the mechanics of attack, drilling is also categorized by the
manner of application. Specifically, the drilling of overburden for the
surface mining of coal can be approached in three manners:
1. Vertical conducted from the top of the highwall down to the
coal perpendicular to the strata bedding planes.
2. Horizontal - conducted from the pit floor parallel to the
bedding planes of the strata.
3. Inclined since the shortest drilling distance to a pitching
seam is perpendicular to strata bedding planes, inclined
drilling is conducted from the top of the highwall normal to the
coal seam's angle of pitch.
Since most coal seams in the East are tabular, the third option is rarely
encountered.
PERCUSSION DRILLING
Penetration of overburden by percussion drilling is affected by succes-
sive impacts of the bit into the rock. The bit is rotated between blows to
aid in the crushing and chipping process. Percussion drilling is used when
small diameter holes must be driven through hard rock where rotary units
are difficult to employ.
There are a number of advantages of percussion drilling. First, small
diameter holes can be drilled through hard rock. The rigs themselves are
light,"less expensive, and are quite maneuverable. Finally, one man can
operate the rig through a wide range of drilling angles (Morrell and Unger,
1973).
Percussive drills are subdivided into two categories, surface-mounted
and down-the-hole. The percussive action of surface-mounted units is trans-
mitted to the bit by a prime mover which remains outside the hole. In
comparison, the prime mover for a down-the-hole drill actually strikes the
bit and follows it into the drilled holes.
The most widely employed percussive drills for surface mining are
surface mounted. Pneumatically powered, these drills also use the air for
flushing the cuttings from the hole. Since the addition of drill steel
can decrease the available drilling energy nearly 50 percent, surface-mounted
percussive drills are used primarily where the drilling depth is less than
30.48 meter (100 ft) (Morrell and Unger, 1973).
Since the bit must be rotated between blows to effectively chisel the
rock, surface-mounted percussive drills are either rifle-bar or independently
rotated (Table 20). The drill steel in a rifle-bar rotation unit is
112
-------
TABLE 20. SPECIFICATIONS OF SOME RIFLE-BAR ROTATION AND INDEPENDENT
ROTATION PERCUSSIVE DRILLS (AFTER MORRELL AND UNGER, 1973)
Specifications
Rifle-Bar Rotation
Independent Rotation
Hole size, in ,
Piston diameter, in ,
Drill weight, Ib
Drill length, in
Blow frequency, bpm
Energy per blow, ft-lb
Energy output, ft-lb per min.
Air consumption, cfm
2
112
30
1,760
135
238,000
195
2 ~ 3 T
*it
155
29
1,975
160
316,000
340
— - 5 —
2 2
2
385
40
1,616
275
444,000
9003
— - 2 —
2 2
1,970
117
230,000
3252
2f-4
**
290
39
1,640
228
374,000
5603
*£
690
55
1,250
375
469,000
1,0503
1 23
At 100-psi operating pressure. Includes air motor. Includes blow air for flushing.
-------
automatically rotated by a fluted rifle bar and nut on the block stroke.
Since rotation is automatic, the ability to vary the rotational speed is
minimal. Independently rotated drills are, as the name implies, rotated by
an independent motor. This adds flexibility to the system when rock of
varying strength is encountered. For example, harder rock requires a
decrease in rotational speed and an increase in the impacting force. Rota-
tional speeds of 200 to 400 rpm are common for 50.8 mm (2 inch) bits, while
152.4 mm (6 inch) bits are usually operated at less than 100 rpm.
Surface-mounted drills offer the following advantages (Morrell and Unger,
1973) :
1. Surface-mounted drills can be more powerful and can penetrate
faster in shallow holes than down-the-hole drills because, unlike
the latter, the drill-piston diameter can be greater than the
hole diameter.
2. The minimum hole diameter for a down-the-hole drill is 101.60 mm
(4 in), whereas a surface-mounted drill is capable of even smaller
diameters.
3. Unlike a down-the-hole drill, the loss of the drill unit due to
hole caving is not a problem for surface-mounted units.
However, surface-mounted drills have the following disadvantages:
1. Total penetration depth is limited to 30.48 meters (100 ft)
because the loss of drilling energy, due to the addition of five
joints of drill steel, can approach 50 percent.
2. Holes greater than 152.4 mm (6 in) in diameter are not common.
3. Unlike down-the-hole drills, noise is not facilitated by hole
penetration.
4. Drill rod life is reduced due to the continued transmission of
impaction energy.
Down-the-hole percussive drills are employed where drillholes greater
than 30.48 m (100 ft) are required and hole diameters of up to 304.8 mm
(12 in) are preferred (Table 21). Since the percussive unit follows the
bit into the hole, a clearance of approximately 12.7 mm (0.5 in) must be
provided by the hole diameter for the passage of cuttings around the drill
unit. The rotational motor is the only part of the operation that takes
place out of the hole. A typical rotation range is 10 to 100 rpm (Morrell
and Unger, 1973).
Down-the-hole drills have the following advantages:
1. Penetration rates are maintained because the drill-to-bit distance
is constant, unlike that of a surface-mounted drill, and percussive
energy is not lost.
2. The drill noise is dampened as soon as the unit enters the hole.
3. Since they do not transmit impact energy, drill rods have a
longer life than those employed with surface-mounted drills.
114
-------
TABLE 21. DOWN-THE-HOLE DRILL SPECIFICATIONS
(AFTER MORRELL AND UNGER, 1973)
Drill
Specifications
Overall diameter, in.
Tool weight, Ib
Tool length, in
Piston diameter, in
Operating pressure, psi
Air consumption, cfm
Hole Diameter
4 In.
H
90-100
40-45
^
100-250
100-500
6 In.
^
180-200
40-60
-------
Qown-the-hole drilling has the following disadvantages:
1. Since the minimum hole diameter is dictated by the drill diameter,
the minimum hole size is four inches.
2. Hole caving can make drill recovery difficult, if not impossible.
3. Since larger bit diameters are required for each drill size, the
amount of energy that can be applied to a bit is limited.
Percussive drill bits are divided into two categories for surface
mining applications, cross bits and button bits. The tungsten carbide cross
bits are either of an "X" (at an angle nearly 90 degrees) or a cross (90°)
configuration. Button bits, in sizes from 50.8 to 165.1 mm (2 to 6.5 in),
are flat-bottomed with cylindrical tungsten-carbide inserts. Cross bits
can usually be sharpened 20 times, while button bits do not require sharpen-
ing.
ROTARY DRILLING
Rotary drilling, the most versatile form of overburden penetration
for surface mining of coal, is conducted by rotating a rigid, tubular string
of rods to which is attached a rock-cutting bit. Thus, the drilling energy
is supplied by rotation as well as thrust. The cutting action consists of
either abrasion, scraping, spalling, or chipping, as well as a combination
of the above. The size range for rotary-drilled blasthole diameters is from
101.6 mm (4 in) to 381 mm (15 in) with 228.6 mm (9 in) and 304.8 mm (12 in)
being common sizes.
For the surface mining of coal, most rotary drilling rigs are truck-
mounted, although crawler-mounted units are also used. All drilling rigs
are equipped with jacks for leveling on uneven ground. The power source
for the rotary drill rigs is either electric (for economy) or diesel (for
flexibility in tramming).
Drill steel, of up to 15.24 m (50 ft) in length, is stored on a circular
rod carrier which rotates in a mast for the automatic changing of drill
steel. Although most drilling is vertical, the mast can be tilted up to
30 degrees off vertical for angled drilling.
There are two types of bits that are used for rotary drilling, drag
or rolling cutter. Drag bits are applicable in soft to medium-soft shales
or their equivalents. Since their effectiveness is dependent upon a
successful scraping action, drilling thrust is lower for drag bits than for
rolling cutter bits, but the opposite is true for torque requirements. The
bits are usually fitted with replaceable tungsten-carbide cutters.
Rolling cutter bits, though available in the two-cone configuration,
are usually of the three-cone style (Table 22). The rolling cones have
either large, steel teeth (for drilling through soft formations) or small,
tungsten-carbide inserts (for drilling through hard formations). Table 23
relates bit size and weight to the strength of the rock.
116
-------
TABLE 22. THREE-CONE ROLLING CUTTER BIT DIMENSIONS AND HOLE
VOLUME PRODUCED (AFTER MORRELL AND UNGER, 1973)
Bit
Diameter,
In.
5
8
6 —
4
7 3
.8
7
7 —
8
8 —
2 • • • •
3
Q J
4
9
9 —
8
5
10 4
8
12 T-
4
15
Pin Size
API* Reg.,
In.
1
2
3 —
2
o 1
2
1
Q .
2
/ 1
2
-i
4 —
2
A 1
2
s 5
8
c
fi —
8
r 5
8
fi 5
8
Bit Weight,
Lb
26
35
49
60
73
75
77
122
141
200
306
Volume of
Hole Drilled,
Cu Ft per Ft
0 173
0.249
0 279
0 338
0.394
0 418
0 442
0 532
0 616
0.818
1.227
* American Petroleum Institute.
117
-------
TABLE 23. BIT WEIGHT AS A FUNCTION OF ROCK STRENGTH AND
BIT SIZE (AFTER MORRELL AND UNGER, 1973)
oo
Rock Strength
Very soft formations: overburden, soft
shales, limestone and evaporites
Medium formations: limestone, dolo-
mites and sandstones
Hard and very hard formations: basalt,
granit, quartzite, caconite
Weight
per In. of
Diameter,
Lb
1,000-
4,000
3,000-
5,000
4,000-
8,000
Recommended Bit Loading for
Bit Size Indicated, Lb
4-6 In.
4,000-
24 , 000
12,000-
30,000
16,000-
40,000*
6-9 In.
6,000-
36,000
18,000-
45,000
24,000-
60,000*
9-12 In.
9,000-
48,000
27,000-
60,000
36,000-
100,000
* Maximum weight for largest bit size indicated.
-------
The major advantage of rotary drilling is its rapid drilling rate.
This is due to the fact that, unlike percussion drilling, rock penetration
is uniform with depth (Worrell and Unger, 1973).
VERTICAL AND HORIZONTAL DRILLING
There are basically two approaches to drilling overburden for the
surface mining of coal, vertical and horizontal drilling. Although vertical
drilling appears to be the overwhelming choice for surface mining applica-
tions, the horizontal method can be used effectively under certain condi-
tions .
Vertical drilling consists of drilling from the top of the highwall
perpendicular to the coal seam. This makes the use of bulk explosives
possible. It is also the preferred approach where overburden, which con-
sists of competent strata, exceeds 15.24m (50 ft) in depth or massive stone
occurs at an elevation higher than that which can be reached by horizontal
drilling. Because common practice in vertical drilling is to penetrate
to the coal and then backfill approximately four feet, there is less
shattering of coal with this approach.
Horizontal drilling, a relatively new development -for the surface
mining of coal, is usually conducted by a twin-mast rotary drill. Both
holes, which can be drilled either simultaneously or independently, can
be varied from 8.53 m (28 ft) to 11.58 m (38 ft) centers and can be drilled
from 10 degree above to 12 degree below the horizontal. Typical hole
diameters vary from 171.45 mm (6-3/4 inches) to 269.87 mm (10-5/8 inches).
Where the only portion of the overburden that must be blasted is
located directly above or near the seam, horizontal drilling is applicable
since less footage is usually required for the same pit width. Horizontal
drills tram on the coal floor which, by eliminating the surface preparation
required for vertical drills, allows for faster rates. Although applica-
tions are usually limited to a maximum of 30.48 m (100 ft) of overburden,
45.72 m (150 ft) deep holes (approximately two pit widths) can be conven-
iently drilled at one time (Nordness, 1976).
BLASTING
Effective bank preparation not only deals with the fragmentation of
overburden, but also with the minimization of incurred costs. As such,
the target for proper overburden preparation is efficiency with regard to
the number of holes drilled and the amount of explosives utilized.
It has already been stated that operators must continually evaluate
their fragmentation procedures. The amount of explosives employed must be
adequate for fracturing the rock to a size which is acceptable for equipment
removal. Since the character of the overburden in a large surface mine can
change over a wide area, economics alone is a sufficient reason for the
emphasis on studying and selecting good fragmentation practices.
119
-------
SYSTEM COMPONENTS
Blasting systems are constructed of the following components:
1) detonators, 2) primers and boosters, and 3) explosives.
Detonators are devices for producing an explosive decomposition in a
high-explosive charge. Initiation of the detonator is by either a safety
fuse or by electricity. Detonators consist of two styles, blasting caps or
detonating cord.
The most commonly used detonator is the electric blasting cap. It
consists of a copper shell, closed at one end, which contains a detonating
compound. Because it is placed in the explosive charge, the ignition of
the cap, by a current or spark, sets off the entire reaction. The advantages
of electric blasting caps are as follows (Dick, 1973): 1) can be safely
handled, 2) accuracy of delay periods, and 3) controlled firing time.
The following are hazardous conditions which dictate caution when using
electric blasting caps (Dick, 1973): 1) lightning, 2) radio frequency
energy, and 3) mixing of different brands of blasting caps.
Blasting caps can also be activated with a safety fuse. A safety fuse
consists of a potassium-nitrate black powder core enclosed in a covering
of textile and waterproofing compound. The fuse is inserted into the open
end of the cap, which is then attached to the charge. Burn rates for fuses
vary between 98.42 and 131.23 second per meter. Care must be taken when
using a safety fuse because of the following conditions (Dick, 1973):
1. Since the fuse cap is open at an end, exposing the explosives,
care must be taken in handling.
2. Altitude changes affect burning rates.
However, the following are advantages for using safety fuses:
1. Elimination of the hazard of stray electricity.
2. No need for a power source or associated lead wires.
The detonating cord is a fuse, consisting of a high-explosive core of
pentaerythrital tetranitrate (PETN), enclosed in tape, and wrapped with
textile yarns, that fires the charge .without the assistance of any other
detonator between it and the charge. However, an electric blasting cap is
needed to start the reaction in the detonating cord. Although blasting
caps with fuses can be used, electric blasting caps are preferred because
of their ability to control the instant of firing. In fact, the use of
two caps at each initiation point is common to guarantee proper firing.
Since a detonating cord is insensitive to shock, friction, and lightning,
it is quite applicable as a safe and convenient trunkline for firing
blasts.
120
-------
Initially, applications of detonating cord were instantaneously fired
by means of a trunkline. To achieve a delay, electric caps were attached
in each hole. Millisecond (M.S.) connectors were then developed to provide
delay firing by surface, rather than downline, initiation.
With instantaneous firing, the layout must be designed with attention
to the direction of the blast movement. For example, the charges nearest
the initiation point should not "cut off" the firing in later holes due to
ground movement. Because of this, the trunkline should run parallel to a
free face and should also have an open end (Figure 58). With this config-
uration, detonation should be from the open end. By looping the trunkline
("double trunkline"), a second means of detonating the holes is available if
a break occurs in the main trunk. In multiple row blasts (Figures 59 and
60), availability of secondary detonation is provided by cross ties, spaced
on intervals of less than 60.96 m (200 ft).
Delay firing, resulting in reduced vibration, improved fragmentation,
and less backbreak, is considered to be more desirable than instantaneous
firing. Delay intervals where detonating cord is used vary between 5 and
35 milliseconds. For surface mining applications where large-diameter
boreholes are spaced on approximately 9.14 m (30-foot) centers, 25-milli-
second intervals are used to avoid cut-offs. In fact, to initiate detonat-
ing cord downline with millisecond delays attached at the surface, the
minimum available interval is 25 milliseconds.
Millisecond delays were quite widely employed prior to the introduc-
tion of millisecond connectors. These short-interval delays are attached,
on the surface, to each detonating cord downline at the individual holes
or to groups of holes by means of a trunkline. With M.S. delays and the
absence of a trunkline, noise is reduced; however, precautions must be
taken because of the large amount of electrical wiring required.
Millisecond connectors are tied directly into the detonating cord
trunklines. Unlike M.S. delays, M.S. connectors come in a variety of
delay intervals, 5, 9, 17, and 25 milliseconds. Figures 61 through 65
show sample configurations with M.S. connector placement marked by an "X".
Special attention should be drawn to Figure 65 because it shows a con-
figuration for horizontal drilling.
A primer is usually the combination of an explosive cartridge and a
detonating cap. By detonation, it sets off the remainder of the charge.
The cartridge is placed at one end of the charge with the embedded detonator
pointing towards the charge, not away from it. Detonating cord can also be
used as a primer. The following precautions should be taken to assure the
proper application of primers (DuPont Co., 1967):
1. The blasting cap should be securely inserted into the primer
cartridge (deeply into the center) to guarantee that it cannot
be pulled out.
2. The wires of electric firing devices or fuses should not be
subject to harmful strains.
121
-------
Figure 58. Single row of holes connected by "primacord" and detonated by
electric blasting caps or caps and fuse attached to "primacord"
at A. Also, note that the "primacord" loop is tied to the
trunkline with a right angle connection (After DuPont Co.,
1967).
Figure 59. Double row of holes connected by "primacord"
(After DuPont Co., 1967).
Figure 60. Three rows of holes connected by "primacord"
(After DuPont Co., 1967).
122
-------
POINT OF INITIATION
WITH 2 EBC'S
X—•—X—•—X—•—X—
S^\l§^^
OPEN FACE
Figure 61. Single row shot fired in sequence. "Primacord"
M.S. connectors inserted at points "X" (After
DuPont Co., 1967).
123
-------
X—9—X— • —X—•—X—•—X
^
^g^<3$^as«$0E88i
OPEN FACE
POINT OF INITIATION
WITH 2 EBC'S
Figure 62. Multiple row shot with holes and rows fired in sequence
from one end (After DuPont Co., 1967).
POINT OF INITIATION
WITH 2 EBC'S
OPEN FACE
Figure 63. Multiple row shot in sequence initiated at center
front (After DuPont Co., 1967).
*"\
u
POINT OF INITIATION
WITH 2 EBC'S
OPEN FACE/
Figure 64. Multiple row shot entire row in sequence (After DuPont Co., 1967)
124
-------
I
POINT OF INITIATION
WITH 2 EBC'S
Figure 65. Single row horizontal hole shot in bituminous stripping
(After DuPont Co., 1967).
125
-------
3. When necessary, the primer should be water resistant.
4. A dynamite punch should be used to make the hole for insertion
of the blasting caps or detonating cord.
5. Since detonating cord will initiate every cartridge it comes
in contact with, care should be taken that it is attached to
the first cartridge loaded into the hole.
6. Primer assemblies should be made just before the holes are to
be loaded and at the blast area, to reduce the exposure time to
hazards such as fire, impact, or electricity.
A booster is an explosive, usually encased in a small metal shell, to
improve the performance of another explosive. They are, quite often,
placed at intervals along the borehole to reinforce the detonation of the
column of blasting agent. Boosters differ from primers in that they do not
contain an initiating device. Since boosters are placed between the
detonators and blasting agents, they are essential for the ignition of
cap insensitive blasting agents.
An explosive is a compound or mixture of compounds which undergoes a
rapid, self-propagating exothermic reaction. This reaction is induced by
either heat, impact, or shock. When the reaction moves through the
explosive at a faster rate than the speed of sound in the unreacted
explosive, it is called a detonation. If it is slower, it is called a
deflagration. In any event, the products of the reaction, usually gases,
exert tremendous pressure during expansion. Since the resultant condition
created by the shock front is also dependent upon both detonation velocity
and the sonic velocity in the rock, the reaction should be designed so that
those velocities are equal.
Figure 66 shows a standard cartridge of explosive in the process of
detonating. The right side is undetonated explosive, the left is expanding
byproducts of the reaction, and the center represents the area where the
explosion, a chemical reaction, is taking place. The explosion is a rapid
oxidation reaction in which all elements necessary for the reaction are
present in the explosive. The reaction takes place in a very definite
zone though, in most cases, the zone is not as wide as is indicated in the
figure. The actual width of the zone varies with a coarseness of the
explosive particles, from small fractions of a millimeter in the fine-
grained high explosives, to several centimeter in bulk, granular blasting
agents such as ammonium nitrate/fuel oil (ANFO) mixtures.
Just as rock properties, such as density, porosity, strength, energy
absorption, etc., affect fragmentation, there are various explosive prop-
erties which also affect the outcome of the fragmentation process. Among
these are detonation velocity, density, detonation pressure, borehole pre-
sure, and water resistance.
126
-------
SHOCK FRONT
EXPANDING
GAS PRODUCTS
UNDETONATED
EXPLOSIVE
DETONATION
Figure 66. A cartridge of explosive undergoing the process of
detonation (After Boddorff, 1974).
-------
Detonation velocity is the speed at which the reaction front moves
forward through a cylindrical charge. The velocity ranges between 1676.4 m
(5500 ft) and 7620 m (25000 ft) per second. Faster velocities are appro-
priate in strong rock where a shattering effect is desirable. Slower
velocities are applicable where less demand is on rock shattering, whereas
a heaving effect is warranted. The following are steps that can be taken
in the field to increase detonation velocity (Dick, 1973):
1. Use a larger charge diameter.
2. Increase density.
3. Decrease particle size.
4. Provide good confinement in the borehole.
5. Use a large initiator or primer.
Density refers to the ratio of explosive density to that of water
(specific gravity). The ratio for commercial products varies between 0.5
and 1.7. By relating the specific gravity of the explosive to the borehole
diameter, the amount of explosive per foot of borehole can be determined.
This is particularly useful when matching blasting requirements to the
economics of blasting. Figure 67 shows a nomograph to aid in this calcu-
lation.
Detonation pressure is the pressure of the detonation wave propagating
through the explosive column. It is measured at the rear of the reaction
zone (Chapmen-Jouget plane). Detonation pressure can be calculated from
the following equation:
P = 4.18 x 10 7 DC2/(1 + 0.80D)
where,
P = detonation pressure in kb (1 kb (kilobar) =
100 x 107 pascals)
D = specific gravity
C detonation velocity in meter per second
Figure 68 also shows the relationship by use of a nomograph. Although the
detonation pressure range for commercial explosives varies between 5 and
150 kb, the higher pressures are applicable where a greater shock wave is
needed, such as in competent rock.
Borehole pressure is that which is exerted perpendicular to the deto-
nation pressure. Since it is exerted on the borehole walls, it is important
in displacing, as well as breaking, the rock. Borehole pressures range
from 10 to 60 Kilobar in commercial explosives, where borehole pressures
common to ANFO are greater than detonation pressures. In high explosives,
however, the converse usually hold true.
128
-------
Specific
gravity
1.8 ~r
1.6 -•
1.4 - -
1.2 - -
1.0 - -
.9 - -
.8 - -
.7 - -
.6 --
.5 --
Loading
density,
Ib/ft
z 200.0
i: 180.0
-- 160.0
-: 140.0
:; 120.0
-_- 100.0
— 80.0
-- 60.0
40.0
30.0
20.0
15.0
10.0
8.0
6.0
4.0
3.0
2.0
1.5
1.0
.80
.60
.40
.20
Charge
diameter,
in
T '8
T '6
- - 14
-- 12
10
9
8
7
-- 4
-- 3
-- 2
Figure 67. Nomograph for determining loading density (After Dick, 1973)
129
-------
Detonation
velocity,
ft per sec * 10
20
15
10
Detonation
pressure,
kbars
300
200
150
100
50
40
30
20
15
10
Specific
gravity
I .6-3
L3~i
1.0 -E
0.8 —
0.6 —
Figure 68. Nomograph for finding detonation pressure (After Dick, 1973)
130
-------
Water resistance is the ability of an explosive to withstand exposure
to water without losing its sensitivity or efficiency. Gelatin dynamites
and slurries have good resistance characteristics. In between these two
extremes are the nongelatinized high explosives.
For surface mining applications, there are various types of explosives
on the market. However, they are categorized into three groups: 1) high
explosives, 2) dry blasting agents, and 3) slurries or water gels. Of all
these varieties, dry blasting agents are the most common in coal applica-
tions because of cost considerations and safety, and dry blasting agents
are cheaper than other forms of explosives and need a primer for detonation.
Ammonium nitrate fuel oil, which has become the most common dry blasting
agent, accounts for over 75 percent of the nation's consumption. Fifty
percent of this figure, 453.6 x 106 Kg (109 Ibs), is used annually for the
surface mining of coal (Condon and Snodgrass, 1974). Due to its overwhelm-
ing acceptance, further discussion in this section will be dedicated to dry
blasting agents.
A blasting agent is a mixture whose individual components are not
classified as explosives and contains a fuel as an oxidizer. Ammonium
nitrate fuel oil, though widely used, has only been commercially available
since the early 1950's. It was discovered when detonation was observed
after mixing ammonium nitrate prills with a certain proportion of a hydro-
carbon fuel (typically, 3.78 x 1(P cu m (4 qt) of oil to each 45.36 Kg
(100 Ibs.) ammonium nitrate). This led to the adoption of ANFO as a blast-
ing agent with 60 percent of the effectiveness of dynamite with a cost,
presently, of approximately $0.22 per Kg. Even though it represents an
added cost, aluminum chips (5 to 30 percent by volume) are sometimes
added to the mixture to increase explosive output. Ammonium nitrate fuel
oil usually comes to the mine site in one of four manners:
1. As separate ingredients in bulk form for on-site mixing.
2. Premixed in bulk.
3. In paper or polyethylene bags.
4. In cartridges.
Because ANFO is usually mixed on site, there are numerous parameters
which must be understood to maximize detonation velocities. Condon and
Snodgrass (1974) list them as follows: fuel oil content, confinement,
blasthole diameter, density, prill particle sizing, water content, critical
diameter, borehole coupling (the percentage of the borehole diameter which
is filled with explosives), temperature, and priming system. They examined
the impact of primer type and borehole diameter on detonation velocity at
two large surface mines in Indiana, since these variables are the easiest
to control by mine operators. They concluded:
1. Larger diameter boreholes provide for you more efficient detonation
of ANFO and greater energy per pound of powder.
2. The same steady-state velocities were achieved with different
types of primers.
131
-------
Although the application of ANFO is relatively safe, there are a
number of precautions which must be taken into account (Coal Age, 1972):
1. Ammonium nitrate and fuel oil should be mixed in special plants
under closely controlled conditions.
2. Locate mixing plants in an isolated area and build them with
fireproof materials.
3. Mixed material should be removed from the plant as quickly
as possible.
4. No welding or cutting should be done in the plant until cleared
of explosive materials.
5. Primers, detonating cord, and caps should be transported in
separate vehicles from mixed nitrate and oil. They should be
brought together only at the point of use.
Design of Blasting Rounds
Figures 69 and 70 show some of the characteristics and parameters used
in the design of blasting rounds. Figure 69 shows those dimensions which
are considered for a single hole, and Figure 70 for a. total round including:
D Diameter of the explosive in the borehole
B Burden, or distance from the charge to the nearest free face
S Spacing, or distance between two holes
H Length of hole
J Subdrilling length, or the length the hole is drilled below
the floor of the working pit
T = The portion of the hole that does not contain explosive, or
collar length
L = Bench height
b = Spacing between rows of holes as measured perpendicular to
the free face
s = Separation between adjacent holes in a given row
These abbreviations are used by both Pugliese (1972) and Ash (1970). As
both authors point out, b and s are terms used by operators in referring
to the drilling grid, and should not be confused with B and S, the burden
and spacing which are measured with reference to free faces.
Ash (1963) and Pugliese (1972) concur in the suggestion of five
ratios for use in the design of rounds. Assuming that all measurements
of length and diameter are expressed in the same units (cm, m, ft, or in.),
the ratios relate the dimensions to be designed through the following set
of equations:
B K^D : 1C Burden ratio (1)
S = KCB ; K = Spacing ratio (2)
S i>
132
-------
Charge
Cool
Figure 69. Bench cross section view showing D , B, H, T and L
(After Pugliese, 1972). e
133
-------
t
t
• J
»J
B
±
B
JL
«J •
-------
H K,,B ; K_T = Hole length ratio (3)
J = KB ; K Subdrilling ratio
J J
T = IC,B ; 1C, Collar distance ratio
(5)
From these relationships, all of the critical dimensions of blasting
round design can be calculated if appropriate values for the various K
ratios are selected. The work of Pugliese and Ash is primarily based on
studies in limestone quarrying. While limestone quarrying is conceded as
different from overburden removal in surface mining of coal, the relation-
ships still hold and the K values which they have developed and assumed
are valid for first assumptions. As is the case with any empirically
based design criteria, empirical relationships from literature are usually
accepted as a first approximation only, and succeedingly more exact
relationships can be developed as actual practice of these design steps
is followed.
For calculation of the burden, given a particular D , both authors
found the following ratios appropriate for limestone and dolomite:
30, under average conditions
K_ = 25, for low-density explosives such as ammonium nitrate
fuel oil (ANFO)
1C 35, for high density slurry explosives and gelatins
The actual K,. ratios are related also to the density of the rock. Lime-
stone and dolomite have a density of approximately 2.7 grams per cubic
centimeter. Lower values of K_ should be used for rocks of greater density
than 2.7 (e.g., 2.85 and up) or for lower density (e.g., 2.55 and less).
The calculation of spacing is based on Kq values between 1.8 and 2
for simultaneously initiated holes in a given row. Most favorable con-
ditions are with staggered drill patterns and for rows in which all charges
are not detonated simultaneously. If exact timing can be achieved,
simultaneously fired holes can be designed using Kq values of 3 to 5.
These charges must be sufficiently long and exactly timed to allow the
proper enhancement of stress effects.
135
-------
Spacing is also determined with respect to height and burden relation-
ships, specifically the ratio H/B. Ash (1970) has shown that the
relationships
1/2
S (B H) for 2B < H < 4B
S ~ 2B for 4B < H
will be sufficient. In this situation, the K of the 1.8 to 2 is satis-
factory, although lower K values can be assigned with H/B less than 3.
If delays are sequenced in a given row, the Kq should range between
1 and 1.2. Square drilling patterns are preferred, and rock will expand
in a wedge or in a 45 degree pattern to the original open face (as in
Figure 70, Plan A). Considering the ratio of s and b, where b = 1.4B
for the regular case, the design relationship
S = Kg b = KS (1.4B)
is appropriate where K 1 to 1.2.
O
The value of the ratio K may be adjusted between 1 and 2 to meet local
conditions, particularly the period of delay between charges. Large ratios
create spacing so much greater than the burden that the vertical face
craters and lumps are left on the pit floor. Small values of K,,, less than
1, create excessive breakage, premature breakage between holes, and pro-
gressive loss (rarefaction) from the hole, creating slabs, boulders, and toe
problems. The nature of these particular relationships is such that they
assume ideal energy balance between charges, so further experimentation
with the balance of individual holes may be indicated if such problems are
not resolved by varying the Kq ratio.
L)
Hole length calculations are generally based on K^ ratios of 1.5 to 4.
Pugliese (1972) cites 2.6 as the typical KH value. The depth should almost
always be greater than or equal to the burden to reduce overbreak and
cratering tendencies. Conversely, high values, that is, greater than 4, may
create underbreak or bootlegging if the hole is single-primed. Staged
priming may allow lengths greater than those indicated by K^ values of 4.
No definite limit to K^ can be established without a definite examination
of the cratering properties of the specific material being blasted.
The subdrilling length, which is applicable only to quarries and not
to coal mines, is generally indicated by K values greater than 0.2.
Pugliese (1972) suggests 0.3 as preferable to insure a full face and quarry
floor. In the case of overburden drilling, however, care is to be taken to
limit fracturing of the coal seam and subsequent unnecessary removal of
136
-------
coal with the overburden. In some instances, negative Kj values may be
applied. Common practice is to drill to the coal and then backfill by 1 to
1.3 m (3 to 4 feet).
The collar distance can be calculated by using a KI> value of 0.7. The
uncharged portion of the hole is usually filled with stemming material and
is left uncharged to prevent airblast and promote gas confinement within the
borehole. In solid rock, Pugliese (1972) suggests that values of Kj less
than 1 can lead to cratering, backbreak, and violence, particularly with
collar priming.
The various calculations are presented in worksheet form in Table 24.
Once the original face and the shape of the cut are selected, the geometrical
relationships necessary to work with can be determined from the plan layouts
of the cuts. For the given explosive type, blasthole diameter and rock
density, the burden is calculated from equation 1. Equation 2 is then solved
for the spacing. The hole length can then be checked according to equation
3 and the timing, delay, and sequencing constraints which the operator has
established. Finally, the subdrilling length (equation 4), if necessary,
can be calculated and the collar distance determined (equation 5).
Blasting Pattern Problems
In the previous section, the design of blasting rounds was viewed from
the standpoint of optimum fragmentation of solid rock. However, from a
practical standpoint, there are many in situ properties of the overburden
and the surrounding environment which also have a great impact on proper
design. The major problems which result from these conditions are ineffec-
tive fragmentation, vibration, and noise.
Jointed ground complicates the blasting pattern layout. This is due to
the fact that the joints act as avenues for the escape of explosive energy.
This weakening of the explosive's effectiveness is costly and requires a de-
sign to limit the amount of escape. Since blasted material fragments along
its structural pattern, it is natural to design the pattern to incorporate
this tendency. For example, Figure 71 shows a typical condition in a sedi-
mentary deposit. A rhombic structural pattern is shown with jointing angles
at 75 degrees and 105 degrees. If the direction of the blast is designed
to be out from the lesser angle (i.e., either N or S), overbreak, coarse
fragmentation, and excessive ground vibrations are usually experienced.
However, if the blast is designed to be out from the open angle (i.e., W),
better fragmentation, a uniform particle size, and a straighter face will
result. Therefore, in jointed ground, blasts should be directed out of the
open angles, wherever possible.
Ground vibration is an important consideration because of public
demands. Damage to surface structures, which is visible evidence of blasting
vibration effects, requires constant monitoring of blasting practices. With
this in mind, Andrews (1975) developed the following timing guidelines:
1. The minimum delay between holes or corresponding decks in adjacent
holes should be 1 millisecond per ft (ms/ft) of burden.
137
-------
TABLE 24. WORKSHEET FOR DESIGN OF BLASTING ROUNDS (AFTER PUGLIESE, 1972)
D =
B V,
B Ve
S = KgB
S KSB
s KQ b = Kc (1.
O ^
s KQ b Kc (1.
O D
H V
H = KRB
HMin = V
"Max = V
J = KB
J
J = KjB
T = I^B
T = K^JB
L H
in; Cut ; Desired
712 ft for K^
/12 = ft for Kj, -
rock progression
30 (average)
(alternative)
(Eq. 1)
For staggered pattern, simultaneous timing
ft for K0
ft for K0
2 (average)
1.8 (alternative)
(Eq. 2)
i>
For square pattern, sequence timing
4 B)= ft for K0
4 B)= ft for KP
v>
ft for K^ ~
ft for K^ ~
ft for KJJ
ft for KJJ
ft for KT
ft for KT ~
ft for 1^,
ft for K,_ -
-J ft (average
-J ft (minimum)
- J ft (maximum)
1 (average)
1.2 (alternative)
2.6 (average)
(alternative)
1.5 (minimum)
4 (maximum)
0.3 (average)
(alternative)
0.7 (average)
(alternative)
or alternative)
(Eq. 3)
(Eq. 4)
(Eq. 5)
138
-------
\ / \
A
Figure 71. Representative sedimentary vertical jointing
system (After Ash, 1968).
139
-------
2. The minimum delay between rows should be greater than the
average delay between holes in the same row, to allow
sufficient forward and lateral relief for every hole.
Figure 72 portrays the results which led to the first recommendation. In
this study (Bergmann, et al, 1974), a low delay ratio produced a large
fragment size. However, all the curves tended to flatten out as the delay
approached 1 ms/ft of burden. This implies that a greater delay than 1 ms/ft
of burden produces little change in fragmentation. The second guideline was
established by the comparison shown in Figure 73. If the delay time between
rows is equal to the time between holes in the same row (Figure 73), high
lateral confinement and poor fragmentation results. In Figure 73, however,
a delay time between rows that is twice that between holes in the same row
produces a jagged internal free face. This provides both forward and lateral
relief for better breakage and reduced ground vibration.
Since any excessive energy from a detonation travels outward through
ground, the intensity of the ground vibration is also a function of the
distance traveled by the shock wave (i.e., there is a decrease in intensity
with an increase in distance). Studies in this area (Nicholls, et al, 1971)
reveal that damage to structures by these seismic waves is caused by particle
velocity, rather than the displacement or acceleration, of the ground under
a structure. Particle velocity is the speed at which the individual unit
particles of the ground move as the shock wave is propagated. Figures 74
and 75 show some of the results of the aforementioned studies. Figure 74
shows data for the conclusion reached by the U.S. Bureau of Mines that
major damage occurred above a velocity of 19.30 cm/sec (7.6 inch/sec), while
minor damage occurred above 13.71 cm/sec (5.4 inch/sec). The Bureau then
separated its findings into damage and safe zones, with 5.08 cm/sec
(2.0 in/sec) set as the maximum value for the safe zone. Ninety-four per-
cent of the minor damage occurred above this figure, which proved to be
statistically acceptable. The value of 5.08 cm/sec (2.0 in/sec) was to be
monitored in all three of the orthogonal components of the ground next to
the structure. Figure 75 shows, however, that the value is not truly pre-
dictive insofar as damage is concerned. It shows that, in some instances,
no damage was observed where the particle velocity was greatly in excess
of the safe blasting criterion. However, the U.S. Bureau of Mines (USBM)
merely feels that damage occurrence increases or decreases as the particle
velocity increases or decreases with respect to the value of 5.08 cm
(2.0 inch). Therefore, a safe blasting limit of 5.08 cm (2.0 inches) per
second for the particle velocity, as measured from any of three mutually
perpendicular directions in the ground next to the structure in question,
has been used (Grim and Hill, 1974). The maximum peak particle velocity
is under reevaluation (see note on Table 25).
For the control of ground vibrations from blasting, a. simple formula
has been devised relating the blasting charge and proximity to structures
(Grim and Hill, 1974). It is given by the equation: W (D/50)2. In this
relationship, W equals the weight of explosives (in pounds) which are
detonated at any one instant, and D equals the distance (in feet) from the
nearest structure. Instantaneous detonation, in this case, is considered
140
-------
E
u
N
"35
a>
o
0"
o
65
60,
55
50
45
40
35
30
25
20
O
A
Square pattern, 13-in. burden, 13-in. spacing.
Rectangular pattern I, ll.0-in. burden,
15.5-in. spacing.
Rectangular pattern II, 9.2-in. burden,
18.4-in. spacing.
Best commercial delays
I I I I
J I
(according to Langefors)
I I I I i I I
0.2 0.4 0.6 0.8 1.0 1.2 1.4
Delay ratio, milliseconds per ft of burden
1.6
1.8
Figure 72. Effect of delay time between shotholes on average
fragment size. The same powder factor was used
for all shots (After Andrews, 1975).
141
-------
Echelon Firing/l.'l slope
Echelon Firing/2'.l slope
8
10 II
ORIGINAL FREE FACE
7 8 9 10 II 12 13 14
ORIGINAL FREE FACE
Figure 73. Comparison of multiple row patterns using
echelon firing on 1:1 and 2:1 slopes
(After Andrews, 1975).
142
-------
I^
K
O 4
v^» «^
O
_J
UJ
>
UJ o
_l ^
O
»—
or
<<
2
i
i
8
.6
.4
.2
i
_ i i i — i 1 r
o o
o <^SA
— r-Mo}or domage ©
, v=7.6 in/sec o
1 fMinor damage ^£>«
J J v^5.4 in/sec •£>*
1 1 0 . o
= Minor damage
A Edwards and Northwood |
O ASCE-BuMlnes Test J
1 III 1 1
II 1 1 1
—
—
—
D
n _, • Q
-B J °B Q -—
n CBfl • G^ _ Q
5r***- - ~
•• r^1,.'
~ ^" ^
•
^^" ^^Ml «^V ^^^MB ^^^^ J^^B^ ^BBl
^
1
* ^ *
Safe blasting criterion ! !
i j
Crondall ER = 3.0, v=3.3 in/sec- [ I
Langefors v=2.8 in/sec--1 _[_
Edwards and '
Northwood v = 2.0 in/sec ' i
Bureau of Mines v=2.0 in/sec =^—
_
iota
data —
1 I 1 II
10 20 40 60
FREQUENCY.eps
100
ZOO
400 600 1000
Figure 74. Particle velocity versus frequency with
recommended safe blasting criterion
(After Nicholls, et al., 1971).
143
-------
100
80
60
40
20
o 10
Q
(/) O
v.
c
o 4
O
UJ
LJ
O
I-
°~ 10
.8
.6
.4
O
o e °
P) O
e
o
Q Q
£, A
OC2>
O O O O A o
o a, o
'
o o
Safe blasting criterion--
o Bureau of Mines
Q Langefors
^ Edwards and Northwood
<3> ASCE-BuMines Test
I I I
No damage data
I
I
I
10 20 40 60 100
FREQUENCY.cps
200 400 600
Figure 75. Particle velocity versus frequency for no
damage data (After Nicholls, et al., 1971).
144
-------
to be all detonation within 8 milliseconds of each other. According to the
Federal standards, the formula is revised as W = (D/60) . The effect of this
revision is to reduce the weight of explosive that can be detonated for the
same distance, (D). Table 25 and Figure 76 summarize the application of this
means of vibration control. Other recommended or practiced blasting patterns
are provided in Figures 77 and 78 (McGraw Hill Inc., 1970).
The reduction in noise is beneficial, not because noise creates property dam-
age but because its reduction lessens the social problems faced by operators. Most
complaints received by surface mine operators start, initially, because of
noise. Although vibrations cause visible physical damage, noise has more of
a psychological effect on the nearby population. Since that effect is so
difficult to predict, measures must be taken to control the situation as
best as possible. As far as noise control is concerned, solutions are
generally twofold: 1) attention to blasting techniques and 2) proper
evaluation of weather conditions.
Andrews (1975) concluded that, to minimize air blast intensity, the
average rate of blast progression along the free face should be less than
the velocity of sound in air. If t is the delay time between holes, the
ratio d/t, when compared to c, is of extreme importance. When the ratio
d/t is less than e, separate shocks radiate in all directions. The resulting
air blast behind the face is, therefore, normally weaker and less noisy.
When the ratio is equal to or greater than c, a strong air blast is produced
due to super-position.
On-site noise control can also be practiced with detonating cord. When
cord is used for trunk lines, covering it with 25.40 cm (10 in) of dirt can
aid in noise abatement (McGraw Hill Inc. , 1970) . Further, a new fo-rm of
low-energy detonating cord has been developed. One-hundred-fifty feet of
this cord produces no more noise than 5.08 cm (2 in) of conventional cord
(McGraw Hill Inc., 1970).
Weather can also affect noise propagation. Blasting should, when
possible, be avoided on foggy, hazy, or smoky days as these conditions are
prevalent during temperature inversions. Sound propagation is increased
due to these inversions. The following weather conditions, when noted by
mine operators, indicate unfavorable conditions for noise control (McGraw
Hill Inc., 1970):
1. Relatively high atmospheric pressure that remains static more
than 25 hours.
2. Wide daily temperature variations.
3. Poor visibility and light winds early in the morning, followed
by cloudiness.
4. Clear, somewhat hazy days with little wind and fairly constant
temperatures.
145
-------
TABLE 25. MAXIMUM EXPLOSIVE CHARGES USING
SCALED DISTANCE FORMULA
Distance to Maximum explosive
nearest residence charge (to be
building or other detonated)
structure (ft) (lb)
300
350
400
500
600
700
800
900
1,000
1,100
1,200
1,300
1,400
1,500
1,600
1,700
1,800
1,900
2,000
2,500
3,000
3,500
4,000
4,500
5,000
25
34
44
69
100
136
178
225
278
336
400
469
544
625
711
803
900
1,002
1,111
1,736
2,500
3,403
4,444
5,625
6,944
Note: In all blasting operations, except as otherwise autho-
rized, the maximum peak particle velocity shall not
exceed 1 inch per second at the location of any dwelling,
public building, school, church, or commercial or insti-
tutional building. Peak particle velocities shall be
recorded in 3 mutually perpendicular directions. The
maximum peak particle velocity shall be the largest o-f
any of the three measurements. The maximum peak particle
velocity allowed may be reduced if it is determined that
a. lower standard is required because of density of popu-
lation or land use, age or type of structure, geology or
hydrology of the area, frequency of blasts, or other
factors.
146
-------
ELEVATION VIEW
O is- O
eo
0 0
0 0
O O
0 O
0 0
0 O
0 0
O 0
0 O
,
c
t/
a
o
J
)
•)
1
4
DRILL BENCH
9 DIAMETER
HOLES
•—18-25' »
\
\
COAL SEAM
TYPI
CAL L
•
1
DAD
Stemming
PLAN VIEW
Distance to nearest dwelling = approximately 5000 feet.
Amount of explosive per hole = 340 Ib.
Figure 76. Proposed blasting plan (example) (After Grim and Hill, 1974)
147
-------
6 1/4" Holes
soft shale
30'-4O
Sandstone
125 Ib
— explosive
100 Ib /|0
explosive
Irreg
shale
36"cool
75 Ib
^£/t "explosive /OMS
^/w/&s&/WW/&sw&M!wt$i0i!igw^
2l'—Jlnst Inst Inst
<*> <"> **
IOMS
IOMS IOMS
Inst
-O
and
Sandstone
Stem8mir1g|-AN"-o7l J «l50grain
2 Bock fill
5 Holes, 55 deep
on 21-24'centers
charging ratio I Ib
to 8-10cu yd of
overburden.
17MS delay*
between hole*
Shale 0-24"
Figure 77. Some commonly used drilling and blasting patterns
(McGraw-Hill, Inc., 1970).
148
-------
SURFACE
8'
SHALE
20'
GRAY
SANDSTONE
90'
COAL
94'l
1
3*9
''•* ', •
\ ', .' \
•'• ',*'•
*•'*••»'
"* l *'
"»*.".
• \ .-
"^
',
96' J
^
^
71' 77
\
^
• 7'
DETONATING FUSE
;'- SURFACE
'•:•. "'
^'. STEMMING LIMESTONE
SHALE
29'
150-LB ANFO
1 PRIMER
BROWN
STEMMING SANDSTONE
2SO-LB ANFO
2 PRIMERS
'••'. STEMMING 48'
V; RIDER COAL
) 430 -LB ANFO
3 PRIMERS
77
;>'. STEMMING
£ GRAY
SANDSTONE
^2 PRIMERS
90 LB. MS-80-3O
.'•I SLURRY
J
650 -LB ANFO ,„•
1 2 PRIMERS LIMESTONE
SOLB.MS-80-30 gg'
BLACK 64'
SLATE
«*»'• •
|
•»•'•*•
*•*«**
*•-,***.-
*'*•)
*.v» , .
•.V*ti
I3l
la
»-
37'
44.5'
51'
0
;«
;;
r"
, r
V.
.-•
V
u
)
T*
1
9
^
&
sM
V;:. ... /."'M STEMMING uwti^ 68'^^
^M 90 L^J es' °°
^1 ^1 COAL
•^ 94' 1^
ETONATING FUSE
STEMMING
2O -LB ANFO
1 PRIMER
STEMMING
80-LB ANFO
1 PRIMER
STEMMING
200-LB ANFO
1 PRIMER
STEMMING
200-LB ANFO
1 PRIMER
STEMMING
200-LB ANFO
1 PRIMER
STEMMING
S5O-LB ANFO
2 PRIMERS
SO LB. MS -80 -30
STEMMING
65O-LB ANFO
2 PRIMERS
SO LB. MS-aO-30
STEMMING
COAL
•^W
-ofi
rt-so1-
79MS | I7MS I7MS I7MS I7MS
6—X—O— X— O—X— O—X— O — X—O
XSMS
SMS I7MS I7M3 I7MS I7MS /
•X—O—X—O—X—O—X—O—X—O
CAP
Open Face
Figure 78. Metalized slurry is used to improve blasting
results with ANFO (McGraw-Hill, Inc., 1970).
149
-------
SUMMARY
Tliis chapter has briefly reviewed the fragmentation practices in
surface coal mining. Vertical rotary blast hole drills are the most common
for overburden drilling. Ammonium Nitrate Fuel Oil is the explosive that
is widely used. It is common in area mining for the drilling and blasting
unit operation to be in advance of stripping by one or more cuts. In
contour mining, where small loaders load onto trucks, the size of the blasted
overburden becomes important. Therefore, the powder factor in contour
operations is very low (varying from 1.68 cu m per Kg (1 cu yd per Ib) of
ANFO to 4.21 cu m per Kg (2.5 cu yd per Ib) of ANFO). Important environmental
considerations on which blasting designs must be based are noise, damage
to surface structures, and protection from fly rocks.
150
-------
SECTION 6
SURFACE MINE HAULROAD DESIGN
INTRODUCTION
For the purposes of this report, coal mine haulroads are defined as
roads constructed and/or used by a mine operator, which end at the pit or the
bench. These roads may cover large areas of ground, and, according to a
U.S. Department of the Interior report (1967), constitute approximately 10
percent of the total area directly disturbed by the surface mining operation.
According to the Surface Coal Mining and Reclamation Operations, Perma-
nent Regulatory Program (Public Law 95-87, Title 30, Chapter VII, Subchapter
A), road means a surface right-of-way for purposes of travel by land vehicles
used in coal exploration or surface coal mining and reclamation operations.
A road consists of the entire area within the right-of-way, including the
roadbed, shoulders, parking and side area, approaches, structures, ditches,
surface, and such contiguous appendages as are necessary for the total struc-
ture. The term includes access and haulroads constructed, used, recon-
structed, improved, or maintained for use in coal exploration or surface coal
mining and reclamation operations, including use by coal-hauling vehicles
leading to transfer, processing, or storage areas. The term does not include
pioneer or construction roadways used for part of the road construction pro-
cedure and promptly replaced by a Class I, Class II, or Class III road
located in the identical right-of-way as the pioneer or construction roadway.
The term also excludes any roadway within the immediate mining pit area.
(a) Class I Road means a road that is utilized for transportation
of coal.
(b) Class II Road means any road, other than a Class I Road, planned
to be used over a 6-month period or longer.
(c) Class III Road means any road, other than a Class I Road, planned
to be used over a period of less than 6 months.
Subchapter K of the above regulatory program, specifically sections
816.150 to 816.175, details the general considerations, location, design and
construction, drainage, surfacing, maintenance and restoration of the Class
I, II and III roads. Movement of coal, equipment and personnel within the
mine plan area may require, other than roads, railroad loops, spurs, sidings,
surface conveyor systems, and other transportation facilities. General
standards are set forth in section 816.810 to ensure the minimization of the
adverse effects to hydrology, fish and wildlife and their habitats, and public
151
-------
and private property as a result of the design, construction and utilization
of these transportation facilities. It is not the purpose here to detail all
the provisions of the regulations but a brief summary of the important re-
quirements for Class I roads will be provided.
CLASS I ROADS
General
Class I roads should be located, insofar as possible, on ridges or on
the most stable available slopes to minimize erosion. No part of this road
may be located in the channel of an intermittent or perennial stream, and
excepting for approved stream fords, all stream crossings shall be made using
bridges, culverts or other structures which are designed, constructed and
maintained as per the provisions of the law. All Class I roads shall be
removed and the land affected regraded and revegetated unless the retention
of the road is approved as part of the post mining land use or as being
necessary to control erosion.
Design and Construction
The design and construction of Class I roads shall be approved by a
registered qualified professional engineer and shall incorporate the demand
for mobility and travel efficiency, based on geometric criteria (vertical
alignment, horizontal alignment, road cuts, and road embankments) appropriate
for the volume of traffic and the weight and speed of vehicles that will be
used for coal haulage. The overall grade shall not exceed 10 percent, and
the maximum pitch grade shall not exceed 15 percent. Considerable importance
has been attached to ensuring that adequate drainage facilities are included
in the design and construction of these roads. Specifically, ditches, cul-
verts, and bridges must be provided so as to minimize 1) the erosion from the
roadway, 2) the siltation of adjacent streams, 3) the adverse impact on fish
and wildlife habitat and 4) the instability of the roadway to protect the
welfare of the public. In general, the water control system shall be designed
to safely pass the peak run-off from a. 10-year, 24-hour precipitation event or
a greater event if required by the regulatory authority.
Maintenance
Class I roads should be maintained so as to meet the approved design
standards at all times. Maintenance will include repairs to the road surface,
minor reconstruction of road segments as necessary, brush removal and revege-
tation. Roads that have been damaged by floods or earthquakes will not be
used until the damaged areas have been reconstructed.
OTHER TRANSPORTATION FACILITIES
These facilities shall be designed, constructed, and maintained, and
the area restored to 1) prevent, to the extent possible, damage to fish,
wildlife, and other environmental values, 2) prevent contribution of suspended
152
-------
solids into streamflow or runoff outside the permit area, 3) control water
quality and quantity degradation, 4) minimize erosion, siltation, and air
pollution, and 5) prevent damage to public and private property.
This chapter will briefly describe the steps necessary for the design
of haulroads, since a well-designed haulroad will minimize the problems
enunciated before.
SURFACE MINE HAULROAD
The distinguishing feature that sets a mine haulroad apart from any
other road is the duration of its usefulness. In general, most roads are
designed with a specific lifetime in mind or with perpetual existence
assumed. Haulroads may be in use for a period of days to many years, depend-
ing upon the mine plan and other related factors. Therefore, it is important
to know the estimated time a proposed haulroad will be in use. The use of
very expensive surfacing material on a short-duration haulroad would not be
justified if the road could be maintained with an additional road grader
for less total cost. In fact, in most surface coal mines, haulroads are
maintained in mined-out pits and through spoil banks for considerable
distances before the permanent section of the haulroad is reached. For this
reason, the total mine plan should be considered when planning haulroads.
HAULROAD DESIGN
The design principles used for haulroads differ little from that for
any other road; however, the roads for many large surface mines today have
to be designed to carry loads far in excess of the loads found on such high-
ways as the interstate system. With the introduction of larger equipment,
such as off-highway trucks, there presently seems to be no weight limit
with trucks weighing 355.6 x 10^ Kg (350 tons) now in existence. Unlike
the interstate highway system which limits the loads on its roads, haulroads
may be subjected to larger loads simply because the mine operator is forced
to purchase larger equipment in order to remain competitive. The ability
to increase the bearing capacity of hualroads with minimum cost is a design
feature that warrants some thought during the initial design stages.
Often, the road-building material is a mine product, crushed overburden
and/or bottom ash. Therefore, the construction of subbase should be on the
basis of the results of the soil survey and the design of traffic loads. In
area mining, since production from a mining site lasts for several years,
the topography is generally gentle, the hauling equipment used is usually
large compared to that used in contour mines, and the roads are usually
better designed. In contour mines, the roads are long and usually have to
be laid on severe grades. Water seepage from the hillsides adds to the
maintenance problem. It is also not uncommon for haulroads to be maintained
through spoil piles and old pits. Even without these, the fugitive dust
problem with vehicular traffic demands that the roads be wetted and main-
tained in good condition to keep ahead of grading and surfacing problems.
153
-------
Regular checking of ditches and culverts for drainage and incorporation of
CaCl2 or NaCl2 into the road surface will add to the mechanical stability
of the road surface. Adequate facilities for watering and water trucks must
be available in the mines.
When a haulroad is abandoned, steps must be taken to minimize erosion
and establish a vegetative cover. For complete abandonment, culverts and
other structures are removed, and the natural drainage pattern is restored.
Side ditches should be obliterated, and properly spaced grade dips or
water bars should be constructed to handle roadway cross drainage. A water
bar must be placed at the head of all pitch grades, regardless of spacing.
All road surfaces must be ripped, treated with soil amendments, seeded with
grasses, legumes, and trees, and mulched. Seeding will help stabilize the
abandoned roads, provide food for wildlife, and improve the aesthetics.
ALIGNMENT
In general terms, alignment refers to the design of the geometric
elements of the haulroad. Such things as horizontal and vertical curves,
site and stopping distance, superelevation,and maximum grade all come under
this general term. Any standard text on highway construction can be
referred to for more information (Meyer, 1969) . Specifically, with regard
to surface mines, a recent study by Skelly and Loy (1976) and the U.S. Bureau
of Mines Information Circular 8758 (1977) are useful references. Murray
(1978) provides a computer program to calculate the best alignment of a
road (i.e., road grade) for a given topographical profile, taking into
consideration the truck capacity and speed-rimpull characteristics.
Superelevation refers to the sloping of the road toward the center of
the curve. Superelevation must be provided on curves to reduce the side
friction between the tires and the road. It is recommended that the
superelevation should never be less than 0.20 meter per meter because most
road surfaces are crowned that much to provide drainage of the road from
side to side. Because of the danger of sliding off the inside of a steeply
superelevated highway when traveling slower than the design speed, the
superelevations in Table 26 are recommended.
Horizontal curves should be well laid out. When horizontal curves
are excessively sharp, the road surface should be widened to account for
the added width of passing-vehicles (Figure 79).
It would be impractical to build a haulroad with an abrupt change in
grade. When negotiating changing grades (i.e., vertical curves), one must
consider two other aspects of vertical design , namely sight distance and
stopping distance. It is recommended that a vertical curve should never
be less than 30.48 meter (100 feet) in length.
By keeping haulroads at their minimum slopes, cycle times will be
increased and wear and tear on the equipment will be reduced. Although
154
-------
TABLE 26. RELATIONSHIP BETWEEN SPEED AND RADIUS OF A
CURVE AND THE SUPERELEVATION RECOMMENDED
SPEED/MPH
RADIUS/FT
50
100
150
250
300
600
1000
10
0.04
0.04
0.04
0.04
0.04
0.04
0.04
15
0.04
0.04
0.04
0.04
0.04
0.04
0.04
20
0.04
0.04
0.04
0.04
0.04
0.04
25 30 35 40
0.05
0.04 0.06
0.04 0.05 0.06
0.04 0.04 0.05 0.06
0.04 0.04 0.04 0.05
50
0.05
155
-------
8+50
Ln
r
150'
10
-1-00
50'
H
fi
d
«
1
0.0
0.0'
1
\
I5'/FT
I'/FT
\
FLAT
0.04V FT
1
RUNOUT
75% OF
O.OZ'/FT
| <•
0.04'XFT
pi IMOI IT »
IMJIMUU 1 w
U
0:
4
0.04'
_1
•
^ —
8-1-50
PLAN VIEW
10+50
LEFT EDGE
RIGHT ft LEFT EDGE
SIDE VIEW
Figure 79. Superelevation runout considerations.
-------
it would be impractical to design haulroads so all vehicles could operate
at their maximum speed at all times, some compromise between percent grade
and cycle time can be reached.
The stopping distance (SD) for various size vehicles can be computed
using the following formula:
SD = [1/2 gt'
V]
[
(gt Sin 0 + V )
2g(U MIN - SIN°0)
where
SD
g
t
0
Vo
U MIN
Stopping distance in meter.
Acceleration of gravity, 9.81 m/sec" (32.2 ft/sec^)
Reaction time of the driver
Angle of descent in degrees
Initial speed before breaking
Coefficient of friction between break shoe and drum
Sight distance is the extent of peripheral area visible to the vehicle
operator. As stated earlier, sight distance should never be less than stop-
ping distance. The easiest and possibly the most used method of determining
sight distance is to graphically construct the vertical curve, and then from
the known stopping distance, see if the sight distance is adequate.
ROADWAY BEARING CAPACITY
Once the geometric parameters of the haulroad have been established
and the route of the road known, it will be necessary to determine the load-
bearing capacity of the various materials used in constructing the road.
Most soil materials are subject to three types of deformation, namely
elastic, consolidation, and plastic (Figure 80). The elastic deformation
will rebound to its original shape when the load is removed. The consolida-
tion deformation will not rebound, but leave the loaded area denser than the
unloaded area. Plastic deformation occurs when the water and air in the
pores of the soil combine with loading forces to displace the roadway
material (see Figure 80).
PLASTIC DEFORMATION
Plastic deformation is the most damaging type of deformation to the
haulroad surface, and is the deformation that the road must be designed
to withstand.
The initial step in this design is to determine the loads that the
road is expected to withstand. This is done by computing the wheel loads
157
-------
BEFORE LOADING
LOADED
AFTER LOADING
i
LOAD
I
LOAD
LOAD
ELASTIC DEFORMATION
I
LOAD
I
CONSOLIDATION DEFORMATION
Figure 80. Plastic deformation.
158
-------
for the heaviest vehicle that will use the road. From the manufacturer's
specifications, the axle load is divided by the number of tires on the axle
to determine wheel load. If dual wheels are on the axle, the wheel load
must be increased by 20 percent to account for the overlapping stresses
(see Figure 81).
For example, suppose there is an axle loading of 45360 Kg (100,000 Ibs) on
the rear axle of a truck with dual wheels. It is known that the wheel load
is 45360 v 4 or 11340 Kg (25,000 Ibs), but because of the dual wheels, one
must add 20 percent, or in this case, 2236 Kg (5,000 Ibs) to the wheel load
to account for the stress overlap under the wheels. The road will now be
designed to withstand wheel loads of 13608 Kg (30,000 Ibs).
The wheel loading of today's off-highway vehicles is seldom less than
78122.7 Kg per square meter (16,000 Ib per square foot), which is the
approximate bearing capacity of soft rock. Table 27 shows the approximate
bearing capacity of various materials encountered in haulroad construction.
As indicated by Table 27 few materials can support large vehicles' wheel
loads in their natural state. For this reason, designers must compute the
amounts of additional materials that must be placed over the original
surface material in order to provide the desired bearing capacity. One of
the most widely used methods for achieving this is an empirical method
developed by the California Department of Highways in 1942. This is the
California Bearing Ratio (CBR) which is the ratio of the load carrying
capacity of the test soil to that of crushed rock. The curves shown in
Figure 82 were developed for designing road surfaces with known CBR. It
should be emphasized that the CBR for any material can be determined by
taking a sample of the material to a testing lab. If the CBR is to be
assumed, some conservatism should be employed. The use of the chart,
Figure 82, is as follows. If the wheel loading is 195306 Kg per square
meter (40,000 Ib per square foot) and it is desired to build a road over
silty clay with a CBR of 4, it can be seen from the chart that the road
surface will be required to be 81 cm (32 in) above the original soil (see
Figure 83). This will allow the 18144 Kg (40,000 Ib) load which is
distributed over possibly 0.0929 square meter (1 square foot) at the surface
to be distributed over about 3.716 square meter (40 square ft) at a depth
of 81 cm (32 inches). A gravel and clay mixture can be used as material
with a CBR of 25. From the chart, it is clear that the top of the sub-
grade must be 25.4 cm (10 inches) from the surface. Crushed stone base
with a CBR of 80 can be planned. Crushed stone must be 12.7 cm (5 inches)
from the top, as can be seen from the chart.
This leaves 12.7 cm (5 inches) of material still not accounted for
between the surfacing and the top of the base. This would normally be
filled with a similar base material with a CBR of 80 or some paved surface
such as blacktop or concrete.
Materials other than crushed stone for haulroad surfaces have been
employed at some mines; however, the advantage of such surfaces is difficult
to justify economically. When roads are paved, the surface must be
protected when running tracked vehicles such as bulldozers on the road.
Spillage is always a problem on haulroads and requires the use of a grader
159
-------
TABLE 27. BEARING CAPACITIES OF MATERIALS
MATERIAL BEARING CAPACITY
LBS/FT2
HARD SOUND ROCK 120,000
MEDIUM HARD ROCK 80,000
HARD PAN OVERLYING ROCK 24,000
COMPACT GRAVEL 20,000
SOFT ROCK 16,000
LOOSE AND SANDY GRAVEL 12,000
HARD DRY CLAY 10,000
LOOSE COARSE SAND 8,000
LOOSE FINE SAND 4,000
160
-------
100,000
w
ROAD
OVERLAPPING
STRESSES
SURFACE
OVERLAPPING
25,000 25,000 25,000 25,000 STRESSES
Figure 81. Dual wheels on an axle.
161
-------
CALIFORNIA BEARING RATIO
(5 20 25 3O 4C 30 60 TO 90 tOO
5MID- :^t
Wl
5o:~5 J~"C J1
sror -y.jfci£._.
^=l,'»'0-S r"1= J1- - ''^^ i-~"—
If-
-------
SURFACE
40rOOO lt>s
I FT
Z2'
CBR = 80
CBR = SO
CBR = 25
( SAND 8 GRAVEL)
33'
!
i
»
'45°
40 FT'
CBR = 4
Ci_AY)
-------
for cleaning, even when the road is paved. The grader is used to continually
dress and smooth a crushed stone surface so the advantage of paving is
unwarranted. The major advantages of surfacing are the reduction of dust
during dry weather and the excellent drainage provided during wet weather.
EMBANKMENTS
When designing embankments upon which the load bearing subgrade and
base materials are to be placed, it is necessary to understand some charac-
teristics of the various soils to be used. There have been several methods
of soil classification of which two are used extensively, namely the
American Association of State Highway Officials' (AASHO) method shown in
Table 28 and the Unified Soil Classification. The AASHO method seems to be
more popular, possibly because it has been used for a longer period of time.
There are two other characteristics of cohesive soils that play an important
part in describing soils, liquid limit, and plastic limit. When a soil
mixture passes from a viscous liquid state to that of a plastic soil, the
moisture content at which this transition occurs is called the Liquid Limit
(LL). With further reduction in the moisture content the material becomes
more stiff and less plastic until it finally reaches a semi-solid state.
The moisture content at this point is known as the Plastic Limit (PL). The
numerical difference between the LL and the PL is known as the Plasticity
Index (PI). Granular soils have a PI = 0. Table 28 shows the various PI
for the different classifications.
Any embankment is made up of three components, soil, air, and water.
The relationship between moisture content and density is particularly
important. All soils have a moisture content at which the soil is at its
maximum density. If embankments are placed with the soil moisture content
not at this optimum value, the embankment cannot be fully compacted. This
optimum moisture content can be determined with a test known as the AASHO
density or Proctor density test. Examples of different soil types and their
moisture density relationship is shown in Figure 84.
The following tables are included as a guide in selecting embankment
material. Table 29 is a selection of soil on the basis of anticipated
embankment performance, and Table 30 lists the recommended minimum require-
ments for compaction of embankments.
When using rock as an embankment material, the size of the individual
rock pieces should not be greater than 12 inches in any dimension, and the
rock embankment should be placed in layers not to exceed two feet. The top
0.91 m to 1.52 meter (3 to 5 feet) of rock embankments should be placed in
soil. The final 30.48 cm (12 in) of embankment should be of material with
a dry density of at least 1633.7 Kg/meter cube (102 lb/ft3) at optimum
moisture. When placing earth embankments, the layers should not exceed
22.8 cm (9 inches) loose before compaction for fine-grained cohesive soil
and 30.48 cm (12 inches) for granular cohesionless soil.
The final step before actually building the haulroad is to design
adequate drainage to keep the road from deteriorating.
164
-------
TABLE 28. CLASSIFICATION OF HIGHWAY SUBGRADE MATERIALS
General classif ication
Sieve analysis,
percentage passing:
No . 10
No . 40
No . 200
Characteristics of fraction
pass ing No . 40:
Liquid limit
Plasticity index
Usual types of significant
constituent materials
General rating as subgrade . .
Granular materials
(35% or less passing No. 200)
A-l
A-l-a
50 max.
30 max.
15 max.
A-l-b
50 max.
25 max.
6 max.
Stone fragments,
gravel, and sand
A- 3
51 min.
10 max.
NP
Fine
Sand
A- 2
A-2-4
35 max .
40 max.
10 max .
A-2-5
35 max.
41 min.
10 max.
A- 2-6
35 max.
40 max.
11 min.
A-2-7
35 max.
41 min .
11 min.
Silty "or clayey gravel and sand
Excellent to good
(continued)
-------
TABLK 28. (continued)
a 1 C ] aS.H i f |ca t i Oil
Croup el ass i f It at Ion
Sieve analysis,
percen Lage pass 1 ng:
No. 10
No. 40
No. 200
Charac ter 1st. Ics of fraction
pass Ing No. 40:
Liquid limit
Plasticity Index
Usual types of significant
constituent materials . .
General rating as subgrade ..
S i 1 t-c I ay ria t e r i a 1 s
(more than 35% passing No. 200)
A-4
36 min
4 0 max
10 max
A- 5
36 min,
41 min.
] 0 ma x .
SI 1ty soils
A-6
'36 min.
40 max.
11 min.
A-7
A-7-5,
A-7-6
36 min
41min.*
11 min.
Clayey soils
Fair to poor
Classification procedure: With required test data available, proceed from left
to right on above chart, and correct group will be found by process of elimination,
The first group from the left into which the test data will fit is the correct
classification.
^Plasticity index of A-7-5 subgroup is equal to or less than LL minus 30.
Plasticity index of A-7-6 subgrade is greater than LL minus 30.
-------
TABLE 29. GENERAL GUIDE TO SELECTION OF SOILS ON BASIS
OF ANTICIPATED EMBANKMENT PERFORMANCE
HRB
classifica-
tion
A-l-a
A-l-b
A- 2 -4
A-2-5
A-2-6
A- 2- 7
A- 3
A-3a*
A-4n*
A-4b*
A-5
A-6a*
A-6b*
A-7-5
A- 7-6
Visual description
Granular materials
Granular materials with soil
Fine sand and sand
Sandy silts and silts
Elastic silts and clays
Silt-clay
Elastic Hllty clay
Clay
Max. dry-weight
range, Ib per cu ft
115-142
110-135
110-115
95-130
85-100
95-120
85-1.00
90-115
Op t imum
Moisture
range, %
7-15
9-18
9-15
10-20
20-35
10-30
20-35
15-30
Anticipated
embankment
performance
Good to excellent
Fair to excellent
Fair to good
Poor to good
Unsatisfactory
Poor to good
Unsatisfactory
Poor to fair
* Ohio mod J.f Lcat Ion.
SOURCE: OliLo Department: of Highways.
-------
TABLE 30. RECOMMENDED MINIMUM REQUIREMENTS FOR COMPACTION OF EMBANKMENTS
Class of
soil
(AASHO
M 145-49)
A-l
A- 3
A- 2- 4
A- 2- 5
A- 4
A- 5
A- 6
A- 7
Condition of exposure
Condition 1
(not subject to inundation)
Height of
fill, ft
Not critical
Not critical
Less than 50
Less than 50
Less than 50
Less than 50
Less than 50
Slope
1-1/2 to 1
1-1/2 to 1
2 to 1
2 to 1
2 to 1
Compaction (% of
AASHO max. D)
95+
100+
95+
95+
90-95*
Remarks: Recommendations for condition 2 depend u
Condition 2
(subject to periods of inundation)
Height of
fill, ft
Not critical
Not critical
Less than 10
10 to 50
Less than 50
Less than 50
Slope
2 to 1
2 to 1
3 to 1
3 to 1
3 to 1
pon height of fills.
Compaction (% of
AASHO max.
density)
95
100+
95
95 to 100
95 to 100
95 to 100
iigher fills of the
00
order of 35 to 50 ft should be compacted to 100 per cent, at least for part of fills subject
to periods of inundation. Unusual soils which have low resistance to shear deformation should
be analyzed by soil-mechanics methods' to determine permissible slopes and minimum compaction
densities.
*The lower values of minimum requirements will hold only for low fills of the order of 10 to
15 ft or less and for roads not subject to inundation nor carrying large volumes of very heavy
loads.
SOURCE: Highway Research Board (4).
-------
No.
I
2
3
4
5
6
7
8
10
12 14 16
Moisture, %
8
22
Soil Texture and Plasticity Data
Description %Sond %Silt
Well-graded loamy sand 88 10
Well-graded sandy loam 72 15
Med -graded sandy loam 73 9
Lean sandy silty clay 32 33
Loessial silt 5 85
Heavy clay 6 22
Very poorly graded sand 94 *-
'oClay
2
13
18
35
10
72
-*.
LL.
16
16
22
28
26
67
NP
P.I.
NP
0
4
9
2
40
NP
Moisture-density relationships for seven soils each compacted by the
AASHO standard method
Figure 84. Relationship between moisture content and density,
1.69
-------
DRAINAGE
Once the haulroad has been designed, it is necessary to provide the road
with ditches and culverts of sufficient size to remove all surface water
before it can damage the road embankments and surface.
If rainfall were applied at a constant rate to an impervious area, the
runoff from this area would eventually equal the rate of rainfall. The time
required to reach this equilibrium is known as the time of concentration
(tc). This value is the basis for a method of estimating runoff known as
the Rational Method. Although there are many methods employed for estimating
runoff, this method is simple and does provide reasonable results for small
drainage areas which are usually encountered in haulroads. The Rational
Method can be reduced to the following empirical formula:
Q = K i A
(1)
where
Q -
K
i
A =
Runoff in cu ft/sec
Runoff coefficient, a function of drainage area (Table 31)
Intensity of rainfall, in/hour for a duration equal to the
time of concentration t
Drainage area in acres
The time of concentration can be calculated from the following formula:
c
0.385
60 (11.9 L )
H
mins
(2)
where
L = Length of channel from headwater of watershed to the end of
the channel (miles)
H = The fall in the channel in ft
Once the time of concentration has been established the intensity i
can be calculated from the intensity frequency curve shown in Figure 85.
This type of curve can be constructed for any location. Each area will
have its own unique set of rainfall criteria. The curve shown in Figure 85
is for southern Ohio.
The use of the Rational Method is illustrated for a drainage area
adjacent to a haulroad located in southern Ohio (Figure 86). The hatched
area of 102 acres will all drain to a new culvert pipe that is to be
installed. The values of L, H and K are shown in Figure 86.
170
-------
TABLE 31. RUNOFF COEFFICIENTS
Type of Drainage Area Coefficient of Runoff, K
Concrete or bituminous pavements 0.8-0.9
Gravel roadways, open 0.4-0.6
Bare earth (higher values for steep slopes) 0.2-0.8
Turf meadows 0.1-0.4
Cultivated fields 0.2-0.4
Forested areas 0.1-0.2
171
-------
15 20 30 40 5060
Rainfall duration, minutes
Figure 85. Rainfall intensity, duration, and frequency for
the southern one-third of Ohio.
172
-------
DRAINAGE
AREA LIMIT
PIPE
A = 102 ACRES
L 3/4 MILE
H = 10'
K - 0.2
Figure 86. Drainage area adjacent to a haulroad.
173
-------
.. 0.385 0.385
60 (11.9 L ) 60 (11.9 x 0.75 ) ., .
g = >- = 46 BODS
From Figure 85, for a 10-year frequency at a t of 46 minutes, the value
for i is 2.3 inches per hour. Therefore, the value of Q can now be
calculated.
K i A
= 0.2 x 2.3 x 102 47 cu ft/sec
There are other more elaborate empirical methods for computing runoff.
Regardless of the method, all ditches and culverts must be designed to
accommodate the expected runoff.
DITCHES
The use of V ditches for controlling and channeling runoff is
economical. As the velocity of water increases with slope, it is advisable
to keep all V ditches as flat as possible and yet allow the water to run.
Several anti-erosion steps are to be taken within the ditch as the percent
slope increases. Rock lined channels are required when slopes exceed 8
percent. The rock lining controls erosion and helps dissipate the velocity
energy of the water.
CULVERTS
The size of culverts necessary to take the runoff must be adequate.
The most common type of culvert is corregated metal pipe, mainly because of
its ease of handling and low initial cost.
When placing pipe culverts in areas of traffic, it is advisable to
have a minimum of 60 cm (2 feet) of cover over the pipes; where vehicles
of 90718.47 Kg (100 tons) or greater cross the culvert regularly 91 cm
(3 feet) would be advisable. When installing pipe culverts the major
concern should be to provide lateral confinement to the pipe. The common
practice is to excavate a trench 60 cm (2 feet) wider than the outside
diameter of the pipe, place the pipe, then backfill in 15 cm (6-inch)
maximum compacted layers.
Pipes should never be placed under an embankment prior to placing the
embankment. The embankment should be placed first, and then a pipe trench
should be excavated through the embankment. The graph shown in Figure 87
shows the size of pipe required for various flow rates. It may sometimes
be advisable to place a pipe smaller than that required to take the maximum
174
-------
600
400
350
300
250
200
150
v>
- 80
o
^ 60
t 50
o
Q.
O
LU
O
z
o:
30
25
20
I 5
10
8
6
5
4
*> •*
-^
x-
s
/
x'Xx
x
XX X
X
Xx
'X
^L
X
X
X
*
x
x
-X.
7
12 15 18 21 24 30 36 42 48 54 60 72 84 96
PIPE DIAMETER, inches
Figure 87. Pipe culvert capacity graph.
175
-------
flow and let the xcater pond on the upstream side of the culvert. As an
example, suppose the flow rate is 0.2832 cu m/sec (10 cu ft/sec). From
the graph it can be seen that a. pipe larger than 53 cm (21 inches) in
diameter will be required. Therefore, a 60 cm (24-inch) diameter pipe is
selected. However, if the water could be allowed to pond a maximum of 90 cm
(3 ft), one can use a 38 cm (15-in) diameter pipe. Even 60 cm (2 ft)
ponding would allow the use of a 45 cm (18-in) diameter pipe.
SUMMARY
This chapter has briefly reviewed the design of surface coal mine
haulroads, and the environmental problems associated with them. Good haul-
road design for coal mines can hardly be over-emphasized in view of the very
heavy loads. Considerations from highway construction, logging roads, etc.,
can be transferred wherever applicable.
176
-------
SECTION 7
SLOPE STABILITY
INTRODUCTION
The earth's crust, particularly that portion termed continental,
contains a variety of rock types which are arrayed in a wide display of
geologic structures and subjected to a great variety of stresses. Some of
these structures are very complex, such as overturned and faulted folds,
others, such as undisturbed sediments, are very simple. The alteration of
ground stability by the introduction of an excavated hole will depend upon
the extent of this complexity plus a variety of other factors. The size of
the excavation complexity of geologic structure, presence of hydrostatic
pressure, and presence of tectonic stresses can all affect the stability of
the slopes surrounding an excavation. Luckily, when consideration is given
only to coal mines within the regions of interest to this report, the
factors simplify themselves greatly. Thus, this slope stability section
discusses only the structures of concern to coal mines: highwalls, low-walls,
storage piles, and to some extent, impoundments. In this section, the word
rock is used generally to cover all crustal material that is found naturally.
For the purpose of mechanical analysis, soil is merely rock with a low or
absent tensile strength.
The section is divided into three major parts. The first part, the
introduction, defines structures of concern, gives a qualitative description
of the elements of stability, and ends with a list of helpful references.
The middle part, on the theory of stability, gives a quantitative descrip-
tion of the stability elements. The final part of the section gives
practical suggestions for constructing a safe slope or for remedying a
failing one. The mine engineer should read all three sections, while the
operating manager may be content to skip the middle section.
Generally speaking, coal is overlain by sediments possessing relatively
simple and flat-lying structure. Although these strata may contain some
faults and jointing, the principal mechanical discontinuity is that created
by bedding. Since there is but little tectonic disturbance, the stresses
in the overburden are primarily due to body weight, that is,the pull of
gravity on the rock itself. Therefore, an analysis of slope stability in
these coal mines generally requires knowledge of the excavation geometry,
rock type and strength, density of bedding, strength across bedding, and
the amount and pressure of included groundwater. Many of these data should
be determined by careful examination of exploration results, before mining
begins.
177
-------
Structures of Concern
Man-made slopes in surface coal mines abound: generally, however, they
may be classified into four distinct types. Thus, the structures of concern
are highwalls, overburden or spoil piles, waste or refuse piles, and
impoundments.
Highwalls—As explained elsewhere in this report, highwall heights are
usually constrained by economic considerations, with the dimensional limita-
tions of the excavator sometimes acting as a bound. In other words, high-
wall heights are defined, through stripping ratios, by the price of the
coal and the cost of removing it. However, in dragline mining particularly,
d maximum digging depth of 50 to 65 m (150 to 200 ft), depending on machine
size, is imposed. Rarely does the mechanical stability of the highwall
itself enter into these constraints. Generally speaking, this situation ex-
ists because coal surface mine slopes are stable with little hydrostatic head
in the sediments, and because draglines mine with such speed that they can
remove an incipiently failing slope before it becomes catastrophic.
However, several specific items enter into the present consideration of
highwall stability.
1. No mine is ever totally free from wall failure. Although
small, from a mechanical viewpoint, falls of rocks can be disastrous
to personnel in the pits. Some major failures have occurred; these
can be extremely disruptive to stripping shovels and coal loading
machinery working on the pit floor, as well as equipment located on
the surface within the shear plane of the failure zone.
2. In times of expensive capital goods, lost time on excavators
due to the need to clean up slope failures can be very costly. Human
injury and fatality costs are impossible to assess realistically.
3. High-priced coal is leading to ever greater stripping ratios
and, hence, higher highwalls. Some wall heights may soon reach the
limits of equilibrium.
4. Multiple benching, either to accommodate multiple seams or
to allow for more flexible excavation of deep seams, puts large
machines on small working platforms. There is limited room available
on an intermediate bench to which a machine can be moved if a slope is
failing. Although items 1, 2, and 3 refer to limited-height dragline
benches, depths of 304.8 m (1000 feet) or more can be achieved by
truck-shovel combinations on multiple benches.
Overburden Piles—Spoil piles, often called low-walls, present another
version of the stability problem. Mechanically, the material is excavated
in discrete particles; it is more homogeneous than when in situ and
possesses little or no tensile strength. Since its strength is dependent
upon confinement, rather than using a convenient measure of strength such
as unconfined compressive strength, more complex measures such as yield
criteria and the angle of internal friction are required. Problems in
existing mines point to situations which will worsen with increasing depth
of mines.
178
-------
1. Planned stacking space may be inadequate if the swell factor
changes at different sites, or if it has been underestimated. To
prevent rehandling or covering of coal when this occurs, the temptation
is to steepen the slope of the spoil pile. While this steeper slope
may be temporarily stable, changes from increased loading or changed
weather conditions may cause rapid failure.
2. Mining around an inside curve causes increased amounts of
spoil to be deposited in one place. Again, the temptation is to
steepen the slope. This is because the fixed casting radius and
heights of most stripping machines prevent a safe spreading of the
excess material.
3. Maintaining access ramps through spoil causes excess
material to be stacked. As with curves, the temptation is to
steepen slopes.
4. The presence of water can materially alter a slope's
stability. Slopes constructed on a dry day may fail, either by
surface erosion or by massive translation, on a wet day. Obviously,
the predominant rock type will have a major control on this occurrence.
Sandstone blocks, with little recompaction and a great deal of inter-
particle porosity, will drain quickly and well and remain stable
during wet conditions. On the other hand, poorly-cemented sandstones
or shales, particularly shales with water-sensitive clays such as
bentonite or montmorillonite, may well fail during wet periods.
Storage piles—Although preparation-waste or refuse piles are not
peculiar to surface mines, they represent a surface structure which can
fail. Inasmuch as their instability may evidence itself years or even
decades after the pile was begun, as with the Aberfan disaster in South
Wales, great care must be taken in their design and construction. Similar
piles that fall into this category are those for the temporary storage of
topsoil and for the permanent storage of excess overburden in head-of-hollow
or valley fills.
Impoundments—Because of the additional problems created by the storage
of water, a separate category is given for those waste piles used as
impoundments. The dam at Buffalo Creek, which failed before it was com-
pleted, is testimony to the need for consideration of and good design for
impoundments.
Elements of Stability
Although a more explicit review of the theory of stability is presented
later, a review of the elements of stability is in order in this introduc-
tory material. These elements are often grouped, conveniently, into sliding
forces and resisting forces; that is those elements which tend to cause
failure and those which tend to prevent it.
Sliding Forces—There are two principal sources of sliding forces.
The first is the force of gravity on the rock itself or on any equipment
resting on the slope; in other words, this force is due to body weight of
the slope and the weight of any equipment on top of the slope. The second
179
-------
force is that due to any water contained within the material of the slope.
Frequently, groundwater pressure is the major contributor to slope failure.
A third force, rarely seen in the relatively undisturbed strata in the coal
fields, is that of residual stress remaining from mountain-building
activity.
Resisting Forces—Those forces that resist failure are unconfined
tensile and shear strengths and the shear strength resulting from internal
friction. Since the weakest part of an overall slope is often at bedding
or jointing planes, the unconfined strengths of concern are those measured
across these discontinuities. In addition to its pressure, water can serve
to diminish the strength or friction of some rock types. Thus, where clays
are present in the rock, water can be doubly damaging to stability.
Sliding Surfaces—The above forces act along pre-existing surfaces
such as bedding planes or joints. In slopes of homogeneous material, either
solid without pre-existing fractures or granular such as in spoil piles,
the forces will act along a curved surface that develops in such a way as
to minimize the expenditure of energy during failure. In granular material
with no internal friction, this arc will be circular; where friction exists,
the arc takes the form of a logarithmic spiral. Apparently, the use of
only circles, and not spirals, in the analysis of slopes does not materially
affect the prediction of failure. The same cannot be said for the substitu-
tion of a plane, particularly when the slopes deviate much from the vertical.
References
Rather than reference each statement in this section, a list of useful
references is given. These references encompass the statements made here,
but also provide the design engineer or consultant with a useful library.
"Slope Stability" is discussed in Section 12.1 of Surface Mining
[Pfleider, 1968]; essentially the same analysis is presented in Section 7.1
of the SME Mining Engineering Handbook [Cummins, 1973], Rock Mechanics
and the Design of Structures in Rock [Obert and Duvall, 1967] gives back-
ground work theory, but has little specific slope stability work.
A recently-released design manual, Engineering and Design Manual:
Coal Refuse Disposal Facilities [MESA, undated], is one that no surface-
mine designer should be without. It gives a thorough review of slope
stability, including extensive references, before embarking on the design
of refuse disposal facilities, both piles and impoundments.
Finally, slope stability is discussed in most soil mechanics textbooks.
Because the authors did extensive research in stability, a favored reference
is Soil Mechanics in Engineering Practice [Terzaghi and Peck, 1967]. One
other text, used by this author is Principles of Soil Mechanics [Scott,
1963], Several other papers, referred to in other parts of this chapter,
are also listed.
180
-------
THEORY OF STABILITY
Mechanics
The theories of fundamental mechanics are, in their most comprehensive
forms, complex and cumbersome. An attempt will be made here to extract
only that material needed to understand subsequent analyses.
It is reasonably well understood in mechanics that relationships among
the three principal stresses can be expressed which define a material's
yield limit. When the stresses are less than the defined relationship, the
material is stable. When the stresses reach their limit, or when conditions
attempt to push these stresses beyond this limit, the material responds by
yielding. This yielding prevents the stresses from exceeding the set limit.
At the yield point, further applications of stress on the material result
in further deformation but not in a stress rise within the material.
Although yield is sometimes called failure, it should not be confused with
rupture or collapse. Yield means deformation beyond the elastic limit; the
physical manifestation of yield may take a variety of forms from crystal
deformation to intergranular slippage to the creation of visible cracks and
failure zones. Since failure is energy conservative, microscopic evidences
of incipient failure will occur within predictable planes or zones. Further
applications of stress will cause the micro-cracks to coalesce into one
large crack. Even in granular material, high differential deformation will
occur in narrow zones that surround undisturbed blocks. This can be seen
in failing slopes, where large blocks will move as one piece along a failure
surface.
Although the subject of yield criteria is extensive and yet evolving,
particularly for three-dimensional analyses of anisotropic material, the
theory for a two-dimensional analysis such as the slip-circle analysis, to
be given later, is not that involved. In a general soil, the criterion for
stability, and thus failure, was given by Coulomb in the year 1773 as
T| <_ a tan + c (1)
That is, the shearing stress (T) on a section of isotropic, cohesive soil
must be less than the cohesion (c) plus the product of the normal stress
(a) on that same section with the coefficient of internal friction (tan).
Failure occurs when the shear stress equals the cohesion plus this value
of the normal stress. The geometry of this situation is shown in Figure 88.
In soils, cohesion can be thought of as unconfined tensile strength; in
rock, however, the tests for shear strength give a better value for cohesion
than do those for tensile strength. Since it is difficult, if not impossible
to test soils directly for their tensile strength, cohesion is usually
calculated from a series of shear tests.
181
-------
c
o; H
OJ+CT,
Figure 88. Coulomb'« failure surface.
Mohr adapted this yield condition and represented it by means of * graphical
plot of principal stresses upon stress co-ordinates, where the normal stress,
a is the abscissa and the shear stress, lf is the ordinate. Injure 88,
a', the major, and a3, the minor, are the principal stresses, c is the
cohesion, c cot<0 is ?he tensile strength, and * is the ^ '
angle of internal friction, made by the envelope surrounding and
to the stress circles and the abscissa. The envelope describes the Una ting
yield conditions of the principal stresses and satisfies the equality of
equation (1); therefore a circle that is tangential to the envelope repre-
sent a stressed soil sample that is yielding, and a circle that lies wxthxn
the envelope represents a stressed soil sample that is rigid. Under the
stated condltioL. it is impossible for the stresses %b;« '1
that a Mohr circle intersects the envelope. In terms of the
stresses, failure occurs when
2c cos*
(a., + a3)sin<|>
(2)
Considering
reduces to
cohesionless soil, c cos* must equal zero and equation (2)
^ (3)
182
-------
Equation (3) represents the maximum value of 0]_, for any 03, and any ;
however, if Figure 88 is examined, it is seen that a circle can be drawn to
the left at 03 so that aj_ is smaller than J-. This is the minimum value
of a-^ and is represented by
03/03 = (1 sin
with the principal or algebraically greater stress direction. Only in
homogeneous and isotropic material can it be shown that stress sliplines
are coincident with strain characteristics.
The implication of a positive value for $ in a material is that there
is an increase in shear strength for each increase in normal stress. Put
differently, a confined soil will be stronger than an unconfined soil. The
weight of a structure will provide the confinement necessary to create
sufficient foundation strength to support that structure, presuming, of
course, that yield is not exceeded. A zero value for <|> , as seen in cohesive
materials such as clays and most metals, means that, derived from Equation
(1),
IT] 1 c (5)
and, in terms of Equation (2), failure occurs when
k, - 0J = c (6)
A purely cohesive substance is weakened, not strengthened, by confining pres-
sures. Equation (6) is known as the yield criterion of Tresca, and is often
applied in metal plasticity.
Demonstration of the notion of yield criteria and establishment of
values for c and is done with a shear tester. Either a direct-shear
(shear box) or triaxial tester may be used. The shear box is preferred
when drained tests are to be performed and where strengths across pre-
determined sliding surfaces are desired. The triaxial tester is preferred
for homogeneous material which is allowed to form its own failure plane.
183
-------
The triaxial tester can also be arranged to manipulate pore pressure. Thus,
the rise of pore pressure due to deformation can be measured or the effect
of increased pore pressure (effective stress: see the subsequent section
on effect of water) can be simulated. Both tests are standard and can be
obtained commercially.
Effect of Geology
In the previous discussion on mechanics, reference is made to homo-
geneity and isotropy. However, earth materials, especially undisturbed
overburden, are unlikely to be uniform. Therefore, some consideration must
be given to the effect of geology. Namely, strengths and angles of internal
friction vary among the rock types; also, within a single rock type these
variables may change according to direction of measurement. Finally, the
presence of discontinuities, such as bedding planes, faults, and joints, may
reduce unconfined strengths to zero. Since friction exists between any
two surfaces, angles of internal friction are often unaffected by discon-
tinuities. This statement is modified by the fact, that many discontinui-
ties, particularly faults, are filled with gouge or other clay-like material
which acts as a lubricant and reduces the angle of internal friction.
The effect of geology, therefore, is to take the homogeneous system
discussed previously and impose upon it predetermined sliding surfaces and
highly variable strength factors. The mine engineer must analyze each
slope according to both its geometric and geologic factors. A purely homo-
geneous and cohesionless material can be analyzed on geometric factors only.
For instance, in an undisturbed highwall the sliding surface will be strongly
influenced by the strength across bedding planes. If, as is often the case,
the contact between the coal and its immediate superjacent stratum (roof)
has low or even zero strength, and if the wall fails, it probably will be
by sliding (translation) rather than by rotation. Should rotation occur,
it will undoubtedly intersect this contact plane. Since coal itself is a
weak foundation material, it would be improbable to find a rotational
failure which involved the strata below the coal. That is, it is unlikely
to find base-circle failures. Similarly, the presence of faults or joints
(cleat) will help to define the sliding surface.
Looking closely at the rocks which compose the slope, it is seen that
a high degree of variation can occur within a single geologic material.
The best example is shale, which may have a shear strength of over 1000
pounds per square inch normal to bedding and a strength near zero parallel
to bedding. Similar variations can occur in the angle of internal friction.
This angle, ranging between 30 and 40 degrees for most soil materials, is
a direct measure of surface roughness and particle shape and is closely
correlated to mineralogy. A smooth spherical particle will have a <£ of
around 20 degrees, while an angular substance can have a <{> as high as 45
degrees. Angles of internal friction over 45 degrees probably represent
interlocking particles rather than true friction. Some soil materials,
those that contain clay and most notably the shales, can have dramatic
alterations to <(> with the addition of a small percentage of water. Sand
and gravel, derived from quartz or unaltered silicate rocks, will demonstrate
cohesionless characteristics (c near zero, <|> between 30 and 40 degrees),
184
-------
while clays and clay stones will demonstrate either cohesionless or cohesive
(a measurable c, $ near zero) behavior, depending upon the amount of water
present. Obviously, shale particles, with little or no cohesive strength,
can lose all strength when suddenly wetted.
Narrowing from a general discussion to the specifics of surface coal
mines, certain simplifying conditions are seen to exist. Undisturbed strata
are flat, and maximum shear strength is parallel to the vertical; as a
corollary, maximum tensile strength is normal to the vertical. This is a
relatively favorable orientation for slope stability. However, of greatest
significance is that most slopes of interest in coal mines are composed of
broken materials. Spoil piles, refuse piles, and certainly, the blasted
portion of the highwall can be characterized as cohesionless soils. Because
the dragline so often sits on blasted material, the last is often of great
concern.
Effect of Water
In the preceeding paragraphs, mention was made of some simplifying
conditions; it is now appropriate to introduce one important, complicating
condition. The presence of water in a slope can reduce its overall stability
or even induce catastrophic failure. As discussed in other portions of the
background reports, water enters a slope as precipitation, as surface
drainage, or as groundwater. Succinctly, water can increase body v/eight,
which is a sliding force, reduce normal stress x^hich contributes to the
resisting force, and, for some minerals, reduce which is tantamount to
lubrication and thus a decrease of resisting force. The following comments
will give some guidelines on how water affects stability.
Water's Weight—All earth material contains holes or empty spaces
termed voids. Even solid rock contains cracks and fissures of inter-
particular spaces in sediments. When the rock contains a large volume of
voids, it is termed porous. When the voids are connected so that a fluid
may flow through the rock, it is termed permeable. The ratio of void
volume to solid volume is appropriately termed the void ratio. The ratio
of void volume to total volume is called porosity. These voids, whether
in rock or in soil, contain either fluid or gas. Commonly the fluid is
water, but may be petroleum liquids; the gas is usually air, but can be
hydrocarbon. For these purposes, concern lies only with air and water.
When the entire void space is filled with water, the soil is called saturated.
Percent saturation is a ratio of water volume to total void volume.
Similar to rock, soils may be permeable or impermeable, porous or
dense, according to their makeup. Soils with particle sizes of sand or
larger (greater than one millimeter) are usually porous and permeable. A
high degree of fines or clays will decrease a soil's permeability and
increase its density. This is because the fines can occupy the void spaces
between the larger particles. Naturally, for one kind of particle, the
densest soil will be well-graded, that is possessing particle sizes from the
finest to the coarsest. Distinction is made between particle density,
which is a function solely of mineralogy, and bulk density, which is an
overall measure and is connected to both particle density and void ratio.
185
-------
Concern lies with the fact that water can add weight to a slope by
filling voids. If a rock, whose density is 2.65 grams/cm3 (165 Ibs/fr3),
has a void ratio of 20 percent, then its bulk density is 2.21 grams/cm3
(138 lbs/ft3). At saturation, its density is 2.38 grams/cm3 (148 lbs/ft3)
which is an 8 percent increase. However, 20 percent void ratio equals 17
percent porosity and is quite low for soils. For a soil, with a 50 percent
void ratio (33 percent porosity) and the same particle density, bulk density
would be only 1.89 grams/cm3 (118 lbs/ft3). Saturation gives a bulk density
of 2.18 (136) or a percentage increase of 15. Since a soil is
much more likely to saturate than intact rock, this weight change should be
of concern to exposed spoil piles, refuse banks, and blasted highwalls.
Water pressure—When the voids of a soil become saturated, it is
possible for the pore water to develop a pressure which is measurably
greater than atmospheric. This pressure is dependent only upon the pore
water head and not upon its volume. Hence, even a low porosity material
can develop a pore pressure. For simplicity, assume that the water is not
flowing and that consideration is given only to the static head. The
development of pore-water pressure can lead to the most disastrous situations.
These range from improbable and feared quicksand to the collapse of the
Buffalo Creek Impoundments on February 26, 1972, which caused an estimated
death toll of 118 people.
The net effect of pore pressure is to reduce directly the normal stress
equation (1). The notion of effective stress (o~) must be introduced,
(7)
Equation (7) says that the normal stress on a surface should be the stress
due to the body weight of the substance plus any imposed loads, such as the
weight of a dragline, less the pore-water pressure (pw) . This reduced
quantity is the effective stress. Equation (1) can be rewritten
|T | <_ a tan<|> + c. (8)
The natural and logical question is what happens when pw = a? Obviously,
a" = 0, a" tan:}> = 0 and the substance will behave as a liquid if c is inconse-
quential. Since liquids have no shear strength, the effect of having a.
zero value for effective stress is to cause a zero shear strength and to
cause seemingly liquid behavior in a bulk solid. This process, called
liquefaction, can happen precipitately fast, since saturation and then
pressurization of the pores can occur with the addition of only a very small
quantity of water. Indeed, during earthquakes, such as at Hebgen Lake,
Montana in 1959, where 28 people were killed by a landslide, liquefaction
can occur due to the consolidation of the soil from the quake's vibration
and the accompanying reduction in porosity, and hence, a rapid rise in
pore pressure.
186
-------
Mine designers must guard against pore water pressure in all surface
mine structures. Areas of concern are joint fractures in a highwall which
intersect the surface and can intercept surface run-off; blasted areas of a
highwall which collect run-off; fill areas (head-of-hollow or valley) and
spoil piles which saturate due to precipitation, surface run-off, or ground-
water flow; and, most specifically, impoundment structures and their
foundation zones. Water interception structures such as diversion ditches,
free-flowing (French) drains where appropriate, and in the extreme, dewater-
ing wells should be built to keep these structures drained. Pore-water
pressure can be measured by placing packers and a piezometer in specially
drilled wells. Even without precise measurement of pressure, its presence
can be presumed if there is~a~water seep or flow from a slope's surface.
In closing, it should be noted that the so-called lubricating effect of
water is often the liquefaction of narrow failure zones. Liquefaction can
occur when a porous layer is entrapped by impermeable layers. While this is
not often a problem in undisturbed terrain, it immediately becomes so when
an excavation enhances the instability of a constrained and pressurized
layer. Liquefaction can also occur within fault or joint zones, particularly
if they are coincident with failure zones predicted by mechanics theory.
DESIGN OF SAFE SLOPES
The following discussion draws from those preceeding, but can, for ease
of reading be taken on its own. It starts with some computational schemes—
as it were, theoretical analyses—for factor of safety in a standing slope.
It then proceeds into some practical or field considerations for slope
stability. The mine designer is warned against taking the simplest approach
if it is not appropriate for his conditions. Evaluate the site and select
the approach that encompasses the given conditions.
Factor of Safety, First Considerations
Generally speaking, the mechanical stability of any substance can be
found if the stresses within that substance and the substance's strength
(stresses that the substance can withstand at the point of failure) are
known. When failure, in the sense of yield, occurs at a point, such as on
an unconfined surface, where strain may develop, then failure may progress
to the point of complete mechanical rupture. Because of confinement, it
should be remembered that yield failure need not necessarily lead to rupture.
By the use of appropriate yield criteria and computers to handle the
iterative computations, it is possible to perform such a total stress
analysis on a slope.
Because of the complexity of these total solutions, and because of an
inability to perform the iterative computations before the widespread intro-
duction of digital computers, a series of limit-equilibrium-analyses were
devised. Some of these analyses are still appropriate for the field
engineer. Basically, these analyses compare those force components inducing
187
-------
a slide against those resisting this slide. To make the problem tractable,
a rupture or slip surface has to be assumed. The ratio of resisting forces
to sliding forces along any slip surface is the factor of safety for that
surface. This is expressed in equation (9),
Factor of Safety , (9)
Moments
where the moments are the forces multiplied by the radius . Because slip
surfaces are assumed, a series of analyses using differently-located slip
surfaces is performed in order to find the minimum factor of safety for
any slope .
Sliding forces are usually due to weight and water pressure. The weight
of the slope material, its contained water, and any surface loads due to
structures or machines, resolved to the components parallel to the slip
surface, are the prime cause for sliding. Any unbalanced pore pressure can
also add to sliding. Other, more limited, sliding forces might be due to
residual ground stresses or formation pressure resulting from expansion of
wetted clays or shales.
Resisting forces would be from cohesive strength across the sliding
surface and from friction. Obviously, friction forces are a function of
the normal component of weight and the coefficient of friction. Resistance
may also be due to toe loadings, such as those that occur in base slip
circles .
The slip surface itself is often thought to be circular, although other
curves such as a logarithmic spiral are possible. Irregular slip surfaces
develop where discontinuities such as faults and joints exist in the ground.
Often, slip surfaces are a mixture of smooth curves and connecting straight
lines. This can be visualized if pre-existing geology is imposed on the
theoretical circle of Figure 88. If one of the circles were to intercept
a vertical joint which had been opened due to freezing, then the upper part
of the slip surface would be defined by the joint rather than a continuation
of the arc. Figure 89 can also be used to locate the critical circle (lowest
factor of safety) for cohesionless slopes (4) = 0) . The variable D is the
depth to a firm foundation and is expressed as a function of the slope
height (H) . Since the coal is assumed to be the solid foundation, most
surface coal mine failures will be toe failures and D H. It should be
emphasized that this simplifying condition does not exist for impoundments,
refuse piles, and fill areas that are built on soil or poor foundation
materials .
Slip-circle Analyses
A calculation of a slope's factor of safety requires some manipulative
method to simplify the determination of forces. This calculation is
delineated in the paragraphs which follow and the steps to follow are listed
188
-------
z 2
0 25 0 50 0 75 1 00
60° 60° 40° 20° 0°
SLOPE ANGLE, i
CENTER COORDINATE FOR CRITICAL CIRCLES
f = 1 5 H
5 20c
= 09H | 1a = ' 65M
€ IOC
[01 TOE CIRCLE
BASE CIRCLE
Figure 89. Critical slip circles for purely cohesive slopes, =0.
in Table 32. Once a rupture surface on which failure will occur is assumed,
the slope is divided into slices and each slice is assumed to have the
chord, defined by its sides, as its base rather than the intercepted arc.
Ten or fifteen such slices give almost as accurate a solution as using the
arc itself. Five or s±x such slices often suffice, for the straight-based
slice method gives a more conservative factor of safety than a circle-
based slice. The method of slices is shown in Figures 90 and 91. Figure 90
is an overview of the method and represents the two-dimensional plane normal
to the slope face. One assumption in this method is that the third dimension
has no effect on the analysis. This assumption is called plane strain.
Figure 91 shows the force on an individual slice.
189
-------
TABLE 32. STEPS IN COMPUTING SLOPE SAFETY FACTORS
1. Draw a true cross-sectional profile of the slope to be analyzed.
If the location of bedding, faulting, jointing, or other mechanical dis-
continuities is known, then place these on the cross-section.
2. Impose a slip circle on the slope. For a first trial, consult
Figure 87, and use the following configuration.
a) Highwall over coal, toe circle.
b) Overburden or refuse pile on solid foundation: slope
circle or toe circle.
c) Overburden or refuse pile on weak foundations: base circle.
d) A bench in homogeneous material (highwall and foundation
are in the same material): base circle.
3. If groundwater is known to be present, impose the phreatic surface
(groundwater table) on the slope.
4. Divide the slope into equal-thickness slices as shown in Figure 90,
Determine the weight, including groundwater of each slice. Assign an angle
of internal friction and a value of cohesion for each slice base. Be
conservative and consider worst-case conditions such as in wet or wintery
weather. Measure a for each slice.
5. Determine or assign pore-water pressures, if any, for each slice.
6. Calculate the sliding force according to equations (10) and the
resisting force with equation (11) for each slice.
7. Summarize the forces and determine the factor of safety according
to equation (12).
8. Repeat this procedure starting with step 2. However, alter the
position of the slip circle. At this point, experience has been gained
with the analysis, it is wise to try other slope slip surfaces. Instead
of a complete circle, for instance, a partial circle truncated by fault
or joint cutoffs can be tried. If a plane sliding surface is observed in
the slope face, then this should be used also as the base for slices.
Continue repeating the analysis until confident that the lowest factor
of safety has been calculated.
9. If this factor of safety is less than 1.5 for a working slope,
or 4.0 for a permanent slope, then it is necessary to redesign the slope,
altering either slope height or overall slope angle or both.
Looking at the idealized slice, it is seen that the sliding force is
the component of the total weight, Wt (Wt = weight of solid, Wg + weight
of water, Wy + weight of overlying equipment or structures, We) which acts
parallel to the base or
F W sin u, (10)
s t
190
-------
23
6 = Slope Angle
Figure 90. Slices on a toe circle for use in slip-
circle analysis of slope stability.
*el
Figure 91. The forces on slice 5,
191
-------
where a is the angle between the normal and the vertical. For the purposes
of calculation, the slice can be considered to be one foot in depth; this
allows the multiplication of the slice area by its assigned density to
obtain W .
The resisting forces are the combination of shear strength due to
friction and cohesion as defined in equation (8) . It should be noted that
the method of slices compares forces, not stresses, and that the equations
in this section refer to the forces on a slice, not stresses (force per unit
area) . Since the friction or resisting force is a function of the effective
normal force on the sliding surface, equation (11) can be written,
FR (Wt kl V cosa tan*kl + c>Lkl
where W is the total slice weight, Pw^ is the pore-water pressure acting
on line kl, L^-j is the length of line kl, (ji^i is the angle of internal
friction along kl, and r' is the effective cohesion along the line kl .
Effective cohesion refers to the cohesion that exists when wetted. If
there is no groundwater and the slope is dry, then <•* ' c. Substituting
in equation (11), using moments instead of forces, gives the analysis for
any one circle .
RE (W PwL)cosa tan* + RZc'L
Factor of Safety = RI Wfc sina (12)
An equivalent value for the denominator is
Sliding Moments = ZW r = RE W sina.
When performing the analysis, it is often helpful to tabularize the
variables. It must be remembered that the analysis should be repeated
with different circles until a minimum factor of safety is found. Both
the circle's radius and the circle's center should be varied. If it is
impossible to obtain a factor of safety greater than that required by
regulation, then the mine engineer must consider changing the slope itself
by altering either H or 0 or both.
Factor of Safety, Additional Considerations
The analysis just presented has several simplifying conditions implicit
in it. However, the method of slices will accommodate a more complex
approach. For instance, the weight of a slice was given as if the density
192
-------
were uniform throughout the slice. It is relatively easy to impose layered
geology over Figure 90 and calculate Wt as the aggregation of the weight
of each layer present in the slice.
No mention was made of the net effect of pore water pressure on the
sides of the slice and its contribution to sliding. Returning to Figure 91,
the average pore-water pressure acting on side dk (Pw^) times the length
of dk that is pressurized (the pore-water pressure need not act along the
entire side) minus the similar force on side el gives an additional component
of sliding. Converting this force to a moment and adding it to the
denominator has the effect of lowering the factor of safety. This calcula-
tion for side pressure is not precise because it assumes that the side
pressures act on the same horizontal plane and do not create a moment
couple. A computerized analysis of slices has been developed to handle
these unbalanced forces, but its use goes beyond the level of this report.
Additionally, the analysis given assumes that each slice stands alone
and shares no strength with its neighboring slice. This assumption is
valid for broken material but not for intact geologic material. The more
sophisticated computer analysis (Morgenstern and Price, 1965) mentioned
above is capable of handling this additional strength variable. The effect
of including this side strength is to increase the safety factor so that
ignoring it tends to be a conservative move. Finally, the analysis ignores
residual ground stresses due to mountain-building activities. It also
ignores poro-water pressures due to the movement of groundwater.
The U.S. Bureau of Mines has published a report (Wang et al, 1972),
RI 7685, which describes a computer program for slope stability that combines
a total stress analysis with a limiting equilibrium analysis. Basically,
the program determines the most likely sliding surface by using a finite
element stress analysis. It then determines the factor of safety for that
surface. For the computer-oriented engineer, this approach gives the most
realistic and rapid solution to stability determinations.
It should be noted and emphasized that the method of slices is not
synonymous with slip-circle analysis. Although, in the example in Figures
90 and 91, slices were used with a slip circle, the method of slices is
equally appropriate for any sliding surface. For a plane surface, such as
overburden stacked on the outslope of a contour mine, slice analysis can
be done, but forces are used instead of moments. Equally, slices can be
used with a sliding surface determined from existing geology rather than
with an assumed slip circle. If the presumed sliding surface is based on
existing conditions, only one factor-of-safety analysis need be performed.
As a precaution, sliding surfaces other than the obvious one should be
analyzed to be certain that the lowest factor of safety for the slope is
known.
Constructing a Safe Slope
In review, the following elements, when varied, contribute to or
diminish slope stability: slope geometry (height and angle); the physical
properties of the slope material (weight, angle of internal friction, and
193
-------
cohesion); geology (structure and presence of discontinuities); the
presence of water (its added weight, pore-water pressure which diminishes
shear strength, and the possibility of liquefaction); and the strength of
the foundation soil or rock which underlies the slope. Slope stability is
obtained by varying those factors in the control of the engineer in response
to weaknesses or, if lucky, strengths over which there are no control. Thus,
a safe slope is one in which the properties of the slope are known and the
slope's height, angle, and method of construction are controlled in response
to these properties. The next few paragraphs deal with those elements of
stability that can be controlled realistically in coal mines.
Slope Height—In many cases the height of a slope is fixed by premining
conditions. However, heights can be reduced by use of top benches or pull
back of spoil piles. In multiple-seam mining, it may be necessary to remove
an upper seam and its overburden completely in order to limit eventual
higtwall heights. Concern should be felt when an analysis shows that the
factor of safety is below 1.5. Obvious signs of impending failure (see
section on monitoring) should lead to immediate corrective action. However,
it must be borne in mind that slope stability results from a combination of
factors. Since slope height and slope angle both affect stability, it is
better, for permanent design changes, to change the slope angle. Nonethe-
less, the following conditions should be watched.
1. Continuing incipient failure of a dragline bench (tension cracks
and slumps) may be an indication that the bench is too high. Since other
factors such as dragline productivity through cycle time reduction may be
improved, an advance bench may be beneficial. This advance bench, cut by
mobile machinery or dragline chopdown, lowers the working highwall.
2. Spoil piles which are allowed to become too high should be watched
and repaired. A particular problem is the height of spoil on an inside
curve in area mining. Because the dragline continues to dump at relatively
the same point as it moves around the curve, the pile builds up above normal
height. The excess material should be pulled back by mobile equipment
because slope failures would cover valuable coal. The same type of problem
exists where a haulroad ramp goes into the pit through the spoil. The spoil
at either side of the ramp is higher than normal. Slope failure here could
close the haulroad.
3. Refuse piles or overburden storage areas should never be allowed to
go above their design height. Regulation may mandate maximum size of these
piles; but even if not regulated, their permanence demands a cautious
approach to factor of safety.
Slope Angle—Many of the permanent remedies for failing slopes involve
reducing the slope's overall angle. Most slope angle reduction methods
have the effect of increasing the horizontal dimension of the mine. Thus,
a reduction in highwall angle will require an increase in the casting radius
of a dragline. A reduction in slope angle of overburden piles, either
stacked on a slope or in fills, will require more land area to dispose
equal amounts. The following methods elaborate on slope angle reduction.
194
-------
1. Benches cut into a slope have the effect of reducing a slope's
angle. As each bench's width is increased, or as its individual height is
reduced, the overall slope angle is decreased. Thus, the use of benches,
regularly spaced, is an important way of maintaining stability in deep-haul
pits. As stated before, many strip mines use at least one bench, the
advance dragline bench, and its design and location should consider stability
factors. Multiple-seam coal strip mines give the opportunity to use
multiple benches for increased slope control.
2. Even though it is not needed for overall slope reduction, all high-
walls which have personnel or vehicles under them should have a safety
bench to catch falling rock. Even the most stable slopes will have some
loose rock on its face. Loosening can be due to prior blasting or to
freezing and weathering. Leaving a safety bench of some five meters (15
feet) in width, no more than one third of the way up the slope, is a good
way to safely catch falling rock. In this case, the safety bench is most
effective for steep slopes because the falling rocks have little horizontal
movement when falling and do not bounce or bound over the safety bench. In
the most protected cases, such as public highways, a chain-link fence is
provided to catch falling rock.
3. Sometimes a bench is built up at the base of an unstable slope.
This transported material increases toe stability, and decreases the overall
slope angle. Such built up toes can be used for unstable refuse or over-
burden fill piles. If the transported material is taken from the top of a
slope and placed at its toe, both slope angle and slope height are reduced.
4. The twenty-degree limitation (Public Law 95-87) on out-slope
stacking and the current requirement for approximate original contour, have
greatly reduced the concern for slope stability of spoil on sloping founda-
tions .
Stabilization of Slope and Foundation—Until now, the physical
properties of the slope material have been taken as a given. However,
when constructing an overburden or refuse pile, there are many steps that
can be taken, other than changing the pile geometry, which enhance the
pile's stability. These steps improve the foundation and strengthen the
fill.
1. The fill should not go on top of vegetative debris such as brush,
shrubs, branches, or general forest matter. Vegetation rots or burns and
eventually looses all strength which can create settlement and piping;
rotting debris is slippery and can act as a lubricant. Finally, rotting
vegetable matter can lower effluent water quality below acceptable limits.
2. Top soil is too valuable to bury with refuse or spoil. Addition-
ally, because of its organic, clay, and water content it makes a poor
foundation. Material piled on top soil can fail in the base-circle mode.
Top soil should always be stripped from a pile site.
3. Some sub-soil may also be poor foundation material, but yet be
too thick to be stripped. If the pile cannot be built elsewhere, some steps
195
-------
can be taken to improve the bearing capacity of the subsoil. Compaction of
the soil will improve it, but maximum compaction occurs only at optimum
water content. Dry soils may have to be wetted, or, conversely, wet soils
dried. Wetting is by spraying, but drying depends both upon the weather
and mechanical exposure (plowing, scarifying, etc.) of the soil to the sun.
Proctor compaction tests on the soil with varying degrees of contained water
will identify the optimum water content. Sometimes a poor soil can be
improved by adding crushed stone. In all cases, good drainage must be
provided. When the foundation material is less strong than the pile's,
more sophisticated computation of bearing stability should be undertaken
than has been described here. A soil mechanics consultant is suggested.
4. It is possible to increase the density and cohesion and decrease
the porosity and permeability of cohesive soils and rocks that are put into
piles by placing them in horizontal layers and compacting these layers as
they are placed. Although compaction will have little effect on purely
cohesionless material, most refuse and strip mine overburden has a high
proportion of shale, mudstone, or other clay-containing materials. These
cohesive materials should be dumped on a horizontal surface and then spread
into two-to four-feet thick layers. The layer thickness limitation is so
that the weight of the spreading machine will be effective throughout the
layer. Layers that are too thick will not be compacted as the stresses
from the machine's weight will be too diffuse. High-bearing load spreaders
such as rubber-tired dozers or even spreader-compactors should be used in
favor of low-bearing machines such as track-type dozers. This form of
construction is imperative for refuse piles and strongly suggested for over-
burden fills where the overburden could be classed as cohesive in nature.
Layering and compacting has decided advantages in reducing spontaneous
combustion in refuse piles. It also reduces water infiltration and associ-
ated pollution from these piles. Finally, it is a good way of reducing
the potential man-made sliding surfaces within a pile. By insuring that
all layers are horizontal and by locking one layer to another through com-
paction, the natural stability of flat-lying sediments is emulated.
Drainage—Generally speaking, a permanently-stable slope is one that
does not contain excessive water. Mine slopes, either in the highwall or
in the overburden, should be protected from surface water by diversion
ditches. Built-up slopes, such as refuse piles or valley fills, especially
if they are on sloping ground, should have foundation drains. Existing
slopes that show signs of saturation—seepage can be partially dewatered by
horizontal wells or drains that are drilled into the base of the slope.
These are particularly effective if they intercept vertical fractures or
joints. Although good drainage is of paramount importance, it is a complex
subject that is handled well in other manuals.
Other Stabilizing Actions—There are a variety of ways to stabilize a
slope that have not been discussed. These include cable or steel-bar
reinforcement, grouting, retaining walls, and so forth. Most of these
methods are expensive and not often used in mining. However, there may be
occasions when they are required. For instance, if the slope above a
196
-------
major haulroad or other permanent facility, such as a tipple, begins to
move, it may be necessary to stabilize it by some complicated means, such
as tie rods; there is no room or no time to take the simpler steps that
have been described. Again, a consultant should be hired.
For convenience, a variety of slope stabilizing methods are listed in
Table 33 (Kimball, undated). Both methods that have been described and
those that have only been alluded to are included.
Monitoring
The final portion of this section deals with the monitoring of existing
slopes to determine their stability. This monitoring can range from
casually collected visual clues to the most intensive observations by
scientific equipment. Three categories are examined: visual clues, field
surveying, and special slope monitoring equipment.
Visual Clues—Frequently, incipient slope failures are spotted by
recognition of surface clues by knowledgeable and responsible mine workers.
These people, who should be taught to recognize the signs of impending
failure, are encouraged to report their concerns immediately. Optimally,
corrective action can be taken to repair the slope; at a minimum, personnel
and equipment can be removed from the dangerous area.
Water in a bank is a potential hazard. Therefore water seepage from
the face of a bank or from the toe of a bank, unless from a designed drain
point, is an early warning sign of problems.
Cracking is a sure sign of slope movement and potential massr-e failure.
Look for tension cracks at the upper part of the slide. If failure is soon
to occur, the downhill side of the tension crack will be lower than the
uphill side. En echelon shear cracks will appear at the sides of the
slumping material, and there will be a zone of uplift or bulging at the
base of the slump. The uplift zone, as it spreads, can also show signs of
tension cracks. Unlike the crest cracks which are normal to the slide
direction, the toe cracks may well be oriented parallel to the slide
direction.
If vegetation covers a slope, then these cracks may not be seen readily,
However, leaning trees or bare patches of soil, called soil stretching, may
indicate movement. Terracettes, which look like regular cattle traces
around a hillside, are signs of soil creep. Other indicators of movement
are surface features that were installed vertically, or in a row, or both,
and that are now out of alignment. Fence lines, stone walls, roads, pole
lines, or pipelines that are out of alignment toward the downhill slope
may be indications of creep or slumping.
Field Surveying—Actual displacement of sliding can be measured by a
variety of field surveying techniques. The precision of measurement is
limited only to that of the measurement system itself; there are no
peculiarities, in other words, that reduce a method's effectiveness. The
197
-------
TABLE 33. SUMMARY OF METHODS FOR PREVENTION AND CORRECTION OF LANDSLIDES
Effect on
Stability of
^andslide
Not affected
Deduces
shearing
stresses
Reduces
shearing
stresses and
increases
shear resis-
tance
Method of
Treatment
I. Advoidance
Methods
A. Reloca-
tion
II. Excavation3
A. Removal
of head
B. Flatten-
ing of
slopes
C. Benching
of slopes
D. Removal
of all
unstable
material
III . Drainage :
A. Surface:
1. Surface
ditches
2. Slope
General Use
Pre-
ven-
tion
X
X
X
X
X
X
X
Cor-
rec-
tion
X
X
X
X
X
X
X
Frequency of
Successful Usel
Fall
2
N
1
1
2
1
3
Slide
2
1
1
1
2
1
3
Flow
2
N
1
1
2
1
3
Position of
Treatment on
0
Landslide'1
Outside slide
limits
Top and head
Above road or
structure
Above road or
structure
Entire slide
Above crown
Surface of
moving mass
Best Application
and Limitations
Most positive method
if alternate location
economical
Deep masses of
cohesive material
Bedrock: also en ten-
sive masses of co-
hesive material
where little material
is removed at toe
Relatively small
shallow masses of
moving material
Essential for all
types
Rock facing or pre-
vious blanket to
control seepage
I-1
VO
00
(continued)
-------
TABLE 33. (continued)
Effect on
Stability of
Landslide
Reduces
shearing
stresses and
increases
shear resis-
tance
(cont'd)
Method of
Treatment
3 . Regrad-
ing
surface
4 . Sealing
cracks
5. Sealing
joint
planes
and
fissures
B . Sub drain-
age
1. Horizon-
tal
drains
2 . Drain-
age
trenches
3. Tunnels
General Use
Pre-
ven-
tion
X
X
X
X
X
X
Cor-
rec-
tion
X
X
X
X
X
X
Frequency of
Successful Use
Fall
1
2
3
N
N
N
Slide
1
2
3
2
1
3
Flow
1
2
N
2
3
N
Position of
Treatment on
Landslide^
Surface of
moving mass
Entire, crown
to toe
Entire, crown
to toe
Located to in-
tercept and re-
move subsurface
water
Best Application
and Limitations
Beneficial for
types
Beneficial for all
types
Applicable to rock
formation
Deep extensive soil
mass where ground
water exists
Relatively shallow
soil mass with
groundwater present
(continued)
-------
TABLE 33. (continued)
Effect on
Stability of
Landslide
Increases
shearing
resistance
Method of [
Treatment
IV. Restrain-
ing
Structures :
A. Buttres-
ses at
foot
1 . Rock
fill
2. Earth
fill
B. Cribs or
retrain-
ing
walls
C. Piling:
1 . Fixed
at
slip
surface
General Use
re-
en-
ion
X
X
X
Cor-
rec-
tion
X
X
X
X
Frequency of
Successful Use
Fall
N
N
3
N
Slide
1
1
3
3
Flow
1
1
3
N
Position of
Treatment on
Landslide^
Toe and foot
Toe and foot
Foot
Foot
Best Application
and Limitations
Bedrock or firm soil
at reasonable depth
Counterweight at toe
provides additional
resistance
Relatively small
moving mass or where
removal of support
is negligible
Shearing resistance
at slip surface in-
creased by force
required to shear or
bend piles
O
O
(continued)
-------
TABLE 33. (continued)
Effect on
Stability of
Landslide
Increases
shearing
resistance
(cont'd)
Primarily
increases
shearing
resistance
Method of
Treatment
2. Not
.fixed
at
slip
sur-
face
D. Dowels
in rock
E. Tie-rod-
ding
slopes
V. Miscellan-
eous Methods
A. Hardening
of slide
mass:
1. Cementa-
tion or
chemical
treatment
(a) At foot
(b) Entire
slide
mass
General Use
Pre-
ven-
tion
X
X
Cor-
rec-
tion
X
X
X
X
X
Frequency of
Successful Use-*-
Fall
N
3
3
3
N
Slide
3
3
3
3
3
Flow
N
N
N
3
N
Position of
Treatment on
Landslide
Foot
Above road or
structure
Above road or
structure
Toe and Foot
Entire slide
mass
Best Application
and Limitations
Rock layers fixed
together with dowels
Weak slope retained
by barrier which in
turn is anchored to
solid formation
Non-cohesive soils
Non-cohesive soils
to prevent movement
(continued)
-------
TABLE 33. (continued)
Effect on
Stability of
^andslide
Primarily
Increases
shearing
resistance
(cont'd)
Method of
Treatment
2. Freezing
3. Elector-
osmosis
B. Blasting
C. Partial
removal
of slide
at toe
General Use
Pre-
ven-
tion
X
X
Cor-
rec-
tion
X
Frequency of
Successful Use-*-
Fall
N
N
N
N
Slide
3
3
3
N
Flow
3
3
3
N
Position of
Treatment on
Lands 1 ide^
Entire
Entire
Lower half of
landslide
Foot and Toe
Best Application
and Limitations
temporarily in
relatively large
moving mass
Effects hardening of
soil by reducing
moisture content
Relatively shallow
cohesive mass under-
lain by bedrock.
Slip surface disrupt-
ed; blasting may
also permit water to
drain out of slide
mass
Temporary expedient
only; usually dec-
reases stability of
slide
ho
o
N)
1 1 = frequently; 3 = rarely; N = not considered applicable
2 Relative to moving or potentially moving mass. 3 Exclusive of drainage methods.
-------
selection of any particular method will depend upon the conditions on the
slope and the equipment available to the engineer. Some of the more useful
methods are listed.
1. The vertical descent of permanent markers can be obtained by running
a line of levels from a stable zone across the slump to the next stable zone.
Short lengths of pipe or concrete are sunk into the soil; periodic checks on
their level gives their vertical movement. This survey is quick, cheap, and
possible with normal mine equipment. It does require a man to cross the
slump zone and does not give horizontal displacement.
2. Several surveying methods which use a transit can improve leveling.
Again, permanent markers are installed. To get total displacement, vertical
angles are taken as well as horizontal angles from a measured base line.
Although triangulation computation is more complicated, nobody is required
to traverse the slump. Should the slump be stable enough to cross, however,
then stadia measurements can reduce computational work. Although a high-
order transit traverse can be done across a slump, triangulation is
preferred.
3. Electronic Distance Measurers (EDMs), particularly those that embody
a transit for angular measurement, substitute well for transits and expedite
the survey. However, the targets are not disposable, and thus, a worker is
required to cross the zone. Modern EDM's are reliable, replicable, fast
and easy to use. They are, however, relatively expensive and subject to
deterioration if not used often.
4. Photogrammetry is an excellent way to observe ground motion. The
camera records the position of permanent markers; displacements are scaled
from the photograph by measuring distances between markers in the slump and
one on stable ground. Although aerial photography gives horizontal dis-
placements more precisely than vertical, oblique stereo photography will
give both. Actually, photogrammetry does not require an airplane and good
results can be obtained with a tripod-mounted camera.
Special Monitoring Equipment—If more sophisticated (although not
necessarily more accurate) measurement of slope failure is desired, there
is a variety of instruments capable of giving displacements, strains, tilts,
or record of acoustic emissions. Home-made extensometers and the most
sophisticated telemetric monitors have been used on moving slopes. Some of
the reasons for using this equipment over field surveying are continuous
output, unattended performance, information from depth, or scientific
faddishness. A slope stability consultant will be glad to advise whether
any of this equipment is appropriate. Frankly, because much of it is fragile
and prone to failure from tampering, it should be used only when normal
surveying appears to be inadequate.
Assessing Displacement Information—To close this section, it is
necessary to suggest what can be done with monitoring information, without
requiring a textbook on rock mechanics. It is suggested that one step is
most important. Displacement data should be plotted against time so that
total displacement from the beginning of measurement and incremental
203
-------
displacements can be observed. The mathematical slope of such a plot is
the displacement rate. If the rate is increasing over time, then the mine
slope is surely failing. If the rate is decreasing—the graph's slope is
flattening—, then the mine slope is stabilizing.
By reasonable extrapolation of a displacement-time plot, predictions can
be made of when a slope will fail disastrously. Although continuing but
small displacements can be considered a failure, they are of no consequence
to a working slope that may soon be removed or stabilized. On the other
hand, continuing displacements in a permanent slope are signs of need for
immediate, corrective action. If these small displacements get larger day
by day, then a day will come when the displacement is total.
SUMMARY
This chapter has given a three-part analysis of slope stability. The
introductory and concluding sections can be read independently to get a
descriptive review of slope stability elements and the diagnosis of slope
failure. If engineering calculations of safety factors are required, then
the middle section should also be read. For the engineer who is deeply
involved with stability analyses, further references are suggested.
204
-------
SECTION 8
OPERATING CONSIDERATIONS
INTRODUCTION
In judging the performance of stripping equipment, several material and
equipment characteristics need to be taken into account. Given the best of
conditions, a piece of equipment may not achieve the optimum level of pro-
duction due to operational conditions. In practice, the mechanical effi-
ciency of machines never equals one hundred percent, and designed production
is rarely realized. Therefore, several correction factors must be applied
to adjust equipment performance from a theoretical maximum to that realiz-
able in practice.
CORRECTION FACTORS
The correction factors arise from machine design, material character-
istics, and modes of operation.
Swing Factor
The superstructure on which the dragline and shovel revolve can swing
360° in either direction. The cycle time of the machines is affected by
the swing angles. The factor that determines the effect of swing angle
variations on cycle time, with 90° taken as the normal, is called the swing
factor. The smaller the angle of swing, the less the cycle time and the
more cycles of operation completed per hour. For example, an angular
increase from 90° to 120° for a shovel will increase the cycle time
approximately 10 percent (Tables 34 and 35).
Depth Factor
There is an optimum cutting depth for a dragline or a shovel. When the
equipment operates at any depth other than this optimum, machine production
decreases. The correction for operating at different depths is usually
made in the cycle time calculation. Tables 34 and 35 incorporate the
effect of depth.
Operating Time Per Hour (MPH)
This factor is applied to derive the machine working time from the
theoretical 60 minutes per hour. For example, if a machine works 48
minutes every hour, the MPH factor is 0.80 (48/60) (Table 36).
205
-------
TABLE 34. EFFECT OF DEPTH OF CUT AND ANGLE OF SWING
ON SHOVEL CYCLE TIME (DREVDAHL, 1961)
Depth of
cut in %
of Optimum
40
60
80
100
120
140
160
Angle of Swing
45
1.
1.
1.
1.
1.
1.
Production
given
depth
o
93
10
22
26
20
12
03
at a
and
60
1.
1.
1.
1.
1.
•
0
89
03
12
16
11
04
96
75
*
1.
1.
1.
,
•
0
85
96
04
07
03
97
90
90
.
.
1.
.
.
•
Production
= optimum
angle of swing
and
90°
o
80
91
98
00
97
91
85
at
depth
in Degrees
120°
.72
.81
.86
.88
.86
.81
.75
X
swing
150°
.65
.73
.77
.79
.77
.73
.67
Factor
for swing
and depth
180°
.59
.66
.69
.71
.70
.66
.62
206
-------
TABLE 35. EFFECT OF DEPTH OF CUT AND ANGLE
OF SWING ON DRAGLINE CYCLE TIME
Depth of
cut in %
of Op t imum
20
40
60
80
100
120
140
160
180
200
Angle
30°
1.06
1.17
1.24
1.29
1.32
1.29
1.25
1.20
1.15
1.10
Production
given
angle
depth
45
1.
1.
1.
1.
1.
1.
1.
1.
1.
at
o
99
08
13
17
19
17
14
10
05
00
a
60
1.
1.
1.
1.
1.
1.
1.
•
•
o
94
02
06
09
11
09
06
02
98
94
75
.
1.
1.
1.
1.
1.
.
,
•
of Swing in Degrees
90
90
97
01
04
05 1.
03
00
97
94
90
o
87
93
97
99
00
985
96
93
90
87
Production at
and =
of swing
optimum depth
and
90
0 swing
X
120° 150°
.81
.85
.88
.90
.91
.90
.88
.85
.82
.79
Factor
for swing
and depth
75
78
80
82
83
82
81
79
76
73
180°
.70
.72
.74
.76
.77
.76
.75
.73
.71
.69
207
-------
TABLE 36. MINUTES OF OPERATING TIME PER HOUR AND
EQUIVALENT OPERATING EFFICIENCIES
Minutes per hour
60
55
50
45
40
35
30
25
20
15
Efficiencies, %
100
92
83
75
67
58
50
42
33
25
208
-------
Bucket-Fill Factor
The bucket-fill factor is sometimes referred to as carry factor or
dipper factor for shovels and bucket factor for draglines. This factor
expresses the actual quantity carried as a fraction of the nominal dipper
or bucket capacity, and is a function of the digging conditions (Table 37) .
These conditions are classified as easy, medium, or hard digging. Since
the values describing the conditions are relative, detailed field tests
should be conducted to determine the fill factors.
Load and Swell"Factors
The load factor (LF) relates the bank volume to the loose volume handled. This
factor is particularly important in contracted overburden removal, since
payments are based on either bank or loose volume. Mathematically, the
load factor is equal to loose density divided by bank density (Table 38).
It is expressed as a fraction. On the other hand, the swell factor (SF) is
the percentage increase in volume of material from bank state to loose
state. The relationship between load factor and swell factor is as follows:
LF 100
LF
ioo + sr
Other Factors
There are several other factors such as management and operator effi-
ciencies and climatic conditions that affect equipment performance. These
cannot be accurately quantified. Atkinson (1971) has divided both the job
conditions and management practices into four categories each, from
excellent to poor and has provided a matrix of values for various combina-
tions of job conditions and management conditions (Table 39). Drevdahl
(1961) has also presented a similar table of correction factors for
combined job-management conditions.
For some earth-moving equipment, a more detailed breakdown of these
factors is possible. For example, Caterpillar Tractor Company has provided
correction factors for dozers for grade, visibility, type of materials,
and mode of operation. Since values for these factors are specific to
particular equipment, they are discussed and presented in the report along
with the equipment production calculations.
CYCLE TIME
The cycle time of stripping equipment consists of two distinct
components, the fixed and the variable. For a stripping shovel, crowding,
dump, digging, and retract times are constant or fixed in every cycle.
209
-------
TABLE 37. FILL FACTOR (DREVDAHL, 1961)
Easy Digging
Medium Digging
Shovel Dipper Factor
85% to 100%
Shovel Dipper Factor
80% to 90%
Loose, soft, free running
materials
Close lying, which will fill
dipper or bucket to
capacity and frequently
provide heaped load.
Overload compensates for
sxvrell of material
Harder materials that are
not difficult to dig
without blasting, but
break up with bulki-
ness causing voids in
dipper or bucket.
Dry sand or small gravel.
Moist sand or small gravel.
Loam.
Loose earth.
Muck.
Sandy clay.
Loose clay gravel.
Cinders or ashes.
Bituminous coal.
Very well blasted material.
Clay wet or dry.
Coarse gravel.
Clay gravel, packed.
Packed earth.
Anthracite coal.
(continued)
210
-------
TABLE 37. (continued)
Hard Digging
Rock
Shovel Dipper Factor
70% to 80%
Shovel Dipper Factor
40% to 60%
Materials requiring some
breaking up by light
blasting or shaking.
More bulky and some-
what hard to penetrate,
causing voids in dipper
or bucket.
Blasted rock, hardpan
and other bulky mate-
rials, which cause con-
siderable voids in dip-
per or bucket and are
difficult to penetrate.
Well broken limestone,
sand rock and other
blasted rocks.
Blasted shale.
Ore formations (not of
rock character) requir-
ing some blasting.
Heavy wet, sticky clay.
Gravel with large bould-
ers.
Heavy, wet gumbo.
Cemented gravel.
Hard tough shale.
Limestone.
Trap rock.
Granite.
Sandstone.
Taconite.
Conglomerate.
Caliche rock.
Any of these blasted to
large pieces mixed
with fines and dirt.
Tough, rubbery clay
that shaves from bank.
211
-------
TABLE 38. MATERIAL WEIGHTS, PERCENTAGE, SWELL AND LOAD
FACTORS (CATERPILLAR TRACTOR CO., 1975)
Weights shown are averages, rounded off to the nearest 50 pounds. Since
actual weights are affected by moisture content, grain size, degree of com-
paction and other local conditions, tests must be made to determine exact
material characteristics. The load factor and swell have been determined
from the weights given, using the equations found on Page 209.
Material
Bauxite
Caliche
Cinders
Carnotite, Uranium Ore
Clay - Natural Bed
Dry excavated
Wet excavated
Clay & Gravel: Dry
Wet
Coal: Anthracite, raw
washed
Bituminous, raw
washed
Decomposed Rock:
75% Rock 25% Earth
50% Rock 50% Earth
25% Rock 75% Earth
Earth - Dry packed
Wet excavated
Loam
Granite - Broken
Gravel Pitrun
Dry
Dry l/4"-2" (6-51
Wet l/4"-2" (6-51
Sand & Clay - Loose
Compacted
Gypsum - Broken
Crushed
Hematite, Iron ore
Limestone Broken
Magnetite, Iron ore
Pyrite, Iron ore
Sandstone
Sand Dry, loose
Damp
l^l~u»fx
Wet
Sand & Gravel - Dry
Wet
Slag-Broken
Stone - Crushed
Taconite
Top Soil
Traprock - Broken
lb/bcy_
3200
3800
1450
3700
3400
3100
3500
2800
3100
2700
2500
2150
1900
4700
3850
3300
3200
3400
2600
4600
3650
2850
mm) 3200
mm) 3800
3400
—
5350
4700
4900
4400
5500
5100
4250
2700
3200
3500
3250
4950
4500
7106-9450
7 inn
L J\J\J
4400
% Swell
33
82
52
35
22
23
25
41
11
35
35
35
35
43
33
25
25
27
23
64
12
12
12
12
27
—
75
75
18
69
18
18
67
12
12
12
12
10
?512
49
Ib/lcy
2400
2100
950
2750
2800
2500
2800
2000
2800
2000
1850
1600
1400
3300
2900
2650
2550
2700
2100
2800
3250
2550
2850
3400
2700
4050
3050
2700
4150
2600
4700
4350
2550
2400
2850
3100
2900
3400
4100-5400
1600
2950
Load Factor
.75
.55
.66
.74
.82
.81
.80
.71
.90
.74
.74
.74
.74
.70
.75
.80
.80
.79
.81
.61
.89
.89
.89
.89
.79
—
.57
.57
.85
.59
.85
.85
.60
.89
O f\
.89
.89
.89
.91
.57-58
.70
.67
212
-------
TABLE 39. OPERATING EFFICIENCY FACTORS (ATKINSON, 1971)
JOB . MANAGEMENT CONDITIONS
T
CONDITIONS EXCELLENT GOOD FAIR POOR
Excellent 0.83 0.80 0.77 0.77
Good 0.76 0.73 0.70 0.64
Fair 0.72 0.69 0.66 0.60
Poor 0.63 0.61 0.59 0.54
213
-------
However, the swing and swing-back times can vary from cycle to cycle, and
therefore, constitute the variable component of the cycle time. The
relationship between production and cycle time is shown in Table 40.
Calculation of theoretical cycle times is rather involved. Manufac-
turers usually provide enough horsepower for various motions to maintain
a uniform cycle time for comparable conditions. Every activity and motion
of the machine is dependent on motor horsepower, machine weight, and the
need for proper control. For example, when calculating the swingtime, one
has to know the average translational inertia of the machine and the average
torque from the speed-torque curve. Here, one must also take into considera-
tion the swing-gear ratio. The acceleration and deceleration times, plus
the continuous running time at full speed are also taken into account for
this calculation. In industry literature, on the basis of machine parameters
and design information, standard cycle times have been provided (Table 41).
These ideal cycle times are subject to correction for any nonstandard appli-
cation (chopping) or differences in swing angle, digging height, etc. The
influence of these factors were discussed earlier. Cycle times can also be
obtained through time studies in the field.
Decreases in cycle time can be achieved by increasing the motor horse-
power. For example, maximizing the swing efficiency would not only require
an increase in power, but also a reduction in inertia loads, resulting in
only a modest reduction in cycle time. Swing power can be increased, but
this does not yield a proportional reduction in cycle time. An increase of
50 percent in power, for example, reduces total cycle time of a dragline
by only 10 percent. Such major increases in power will be reflected in
high operating costs.
For draglines,the swing time is the most critical, although in some
cases, hoisting time is important. In cases where the hoisting time makes
up to about 40 percent of the total cycle time, an increase of 33 percent
in power reduces the cycle time by 10 percent. This reduction in cycle
time occurs only in the hoisting time portion of the cycle, unlike the swing-
time reduction which spreads over the entire range of the total cycle. The
average reduction in cycle time, due to decreases in hoisting time, in
actuality, will be less than 10 percent.
The bucket-fill time appears fixed and is not affected by the size of
the bucket, provided the machine operator is efficient.
It should be noted that a chopping operation increases the cycle time
by as much as 40 to 50 percent. An average increase of 45 percent would
be appropriate for calculation over the normal overall cycle time without
extended swing or hoisting time (Table 42). Also, bench digging or exten-
sion increases the dragline cycle time by as much as 80 percent over the
overall cycle time.
Operating Delay
In production calculations, not all the hourly minutes are parts of
the cycles as there are inevitable delays that occur which take some of the
214
-------
TABLE 40. WALKING DRAGLINE MACHINE OUTPUT (LEARMONT, 1975)
Bank Yards per Operating Hours = ——: ——
v * & Cycle Time
x
Operating Delay Factor
Efficiency Factor _
x —.. i,— x Bucket Capacity
Swell Factor v J
2570-W 335 Ft Boom, 38°300' 2570-W 360 Ft Boom, 31-1/2°340'
Radius, 110 Cu Yd Radius, 90 Cu Yd
Case 1, 110 Ft Overburden
Weighted Average Cycle Time, From Table 42
Operating Delay Factor, From Table 43
Efficiency Factor
Swell Factor
Bank Yards per Operating Hour
Case 2, 130 Ft Overburden
Weighted Average Cycle Time, From Table 42
Operating Delay Factor
Efficiency Factor
Average Efficiency Factor
Swell Factor
Bank Yards per Operating Hour
(including 38% rehandle)
Productive Bank Yards per Operating Hour
48.2
0.80
0.85
1.25
40.6 x Bucket Capacity
= 4470 yd per hr
57.7
0.82
0.85 Bank Material
0.95 Rehandled Material
0.88
1.25
36.0 x Bucket Capacity
= 3960 yd per hr
26.1 x Bucket Capacity
= 2870 yd per hr
45.7
0.82
0.85
1.25
43.9 x Bucket Capacity
= 3950 yd per hr
47.7
0.84
0.85
1.25
43.1 x Bucket Capacity
= 3880 yd per hr
-------
TABLE 41. APPROXIMATE CYCLES PER HOUR FOR STRIPPING
SHOVELS AND DRAGLINES (ATKINSON, 1971)
Bucket or Dipper Size Dragline Shovel
Yd3
8-35
36-59
60-200
3
m
6-27
28-45
46-150
58
56
53
69
68
64
* These figures are based on a 90° swing for a shovel and
a 120° swing for a dragline, which approximates most
field conditions.
216
-------
TABLE 42. WALKING DRAGLINE CYCLE TIME CALCULATION (LEARMONT, 1975)
2570-W 335-Ft Boom, 38° 300' Radius
Case 1, 110 Ft Over-
burden
Depth, Ft
Percent of Total
Average Swing
Angle, Degree
Spot Bucket, Sec
Fill Bucket, Sec
Hoist Bucket, Sec
Swing Over, Sec
Swing Back, Sec
Total Time, Sec
Weighted Average
Cycle Time, Sec
Chop Down
20
18
120
3.0
10.2
—
23.2
22.3
58.7
110 Cu Yd
Median Depth
45
41
45
3.0
10.2
—
13.5
12.7
39.4
48.2
Maximum
Depth
90
41
90
3.0
10.2
20.6
—
18.5
52.3
2570-W 360-Ft Boom, 31-1/2° 340 'Radius,
Chop Down
20
18
105
3.0
9.5
—
22.0
21.0
55.5
90 Cu Yd
Median Depth
45
41
40
3.0
9.5
—
13.2
12.4
38.1
45.7
Maximum
Depth
90
41
80
3.0
9.5
—
18.8
17.8
49.1
(continued)
-------
TABLE 42. (continued)
N>
h-1
oo
Case 2, 130 Ft Over-
burden
Depth, Ft
Percent of Total
Average Swing
Angle, Degree
Spot Bucket, Sec
Fill Bucket, Sec
Hoist Bucket, Sec
Swing Over, Sec
Swing Back, Sec
Total Time, Sec
Weighted Average
Cycle Time, Sec
2570-W
Chop
Down
20
11
135
3.0
10.0
—
25.1
2A.2
62.5
335-Ft Boom, 38° 300' Radius
110 Cu Yd
Median
Depth
55
30
50
3.0
10.2
—
14.2
13.4
40.8
57.7
Maximum
Depth
110
30
100
3.0
10.2
24.4
—
19.8
57.4
Extend
Bench
29
180
3.0
10.2
—
30.8
29.8
73.8
2570-W 360-Ft Boom, 31-1/2° 340' Radius,
90 Cu Yd
Chop Down
20
15
120
3.0
9.5
—
23.9
22.9
59.3
Median Depth
55
43
45
3.0
9.5
—
14.1
13.2
39.8
47.7
Maximum
Depth
110
42
90
3.0
9.5
—
20.1
19.1
51.7
-------
operating time that could have been better utilized for useful cyclic opera-
tions. These are called operating delays (Table 43). Such delays can occur
due to: 1.) machine movement, 2.) cleaning coal, and 3.) miscellaneous
minor causes. An estimate of these delays for selected conditions is neces-
sary to arrive at actual production time. The actual cycle time in practice
is a weighted average of the chopping, minimum, maximum, and optimum-depth
cycle times (Table 42) .
SURFACE MINE DESIGN
The topography of the area determines the stripping method to be
employed. For example, mountainous areas call for mountaintop removal or
other contour mining methods, while flat terrain may dictate simple over-
casting by area techniques. Seam thickness, as well as multiple-seam
occurrences, must also be considered in equipment and method selection.
Seam pitch, characteristics of the coal, the amount of groundwater over
the coal seam, and stability of the highwall are also important factors.
Finally, information concerning faults within the deposit, limits of the
deposit, property lines, roads, surrounding private properties, and local
land-use regulations must be carefully evaluated before actual mine design.
The economic life of the mine is related to pit planning by the yearly
average stripping ratio and other factors that could assure profitability
of the venture. Early return on the investment is desired and often pit
designs are planned with this in mind. Equipment selection has direct
financial consequences on pit planning. It may be possible to achieve low
operating and maintenance costs with high initial costs. On the other hand,
low initial capital investment will call for expansion capital in the future.
Pit entrances should be centrally located so that haulage distance
between the center of gravity of the pit area and the dump is near minimum.
Dead heading can be avoided by construction of several pit entrances
preferably perpendicular to the cut.
Environmental concerns about surface mining have been on the increase
in the last decade. Mine planning should include plans for reclamation
procedures prior to the commencement of mining. Regulations require that
reclamation be accompanied by revegetation and burial of toxic material to
avoid groundwater contamination. Acid mine drainage must be treated and
plans for its treatment initiated prior to mining. To gain public and
community acceptance, special consideration should be given, to the immediate
community in hiring unskilled labor. Job training plans should include
preparing those hired for more skilled jobs and positions of responsibility.
ENGINEERING DESIGN AND PIT LAYOUT
Generally, equipment is selected to conform with an existing mine
design or mines are designed based on the anticipated equipment to be
employed. In either case, the equipment performance characteristics and
operating conditions determine the actual production levels. The following
219
-------
TABLE 43. WALKING DRAGLINE - OPERATING
DELAYS (LEARMONT, 1975)
2570-W 335-Ft 2570-W 360-Ft
Boom, 38° Boom, 31-1/2°
300' Radius, 340' Radius,
110 Cu Yd 130 Cu Yd
Case 1, 110 Ft Overburden
Approximate rate of face advance per
hr, ft 10.8 8.8
Movement across face for a 30 ft
advance, ft 80 80
Total dragline movement per hr. ft. 39.6 32.3
Move time per hour (1 min per 8 ft
step including preparation), min 5.0 4.0
Clean coal time per hr, min 3.0 2.4
Miscellaneous delays per hr, min 4.0 4.0
Total operating delays per hr, min 12.0 10.4
Operating delay factor, % 80 82
Case 2, 130 Ft Overburden
Approximate rate of face advance per
hr, ft 6.6 7.4
Movement across face for a 30 ft
advance, ft 160 80
Total dragline movement per hr, ft 41.8 27.1
Move time per hour (1 min per 8 ft
step including preparation), min 5.2 3.4
Clean coal time per hr, min 1.8 2.1
Miscellaneous delays per hr, min 4.0 4.0
Total operating delays per hr, min 11.0 9.5
Operating delay factor, % 82 84
220
-------
is a set of definitions for use with the mathematical equations to be devel-
oped in this section. Since the use of English units is most common, the
derivations here are in English units.
P = Production per month in tons
A = Reject fraction in run-off mine coal
U = Recovery fraction from coal in ground
COAW = Equivalent Ib per cu yd of coal in the solid seam
(Usually = 2000 Ibs)
TT Equivalent Ib per ton of coal
1 ton 2000 Ib for short ton
1 ton 2240 Ib for long ton
1 ton 1000 kg for metric ton
T Thickness of coal seam in ft
H Overburden height in ft
HI Depth of dragline working bench in ft
SMH Scheduled monthly operating hours
CT Cycle time in seconds
Bs Bucket size for shovel in cu yd
Bd = Bucket size for dragline in cu yd
BF Bucket fill factor (fraction)
OF Operating efficiency (fraction)
W = Pit width in ft
SP = Swell of material (percentage usually 25% to 30%)
$ = Angle of highwall in degrees
0 = Angle of spoil in degrees
FACD = MUF factor for dragline in Ib/MUF
FACS = MUF factor for shovel in Ib/MUF
Rs = Shovel reach in ft
Rd = Dragline reach in ft
Rsd Shovel dumping radius in ft
Rdd Dragline dumping radius in ft
Bb Bucket size of bulldozer in cu yd
Bl Bucket size of loader in cu yd
Be Bucket size of scraper in cu yd
S Speed in feet per minute (fpm)
SI Loaded speed in mph
S2 = Empty speed in mph
LF = Load factor (fraction)
MF = Material factor (fraction)
MAT Maintenance factor (fraction)
CF = Carry factor (fraction)
MPH = Available minutes per hour (fraction)
BCYPH = Maximum production per hour in cu yd
HD = Haul distance in ft
TCT = Total cycle time in sec
VL = Loose volume in cu yd
VB = Bank volume in cu yd
HPS = Hours per shift
SPD = Shifts per day
DPW = Days per week
WPM = Weeks per month
221
-------
Overburden Yardage to be Moved
Knowledge of the amount of overburden to be moved is the primary infor-
mation on which the selection of a shovel or dragline is based. For a clean
coal production of P tons per month, the acreage of coal to be exposed and
the amount of overburden to be removed, calculate as follows:
P Production per month in tons
A Reject fraction in run-off-mine coal
U = Recovery fraction from coal in ground
COAW Equivalent Ib per cu yd of coal in the solid
T = Thickness of coal seam in ft
H Overburden height in ft
ACP Amount of run-of-mine coal to be mined in tons
ACE Amount of coal to be exposed in tons
CYE Cubic yards of coal to be exposed
SYE Square yards of coal to be exposed
ACYE = The acreage of coal to be exposed
COB Cubic yards of overburden to be removed
ACP =
ACYE
COB Mipi
Calculation of Bucket Size
Knowing the amount of overburden to be removed each month, a suitable
bucket size can be calculated.
SMH = Scheduled monthly hours
CT = Cycle time in sec
OF Operating efficiency
BF = Bucket-fill factor
NTCPM = Number of theoretical cycles per month
222
-------
N>
CO
datum
Figure 92. Shovel pit with coal fender (Stefanko, et al., 1973)
-------
N3
N3
Figure 93. Shovel pit without coal fender (Stefanko, et al., 1973)
-------
TBC Theoretical bucket capacity
NBSIZE = The actual bucket capacity after considering
the swell percentage, the bucket-fill factor
and the operating efficiency.
NTCPM (SMHK3600)
v» -L
TBC C°B
NBSIZE
NTCPM
TBC
(OF)(BF)
It is now a common practice among operators to restrict the bucket sizes
within the range of 70 to 140 cubic yards. Therefore, if the calculations
should lead to a large bucket size, it is advisable to use two or more
machines in multiple pits to meet the production requirements.
Calculation of Reach
Since the primary equipment can be a shovel or a dragline, the formulas
applicable to both machines are developed here.
1. Shovel Reach—
The usual operating practice with a shovel is to align one set of its
track on the rib of coal adjacent to the top of the spoil. An imaginary
plane passing through this location would contain the point "D" shown above
the front crawler in Figure 92. Projection of a horizontal line from the
imaginary vertical plane to the point of discharge of the shovel dipper
establishes the reach of the shovel. The reach is calculated as follows
(Rumfelt, 1961). Attention is drawn to Figure 93.
ACS Area cut in solid, sq ft
ACB Area cut in broken (or spoil), sq ft
$ Angle of highwall, degrees
0 = Angle of spoil, degrees
h = Height of spoil, ft
H Height of overburden highwall, ft
T Thickness of coal, ft
W Width of pit, ft
b Berm width, ft. This is assumed to be small and
therefore is not included in the calculations.
Fc Width across both crawlers in ft
Rs Reach of shovel in ft
Rsd = Dumping radius of shovel in ft
225
-------
With reference to Figure 93
Tan 0
h2 h Tan 9
ACB W [h (-—) Tan 0] + (~~) (~-) Tan 0 + (W) (T)
w2 w2
-!p) Tan 0 + (-£-)
+ (W)(h) (-7-) Tan 0 + (W) (T) (1)
= (W) (h) - (-=-) Tan Q + (-=-) Tan 0 + (W) (T)
The corresponding area in solid is given by,
ACS = (W)(H) (2)
Assuming that the length is the same in the spoil as in the solid, it
can be deduced that
cp
ACB = (W)(H)(1 + ~) (3)
Therefore, equations (1) and (3) are the same, i.e.,
w
(W)(T) + (W)(h) - (--) Tan 0 (W) (H) (1
and,
op u
h = H(l + -) + (-~) Tan 0 (T)
Hence, from the geometry in Figure 93
Rs = h COT 0
which yields
Ian0-
Defining FC = the width across both crawlers in feet, the dumping radius
of the shovel is given by,
Rsd = Rs + .5FC
226
-------
2 . Dragline Reach —
All definitions used in calculating shovel reach apply to the dragline
case also. While the shovel operates on the top of the coal, the draglines
work on a bench on the highwall or on top of the highwall itself. An
imaginary vertical plane passing through the edge of the highwall would
contain the point "D" shown to the left of the- crawler in Figure 94. Pro-
jection of a horizontal line from the imaginary vertical plane to the point
of discharge of the bucket content establishes the reach of the dragline.
The reach is calculated as in the case of the shovel, x^ith the following
equations
Rd Reach of dragline in ft
Rdd Dumping radius of dragline in ft
h H(l + Y||) + (-|-) Tan 0 - (T)
But the dragline reach as can be seen from Figure 93 is given by,
Rd = H COT $ + h COT 9
Tan * Tan 0
(4-) Tan 0 (T)
Defining E = the outside diameter of the dragline tub, the dumping radius
can be calculated as,
Rdd = Rd + .75E
3. Dragline with a Working Bench —
This case is illustrated in Figure 95. Here the dragline works on a
bench at a depth HI ft below the top of the highwall. This procedure may
be practiced to increase the operating range of the dragline. The overburden
to depth HI is removed by some other equipment, and the reach calculation
with consideration to bench is modified as follows :
Rd • + r «" - H"<1+ iff) +
Rdd = Rd + .75Et
However, if the dragline is used for cutting this bench also, the dragline
bucket size should be larger for the same clean coal production.
227
-------
Figure 94. Dragline pit (Stefanko, et al., 1973)
228
-------
NJ
K>
VO
Figure 95. Dragline pit with working bench (Stefanko, et al., 1973)
-------
Machine Weight Using the Maximum Usefulness Factor (MUF)
Maximum usefulness factor (MUF) is the measure of the capability of a
machine to do a job. For a stripping machine, it is defined as follows:
MUF = Bucket size (cu yd) v Reach (ft)
Rumfelt (1961) investigated a large number of stripping machines and
produced MUF curves for shovels and draglines. The curves of MUF plotted
against machine weights are straight lines for different sizes of shovels
and draglines, which implies that the weight of the machine in pounds
divided by its MUF was a constant. The value for a constant was found to
be approximately 745 for shovels (Figure 96). For draglines, the values
varied between 467 and 575 (Figure 97).
Rs = Reach of shovel in ft
Bs Bucket size of shovel in cu yd
FACS = MUF constant for shovel in Ib/MUF
SMW = Weight of shovel in Ib
SMW = (Rs)(Bs)(FACS)
Rd = Reach of dragline in ft
Bd = Bucket size of dragline in cu yd
FACD = MUF constant for dragline in Ib/MUF
DMW = Weight of dragline in Ib
DMW = (Rd)(Bd)(FACD)
Dragline Digging Its Own Bench
The dragline sometimes digs its own bench in order to increase its
range. In this type of operation, the machine chops material above its
operating bench and therefore, the cycle time is no longer the normal cycle
time but rather a weighted average cycle time. This can be calculated from
the normal arid the bench "chopping" cycle times as follows:
SMH Scheduled monthly hours
NTCPM Number of theoretical cycles per month
COB = Cubic yards of overburden to be removed
TBC Theoretical bucket capacity in cu yd
NBSIZE = The actual bucket capacity in cu yd
OF = Operating Efficiency
BF = Bucket fill factor
CT = Normal cycle time in sec
CTB = Cycle time for "chopping" its bench in sec
CTW The weighted average cycle time in sec
HI Depth to dragline bench in ft
H Overburden height up to dragline bench from coal in ft
Ht Total height of the overburden in ft
HT H + HI
IT IT-l
CTW = CT (-jjj-) + CTB (~~-)
230
-------
12
ro
2 10
X
u.
10*
>•
II
8
GRAPH FOR SHOVEL
SHOWING RELATIONSHIPS OF
GROSS MACHINE WEIGHT
AND MUFs NUMBERS
T
1
I
I
NOTE;
SLOPE OF CURVE= 1/745
8
10 12
GROSS MACHINE WEIGHT, POUNDS x 10
Figure 96. Shovel MUF graph (Rumfelt, 1961).
231
-------
to
5
X
ro 4
n
GRAPHS FOR DRAGLINES
SHOWING RELATIONSHIPS OF 7
GROSS MACHINE WEIGHT '
AND MUFdNUMBERS /
r
NOTE:
SLOPE OF A
SLOPE OF B
1/575
1/467
1
1
1234
GROSS MACHINE WEIGHT, POUNDS x 10*
Figure 97. Dragline MUF graph (Rumfelt, 1961)
232
-------
(SMH)(3600)
CTW
COB (P)(TT)(HT)
TBC
NBSIZE =
(1-A)(U)(COAW)(T)
COB
NTCPM
TBC
(OF)(BF)
If the dragline digs its own bench in an operation, the weighted cycle
time for both the bench digging and the normal operation will be greater
than the normal cycle time. The net result in the selection of the dragline
is one with a larger bucket size. The reach of the dragline must be calculat-
ed for the total overburden height. It is obvious that both of these will
increase the maximum usefulness factor and therefore, the machine weight
(Figure 95).
Multiple-Seam Mining
In multiple-seam mining, the dragline removes the overburden over the
first coal seam, casts the spoil, and the coal seam is loaded out. Following
this, the top of the spoil is leveled to form an operating bench for the
dragline. The dragline is repositioned on this bench to strip the parting.
The dragline action here is a chopping rather than a digging one. In
chopping, the bucket fill factor decreases because of reduced contact angle
between the bucket teeth and the overburden. Also, there is usually an
increase in the bucket swing angle; i.e., when the dragline works from the
leveled spoil it has to swing more than the normal swing angle in order to
dump the parting material behind the original spoil. With reference to
Figures 98 and 99, the weighted average cycle time can be calculated as
follows:
SMH Scheduled monthly hours
BF Bucket fill factor
T Coal thickness in ft
H Height of overburden in ft
P = Production per month in tons
TT Equivalent Ib per ton of coal
COAW = Equivalent Ib per cu yd of coal in solid
A = Reject fraction in run-off mine coal
U = Recovery fraction from coal in ground
COB Cubic yards of overburden removed
CT = Normal cycle time of dragline in sec
CT2 Cycle time for chopping overburden 2 in sec
CTM = The weighted cycle time in sec
PI = Production from seam 1 in tons
233
-------
Plan View
Surface
Leveled Spoil
Spoil
Section View
Figure 98. Multiseam dragline operation with upper seam exposed
(Stefanko, et al., 1973).
234
-------
Dragline
Road on
Leveled
Spoil
Exposed Seam
Plan View
Section View
Figure 99. Dragline positioned on leveled spoil removing parting
(Stefanko, et al., 1973).
235
-------
P2 Production from seam 2 in tons
HI = Height of overburden 1 in ft
H2 Height of overburden 2 in ft
Tl = Thickness of seam 1 in ft
T2 = Thickness of seam 2 in ft
BF1 Bucket fill factor for seam 1
BF2 Bucket fill factor for seam 2
GOBI Cubic yards of overburden removed in 1
COB2 Cubic yards of overburden removed in 2
T Tl + T2
H HI + H2
P PI + P2
CTM CT (-
CT2
NTCPM =
(SMH)(3600)
CTM
BF = BF1 x
HI
H
+ BF2 x
H2
H
PI =
P2 = P (-^|-)
GOBI
(1 - A)(U)(COAW)(T1)
COB 2 =
(P2)(TT)(H2)
(1 A)(U)(COAW)(T2)
COB COB1 + COB2
TBC =
COB
NTCPM
NBSIZE =
TBC
(OF)(BF)
The total effect of multiple-seam mining on dragline selection is that
the weighted cycle time increases, the number of passes per hour decreases,
the bucket-fill factor decreases, and the bucket size increases. Therefore,
the maximum usefulness factor increases. Even if the reach remains the
236
-------
same, the machine weight will increase to meet the required production per
month. The calculation of the reach in this case is involved since the
dumping positions for the two modes of operation are different.
BUCKET-WHEEL EXCAVATOR SELECTION
The design of a bucket-wheel excavator (BWE) for the excavation of
ore or waste is determined by the digging power required and the abrasiveness
of the material to be excavated. A BWE consists of too many assembly groups
and has to fulfill too many simulataneous functions to allow for the calcula-
tion of any individual component by one definite standard formula. Lengths
and dimensions of a wheel excavator are highly interrelated, and their com-
bination is always a compromise between the designed performance of the
machine and its cost. In general, these machines are job-tailored and, as
for every other type of excavating equipment, the cutting, tearing, loosening
and splitting force required to excavate material is the deciding factor in
the design of a BWE. In the United States, wheels are operated in tandem
with shovels and/or draglines thereby necessitating considerations of
equipment matching and working area limitations in pit design. Since bucket-
wheel excavators are essentially German development, the standard units
associated with BWEs are in the metric system. Therefore, in the discussion
here metric units are used.
SOIL CLASSIFICATION
The BWE is an ideal machine for working in soft and medium hard strata,
and it is gradually being developed for use in harder strata. Specifically,
the digging resistance of the soil is the most sensitive parameter affecting
the BWE operation. For excavation purposes, soils can be classified as
follows:
1. The soils that can be excavated by a spade, e.g.,
unconsolidated strata like sand, gravel and silt.
2. The soils that can be loosened by a pickaxe before
excavation, e.g., partly consolidated strata like
clay beds.
3. The soils that can be loosened by blasting before
excavation, e.g., hard shales, sandstone, etc.
The above classification is broad and general and cannot be of specific
assistance to the BWE designer. Classification of soils according to their
diggability is vital; but unfortunately, there is not any clearcut, standard
method for determining this factor. Some common methods of expressing the
digging resistance are:
1. Kg/cm, where the cutting resistance is calculated
on the basis of load per unit length of the cutter
in contact with the soil.
f\
2. Kg/cm , where the load is related to the area of the cut slice.
237
-------
3. Kg/cm3, where the digging resistance is calculated
as a function of the excavated volume per bucket.
The theories that are being advanced and the discussions taking place
on what is considered to be the most suitable reference quantity vary widely.
So far, most design calculations have been based on the cutter lengths or
on the cross-sectional area of the slice cut, i.e., methods 1 and 2; but it
must be recognized that the volume of material excavated per bucket has no
small influence on the digging resistance. At the Fifth International
Earthmoving Conference held in Prague in October, 1963, Professor N. G.
Dombrowsky (1964) gave the following specific digging forces for various
types of ground:
2
a. Light ground 1.8 to 2.5 kg/cm
b. Medium ground 3.0 to 3.5 kg/en/
c. Heavy ground 7.0 to 18. kg/cm^
While it is not uncommon in non-German technical literature to consider the
resistance to digging in relation to the cross-section of the slice, manu-
facturers of wheel excavators in Germany are still largely using the effec-
tive cutter length for determining the cutting forces. This is so because
of the large number of test results available on various soils. Based on
150 large-scale tests of BWE operations, Himmel (1961) has calculated the
specific cutting forces required to dig the types of ground encountered in
open-cut mines. On an average, these were found to be:
1. About 20 kg/cm for light ground such as sand and
gravel.
2. About 30 to 40 kg/cm for medium heavy ground such
as sandy loam, pure loam, loess, lean clay, etc.
3. About 50 to 60 kg/cm for heavy ground such as heavy
compact and plastic clays.
It is not possible to draw any conclusions on the above proposed
values because the cutting power depends on the type of ground, the cross-
sectional area and the shape of the slice cut, the configuration and sharp-
ness of the cutter blades, the shape of the teeth, and the cutting speed.
Hard ground requires a high specific cutting force, high cutting speeds, and
additional cutting blades between the buckets. Reducing the output and the
rate of swing decreases the slice cross-section and increases the specific
cutting force for a given wheel drive rating. This will be clear from the
following formula (Rasper, 1964):
N,, = —— /Q x S x R
G n x c
233
-------
where,
NQ = Power required for cutting in kw
Q = Actual digging capacity in cubic meters per hour
S = Number of bucket discharges per minute
R = Radius of the cutting wheel in meters
c = Constant depending on the bench height/wheel
diameter ratio. For the cutting height 2/3
wheel diameter, its value is 171
n = Efficiency of the wheel drive
k = Specific cutting force in kg/cm
Shown in Tables 44 and 45 are reference values for specific cutting forces
applied to the more important bulk materials and virgin soils. The tables
must be viewed with reservation as to the nonhomogeneity of the different
soils from one geographical location to the other.
Devising methods for measuring the cutting resistance is another
difficulty because the results from such theoretical considerations vary
with the size of the sample tested, the angle of loading, the rate of loading
and other such experimental procedures. The methods commonly employed for
the determination of the cutting resistance are:
1. Simulating equipment pulled through the soil to be
excavated for direct measurement of the force applied.
2. Laboratory tests with tooth and tooth impression
measurements on the rock or soil samples.
3. Determination of the compactness by Proctor needles.
In a method developed by a German manufacturer (LMG), the entry of an
excavator tooth into soil is copied in the laboratory on a soil specimen
(Rasper and Ritter, 1961). This gives some information with respect to the
soil split resistance. If numerous tests have been made with soils on
which the cutting effect of BWE has been studied and analyzed, a correlation
can be drawn between the theoretical laboratory values and the actual digging
resistance. From observations so far, theoretical analysis can at best
serve only as a guideline in the selection of the wheel because of the vast
complex of influences which exist in practice that cannot be considered
theoretically. Practical tests on the site of employment will confirm the
suitability of the machine selected for a particular job.
In tests conducted on glacial till, sandstone and devonian shale, it
was confirmed that earth containing a high percentage of boulders is unsuit-
able for wheels (Cheek, 1966). The capacity of any excavator operating in
earth material with boulders will be affected adversely in proportion to the
frequency of boulder occurrence. Two possible reasons can be attributed to
this poor performance. Some boulders, like basalt and granite, are too
hard for economical cutting. On the other hand, boulders and rocks composed
of materials which the machine can cut in the solid state become loosened
during excavation and cannot be rehandled.
239
-------
TABLE 44. SPECIFIC CUTTING FORCES OF VIRGIN SOILS
FOR BWE EXCAVATION (GARTNER, 1965)
SOIL TYPE
SPECIFIC CUTTING FORCES
KG/CM
Earth
Loess
Sand (fine, coarse, wet or dry)
Clayey sand
Gravel fine
Gravel coarse
Sandy loam and wet loam
Dry loam
Clay wet
Clay dry
Clay schistose
Sandy clay
Clayey slate
Slate
Sandstone (easy digging)
Sandstone (hard digging)
Gypsum
Lime
Phosphate
Marl
Limestone
Weathered granite
Alluvial light consolidation
Alluvial heavy consolidation
Alluvial medium consolidation
Hard coal normal
Hard coal frozen
Lignite
Brown iron ore
10 30
20 40
10 - 40
10 - 50
20 - 50
20 - 80
20 - 60
20 - 80
30 - 65
50 - 120
35 120
20 65
50 160
70 - 100
70 - 160
160 - 280
50 - 130
30 120
80 200
60 140
100 - 180
50 - 100
30 - 60
70 150
50 80
50 100
100 160
20 70
190 210
240
-------
TABLE 45. SPECIFIC CUTTING FORCES OF MATERIALS SUITABLE
FOR BWE LOADING OPERATIONS (GARTNER, 1965)
MATERIAL TYPE SIZE
mm
Sand
Gravel fine 0 - 100
Gravel coarse 100 - 400
Earth
Phosphate
Bauxite
Iron-Mn-Cr 0 150
Ores 150 300
300 450
Loose hard coal
Solidified hard coal
Lignite
Coke
Pellets (ore; cement)
Limestone
Slag
SPECIFIC CUTTING FORCE
KG /CM
10 -
20 -
10 -
10 -
10 -
10 -
20 -
20 -
20 -
10 -
10 -
10 -
10 -
10
10
10 -
20
25
25
15
20
20
40
50
80
30
40
25
30
25
30
25
241
-------
BWE OUTPUT CONSIDERATIONS
The theoretical output of a BWE is based on the bucket size and the
number of bucket discharges per minute. If
I = Nominal bucket capacity in cubic meters
Z = Number of buckets in the wheel
V]_ = Cutting speed of the wheel in meters per second
D = Diameter of the wheel in meters,
then
Ss - V x Z/trD
where
S = Number of bucket discharges per second
Q = I x S x 3600
xt s
and
Q Theoretical capacity of the excavator in cubic
meters per hour.
As can be seen from the above equations, the number of bucket dis-
chargers is dependent on the peripheral speed. The peripheral speed of a
bucket wheel is limited by the ability of the wheel to discharge its bucket
content on the chute against the counteracting centrifugal force. In theory,
the maximum peripheral speed must be such that the bucket discharge will be
ensured. Mathematically,
M x g = M x V 2/ R
where
M = Mass of material in the bucket, kg
R = Radius of the wheel in meters 2
g = Acceleration due to gravity, meters per second
which yields the following expression:
V
max
Practical values of speed lie between 0.4 Vmax to 0.6 V^^. To keep
the wear on bucket's cutting knives or teeth at a minimum, speeds do not
exceed 5 meters per second. The peripheral speed selected will also depend
greatly on the nature of material to be excavated. In principle, however,
a higher peripheral speed will be decided upon if hard material is to be
242
-------
cut, in which case maximum output may not be attained. Based on a constant
output, the doubling of the peripheral speed will halve the amount of
material excavated by each bucket; thus, cutting performance will be
reduced.
Yet another factor that affects the output of a BWE is the bucket-
filling capacity. It has been found that, in hard ground, the bucket-
filling is around 30% to 40% of the nominal capacity. The relationship
between digging resistance and the hourly capacity of the BWE is
Q./Q2 k22 / k'
where
Q BWE hourly capacity in cubic meters in soil with
specific cutting resistance k^
Q2 BWE hourly capacity in cubic meters in soil with
specific cutting resistance \f-2
Thus, the actual capacity of the BWE in any soil is given by:
Q = I x B. x S x 3600
xa f s
where
B Bucket-filling capacity in the soil expressed as a
fraction of the nominal bucket capacity
S Number of bucket discharges per second
Q = Actual capacity of the BWE in cubic meters/hour
3.
In soils with high cutting resistance, higher cutting speeds with lower
bucket-filling will result in a very small Qa as compared to Qt. The
ratio may be as small as 0.2. One can visualize that in hard soils the
excavating operation of the BWE has changed to a milling operation.
In most BWE calculations, the ratio of the bench height to the wheel
excavator diameter is taken at two-thirds because the machine performance
is optimum at this ratio. Obviously, with a lesser bench height full
advantage of the wheel capacity is not taken. Also, increases in bench
height above this limit result in undercutting which may lead to excessive
spillage to be rehandled; and in soft soils sliding of the burden over the
whell may result. The specific cutting resistance of the soil which can
be calculated for a given digging power rating from the following formula,
Np x n x C
K = G
/Q x S x R
243
-------
where
K Specific cutting resistance in kg/cm
S Number of bucket discharges per minute
Q = Actual capacity in the soil in cubic meters/hour
R = Radius of the bucket wheel in meters
NQ = Digging power rating of the drive in kw
n Efficiency of the drive
C Constant of the bench height to wheel diameter ratio
is found to be increasing with decreasing bench height for a given wheel
diameter. The change in the value of "C" with a change in the height/
diameter ratio can be observed in the following table (Rasper, 1964).
Height/Diameter
Ratio 0.1 0.2 0.3 0.4 0.5 0.6 0.67 0.7
Value of "C" 295 248 222 203 189 178 171 168
BUCKET-WHEEL DESIGN
At present,two types of bucket design are recognized, cellular and
noncellular. In noncellular wheels, a ring chute isolates the bucket from
the material flow behind it during the digging and lifting motion. Material
is carried in the bucket to a point where the ring chute clears the way for
the flow of material to the ladder conveyor. In the cell-type wheels, a
buffer or cell is arranged behind each bucket for the excavated material.
The cells side discharge this material onto a ladder conveyor. A cell-type
wheel is usually stronger and more resistant to wear; but with the applica-
tion of BWEs to harder formations which require higher cutting speeds, the
cell-less wheels have superseded the cellular wheels. This preference for
cell-less wheels arises because the discharge of excavated material from
cells is governed by the rotary motion of the wheel. At high speeds,
cellular wheels do not discharge completely.
The most important factor in bucket-wheel design is the diameter of
the wheel. Much stress is placed on the correct size of the wheel since
the service weight of the machine varies as the square of the diameter;
cutting torque varies inversely as the diameter; and the cutting speed
should never exceed /O.2gD, where g is the acceleration due to gravity and
D is the wheel diameter. The diameter should also be sufficient to accom-
modate the chute and roller bar and to transfer material from the buckets
to the ladder-conveyor. Since the ladder-conveyor is taken beyond the
axle of the wheel, it is a controlling factor in wheel design.
244
-------
OTHER CONSIDERATIONS
Wheel-boom and discharge-boom lengths have to be derived from operating
conditions. The length of the wheel boom depends on the specified cutting
height above and cutting depth below the grade and on the block width. The
gradient limitation on the ladder-conveyor is an important factor which is
decided by the type of material being handled. The length of the discharge
boom depends on the amount of conveying necessary and the height, or
possibly, the depth of discharge below grade. In an open-cut mine, the
length of the discharge boom should be such that "spoiling" is properly
affected on the dump zone and not within the pit limits.
Restrictions on the size of a modern-day BWE are imposed not by the
design of the wheel but by the conveying of material from the bucket to
the discharge point on the stacker boom. Also, with the design of BWEs for
deep mines, a dependable steep angle conveyor is necessary. Belt speeds
of 300 to 400 meters per minute and belt widths of 200 centimeters are not
uncommon. In cases where the inclination angle of the ladder boom is more
than 35 degrees, sandwich belt conveyors have proved to be highly satisfac-
tory. Because of the types of material that the BWE handles, the conveyors
must be liberally dimensioned and properly maintained.
MODE OF OPERATIONS
Modern BWEs generally excavate in blocks. Figure 100 shows a BWE
working in an established cut. The wheel is positioned to travel on the
pit floor in line with the top edge of the old highwall. As it advances,
a new highwall is exposed in the direction of excavation. Digging is
done by rotating the wheel, swinging it from side to side in long
parallel arcs, and "crowding" into the bank, by advancing the entire
machine, or by the travel of the digging boom if an automatic crowd is
available (Figure 100). A second way by which the wheel can be advanced
into the bank is by the falling cut method. A brief description of each
of these methods is given below.
Cut with a Crowding Machine
At the end of every swing, the digging boom can be extended by the
thickness of cut desired and the boom swung back in the reverse direction.
Obviously, the thickness of bank excavated does not vary with the boom
position; therefore, the slewing motion of the boom is fairly constant for
uniform output. The thickness through which the digging boom can be
advanced into the bank is theoretically calculated from the formula
(Mani, 1966):
t 0.133 x /Q /R x S
3,
245
-------
Direction of Mining
Direction of Digging
H Height of face
D Depth of block
W = Width of block
Bench Cut
1
-»
T
h
J,
T
I
I
1
T ~ Thickness of slice
W - Width of slice
h - Height of bench
C,- Center of the wheel
Figure 100. Block digging operation with BWE.
246
-------
where
Q Actual capacity of the excavator in cubic
meters/hour
S Number of bucket discharges per minute
R Radius of the wheel in meters
t Thickness of slice in meters
Cut with a Noncrowding Machine
The essential difference between this cut and the former is that here
the machine is moved forward bodily in the direction of the bench by the
thickness of the cut. The slew axis progresses by successive distances,
t, along the line of advance; and the successive segments that are cut are
not parallel but sickle-shaped. The thickness of slice varies constantly
with angle 0, both to the right and to the left of the direction of advance.
A very close approximation of slice thickness at any point P is given by:
2
,„ ,„. Sin 0 x
t = t (Cos(0) + ~2p— )
where
P! sr/t
Sr Slewing radius in meters
t Thickness of slice in meters
0 Angle between the direction of advance and
direction of P
Therefore, to maintain a uniform output as the boom slews away from the
center, the loss in the thickness of slice must be compensated for by
increasing the width of the slice cut. This is achieved by increasing the
slewing speed such that V V/cos(0), where V is the slewing speed at
the point P, and V is the starting slewing speed in the direction of the
advance.
Two factors limit the angle of slewing. At angles more than 70 degrees,
the slewing speed increases disproportionately to the increase in output;
therefore, usually with noncrowding machines, the width of the block is
such that the swing on either side of the direction of advance is maintained
within limits. The width of the bench is given by:
where
S (Sin 0T + sin 0n)
r L K
Sr = Radius of slewing in meters
0R = Angle to the right of the direction of advance < 70
0T = Angle to the left of the direction of advance < 70°
JL —
o
24/
-------
Falling Cut Method
Depending on stratification and fissuring, it is often advantageous in
firm ground to operate by the drop or falling cut method. At the commence-
ment of a block of maximum face height, a top cut about half the wheel
diameter is taken by the bench cut method, the depth of the cut being great
enough to leave a face formed for the top cut of the subsequent block. After
this initial preparation, the excavator retreats a short distance and the
rest of the block is excavated in segments which are proportioned by lowering
the wheel to the desired thickness (t) . The slice taken here is very similar
to that of the bench cut. The geometry of the cut is explained in Figure 101
including the slice cross-section at the line of advance. If
A, = Depth of slice normal to the working face at
0° slew in meters
A = Depth of advance in meters
the variation in the sickle depth in the falling cut method can be expressed
as :
where
S Radius from slew axis to outer edge of segment
A Depth of advance = A^ Cosec u
a = Angle of frontal batter
G = Angle slewed from the direction of advance
The slew angle in the falling cut can be considerably more than that
of the bench cut because A^0 does not become negligible until 0 exceeds 85°.
Theoretical calculations will show that the slewing speed is to be increased
in the same relation as before and that the output is identical by either of
the two cuts. However, this conclusion can be erroneous if the material is
not homogeneous in all the planes, which is usually the case with soils.
BWE OPERATIONS IN USA
Reference has already been made to the introduction of wheel excavators
in Illinois strip mines during the 1940's. There are only a very few wheels
in operation in the United States. Except for one operation each in the
states of Washington and Wyoming, there are six wheel excavators in the
Illinois.coal fields. In American strip mining practice, bucket wheels have
been used to remove loose top soil and soft beds, whereas the harder beds
and the coal itself are handled by large stick shovels or draglines. In
order to fully utilize the capacity of the two units in a tandem operation,
the output capacity and dimensions of the wheel and shovels must be matched.
248
-------
Top cut
taken by
bench cut
CL Batter angle
T Thickness of slice
W Width of slice
Cj Center of the wheel
Figure 101. Fall cut block digging operation with BWE.
249
-------
The Kolbe wheel excavator can operate only by the frontal block digging
method on benches of limited width. Overburden is stripped with the exca-
vator operating in a long trench about 30 to 32 meters wide at the bottom and
transferring it directly to the spoil heap. With burdens of 30 meters, the
width of the cut is only about 15 meters (Figure 31). Since there will be
occasions when the wheel and the tandem shovel have to cross each other, the
pit must be designed wide enough to allow this.
BWE SELECTION PARAMETERS
The required data for input are divided into four categories:
1. Mining Dimensions; These include (a) the length of
the mining property; (b) the number of benches; and
(c) the elevations of the benches.
2. Soil Characteristics: The data required here are
(a) the various soil types; (b) their specific cutting
resistances; (c) the operating speed of the wheel in
the soil types; (d) the bucket filling capacity; and
(e) the swell factor.
In soils with higher specific cutting resistances, the wheel is speeded
up and the successive slices that are taken are thinner. This seriously
affects bucket filling capacities and in some cases less than 25 percent of._
the theoretical output is mined. Swell factors for sand and similar loose,
broken material is not very significant, whereas a compacted material may
increase in volume by 30 percent to 40 percent on fragmentation.
3. Wheel Specifications: The data input here include (a)
the diameter of the wheel; (b) the number of buckets in
the wheel; (c) the length of the slew axis; (d) the
overall weight of the machine, and (e) the width of
the machine.
4. Miscellaneous: Realistic efficiency figures are
necessary for the mechanical and electrical components
of the BWE to arrive at proper production figures.
Much research and trial-and-error experience have gone into the most
recent designs of the digging wheel, its bucket and other appurtenances.
The result is a device that will satisfactorily handle the variety of
materials encountered, especially sticky clays, soils containing some float-
ing boulders, and shale of varying hardness and composition. BWEs are
machines designed for continuous digging and discharge. The continuous
steady action requires less specific power and greatly reduces load fluctua-
tions, characteristic of intermittent methods like that of shovels and
draglines.
The range of application of BWE is not limited to mining alone. The
classic method of stacking out and reclaiming bulk materials by conventional
machines has been replaced today by far more economical use of conveyors,
stackers, and bucket wheels.
250
-------
A BWE is an asset to land reclamation. Spoil piles are much more
regular in size and contour than those produced by shovels or draglines. The
hard rock is spoiled at the bottom of the mined area by the shovel or the
dragline, and the original loose top soil is returned to relatively the same
position, making replanting easier. In many cases, the crop and grazing
value of the reclaimed spoil areas is greater than prior to excavation.
DRILLING AND BLASTING
Since this subject has been covered in detail in an earlier section,
here the important considerations will be briefly outlined. Selection of
drilling equipment depends on the types of materials and their strengths.
Thus, penetration resistance to drilling should be studied before selection
of a drill and a method of drilling. Soils and other loose materials do not
require fragmentation, but are drilled in the process of drilling underlying
consolidated materials, when these loose materials overlie the rocks.
Although there are several methods of drilling, only mechanical methods which
utilize percussion drills or roller bit rotaries, or a hybrid of both, are
considered.
Selecting a particular drill involves the evaluation of energy and
power consumed, bit wear and replacement cost, drilling rate, and cost of
drilling.
This evaluation depends on several factors, namely: 1) the design and
operating parameters comprised of drill, rod, and bit which should be
selected to match the rock type; 2) lubrication and cleaning variables,
i.e., circulation fluids usually either water, air, or mud; 3) the hole
geometry, i.e., size and depth, which are exogenously determined; 4)
environmental and geologic factors such as the rock properties, structural
and petrological properties of the rock, as well as the overburden stress,
all oj: which determine the drillability of the overburden; and 5) factors
external to drilling such as cost of labor, job site, scale of operation,
type and availability of power to the site, climatic conditions, and manage-
ment and supervisory efficiency.
Selection of Drilling Equipment
Even after good mathematical analysis, the pit engineer has to use
judgement to arrive at the selection of drilling equipment. The following
procedure would aid in decision making:
1. Determination and specification of job conditions, i.e.,
labor, weather, and site location.
2. Determination or goal setting for the unit operations,
e.g., excavation, haulage, crushing capacity, production
required, pit geometry, fragmentation, and throw required.
3. Determination of number of drill holes, size, design, and
area of influence expected.
251
-------
4. Determination of rock strength and drillability and
selection of feasible drill methods.
5. Specification of operating criteria for each system to be
considered, i.e., rock, bit, and circulation fluid.
6. Estimation of the performance characteristics including
costs for depreciation, maintenance, power, and fluids.
7. Selection of the drilling system and equipment which best
meet the above requirements as well as has the lowest
overall cost (if possible) . Table 46 gives information
for selection of drill type based on rock strength and type.
Number of Drills
In surface coal mining, drilling and blasting is done for removal of
the overburden. Amount of drilling and blasting depends on the types of
strata present in the area. Hard rocks require more drilling and blasting,
and often the soil is not fragmented. The number of drills needed for
drilling can be calculated as follows:
SYE = Square yards of coal to be exposed
COB = Cubic yards of overburden removed
DS = Spacing of the drill holes, ft
DB = Burden of the drill holes, ft
DP = Drilling pattern (ft)2
LDPM = Length of drilling required per month (ft/month)
NH Number of drill holes
DR Drilling rate (ft/hour)
HPS Hours scheduled per shift
SPD Shifts per day
DPW = Drilling days per week
SPM Weeks worked per month
SMH Schedule monthly hours
DPM Number of drills required
DP DS x DB
NH
LDPM (^|) (27)
SMH HPS x SPD v DPW x WPM
LDPM
DPM =
(SMH)(DR)
EXPLOSIVES REQUIRED PER MONTH
The explosive commonly used in stripping is ANFO. Ammonium nitrate
is not readily explosive until the addition of the fuel oil, which acts
252
-------
TABLE 46. APPLICATION OF DRILLING AND PENETRATION METHODS
TO TYPES OF ROCK (PFLEIDER, 1968)
ROCK TYPE
RATING ->
CATEGORY ->-
EXAMPLES -»•
DRILLING METHOD 4-
rotary, drag
rotary, roller
percussion
jet piercing
1
Soft
(shale, weath.
limestone)
X
X
2
Medium Hard
(limestone,
weath. sandst.)
X
X
X
3
Hard
(granite,
chert)
X
X
X
4
Very Hard
(quartzite,
taconite)
V
TT
X
253
-------
like a catalyst to enhance its explosive strength. It is much safer, less
expensive, easier to prepare and more effective for rock blasting than most
comparable explosives. ANFO is available in bulk premixed, on-spot mix and
bagged forms.
Fundamental to effective blasting is the use of correct amount of
explosive for the volume of overburden to be fragmented. This relationship
is usually expressed as the "powder factor," which is defined as the ratio
of cubic yards of overburden fragmented per pound of explosives. It varies
from 2.5 to 9.0 (from hard rocks to soft soils).
To determine the amount of explosive needed per foot of hole, the area
of influence of a drill hole (DP) is divided by 27, and then divided by the
powder factor "PK" .
21 PK
Therefore, knowing the number of holes drilled per month, the drilling
pattern and the overburden depth, the amount of explosive required per
month (AANFO) can be calculated:
DC Length of explosive charge in hole in feet
AANFO = Amount of explosive required per month
DP 1
AANFO = (~-} • (--) * (NH) ' (DC) lb
Alternatively, if the entire overburden is to be blasted, then
where
COB Cubic yards of overburden removed
But it is common knowledge in stripping, that not all the overburden
requires fragmentation before removal. This is particularly true for
topsoil and soft soil. Therefore, actual consumption of explosive must be
calculated on the basis of experience with the overburden, i.e., for
individual application, the fraction of overburden that should be fragmented
must be known.
COAL LOADING
Although there are several methods of coal loading, the method consider-
ed in this report is a shovel-truck combination.
254
-------
Here, shovels are selected to match the advance of the stripping equip-
ment and the capacity of the trucks. Even when matching is achieved, the
general trend is to select a loader with a slightly higher loading ability
than calculations show for normal operations. The calculations for selection
of a loader are as follows:
CTL = Loader cycle time in seconds
BL Bucket size of loading shovel in cu yd
TPB Tonnage per bucket
F Factor for converting 1 cu yd of loose coal to tons
SMH Scheduled monthly hours
SMH HPS x SPD x DPW x WPM
NPH Number of cycles
NPH
TPB (, .) NPH
1 - A
TPB
BL F
A loader with this capacity could easily handle the monthly production.
Truck Fleet Selection
The truck capacity and loader capacity should be matched so that the
truck waiting time on loader and loader waiting time for trucks are
minimized. The calculations for matching the loader x
-------
The travel time and return time of the truck used above should be calculated
using average velocities for loads (SI) and empties (S2). Figure 102 will
aid in the sizing of trucks with regard to shovel bucket capacity.
SECONDARY EQUIPMENT SELECTION
The secondary or support equipment in surface mining ranges from highly
sophisticated construction equipment to small farm machines for reclamation
and revegetation purposes. In this section, the support equipment considered
are bulldozers, scrapers, and front-end loaders. These machines are used
for initial land clearing, mine road construction, rehandling of spoil,
loading, and haulage and reclamation jobs. They are virtually indispensible
in tight spots where larger machines cannot easily maneuver.
As with shovels and draglines, the secondary equipment are associated
with fixed and variable components in their cycle times. For example,
Atkinson (1971) gives the following values for the fixed and variable com-
ponents of the cycle time for front-end loaders (Table 47 and Figure 102).
According to Caterpillar Tractor Company (1975), a fixed time of 0.40 minutes
(24 seconds) for front-end loaders working in loose granular material on a
hard-smooth operating surface is recommended. This fixed time includes load,
dump, four reversals of direction, full cycle of hydraulics and minimum
travel.
Since material type affects the fixed time of the loader, Caterpillar
Tractor Company (1975) recommends the corrections given in Table 48 for
other material characteristics and operating conditions.
For scrapers, Caterpillar (1975) recommends the use of 0.7 minutes
(42 seconds) each for loading, maneuvering, and spreading.
There are several methods by which the secondary equipment can be
selected. They should match the performance of the primary stripping equip-
ment. Two alternative selection methods are outlined here:
A. Procedure I. Briefly, this procedure consists of the following steps:
1. Select a suitable bucket (or blade) size. The capacity
is usually expressed in cubic yards.
2. Calculate the variable part of the cycle time.
3. Determine the fixed times from time study. Sometimes,
this can be obtained from manufacturers' handbooks.
4. Calculate the total cycle time by adding the variable
cycle time to the fixed cycle time.
5. Calculate the number of passes per hour.
6. Determine the hourly production by multiplying the number
of passes (cycles per hour) by load per cycle (capacity).
256
-------
TABLE 47. FIXED TIMES FOR FRONT-END LOADERS (ATKINSON, 1971)
BUCKET CAPACITY
Note:
yd3
5
6
7
10
12
15
The application
FIXED TIME
Easy Digging Medium
32
33
33
37
39
41
of front-end loaders in
SEC
Digging Medium-Hard
33
34
35
39
42
44
hard digging
41
42
44
51
56
60
Digging
conditions is
marginal and requires comprehensive investigation.
257
-------
100
300
80
-------
TABLE 48. MISCELLANEOUS CORRECTION FACTORS (CATERPILLAR, 1976)
MATERIALS MINUTES TO BE ADDED TO 0.4 MINUTES
Mixed +0.02
Up to 1/8 inch + 0.02
1/8 inch to 3/4 inch 0.02
6 inches and over + 0.03 and up
Bank or Broken +0.04 and up
Conveyor or Dozer piled 10 feet and up 0.00
Conveyor or Dozer piled 10 feet or less +0.01
Dumped by truck +0.02
Management Conditions
Common ownership up to - 0.04
Independently owned trucks up to +0.04
Constant operation up to - 0.04
Inconsistent operation up to +0.04
Small target up to + 0.04
Fragile target up to + 0.05
259
-------
To arrive at the actual production, correction factors from the manufacturer's
handbook or from experience are applied.
B. Procedure II. Here, one can use the maximum production per hour given
by the manufacturer for some standard condition. To determine the actual
production, job condition correction factors must be applied to the maximum
production.
Correction Factors. Most of the correction factors discussed earlier for
draglines and shovels equally apply to the secondary equipment. Additional
factors that apply to secondary equipment are discussed here. Some of these
factors are presented in Tables 49 and 50.
Material Factor. The performance of bulldozers, scrapers and front-end
loaders are affected by material constituents and characteristics. Since
these equipment have varied applications, from loading and grading in loose
material to excavating hard frozen grounds, their capacity for production as
given by manufacturers must be corrected by a factor known as material
factor. Table 50 gives such factors for bulldozers.
Grade Factor. Since secondary equipment are mobile, their production
capability is affected by grade of the roadway and the work area conditions.
Factors are applied to production capabilities as given by manufacturers
to correct for the grade effect.
Visibility Factor. Impairment of operator's visibility due to atmospheric
conditions such as snow, fog, rain, etc., affect the performance of the
secondary equipment. Visibility factors are applied particularly to bull-
dozers and scrapers.
Transmission Factors. This factor accounts for the transmission mode of
the equipment: direct or automatic drive. The effect of the transmission
train mode on the equipment's productivity can be corrected by applying
transmission factors to the production capability given by manufacturers.
Dozing Factor. This factor accounts for the mode of operation of a bull-
dozer, whether slot dozing or side-by-side dozing. It applies to bulldozers
only and relevant values are given in Table 49. Factors for operating effi-
ciency and job efficiency, as applied to primary equipment, are also applied
to secondary equipment. Values for these factors are given in Table 49.
1. Bulldozers. Bulldozers are used extensively in surface mining.
They are extremely useful for dozing material through short distances under
severe conditions. Selection of a suitable bulldozer can be accomplished
by calculating,
Bb Bucket or blade size (machine capacity) in cu yd
The choice of blade type depends on the kind of job application and soil
conditions. Table 51 shows bulldozer blade selection chart according to
job application.
260
-------
TABLE 49. BULLDOZER CORRECTION FACTORS (CATERPILLAR, 1976)
Job Condition Corrections:
OPERATOR - Excellent
Average
Poor
MATERIAL -
Type-
Loose stockpile
Hard to cut; frozen-
with tilt cylinder
without tilt cylinder....
cable controlled blade...,
Hard to drift; "dead" (dry,
non-cohesive material or
very sticky material)
Rock, ripped or blasted -
SLOT DOZING
SIDE BY SIDE DOZONG
VISIBILITY Dust, rain, snow
fog or darkness
JOB EFFICIENCY 50 min/hr. . .
40 min/hr. ..,
DIRECT DRIVE TRANSMISSION ,
(0.1 min. fixed time) ,
*BULLDOZER Angling (A) blade,
Cushioned (C)
blade
D5 narrow gauge..
Light material
U-blade (coal).,
Blade bowl
(stockpiles)
Track-
Type
Tractor
1.00
0.75
0.0.60
1.20
0.80
0.70
0.60
0.80
.60-0.80
1.20
.15-1.25
0.80
0.84
0.67
0.80
,50-0.75
.50-0.75
0.90
1.20
1.30
Wheel-
Type
Tractor
1.00
0.60
0.0.50
1.20
0.75
0.80
1.20
1.15-1.25
0.70
0.84
0.67
0.50-0.75
1.20
1.30
*Note:
Angling blades and cushion blades are not considered
production dozing tools. Depending on job conditions,
the A-blade and C-blade will average 50-75% of straight
blade production.
261
-------
TABLE 50. SWELL FACTOR - BULLDOZED MATERIAL (ATKINSON, 1971)
MATERIAL SWELL FACTOR
Broken rock 1.65
Heavy clay (wet) 1.45
Earth with boulders 1.35
Earth 1.25
Sand and small gravel 1.1
262
-------
TABLE 51. BULLDOZER SELECTION CHART FOR RIPPING AND DOZING
APPLICATION (SKELLY AND LOY, 1975)
Job
Clean up
Around
Small
Shovel
Clean up
Around
Large
Shovel
Waste
Dump
Dozing
Capabilities
Ground
Condition
Friable
Hard
Friable
Hard
Jagged
Friable
Hard
Friable
Hard
Jagged
Friable
and Hard
Friable
Hard
Jagged
Wet
Muddy
Rubber
Tired
Dozers
rH
rH
tO
&
X
X
X
Medium
X
X
X
X
X
X
X
X
0)
ao
M
3
X
X
X
Track
Dozers
iH
i-l
cd
CO
X
X
X
X
X
Medium
X
X
X
X
X
X
X
X
X
X
X
OJ
M
i-i
cd
hJ
X
X
X
Remarks
No Ripping
Medium Ripping
Hard Ripping
No Ripping
Medium Ripping
Hard Ripping
No Ripping
Medium Ripping
Hard Ripping
No Ripping
Minimum Ripping
Medium Ripping
No Ripping
Medium Ripping
Hard Ripping
No Ripping
Medium Ripping
Hard Ripping
9 (26 Ton Trucks)
32 (59 Ton Trucks)
>32 (50 Ton Trucks)
9 (26 Ton Trucks)
32 (59 Ton Trucks)
90 to 180 Tons/Hour
270 to 720 Tons/Hour
>720 Tons/Hour
263
-------
FT Fixed time in min
TF Travel loaded time in min
TE = Travel empty time in min
VT = Variable time in min
MT = Maneuver time in min
CT = Cycle time in min
LT Loading time in min
NPH Number of passes per hour
PPH = Production per hour (theoretical) in cu yd
MAXPH = Actual Production per hour in cu yd
HD = Haul distance in ft
SI Loaded speed in mph
S2 Empty speed in mph
Bb = Bucket size of bulldozer in cu yd
LF = Load factor
BF = Bucket fill factor
AO = Operating factor
The selection procedures are outlined here.
FT = MT + LT
TF =
TE
(SI)(88)
HD
(S2)(88)
VT = TF + TE
CT = TF + FT (minutes)
NPH = — (cycles per hour)
PPH (NPH)(Bb)
k
MAXPH = (PPH)(LF)(BF)(AO)(I£I CF±)
where k is the number of other applicable correction factors from the manu-
facturer's handbook, and CF1 is the value for the ith correction factor.
Alternatively, the actual production can be calculated as,
k
MAXPH (MAXPHG) (LF) (BF) (AO) (^ CFj
where MAXPHG is the maximum production obtained from catalog or handbook
under specified conditions.
264
-------
2. Scrapers. There are several types of scrapers, namely, elevating
and conventional scrapers, and single- and double-engine powered scrapers to
mention a few. The major distinction is between wheel-tractor scrapers and
crawler-tractor scrapers. Wheel scrapers, because of their mobility, are
used in surface coal mining for topsoil removal, road building, etc. Shown
in Figure 103 are the scraper cycle times as functions of distance and grade.
Using the same parameters as before and denoting the bucket capacity
by "BS", the following relationship can be established. The calculations
for the cycle time and the number of passes of cycles per hour are the same
as for the dozers,
BS = Bucket size of scraper in cu yd
PPH = (NPH)(BS)
MAXPH = (PPH)(LF)(BF)(AO)(.ir CF.)
Alternatively,
k
MAXPH = (MAXPHG) (LF) (BF) (AO) (^CF )
where k is the number of other applicable correction factors from the manu-
facturer'^ handbook and CF^ is the value if the ±^ correction factor. For
calculating the number of pusher tractors, the pusher cycle time must be
known. The maximum number of scrapers that can be served by one pusher
tractor is given by rounding the following ratio to the next lower integer,
Number of scrapers Scraper cycle time
per pusher tractor Pusher cycle time
Number of pusher Total number of scrapers
tractors Number of scrapers per pusher
3. Front-end Loaders. In selecting front-end loaders, one must avoid
overestimating the capacity of the bucket. Improper bucket-sizing may lead
to instability of the equipment. There are three Society of Automotive
Engineers (SAE) ratings for front-end loader bucket capacities: a) struck
capacity; b) heaped or rated capacity; and c) static tipping load.
BL Bucket capacity in cu yd
DP Dumping time in min
MS Maneuvering time in min
Ft Total fixed time in min
LT Loading time in min
CT Cycle time in min
MAXPH Actual production per hour in cu yd
LF Load factor
BF Bucket fill factor
AO Operating factor
NPH = 60/Ct
PPH = (BL)(NPH)
265
-------
Ml I
J200
7000
1800
1600
2 1400
z
o
o
iTl
s
651B (37.50X39) DISTANCE VS TIME - LOADED
0% 2%
? 1C 2.80
TIME - MINUTES
J.ZO
651 B (37.50 X 39) DISTANCE VS T!ME - EMPTY
z
n
MET
72OO
7000
1800
1GOO
1400
1700
1000
800
600
400
no
fits n
j— 7000
6000
~ 5000
1000
3000
2000
1000
— 0
ET 0% 4% 6% 8%
t t * *?f A
- - f-^^r-^ .
4*2*** ' *•
—jr~ ' ~
)
'->•
'- T. " , - • 7 "
•ri
^
y
Xx
-^
!£-
/
-f^^-
^
X1
/
/
X
x
1
/
/
,x
^x"
-^"
/
s^
f/
X
x^
/
/
^^
^~
/
<^ —
x
X
x-"
^--
/
r
7
/
^
r^~
/
/
/
/
^
^
/
/
71
//
^
^"
/
/
_/*
X*
x^
r^'
y
/
/
X
^"
r---'
/"
x
/
^
^^~
X
X
.x-1
. —
/
X*
x
^.
__^.
VEHICLE
^
x°
^^^
.'"'
EM
X
X'
'^
PTY
jWx
^'~
^
WE
X
—
«^---
IGH
__X
,/-
^-"
T -
(
-^
^
124
562
X
x^
„--
000
00k
-^
-^
\b
g)
8.00 ? in J.BO 3.80
TIME -MINUTES
1054
15%
> 2
o •"
m "*
31 CT
O ID
r- >
Figure 103. Scraper distance versus travel times (Caterpillar, 1976).
266
-------
MAXPH (PPH) (LF) (BF) (AO) ( i^CF )
where k is the number of other applicable correction factors from the manu-
facturer's handbook and CF-^ is the i^1 correction factor.
Alternatively, Table 52 provides manufacturer production capacities for
front-end loaders, from which actual production can be calculated by applica-
tion of the correction factors,
k
MAXPH = (MAXPHG)(LF) (BF) (AO) (j^CFj
In using the correction factors average conditions should be assumed where
performance data is not available. Some correction factors are a combination
of two or more other factors. For example, "AO" the operating factor, is a
product of operator efficiency factor and job management factor. Where
accurate data are available, results would reflect the actual performance.
SUMMARY
This chapter has briefly reviewed the engineering considerations in
equipment selection and pit design. Equipment selection formulas have been
provided to select the bucket size, reach of machines, width of pit, number
of drills, loading shovels, trucks, and bulldozers to achieve a production
target. Formulas have also been provided to select bucket-wheel excavators,
scrapers, and front-end loaders. Important job correction factors were
discussed. In almost all cases, the selection of other equipment is dictated
by major stripping equipment; because of this, great care must be exercised
in their selection.
267
-------
TABLE 52. WHEEL-LOADER PRODUCTION ESTIMATING TABLE (CATERPILLAR, 1976)
CUBIC YARDS/60 MIN. HR.
ESTIMATED BUCKET PAYLOAD - BANK CUBIC YARDS
Estimated
Cycle Time_s_
Bucket Sizes
Hundredths
.35
.40
.45
.50
.55
.60
.65
.70
.75
Cycles
Per Hr.
171
150
133
120
109
100
92
86
80
1.0
150
133
120
109
100
92
1.5
225
200
180
164
150
138
2.0
300
268
240
218
200
184
2.5
375
332
300
272
250
230
3.0
450
400
360
328
300
276
3.5
525
466
420
382
350
322
4.0
530
480
436
400
368
342
4.5
600
540
490
450
416
386
5.0
665
600
545
500
460
430
5.5
730
660
600
550
505
474
6.0
800
720
655
600
555
515
(continued)
268
-------
TABLE 52. (continued)
CUBIC YARDS/60 MIN. HR.
ESTIMATED BUCKET PAYLOAD - BANK CUBIC YARDS
Estimated
Cycle Times
Bucket Sizes
Hundredths
.35
.40
.45
.50
.55
.60
.65
.70
.75
Cycles
Per Hr.
171
150
133
120
109
100
92
86
80
6.5
865
780
705
650
600
560
7.0
840
765
700
645
600
560
7.5
900
820
750
690
645
600
8.0
960
870
800
735
690
640
8.5
1003
925
850
780
730
L 680
9.0
1080
980
900
830
775
720
9.5
1140
1008
950
875
815
760
100
1200
1090
1000
920
860
800
JOB EFFICIENCY
Worktime/Hr.
60 Min. Hr.
55
50
45
40
EFFICIENCY FACTOR
100%
91%
83%
75%
69%
BUCKET LOAD FACTOR
Bucket Size X 1.00
.95
.90
.85
.80
.75
269
-------
SECTION 9
RECLAMATION
INTRODUCTION
Competition for land to be used for agriculture, living space, recre-
ation, and industry is becoming more intense. Mineral exploration, develop-
ment, extraction and processing all have different impacts and conflict
with these essential uses. The very nature of the non-renewability of a
mineral deposit necessarily makes mining a temporary activity, to be com-
pleted sooner or later. The importance of mining, particularly surface
mining, however, can hardly be overemphasized as it is doubtful if there is
any aspect of human life which does not require the products of mining. It
is known that over 90% of the mined products are obtained through the surface
method. In fact, more than 75% of the land affected by surface mining is
the result of the extraction of coal, crushed stone, and sand and gravel.
Therefore, it is necessary to be concerned with the concurrent and sequential
land use for optimum land utilization. The allocation of land for agricul-
ture, forest, wildlife, urban growth, recreation, transportation, and so
forth, is compounded by political, economical, sociological, and ecological
considerations. When some of these lands hold mineral values, the big
question is: How can one mine these riches without impairing the utility
value for satisfying land demands during and after mining? While this is
too frequently interpreted as a problem, it should also be recognized as an
opportunity to remold the surface for a desirable end-use.
LAND EFFECTS
Even though surface raining of coal constituted only 0.3% in 1914 and 7%
in 1937 of the total U.S. coal production, the use of surface mined lands
for other productive purposes was demonstrated and pursued by some coal
companies voluntarily as early as 1920. In fact, prior to 1965, there were
only seven states which had specific laws requiring land reclamation with
West Virginia enacting the first legislation in 1939. The post-1965 period
has been one of rapid growth in laws and regulations with regard to environ-
mental control in general, and surface mining in particular, both at the
federal and state levels. This discussion of land effects is provided only
to serve as an overview of the genesis of the environmental problems
associated with surface mining.
270
-------
Land disturbances from surface mining, in terms of area affected, have
grown to enormous proportions in recent years, for the following techno-
logical reasons: as easily mined and richer deposits are fast being extrac-
ted, and in some cases, completely exhausted, attention has been directed
towards mining larger tracts of lower-grade, higher cost deposits; in surface
mining of coal, the ratio of overburden thickness to that of coal has been
increasing; and current operations are involved with handling more material,
and often from greater depths, than heretofore. The need to exploit vast
reserves with thick cover and heavy relief has resulted in the development
of larger and more powerful equipment which can displace and remove enormous
amounts of materials at a much faster rate.
Land damages that are cited against surface mining are mainly due to
the destruction of surface topographies and of soil conditions that existed
before mining commenced. Often, the potential productivity of the soil for
plant growth is greatly reduced after mining. Soils that are disrupted by
these operations are often chemically active and toxic, thereby becoming a
source for water pollution. Also, if the overburden is a massive rock
formation, huge blocks of rock occur in the graded spoil which make it
difficult for the smooth passage of farm machinery. Much larger areas are
also affected by the unconsolidated spoil heaps and voids because these
conditions affect drainage patterns. Here, the natural processes of erosion
and sedimentation are accelerated, moving huge volumes of soil into receiv-
ing streams. When not properly graded and maintained against dumping waste
and refuse, these sites become potential health hazards. Finally, the steep
walls created as a result of thick burdens and heavy relief, coupled with
the possibility of ground movement with lapse of time, present a safety
hazard in inhabited neighborhoods and are aesthetically unattractive.
Additional areas may also be affected by the associated haulroads and
preparation and processing plants.
A summary of the major environmental effects due to the various
surface mining unit operations is presented in Table 53. A detailed report
on this subject has been prepared by Grim and Hill (1974). While this
section is concerned with surface mining effects on land and land reclama-
tion, the effects on water, air, wildlife, and other resources must also be
considered. In fact, all these effects are not mutually exclusive. For
example, air pollution may eventually lead to water pollution and both may
degrade the land. Surface mining without control may result in serious
blights and hazards to public health, contaminate air and water resources,
adversely affect land values, create public nuisances, and generally
interfere with community development. In fact, the controversy with regard
to surface mining of coal is so intense that some have sought a complete
ban on surface mining of coal and a greater emphasis on underground mining.
RECLAMATION PRACTICES
Rehabilitation of surface-mined lands is, however, a tangible program.
In fact, strip-mined land, abandoned open pits, and spoil and waste piles
are being shaped and graded, stabilized and seeded, and returned to produc-
tive use with proper care and management. Mined-area spoil banks that can
271
-------
TABLE 53. POTENTIAL MAJOR EFFECTS OF SURFACE MINING UNIT
OPERATIONS ON LAND, WATER AND AIR RESOURCES
NJ
•~J
N5
T.ur face
Mining
Unit
Operations
1. Exploration
2 . Area Dewatering
Diversion, Etc.
3. Drilling
4. Blasting
5. Stripping (Over-
burden Removal)
6 . Haulage
7. Top Soil or Other
Soil Storage
8. Maintenance
9. Beneficiation
Land and Soil
Soil Erosion
Overburden Swelling
Toxic Strata
Soil Inversion
Soil Stability
Landslides
Spoil Piles
Oil Spills
Coal Spills
X
X
XX X
XXX XX
X XX
XX X
X
Water
Aquifer Effects
Runoff Alteration
Sediments
Toxic Substances
Groundwater
Contamination
Industrial Effluents
Sludge
X
X X X X
X X
X X
X
X X
X X X X X
Air
Exhaust Emissions
Dust
Noise
Other (Welding, etc.)
Blasting Fumes
XXX
XXX
XX X
X X X X
XXX
X
X X X X
X X
Wild Life
Habitat Altered
Species Diversity
Aquatic Life
Animal Essentials
Accident /Deaths
Soil Organisms
Vegetation Potential
Wildlife Disturbed
X
X X X X X X
X
X XX
X X
X XXX
X XXX
Other
Aesthetic
Dangerous Material
X X
X
X
X X
X
-------
be revegetated present minor problems and have great potential for develop-
ment. Various reclamation programs that are being actively pursued include
returning the ground for agricultural and livestock farming, for reforesta-
tion, for recreation, and for housing and industrial sites (Figures 104,
105, 106, 107, 108, 109). The possibilities for development under these
conditions are limited only by cost benefit and innovation. There are,
however, marginal and problem spoils (acid, toxic, etc.) which require
special attention and even better planning. Concern for good mining
practices and effective land reclamation programs is witnessed by the adop-
tion of new federal regulations and an increased research and development
effort.
On August 3, 1977, the U.S. Congress enacted Public Law 95-87 titled
The Surface Mining Control and Reclamation Act of 1977. It represents a
significant effort on the part of the Federal Government to establish a
nationwide program to protect society and the environment from the adverse
effects of surface mining operations. Among the stated purposes of the
1977 Act, the following are particularly relevant to mine planning and
operation:
1. To assure that surface coal mining operations are so
conducted as to protect the environment.
2. To assure that adequate procedures are undertaken to
reclaim surface areas as contemporaneously as
possible with the surface coal mining operations.
3. To assure that the coal supply essential to the Nation's
energy requirements and to its economic and social
well-being is provided, and strike a balance between
protection of the environment and agricultural
productivity and the Nation's need for coal as an
essential source of energy.
Even without the federal law, new or amended surface mine regulations
were being enforced in 38 states. At the present time, several states are
revising their statutes to be in accordance with the federal law.
Currently 99% of the nation's surface mined coal is being produced
under land reclamation programs (Daley, 1974). In the history of the U.S.,
surface mining has disturbed 0.177 percent of the land and about one third
of that has been judged as reclaimed by the U.S. Department of Interior.
In 1972, there were over two million acres of land requiring reclama-
tion, and about two million not requiring reclamation for a total of
4,004,900 acres of disturbed land.
In the seven-year period between 1965 and 1971 private landowners
reclaimed over a third of a million acres. More than 90 percent of the
four million acres disturbed up to 1972 is privately owned. The USDA
Soil Conservation Service (SCS) began participating in land restoration in
the 1930's. Soil Conservation Service is active in developing new plants
that can survive under the difficult slopes and acid conditions found on
273
-------
hO
^J
.0
Figure 104. On land reclaimed with grass and trees after strip mining, responsible
coal companies have created public recreation areas like this roadside
park and campground in southeastern Ohio.
-------
••. .
i-o
-j
Ul
Figure 105. Charolais cattle and Giant Canadian Geese on mined lands
-------
ON
Figure 106. Corn, as well as forage, can grow on former coal land. The ponds
in the background (former pits) provide watering places for cattle
-------
Figure 107. Charolais calves fatten in a feedlot on strip-mined land in
Indiana. The farm machinery shed in the background was a repair
shop for mining equipment in the days when the area was strip mined.
-------
Figure 108. A golf course on strip mined land,
278
-------
N>
-J
VO
Figure 109.
Various phases of surface mining and reclamation in a Kentucky coal mine: (1) virgin farm
lands, (2) reforested area mined in 1947-1951, (3) reclaimed area mined 1954-1968, (4)
reclaimed areas mined 1970-1971, partially contoured but not yet seeded, and (5) current
mining operation with exposed coal seam.
-------
most surface-mined land. Most research now underway in government and
industry is in six categories: revegetation, chemistry of overburden and
spoils, hydrology, earth movement and placement, haulroads, and land-use
potentials.
RECLAMATION PLANNING
The specific planning problem with regard to reclamation is one of
maximizing production and maintaining environmental quality to be consistent
with the constraints imposed by the regulatory agencies, and the need to
satisfy the profit objectives of the companies. Mining actions, on one hand,
use some resources to produce coal, and at the other extreme, produce waste
products, both of which affect the environment. There are direct costs and
revenues to the company and public, as well as indirect costs and benefits
which must be coordinated in the planning. The necessary planning steps are
(1) to make an inventory of the premining conditions, (2) to evaluate and
decide on the postmining requirements of the region to accomodate the
needs and desires of the affected groups, (3) to analyze alternative mining and
reclamation schemes to achieve the requirements optimally, and (4) to developan
acceptable integrated mining and reclamation scheme most suitable under the
technical, social, and economic conditions.
For planning reclamation, information is needed in several areas. The
initial concern from the environmental standpoint was the ability to reclaim
the land in a minimal manner; today, however the environmental studies prior
to launching a strip mine are much more detailed. The ecosystem factors that
need to be considered today are set forth in Table 54. The listing is not
complete, but points to the recognition of the fact that any attempt to
extract mineral resources involves a change in the physical, biological,
social, physiosocial, biosocial, and psychosocial factors. From the initial
construction of access facilities to prospect to the final completion of the
mining and beneficiation operations, the ecosystem balance is continually
being altered. A discussion of the major factors as they affect the reclama-
tion decision follows.
Physical and chemical attributes of the premining environment are
factors in several areas of design and planning, particularly in the engi-
neering design of excavation, materials handling, and ground support systems.
They are also necessary in assessing water quality control of the operation
and determining the potential future land uses.
Geological aspects of the coal occurrence (e.g., attitude, depth, thick-
ness, stratigraphy, etc.) is the basis for establishing alternative mining
and reclamation schemes including slope stability. The nature of the over-
burden, especially those components as sulfur, having potential environmental
impacts must be studied. The structural and stratigraphic evaluation which
effect the hydrological conditions are critical for effective mining and
reclamation. Figure 110 and Table 55 show the type of analyses that were
done in an exposed highwall in southern Ohio.
280
-------
TABLE 54. INFORMATION NEEDS ON ECOSYSTEM FACTORS
FOR SURFACE MINE PLANNING AND LAND USE
I. SOCIOECONOMIC INVENTORIES
(a) Land demands for mining and associated activities, and
availability, encroachment into critical areas, other
resources affected and foregone benefits.
(b) Past, current and future land use plans for the area-
industry, agriculture, wilderness, etc.
(c) Transportation network
(d) Recreation networks
(e) Archeological and historical interests
(f) Demographic and population characteristics
(i) employment patterns
(ii) tax base
(iii) land value alterations and their
impacts on local government
(iv) attitudes of public, political, and
educational institutions in the
community/region
II. GEOLOGIC ANALYSIS
(a) Topography
(b) Overburden Characteristics
(i) stratigraphy
(ii) physiography
(iii) geomorphology
(iv) chemical nature
(c) Soil and Toplayer Characteristics
(i) physical
(ii) chemical
(iii) limitations for vegetation and future land uses
(iv) limitations for engineering
III. COAL CHARACTERIZATION
(a) Coal seam attitude
(continued)
281
-------
TABLE 54. (continued)
III. COAL CHARACTERIZATION (continued)
(b) Coal seam thickness
(c) Coal seam analyses
(i) proximate analyses
(ii) ultimate analyses
(iii) washability analyses
IV. TERRESTRIAL ECOLOGY
(a) Natural vegetation, characterization, identification of
survival needs
(b) Crops
(c) Game animals
(d) Resident and migratory birds
(e) Rare and endangered species
V. AQUATIC ECOLOGY
(a) Aquatic animals fish, waterbirds, resident and migratory
(b) Aquatic plants
(c) Characterization, use and survival needs of aquatic
life system
VI. WATER RESOURCES
(a) Surface hydrology
(i) watershed considerations
(ii) flood plain delineation
(iii) surface drainage patterns
(iv) amount and quality of surface run-offs
(b) Ground water hydrology
(i) ground water table
(ii) aquifers
(iii) amount and quality of ground water flows
(iv) recharge potential
(continued)
282
-------
TABLE 54. (continued)
VII. CLIMATOLOGY
(a) Precipitation
(b) Wind airflow patterns, turbulence
(c) Humidity
(d) Temperature
(e) Climate type
(f) Growing season
VIII. OTHER
(a) Noise
(b) Paticulates in the air
(c) Aesthetics
283
-------
OHIO POWER HIGHWALL STUDY
Ground Surfo
'•'•'..• •'-• \ Mixed Red
•:-.- No. J4\ shale a Cle
East Pit Hi<
ce _. . ,
Thicknesj
(feet)
,y 4
No. 13 \Brown Shale 1.5
^ No. 12 1^ Limestone 1
• • • f • • - • ' \
• L, '*Uw . \ Broken Shale
^'.No. iOja 1 1\ (ye,|ow _brown) 4
*." •."•:'-•'-•' -'-V ^-Surface of
—r- ' ' 1 i
T i i i i
i i iii
l^J-r'-r1-
1 i ' 1 1 1
7 1 1 T 1
-^="-= = — _
^£No.B :±r — _~
i^No.T'S - ' \
'T No. 6' 1 i 3
'. 1 I T-- - ' J
i i] i ^ 1
- ' •- ' . ' . ' . ' — '
, No. £3 , , . .
i , •
i , . i
::". No. 4- :.•'.; : ;•
I I )
) 1 )
] No. 3 > >
) { ) 1
^"No.2 -I-.' '» i TJ
^Nal=-~^?=:
_ — ^-^ ^ — ^ ~~~ _
~^j=— , —
Bench
Sweet Water 6
Limestone
Green Shale 3
Red Shale 1.5
Yellow
Material 3
(rock-like)
Blue Shale 4
Rock
Yellow ,
Material
• Massive 5
r Sandstone
r Fibrous 195
Cool-Like
Black Shale 6
/Haulroad
jhwall
POUNDS PER ACRE
l pH
Fe Sulfates Soluble
(ferric) Salts
5.3 .4
7.7 .2
8.1 .3
8.2 ,1
8.1 .3
7.8 .5
8.0 .5
8.2 .3
3.8 22.5
3.4 23.5
3.1 25.0
8.4 1.3
Source:
40 320
22 420
30 100
25 360
91 800
82 940
61 520
61 660
106 3,200
289 5,000
3IO 5,500
61 1,200
Form Clinic of U.
Toxic
Aluminum
.42
0
0
0
0
0
0
0
.72
.66
.54
0
S.
Figure 110. Analysis of the east pit highwall, a problem wall
5 years old, Ohio Power Highwall Study (Krause, 1973)
28A
-------
TABLE 55. CHEMICAL ANALYSES OF THE EAST PIT HIGHWALL OHIO POWER HIGHWALL STUDY, 1967
(SOURCE: THE FARM CLINIC, WEST LAFAYETTE, IND.) (KRAUSE, 1973)
l-o
oo
Ui
Sample Organic
No. Matter
% pH
4 0.9 3.8
6 0.8 8.0
10 0.9 8.2
12 0.8 8.1
13 1.0 7.7
14 0.8 5.3
Low
Satis-
factory
Excessive
or toxic
Total ... ^
Sol. Nitr°8en
Salts N03
3200 10.8
520 9.2
360 12.7
180 8.5
420 9.7
320 8.1
<20
— 20-40
>40
NH3 P2
42 1
46 4
43 4
47 2
45 6
49 1
<30
30-50
>50
°5
.1
.4
.2
.6
.1
.5
K2°
91
112
40
44
82
52
Ca
1365
3500
3450
3640
3250
3080
Interpretat
<5 <80 <500
5-20 80-150 500
to
>20
1000
>150>1000
Fer- Fer-
ric rous
Mg Fe Fe Mn
Ib/acre
270+ 22.5 0.2
270+ 0.5 0.1
270+ 0.3 0.1
270+ 0.4 0.3
270+ 0.3 0.2
114 0.5 0.4
:ion
<• i nn
*• J.UU
inn i fin
>180
3.0
4.5
0.6
0.5
1.4
4.5
1-4
>4
>20
0
0
0
0
0
0
B
.34
.36
.40
.39
.42
.37
<0.32
0.32
to
0
>0
>0
.37
.38
.45
Cu Zn SO,
4
0.8 0.65 106
1.0 0.13 61
0.8 0.11 25
0.5 0.09 30
0.8 0.16 22
2.1 0.50 40
<• i <- 1 —
<, _[_ <. j. —
-i /, 1 A
>A >A —
Toxic
Al Mo
ppb
0.72 12
0 62
0 75
0 50
0 58
0.42 25
o s-sn
v. C:A
-------
Soils must be considered to ensure their proper removal, placement, and
postmining disposition. While the definition and identification of the
material to be classified as topsoil is fraught with difficulties, the char-
acteristics of the material to serve as soil after mining must be known. The
characteristics of topsoil should be such that it should ensure an effective
growth of vegetative cover than other materials. Topsoil and subsoil, though
not always the most ideal materials, do have certain advantages over other
materials, including higher organic matter, better moisture-holding capacity,
natural seed source, and soil microorganisms. The removal and storage of
topsoil and replacement of it on top of the graded soil, in many instances
leads to better revegetation potential, which in turn leads to better erosion
control. In many older mining districts, spoil banks are only sparsely
revegetated. This is due in part to the lack of nutrients in the spoil and
the water flow characteristics of the restored strata. In cases where bed-
rock material, shale, sandstone, siltstone, and fireclay have been mixed
with the original topsoil covering, inadequate fine-grained matrix material
and organic matter may be available to provide nutrients or moisture-holding
capacity. Too frequently, the fine grained material sinks through the in-
terstices in the coarse matrix. High infiltration capacity, high porosity
and high permeability are too commonly encountered when no attempt is made
at toplayer restoration and compaction. Color of the spoil has an effect
on the surface temperatures of the spoil banks. Temperatures in excess of
150°F have been reported on spoil surface composed of dark shales.
Regrading of spoils alone may be inadequate to provide a growth medium.
In fact, the mining operation provides an opportunity to modify the soil
strata to include burial of toxic soil strata and complete rebuilding of soil
layers at the top. A few of the soil modification and reconditioning methods
include treatment with lime or limestone, fertilizers, and other additions
such as mulch, fly ash or sewage sludge to adjust pH, nutrient and physical
conditions. Two important factors to be determined here are the amount and
depth of application of these modifiers.
Topography influences the selection of the type of mining, and mining
equipment. These topics were discussed in an earlier section. Severe
terrains also influence the methods of revegetation, and limit alternatives
for future land use.
Extremes ^.n climate affect the selection of equipment. Seasonal fluctu-
ations determine desirable planting periods. In addition to the problem of
permafrost in severe cold climates, operation of mobile equipment becomes
difficult. Heavy rains can cause erosion complicating water quality control,
sediment formation, and the establishment of vegetation.
Hydrology, both surface and underground, together with the physical
and chemical characteristics of overburden and coal, is important for any
reclamation program planning. A knowledge of water movement and quality in
the mine area and modification during mining is essential. In all cases,
erosion and sediment production are immediate concerns in the critical post-
mining, pre-revegetation time period, although their control during storms
is a long-term problem that requires detailed planning. The Surface Mining
286
-------
Control and Reclamation Act of 1977 requires that the surface mining opera-
tion be so conducted as to minimize the disturbance (1) to the prevailing
hydrologic balance at the mine site and in associated offsite areas, and
(2) to the quality and quantity of water in surface and ground water systems
both during and after surface coal mining operations. Among the specific
provisions of the law for the avoidance of acid or toxic mine drainage,
included are such measures as:
1. preventing or removing water from contact with
toxic producing deposits.
2. treating drainage to reduce toxic content which
adversely affects downstream water upon being
released to water courses.
3. casing, sealing, or otherwise managing boreholes,
shafts, and wells to keep acid or other toxic drainage
from entering ground and surface waters.
With regard to hydrologic balance, other provisions of the Act deal
with factors such as suspended solids and recharge capacity of the mined
area. The negative impacts of surface mining are to be minimized by:
1. conducting surface coal mining operations so as to prevent,
to the extent possible using the best technology currently
available, additional contributions of suspended solids
to streamflow or runoff outside the permit area. In no
event shall contributions be in excess of requirements set by
applicable State or Federal law.
2. constructing any siltation structures, prior to commencement
of surface coal mining operations. Such structures are to be
certified by a qualified registered engineer and constructed
as designed and as approved in the reclamation plan.
3. cleaning out and removing temporary or large settling ponds or
other siltation structures from drainways after disturbed areas
are revegetated and stabilized; and depositing the silt and
debris at a site and in a manner approved by the regulatory
authority.
4. restoring recharge capacity of the mined area to approximate
premining conditions.
5. avoiding channel deepening or enlargement in operations
requiring the discharge of water from mines.
6. preserving throughout the mining and reclamation process
the essential hydrologic functions of alluvial valley
floors in the arid and semiarid areas of the country.
7. other such actions as the regulatory authority may prescribe.
Other volumes of the final report (Volumes IV and V) have dealt with the
subject of water quality and quantity in greater detail.
While the scope of the present study does not include a detailed
discussion of the biological characterization of the mining area, such
typification is important in the selection of a reclamation plan. In many
287
-------
cases, the law provides for the protection of unique or endangered species
of plants and animals which require specific consideration. Such provisions
may even prevent the mining of coal.
Sociohumanistic factors, though very important, are another study area
which is outside the scope of the present report. Nonetheless, a reclamation
plan must consider attitudes of the public and current and future uses of
the land. Aesthetic or historic value of the land prior to mining may also
be important enough to preclude mining.
LAND-USE ANALYSIS
The current land-use patterns may provide the best guideline for
reclamation. For example, if there is a need for pasture lands, forested
hillside land adjoining a pasture may be reclaimed for pasture after mining.
Similarly, orphaned mined land in a wilderness area may be more worthwhile
if restored as a forest. Future land-use plans contrary to premining use
are unusual and not the common practice. Acceptance of mountaintop removal
and valley fill mining methods which create new flat areas in regions where
such areas are in short supply reflect the permanent, positive potentials of
the temporary, drastic disturbance of lands due to mining. Identification
of the current and probable future land uses can be summarized into the
following five categories:
1. "Wilderness" or unimproved use
2. Limited agriculture or recreation, with little development,
such as grazing, hunting or fishing or as timber land
3. Developed agriculture or recreation, such as crop land,
water sports, vacation resorts
4. Suburban housing or light commercial and industry
5. Urban housing or heavy commercial and industry
Analysis of alternative long-term uses requires compliance with
socioeconomic and public constraints as well as the economic goals of the
mining company. Discussion on this subject is necessarily limited here
since another report on land-use planning has dealt with this subject in
greater detail. Irrespective of the choice of short-term and long-term land
use, these plans must be linked to the mining plan in space, time and method.
Definition of the long-range plans do achieve a purpose by minimizing distur-
bance unnecessary for the determined future use and delimiting those reclama-
tion activities that are not a part of the normal mining or land-use plans.
Reference has already been made to the enactment of legislation at the
federal and state levels pertaining to surface mining and reclamation.
Several provisions of the Surface Mining Control and Reclamation Act
of 1977 and the interim regulations deal with land-use planning. Specifi-
cally, with regard to obtaining a mining permit, a reclamation plan must be
submitted which shall be of sufficient detail to demonstrate that reclamation
can be achieved as required by Federal or State programs. Such a plan must
cover, among other items, the following important points with regard to the
uses of land:
288
-------
1. the uses existing at the time of the application, and if
the land has a history of previous mining, the uses which
preceded any mining.
2. the capability of the land prior to any mining to support a
variety of uses giving consideration to soil and foundation
characteristics, topography, and vegetative cover.
3. the use which is proposed to be made of the land following
reclamation, including a discussion of the utility and
capacity of the reclaimed land to support a variety of
alternative uses and the relationship of such use to existing
land use policies and plans, and the comments of any owner of
the surface, State, and local governments or agencies thereof
which would have to initiate, implement, approve, or authorize
the proposed use of the land following reclamation.
4. a detailed description of how the proposed post-mining
land use is to be achieved and the necessary support activities
which may be needed to achieve the proposed land use.
5. the consideration which has been given to making the surface
mining and reclamation operations consistent with surface owner
plans and applicable State and local land-use plans and
programs.
6. the consideration which has been given to developing the
reclamation plan in a manner consistent with local physical
environmental and climatological conditions.
Even otherwise, mining companies today have to meet with nongovernmental,
historical, wildlife, and other environmental groups and allay their fears.
The integrated mining, reclamation and land-use planning diagram (Figure 111)
sets out an iterative procedure for the evaluation of mining and reclamation
plans. The feedback loops in the diagram serve to indicate the procedure
which, when the plans are optimal will lead to the mining and reclamation
of the coal seam. Shown in the Figure is also another set of interactions
that takes place before a mining plan is put into practice. The iterative
loop through which the mining and reclamation plans have to pass includes not
only the federal, state and local agencies but the special interest groups
and the general public. Finally, the ecosystem factors to be considered in
the development of a mine today are varied, many, and increasing. The
planning of surface mines, thus, is very involved and thorough, and the
objective of planning is the long-term use of the land as a prime natural
resource.
ECONOMIC FEASIBILITY STUDIES
In a free enterprise society, it is difficult to continue in business
unless there is an adequate rate of return in monetary terms on the invest-
ment. The goals of environmental protection and mineral resource extraction
must therefore be analyzed in terms of both the economic return and the con-
straints and incentives of the public and the government. At the mine level,
it is necessary to establish the direct cash cost of reclamation. The con-
sideration of internal, direct costs, and income from mineral sales, by
more easily quantifiable rate of return analysis, is traditional.
289
-------
VO
o
R EC 0 N N A I S S ANC E
SITE INVESTIGATIONS
BASELINE
DATA
ON
ECOSYSTEM
FACTORS
LITERATURE
SEARCH
'RESEARCH'
COMMENTS
CRITICISMS
REVISIONS
INTEGRATED
MINING PLANS
MINING PLAN
RECLAMATION
PLAN
LAND USE PLAN
ENVIRONMENTAL
___G_ROUPS_
PUBLIC MEETINGS"
FEDERAL, STATE
AND LOCAL
AGENCIES
LOCAL ZONING
STATE MINING PERMIT
H IGHWAYPERMIT
EPA, MESA, ETC.
FINAL
APPROVED
PLANS
M]NJ_NG
RECLAMATION
MONITORING
OTHER
PREDETERMINED
LAND USES
Figure 111. Integrated and interactive surface mining, reclamation and land-use planning.
-------
Grim and Hill (1974) have summarized the results of several studies on
reclamation costs. Mathematica Incorporated (1973) has estimated that
reclamation costs will be about 8% of the total mining cost (Table 56).
Other cost figures in literature vary from $1,236 to $19,768 per hectares
($500 to $8,000 per acre) depending on the severity of the land effects and
the reclamation requirements. Doyle et al. (1974), have prepared a useful
report on the analysis of pollution control costs. Their report addresses
the costs involved in revegetation, acid mine drainage, stream diversion,
strip mine spoil and waste, backfilling and regrading, etc. Table 57 is
extracted from that report to indicate the various factors that affect back-
filling and grading costs.
Since reclamation costs are site specific, the requirements of labor,
equipment and supplies must be estimated. A convenient basis for estimation
purposes may be to identify the acreage disturbed for the recovery of coal
from underneath an acre of ground. The costs of reclamation are calculated
for the disturbed land and then assigned to the tons of coal recovered.
Reclamation costs are preferably indicated both in $/acre and $/ton.
To minimize environmental disturbances and to maximize land use, it
is necessary to include the sociohumanistic benefits and costs. These are
not as easily quantified, defy traditional engineering approaches, and raise
conflicts between the various organizations' objectives. The consideration
of these benefits and costs are outside the scope of this report. The future
of surface mining hinges partially upon the ability to reclaim land in a
satisfactory manner. Where the environmental impacts cannot be clearly
ascertained, it may be that the mineral resource should not be mined.
However, in a majority of the instances, knowledge and technology are
available to achieve acceptable impacts. Results of several studies reveal
that no major breakthroughs are indicated in the near future. Only through
concerted efforts will evolutionary processes result that adapt present
equipment with mining cycles which reflect in greater social concern for
environment.
TECHNICAL FEASIBILITY STUDIES
Technical feasibility is a prime requisite for the adoption of any
reclamation plan. It is preferable to key-in the reclamation plans with
mining plans since nonconcurrent reclamation methods tend to be environmen-
tally less optimal and more expensive. Reclamation efforts initiated long
after the mining activity has ceased generally result in reclamation that
serves a limited land use. Concurrent reclamation with positive controls
aids in the development of well-conceived future land uses with a long-term
program to finish and maintain the reclamation plans. Concurrent methods
allow total spoil manipulation and/or complete spoil profile construction.
Technical requirements of reclamation will vary depending upon the
degree of restoration. In the standard surface mining approaches of today,
selective replacement of material is relatively simple. Contour mining
approaches such as block cut, box cut, haul back, mountaintop removal and
modified area mining are generally flexible to allow segregation and
291
-------
TABLE 56. ESTDIATED AVERAGE PRODUCTION COSTS @ AVERAGE STRIPPING
RATIO = 8:1 and 0.5 ACRES DISTURBED PER THOUSAND TONS
OF COAL PRODUCED* (MATHEMATICA, 1973)
COST ELEMENT
Sediment structure
Acreage fees
Bonding
Scalping
Stripping
Overburden haulage
Coal loading
Coal haulage
Backfilling and grading
Revegetation
Royalties
Severence tax
COST ($/TON)
(1")
0.12V ;
(1")
0.01V ;
0.00
(I)
0.08^ ;
2.40
0.05
0.10
0.50
(I)
0.08U;
(1)
0.03^ '
0.50
0.30
COST (% OF TOTAL)
2.9
0.2
0.0
1.9
57.5
1.2
2.4
12.0
1.9
0.7
12.0
7.3
Total $4.17 100.0
* Acre = 0.40 hectares; short tons = 0.907 metric tons
^ 'Reclamation costs. These costs are equivalent to $0.32 per ton or
7.6% of the total costs
Note The primary assumption underlying these estimates is that the
stripping ratio is 8:1; the cost balance would change somewhat
if the assumed stripping ratio were to change.
292
-------
TABLE 57. STRIP MINE RECLAMATION PROJECT VARIABLES AFFECTING
BACKFILLING AND GRADING COSTS (DOYLE, 1974)
1. Geographic location
2. Topographic setting (original, pre-reclamation and final ground
slopes)
3. Type of strip mine:
a) area, b) contour, c) area-contour, d) other
4. Coal seams mined and thickness
5. Inclination of coal seams in back of highwall:
a) dip, b) rise, c) horizontal
6. Condition of coal seams in back of highwall:
a) not mined, b) auger mined, c) drift mined (entries opened or
caved, d) mine workings exposed by stripping operation
7. The probable hydraulic head that could develop if coal in back of
highwall was mined
8. Strip mine area information:
A. Length, width, and area covered by spoil before reclamation
B. Highwall height (maximum and average height)
C. Highwall length
D. Number of cuts
E. Total area affected during reclamation in acres (including
area above highwall and outside of slopes)
F. Volume of spoil to be moved
G. Average haul distance for backfilling and grading
H. Texture of spoil
I. Amount of large rock and material requiring special handling
of mining timbers, machinery, and debris, junked cars, and
other solid waste
J. Amount and reactivity of pyritic material (minerology and mode
of occurrence. For example finely dispersed; single crystals
or crystal aggregates; coatings on joint surfaces; in form of
lenses, layers or modules; "sulfur balls"; pyritic shales, etc.)
K. Clearing and grubbing requirements
9. Type of backfill:
a) contour, b) pasture-reverse slope, c) swallowtail, d) head of
hollow, 3) submergence, f) other
(continued)
293
-------
TABLE 57. (continued)
10. Physical sealants for covering toxic material
a) none, b) clay, c) bituminous material, d) plastic material,
e) other
11. Compaction desired:
a) none, b) only toxic materials, c) all spoil material with
exception of upper layer up to 0.91 meter (1 to 3 feet)
12. Accessibility factors:
A. Right-of-way problems
B. Ingress and egress construction (include clearing and
grubbing for access and post-construction revegetation)
C. Other factors affecting access
13. Surface and subsurface ownership of strip-mined area. Also,
ownership of properties for ingress and egress:
a) public, b) private, c) in process of being acquired or
line placed on property, d) abandoned, 3) temporary easement,
f) other
14. Time of year reclamation performed
15. Weather conditions during reclamation period(s)
294
-------
selective replacement. Two such plans are illustrated in Figures 112 and
113 for contour mining. Larger scale area methods employing traditional
overcasting are less flexible. It is common practice to handle overburden
by direct overcasting preferably with no rehandle. The problem with simple
overcasting with a large machine is that the spoil can be cast only at a
limited distance from the cut, and then only at a limited range of angles to
the line of travel of the equipment. However, selective placement of spoils
in these cases can also be achieved by a combination of equipment and
methods. For example, the requirement to save topsoil and to spread it over
the regraded land has led the mining companies to the use of scrapers that
remove the topsoil, run around the pit and dump over formerly graded spoil.
Selection of shovel-truck systems, as opposed to that of conventional over-
casting, also provides the needed flexibility in soil placement. In fact,
the overburden conditions can be closely monitored, and the placement of the
overburden so achieved to ensure 1) burial of toxic spoils and less
desirable material such as boulders, or impermeable clays, 2) topsoil re-
placement, 3) desired stratigraphic sequence in the spoil, and 4) flatter
or other desired valley configurations. Additionally, truck shovel systems
permit blending in the pit for coal quality and greater recovery of coal.
The goal for premining planning and scheduling for resource extraction and
reclamation is to develop such steps as an integral part of the mining
method.
The question of restoration versus adaption of contour is one which is
decided by the general mining approach. In strict contour mining, in which
the seam is followed around a hill, the restoration to approximate original
contour with excess spoil properly disposed in the valleys is generally most
appropriate. Where there are flat lands, and the need is for small hills,
terrace and shallow hollows, the indicated approach may be to restore some
degree of relief by leaving the mined-out areas and spoils in such a
reclaimed state that the future potential may be for grazing, wildlife, or
recreation. In areas where the lands are used for agriculture prior to
mining, the plan may be to re-establish all mined areas to a more tillable
topography.
In backfilling, three major areas of long-term concern are 1) slope
stability, 2) ground water control, and 3) water quality maintenance. The
primary source of slope failure is the prolonged action of water, through
erosion, and mass wasting. The primary result of water's actions on slopes
is erosion and sediment production. Long uninterrupted slopes are undesir-
able as compared to ones with terraces and diversions which decrease slope
lengths and direct runoff water to safe outlets. Additionally, while back-
filling, the permeability of the fill for water perculation can be modified.
Such modifications are achieved by sealing or leaving the coal seam open
and by selective placement of the spoil. One of the concerns of water
quality control, in addition to the acid potential, is the sediment. The
interconnection between slope stability and water quality (sediment) is
enhanced by apparently conflicting approaches to these problems. Gross
stability is enhanced by establishing grade and cover to minimize water
infiltration and permeation. Conversely, erosion and sediment control are
achieved by encouraging infiltration and permeation to minimize runoff.
295
-------
UPPER LEACH LAYER
LOWER LEACH LAYER (LL)
TOPSOIL
Figure 112. Movement of spoil within a block to insure segregation
and burial of acid-bearing material (Saperstein and
Secor, 1973).
296
-------
ORIGINAL GROUND SURFACE
BACKFILLED GROUND SURFACE
WATER::;:
MINERAL SEAMfeM|r£l
yjj-r-T^W.i&J
BACKFILLED SPOIL
*/~
ACID SEEP
ACID-FORMING MATERIAL
POOR PLACEMENT RESULTING IN ACID SEEP.
•ORIGINAL GROUND SURFACE
DIVERSION DITCH
IMPERMEABLE
UNDERCLAY
/^BACKFILLED GROUND SURFACE
'^r HIG H WALL ;-y.;i '•: ••:•:•.;/>..
BACKFILLED SPOIL
IMPERMEABLE MATERIAL
ACID-FORMING MATERIAL
•GRADED MATERIAL 0.9 METER (3') MIN.
PROPER PLACEMENT FOR ACID FORMING MATERIAL BURIAL
Figure 113.
Spoil placement for better environmental impact
(Hill and Grim, 1975).
297
-------
The proper approach would entail, first, the establishment of as heavy a
cover crop after backfilling as possible, to encourage vegetative inter-
ception and evapotranspiration as opposed to surface or subsurface runoff.
However, not all of the water can be handled in this manner. Provision must
be made for storms by designing clean-running interceptions and diversions
for excessive rainfall.
Stability and water quality are inherent in the design of fills and
backfills. If a. suitable, large-rock drainage base and contact zone is
established at the base contact of the fill, slippage can be minimized while
rapid drainage of groundwater is enhanced. Such approaches to water quality
and the control of stability are exemplified by the head-of-hollow or valley-
fill concept, explained in the earlier section on mining methods.
Replacement of topsoil is practiced in mining where required by regula-
tions. The dictum of topsoil replacement could be viewed as a requirement
to replace or establish a layer of material of specific thickness, having
certain physical and chemical characteristics, and primarily, capable of
supporting some specific vegetation. In pre-mining planning, attention
should be directed at identifying all strata potentially able to adapt as a
surface covering. If topsoil is not available, it becomes necessary to
consider soil amendments to establish vegetation. Common approaches to soil
amendments involve the addition of lime, mulch, and fertilizer. Many mulches
have been used to assist in establishing vegetative cover. Straw or hay can
be manually applied, but they, as well as wood chips, are often spread
mechanically or hydro-pneumatically. Mulch should be tagged down with tar,
etc. Rapid growing annual plants and small grains can be used to establish
ground cover, to leave organic matter in the soil, and to serve as a mulch
for perennial establishment. The common practice today is to establish such
quick-cover mulch crops prior to topsoiling to assure vegetative stabiliza-
tion and to enhance the organic content and water-holding capability of the
surface layer. Specification of the graded surface, furrowing or harrowing,
is also done to act as seed and water traps. Care must be taken in selecting
species to avoid competition.
Fly ash, the residue of coal-fired boilers, can be applied to increase
nutrients, moisture and alkalinity and to ameliorate toxic materials present.
The trace elements contained in fly ash are often important micronutrients,
while fly ash is also rich in the common macronutrients except nitrogen.
Nitrogen deficiency can be handled by planting inoculated clover, alfalfa,
vetch, or other legumes. The use of fly ash may be particularly attractive
when the power plant is located at or near the mine and the ash must be
disposed of in some fashion (Adams, 1971).
Stabilized sewage sludge is another recyclable amendment to reclaimed-
mined land. Sludge is rich in nutrients and may reduce acidity, as well as
provide stable organic matter to form a humus. A related type of amendment
is composting with municipal waste, which has been practiced in other
countries but has seen limited application in the United States. Care must
be taken, however, to avoid poisoning the soil with too high a level of
heavy metals and salt from the sludge (Grim and Hill, 1974).
298
-------
In surface coal mining, positive control of environmental impacts is
achieved by proper mining and reclamation techniques which include concur-
rent reclamation, overburden segregation, toxic material burial, and top-
soiling. Grading and backfilling should immediately follow mining as the
freshly spoiled material is easier to move, backfill, and grade. Revegeta-
tion should follow grading as soon as possible. The primary purpose of
vegetation is to stabilize the spoil to prevent erosion and to ensure water
quality and quantity control. In most cases, therefore, revegetation is
best accomplished by the establishment of grasses and legumes, as soon after
mining as possible, though the long range plan may call for reforestation
or other land uses. It is possible to develop a time sequence for the
various reclamation activities. Table 58 presents such a sequence. For
operational management, this sequence must have specific dates for a
particular operation.
MONITORING AND MAINTENANCE
Even with proper reclamation, rededication of the land for productive
purposes may require more time. Grazing, for example, should not be allowed
immediately on a new cover crop. In any case, use of the land is not
definitely accomplished at the end of mining. A careful plan of monitoring
and control, and maintenance and development are required.
Monitoring slopes to assure stability and control to maintain water
quality are two important aspects. To minimize these problems initially and
to prevent them from occurring in the future, at the time of grading it is
necessary to keep all surfaces as minimally sloped as possible and under
the angle of repose. Additionally, through regular inspections any failure
or runoff is corrected by refilling, compacting, and rapid establishment
of cover. To control the erosion affects of running water, it is necessary
to place interceptor and diversion channels to reduce downslope flows.
Where the coal mine drainage phenomenon cannot be prevented, a long-
term program to neutralize the discharge is necessary. In most surface-
mining situations, much of the necessary layout is present in sediment-
control ponds. Addition of limestone, lime, anhydrous ammonia (where no
discharge to surface drainage is encountered), soda ash, sodium hydroxide,
or other suitable neutralizing reagents can minimize acid formation.
However, isolation of toxic material and concurrent backfilling is probably
the most desirable approach to minimization of acid formation problems.
One other aspect of maintenance is the sediment and erosion control
installations. With appropriate planning and no unusual storms, such
maintenance of these features may be minimal. However, control structures
may become clogged or filled and should be cleared and cleaned through
regular inspections.
299
-------
TABLE 58. TIME SEQUENCE FOR RECLAMATION ACTIVITIES
I. DURING SITE PREPARATION:
1. Install control measures (diversion, sediment traps and
basins, etc.).
2. Clear and grub, market lumber if possible; stockpile brush
for use as filters; run brush through woodchipper and use
chips for mulch.
3. Stabilize areas around temporary facilities such as
maintenance yard, power station, and supply area.
II. DURING OVERBURDEN REMOVAL:
1. Divert water away from and around active mining areas.
2. Remove topsoil and store it if possible and/or necessary
(Figure 114) .
3. Selectively mine and place overburden strata if possible
and/or necessary.
III. DURING COAL REMOVAL:
1. Remove all coal insofar as possible.
2. Break or prevent damage to the strata immediately below
the coal seam as desired.
IV. IMMEDIATELY AFTER COAL REMOVAL:
1. Seal the high wall if necessary.
2. Seal the low wall if necessary.
3. Backfill—bury toxic materials and boulders, dispose of
waste, ensure compaction.
V. SHORTLY AFTER COAL REMOVAL:
1. Rough grade and contour, taking these factors into consider-
ation (Figures 115, 116 and 117):
a.. Time of grading, specific time limit; tied to advance of
mining; seasonal considerations.
b. Slope steepness.
c. Length of uninterrupted slope.
d. Compaction.
e. Reconstruction of underground and surface drainage
patterns.
(continued)
300
-------
TABLE 58. (continued)
V. SHORTLY AFTER COAL REMOVAL:
2. If necessary, make mine spoil amendments (root zone), taking
these factors into consideration:
a. Type of amendment, fertilizers, limestone, fly ash, sewage
sludge, or others.
b. Depth of application.
u. Top layer considerations—temperature (color), water
retention (size consist, organics), mulching and tacking.
VI. IMMEDIATELY PRIOR TO FIRST PLANTING SEASON:
1. Fine-grade and spread topsoil, taking seasonal fluctuations
into consideration (Figure 118).
2. If necessary, manipulate the soil mechanically by furrowing,
deep-chiseling or harrowing, or constructing dozer basins.
3. Mulch and tack.
VII. DURING THE FIRST PLANTING SEASON:
Seed and revegetate, considering—time and methods of seeding;
choice of grasses and legumes (Figures 119 and 120).
VIII. AT REGULAR, FREQUENT INTERVALS:
Monitor and control: slope stability; water quality, both
chemical (ph, etc.) and physical (sediment); vegetation growth.
301
-------
fL\
Figure 114. A pusher-sr.raper combination for topsoil and overburden removal.
-------
UJ
c
U!
Figure 115. Large crawler tractor in the process of rough grading coal mine spoil banks in hilly
terrain.
-------
o
-L-
Figure 116. A giant bulldozer with an experimental 40-foot blade levels mined land in southeastern
Kansas.
-------
Figure 117. Completed rough grading of coal min* spoil banks in hilly terrain.
-------
•
UJ
o
a--
Figure 118. Bulldozer spreading topsoil on graded spoil.
-------
00
O
Figure 119. Coal mine spoil banks reclaimed to farmland.
-------
•
. ., ..... 4
•
o
CO
Figure 120. Area which was mined and reclaimed in 1975. The stripping equipment
can be seen in the background (Photo taken in July 1976).
-------
SUMMARY
This chapter has briefly reviewed the premining planning for reclama-
tion. Detailed discussion on water-management and land-use planning aspects
were avoided in view of the extensive background reports on these two areas.
A time sequence for reclamation activities was provided, along with examples
of good reclamation practices.
309
-------
SECTION 10
SURFACE MINING COSTS
INTRODUCTION
Economic evaluation' of various mining plans is necessary to make a
comparison of the alternatives. The key to preliminary analyses is the
stripping ratio since, for a given situation, it usually dictates the initial
decision to mine or not to mine by surface methods. Capital cost consider-
ations can hardly be over-emphasized. However, equipment costs vary widely
and are a function of the amount of steel and the fabrication in the design
and construction of the equipment. Cost involved for other surface facili-
ties (e.g., storage, office space, buildings, etc.) are also subject to great
regional variances. Also, the financing and accounting procedures of
companies differ, thereby making it difficult to arrive at meaningful strip
mining costs.
Estimation of the mining costs must be, by necessity, based on the
company's experience. The labor and material cost must be estimated for
drilling, explosives, overburden removal, reclamation, pit cleaning, coal
loading, haulage, road building, fuel, oil, grease, maintenance, supervision,
depreciation, etc. Additionally, costs for transporting, erecting, disman-
tling, and moving the primary stripping and other equipment must be con-
sidered. Since the viability of a project must be determined over the mine
life, these have to be projected into the future, taking into account the
inflationary and productivity trends.
The selling price of coal is an uncertain factor. It is a complex
function of the demand and availability of other energy resources and their
prices. The correlation between the selling price and the mining and prepa-
ration cost, on one hand, and the attractiveness of investment in stripping,
on the other, is strong.
STRIPPING RATIO
The stripping ratio is the proportion of overburden to coal removed.
Though there are several units in which this proportion can be defined, the
most common in the industry is in the in situ cubic yards of overburden
removed per ton of clean coal (1 ton = 2000 Ib). The equivalent unit in
310
-------
the metric system will be cubic meters of overburden removed per tonne
(1000 kg) of clean coal. Simple arithmetic leads to the conversion factor
between these two units:
3 3
of overburden/^ _ (yd of overburden/. , 1 _ >
tonne of coal ' ~ ( ton of coal ' * 4.2075;
The stripping ratio can also relate the selling price of coal with the
costs of mining the coal and stripping the overburden. In technical litera-
ture, the calculations are sometimes based on average overburden depths,
though in reality, the break-even stripping ratio is a point value beyond
which the coal seam cannot be economically stripped. As the overburden
depth increases, more money is spent on exposing the coal seam until a limit
is reached when the value of the recovered material (clean coal) is just
enough to pay for all the cost involved in mining, preparing, and selling
the material.
The removal of overburden, or stripping, is generally regarded as the
most significant component of surface coal mining costs. Variations in
stripping ratio affect the scale of the equipment and the efficiency
of its operation as far as overburden handling is concerned, while procedures
for handling, preparing and marketing coal, and costs associated with these
three steps are, in comparison, fixed from mine to mine.
Distinction is often made between several different ratio calculations,
such as actual or overall stripping ratio, ultimate or break-even stripping
ratio, daily or working stripping ratio, and the stripping ratio developed
on the basis of underground mining costs vs. surface mining costs excluding
stripping costs. The considerations of various ratios are also dependent on
the scale of the operation, the capabilities of equipment and the general
economic structure of the company.
Usually the first analysis is the calculation of a stripping ratio at
which recovery by surface mining is cost equivalent to recovery by under-
ground methods. Once a surface vs. underground ratio is established, the
amount of surface mineable reserves is established. In addition, whatever
resources can be recovered at a lower stripping ratio, it is being mined
more economically than is possible by an underground approach.
For example, assume that the underground mining costs for a 1.0-m
(3.3-ft) seam are $8.50 per tonne ($7.74 per ton). For the same seam,
suppose the removal costs by surface methods are $0.35 per tonne ($0.32 per
ton), and that the overburden can be removed by large equipment for $0.20
per m ($0.18 per yd3). Costs for regrading and reclaiming the mined area
are assumed to be $700 per ha ($0.07 per m^ or $283 per acre). The under-
ground vs. surface-stripping ratio (Rsu) may be calculated as follows,
realizing that there are 16145 tonnes/ha-m (1800 tons/acre-ft.) :
311
-------
Underground mining cost Surface mining and reclamation
R _ per tonne coal cost per tonne coal
o
Surface stripping cost per ™ overburden
$8.50 (($0.35) +
= $8.50 - ($0.35 + $0.04)
su $0.20
3
R 40.55 m overburden/tonne coal
su
Coal should therefore be mined only in that portion of the deposit
where the ratio does not exceed 40.55 m^ overburden per tonne coal. In a
deposit to be area mined, these calculations would show that the deposit
could be exploited by surface mining as long as the overburden covering was
less than 64.88m (197 ft). These conditions would be encountered in the
Ohio eastern regions which incorporate the largest stripping equipment to
such depths of overburden.
Total design of the mine requires that an ultimate break-even stripping
ratio be calculated. This ratio is that which establishes the ultimate
limits of mining, and is not the overall stripping ratio. The break-even
ratio is that ratio at which the total cost of marketing one tonne of coal
is equal to the price of that tonne of coal. The overall ratio must be less
than the break-even ratio or no profit would be realized. The break-even
ratio (RT,P) is calculated as:
Recoverable value per Production cost per unit
unit of coal of coal
Stripping cost per unit waste
f)
Working in common units of tonnes of coal and m-5 of overburden, a family
of analyses and curves may be prepared based on various costs and prices
which can be estimated.
Table 59 shows an analysis accounting for variations in percent reject
and market price of coal.
In reality, RBE (or any stripping ratio) is most commonly a linear
polynomial function in several variables, including those as summarized in
Table 60.
312
-------
TABLE 59. ANALYSIS OF STRIPPING RATIOS
PER CENT REJECT
10
15
20
Recovered Coal
per mined tonne
0.950 tonne
0.900 tonne 0.85 tonne 0.80 tonne
Costs: Mining $2.00 per tonne
Washing 2.00 per tonne
Reclamation 0.05 per tonne
Marketing 6.50 per tonne
Total Cost,
excluding stripping $10.55 per tonne
Given stripping
cost
For market value of
$15 per tonne
Value
Net
Ratio
$30 per tonne -
Value =
Net =
Ratio
$45 per tonne -
Value =
Net =
Ratio =
$60 per tonne
Value
Net =
Ratio =
$0.20 per nf
$14.25
3.70
18.5:1
$28.50
17.95
89.75:1
$42.75
32.20
161:1
$57.00
46.45
232.25:1
$13.50
2.95
11.75:1
$27.00
16.45
82.25:1
$40.50
29.95
149.75:1
$54.00
43.45
217.25:1
$12.75
2.20
11.0:1
$25.50
14.95
74.75:1
$38.25
27.70
138.5:1
$51.00
40.45
202.25:1
$12.00
1.45
7.25:1
$24.00
13.45
67.25:1
$36.00
25.45
127.25
$48.00
37.45
187.25
313
-------
TABLE 60. FACTORS IN STRIPPING RATIO CALCULATIONS
FACTOR
SYMBOL
TYPICAL RANGED
LOW HIGH
Percent
reject ,
Mining cost per
Cleaning
as decimal
tonne mined
cost per tonne
P
C
RE
MI
CrT.
$
$
0
0.
0.
20
00
$
$
60
2.
5.
00
00
(2)
\ *- /
Site preparation and
revegetation cost per
tonne mined
Stripping cost per cubic
meter overburden
Market price of coal, per
clean tonne
Transportation, sales and over-
head per clean tonne
"RE
$ 0.003
$ 0.04
CC
$10.00
0
HD
$0.00
(4)
$ 1.00
$ 1.00
$100.00
(3)
$ 15.00
Notes 1. Although some values may appear high at the present
time, allowances are made for trends in the future.
2. Some mines do encounter long (>10 km) truck haulage,
which is considered as secondary transportation in
this analysis and included in 0 .
rilJ
3. Market price for metallurgical coal could conceivably
inflate to this level in the future.
4. Mine-mouth plants have essentially no 0 costs.
314
-------
The break-even stripping ratio function may be expressed as:
(P _ n 1 v c\
CPCC °HD; X U
•° V
(CMIH
hCCLH
hCHE>
CST
PCC 'W^CC^ (QHD)(PRE)
r r r r r r r
UST ST ST ST ST ST ST
In this second form, it is apparent that the calculation of a break-
even stripping ratio may be performed by determining each term separately
and accumulating the result.
A sample calculation is displayed as follows:
Given: PDT, = 0.10 (i.e., 10% reject)
K-b
C $ 1.00 per tonne coal
C = $ 3.00 per tonne coal
Cij
C p $ 0.35 per tonne coal
IXili
0 $ 8.00 per clean tonne coal
HI)
P = $22.50 per clean tonne coal
C = $ 0.25 per cu m overburden coal
o J.
then,
P
-0
-c
-c
-c
•y ("\ P } C ^99
CC X U> PRE; CST ?ZZ
x (1 - P
HD RE
T C
MI ST
CL T CST
RE * CST
) ' C^ = $
ST
= $
= $
= $
8
1
3
0
.50
.00
.00
.00
.35
x 0
x 0
T 0
V 0
v 0
.9 : 0.25 +81
.9 - 0.25 =
.25
.25
.25
28
4
12
1
.0
.8
.0
.0
.4
_>
m /
™3,
3
m /
3
m /
3
m /
't
't
't
't
't
= Total +34.8 m3/t
It is possible to analyze variations in stripping ratio and sensitivity
with regard to variations in any single factor or combination of factors.
Rather than attempt such an involved analysis, the following figure presents
315
-------
the variation in stripping ratio with respect to the total net value per
tonne of coal, excluding of stripping cost, and stripping cost per cubic
meter of overburden. Figure 121 shows the stripping ratio in cubic meters
of overburden per tonne of coal. It is possible to obtain the stripping
ratio in meters of overburden per meter of coal by multiplying this ratio by
1.34. Of course, not all of these situations are probable in today's con-
ditions. However, even the unlikely costs of $100 for coal and $0.05 for
stripping could be possible if high-grade metallurgical coal continues to
rise in price, while earth-moving technology continues to develop along its
current trends.
STRIPPING RATIO IN MULTIPLE-SEAM MINING
The analysis of stripping ratio may be complicated by considering the
case where several seams are mined from one cut. Multiple-seam planning
requires that both total stripping ratio and the incremental stripping
ratios be analyzed. A necessary condition is that the cost of stripping all
overburden, innerburden, and partings be less than or equal to the total net
value of coal extracted from the seams. However, the total analysis is not,
of its own, sufficient to justify multiple-seam mining. Even if it is
feasible to remove two seams in terms of the overall stripping ratio, the
lower seam will not be mined unless the incremental stripping ratio for the
lower seam is also acceptable.
Consider, for example, a situation where the deposit consists of 10 m
of overburden, 1.0 m of coal, 20 m interburden, and 0.5 m coal. The overall
stripping ratio is calculated as:
10 + 20 m overburden 30 on 1
1.0 + 0.5 m coal = T^ = 20:1»
If the break-even ratio for the company were 25:1, m/m, then the total
ratio would be acceptable, and seemingly both seams should be mined.
However, the incremental stripping ratios are:
Upper seam, — = 10:1 m overburden to m coal
Lower seam, 7—=— = 40:1 m interburden to m coal
0.j m
While the upper seam is quite favorable for mining, the incremental analysis
of the lower seam fails to justify mining both seams.
316
-------
u>
100.00-
97.50-
95.00 -
92.50 -
90.00 -
87.50 -
85.00-
82.50-
80.00-
77.50-
75.00 -
72.50 -
70.00
67.50-
65.00
62.50-
60.00 -
57.50 -
55.00 -
52.50-
50.00-
47.50-^
45.00-
42.50-
40.00 -
37.50 -
35.00-
32.50-
30.00-
27.50 -
25.00^
22.50-
20.00-
17.50-
15.00-
12.50-
10.00-
7.50-
5.00-
2.50-
0.00
Stripping Ratio
250JC225) QOOJ
2000.
1950.
1900.
1850.
180C.
1750.
1700.
1650.
1600.
1550.
1500.
1K50.
1400.
1350.
1300.
1250.
1200.
1150.
1100.
1050
1000
0.00 0.05 0.10 0.15 0.20 0.25
-1 - 1 - 1 - 1 - 1 - T
0.30 0.35 0.40 0 45 0.50 0 55
Stripping Cost, $'m3
0.60
T T
0.65 0.70 0.75
0.80 0.85 0.90 0.95 1.00
Figure 121. Stripping ratio in m /tonne for various coal values and stripping costs.
-------
A complication might exist when considering more than two seams; a
situation where mining the upper two seams is not feasible while mining
three or more seams is feasible. Consider a situation as above, but in
which 10m of interburden lie over a third, lowest seam, 2.5 m thick. The
incremental analysis would become:
10 + 20 + 10 =
1 + 0.5 + 2.5 ~ m/m
Upper seam, 10:1, as above
Middle seam, 40:1, as above
Lowest seam, without considering middle seam
20+0.5+10 30.5 ,
r—^ = • 'r-v 12.2:1, m/m acceptable
Lowest seam, including middle seam as byproduct
20 + 10
3.0
= 10:1
It would then be advisable to mine to the lowest seams. Of course,
since the lowest seam is to be mined, and since the stripping ratio con-
sidering the second seam as "byproduct" is better than mining the lowest
alone, the second seam will also be recovered. Unique cases may exist
when the particular economics of coal handling and cleaning may indicate
that the middle seam should be treated as interburden. With stringent
requirements for the burial of refuse and conservation of natural resources.
the practice may not be justified.
STRIPPING RATIO AND COSTS
The determination of stripping ratio has been shown to be quite
dependent upon and sensitive to a variety of costs. Such a dependence on
costs means that the break-even stripping ratio is related to the scale of
the operation. The scale and type of operation, economic nature of the
company, and reclamation approach will affect the development of the break-
even stripping ratio and the daily or working stripping ratio of the
operation.
318
-------
The working stripping ratio is generally much less than the break-even
ratio. It varies with the depth of cover at the time and provides an overall
ratio less than or equal to the break-even stripping ratio. In general,
lower stripping ratios imply lower production costs. An objective in plan-
ning is often to mine with the contour. Mining with the contour means
starting where the depth of overburden is least or at the outcrop, if the
seam is outcropping. The working stripping ratio then increases with the
progress of mining. Such an approach allows higher unit profits (i.e. low
production costs) in the first years of life, when the time value of money
is highest, and defers high unit production costs to a later point in time.
Additionally, if improvements in technology or a higher price for the product
is attained at a future date, the break-even stripping ratio can be
increased. However, where the topography is severe (steep hill), the break-
even point is reached in a very small width. Therefore, the entire width
is mined to the break-even point and then reclaimed. However, in these
operations, scale and bonding are minimized.
In area mining and mountaintop mining there is usually some coal
recovered under working stripping ratios in excess of the break-even ratio.
To be justified, the incremental gain in total profits attributable to mining
the coal so burdened must still balance to an overall ratio less than the
break-even ratio. This practice is required in the area directly under a
mountaintop or under a small knoll in an area mine where leaving the moun-
taintop or knoll would require that different equipment of smaller scale be
used in the entire operation. The extraction of a small tonnage of coal
which does not, strictly speaking, pay for itself may be justified on the
basis of economy of scale that could not be otherwise realized.
COST ANALYSIS
Determination of mining acceptability is not completely restricted to
stripping ratio. There can also be established a minimal mineable thickness,
a minimal value of coal, and technological limits to exploitation.
Where sufficient tonnage and market are indicated, a more detailed
analysis is performed. In this analysis, specific alternative mining methods
or groups of subsystems will be specified as indicated by the geometry of
the deposit and the stripping ratio. The design problem can be one of
selection of primary earthmoving methods and equipment from several alterna-
tives. For each alternative method proposed, capital and operating expenses
are detailed and more exact stripping limits and recoverable tonnages can
be determined. There are at least three stages of analysis which are
carried out:
1. Mine vs. no-mine analysis, based on stripping ratio,
gross cost, and gross return estimates.
2. Selection and analysis of alternative methods, based
on gross geological aspects.
3. Final design analysis, based on detailed comparison of
specific equipment, production requirements, and legal
and social constraints.
319
-------
Costs incurred by mining companies are broken down in two ways: 1.)
capital costs are large outlays of money for land, equipment, and buildings.
For tax purposes, these costs are not regarded as directly deductible
expenses, but must be either amortized, depreciated or depleted. 2.)
operating costs (or expenses) are for such items as labor and supplies, and
are deductible, in total, in the year of occurrence. Costs are also treated
as direct or indirect, with fixed or variable costs, depending on their
relationship to coal production.
Some capital expenditures are usually expected by the government to be
treated as amortizable items. These include organization costs of the
corporation under certain conditions, acquisitions, or nonmineral leases,
research and development expenses not treated as expenses, amounts other
than purchase connected with acquisition of trademarks, copyrights and the
like, and the cost of certified pollution control facilities placed in
service in the years 1969 through 1974, inclusive. Amortization is calcu-
lated by a straight line method. For example, for an amortizable item of
five-year life, one fifth of the amount is deducted from the gross income
in each year of the project life.
Depreciation is an allowable deduction prior to taxes for the wear and
tear of equipment during their normal use. It is in effect a measure of
the decrease in value of capital equipment items, and book value decreases
as equipment depreciates with time. In any instance, depreciation is
regarded as the yearly decrease in real value of property which is used or
employed in the course of business. Depreciation may be calculated strictly
on the basis of time or may be calculated strictly on the basis of units of
production. By either method, the first step is to determine the depreciable
amount for the item. This is generally the purchase price less the antici-
pated salvage value at the end of a useful life. While mining companies
must establish meaningful and useful lives for buildings, machinery, and
other depreciable items, sometimes guidelines and acceptable lives are
established by the Internal Revenue Service (IRS). The purchase price,
salvage value, and equipment life must be consistent with the IRS guide-
lines and acceptable to. the IRS.
A company exploiting a wasting asset like coal is allowed to deduct a
certain sum from the income it receives from mining coal to reflect the
loss of the capital asset, which is the coal in the ground. This sum is
the depletion allowance. Two alternative methods are available for
calculating this depletion allowance, cost depletion and percentage
depletion. The company can select whichever method gives the greater
deduction and is also allowed to switch from one to another, if desired,
from year to year. Methods of depreciation and depletion are discussed
in several engineering economy texts.
The operating costs can be broken down into direct and indirect, and
fixed and variable costs. Certain items of the mining work can be
directly related to the production of a given number of units of coal: the
amount of work the men in the pit perform with the utilization of a shovel
or dragline. Other costs, such as taxes, sewer, mine plant maintenance,
320
-------
dry-change facilities, etc., do not relate directly to the production of
specific units of coal but accrue indirectly to the operation as a whole.
The former may be regarded as direct costs while the latter are regarded as
indirect costs. In a similar analysis, certain costs are fixed without
regard to the quantity of coal being mined, such as the maintenance of facil-
ities, engineering staff pay, supervisory pay, etc. Other costs, such as
direct labor pay, operation expenses for drills, shovels, trucks, blasting,
etc., are incurred directly by and can be accounted directly to the produc-
tion of given quantities of coal. In most cases, these fixed costs are pro-
rated for accounting purposes to a unit of production basis.
Where appropriate accounting and industrial engineering records have
been kept, the estimation of costs for new methods and new approaches is
greatly facilitated, as the increments of cost involved in the new approach
may well be historically documented at present in previous, traditional
approaches.
COST ESTIMATION
It is not the purpose here to go into details of cost analysis models.
From the discussion thus far, it is obvious that there is no one model or
method that will give correct answers. Even if the models or analysis
procedures are analytically sound, the input data in most cases is proprie-
tary to the company. The U.S. Bureau of Mines has published several models
that address the cost of surface mining of coal.
Boyce (1975) reports on a computerized model to analyze the technical
and financial aspects of large-scale surface mining. The model is intended
to be used for first-order planning of large scale surface coal mines and
for cost evaluation under various conditions. It was developed for the
Energy Research and Development Administration (ERDA) by Flour Utah, Incor-
porated.
Otte and Boehlje (1975) have developed a model for calculating the cost
of mining and reclamation with various materials handling techniques and
procedures in different pit configuration. This computer-oriented model
calculates production and costs for each individual task performed to uncover
and remove coal and to reclaim the mined area. Total costs per ton of coal
removed and the capital outlay required to obtain a specified level of
annual production are also determined by aggregating the cost of each task.
Manual et al (1975) have developed a cost-model for anthracite open-
pit mining which determines a rate of return for a specified realization
(i.e., selling price) or vice-versa through a discounted cash-flow analysis.
The model divides costs into four groups; ownership, labor, supplies, and
miscellaneous costs. The last category includes the charges for welfare,
royalties, etc.
NUS Corporation has developed a cost model for the Electric Power
Research Institute (EPRI, 1977). The model incorporates the essential
features of surface coal mine equipment selection and cost engineering
321
-------
principles to analyze the costs of producing coal from a new surface mine.
It is intended to provide an engineering approach to determination of
production costs to be used in developing coal price-availability relation-
ships .
A computer model has been developed by Ramani, Murray and Manula (1977)
to generate the selling price of coal based on a combination of geologic,
engineering and economic variables, as these are related to area, contour
and open-pit mining. Much of the background information used in model
development was based on the report by the NUS Corporation to Electric Power
Research Institute (EPRI, 1977).
Surface Mining Cost Models
Katell and Hemingway (1972, 1976) have estimated typical capital invest-
ments, operation costs, and selling prices for coal strip mines that will
fuel a 250-million-standard-cubic foot-per-day, high-BTU gas plant in
Eastern, Interior, and Northern Great Plains states of the United States.
These studies are regularly updated; the latest study can be found in the
Bureau of Mines Information Circular 8703 (Katell, Hemingway and Berkshire,
1976).
Toth and Stinnett (1976) present a somewhat simple model of the NUS/EPRI
Model (EPRI, 1977) for a preliminary estimate of surface mining costs. The
model is presented here in its entirety because the principles on which the
model is based can easily be extended to any surface mining operation with
modifications. The accuracy of the model results may be in the range of
+ 30%, typical of preliminary estimates. Since some values used here are
outside the range of the graphs, instead of using the graphs, calculations
are performed to arrive at the parameter values. In general, the approach
presented here parallels that which an engineer would follow in performing
a preliminary new mine cost estimate. The "bottom line" of the analysis is
the minimum, acceptable selling price for the coal (from the seller's view-
point), but certain other important figures are generated as well. The
model centers around a set of nine graphs which allow for gross variances
that are usually encountered among coal deposits, and some input, therefore,
is required from the user.
The first graph (Figure 122) illustrates the relationship between the
designed mine capacity and the volume of overburden which must be removed.
The variable relating the two is the stripping ratio. In actual practice,
a mine's stripping ratio will vary from year to year and likely will approach
some upper limit late in life, beyond which the mine can no longer be
operated profitably. The strip ratio suggested for this analysis is the
average effective strip ratio (including rehandle), although a rough sensi-
tivity analysis can be performed by simply varying the ratio selected and
noting the effect on coal selling price. A 1,696,720 tonne/year (1,670,000
ton/year) mine located in Illinois and characterized by a strip ratio of
twelve cubic yards of overburden per ton of coal is selected for illustrative
purposes. The resulting overburden figure, 15,290,000 cu m (20,000,000
cubic yards), is used as input to the next graph (Figure 122).
322
-------
u
o
E
-------
In Figure 123, the annual overburden to be removed is translated into
total machine capacity needed to affect removal. The basic overburden
removing equipment in area stripping operations is assumed to be either a
shovel or dragline. The variable in Figure 123 is machine "efficiency"
which is dependent upon a number of operating parameters including cycle
time, operating hours, operator efficiency, maintenance reliability and so
forth. Additionally, if the machine is used for road building or encounters
excessive rehandling or chopping down as a result of multiple-seam mining,
the efficiency will be lowered considerably.
In the example, overburden digging conditions are presumed to be good—
blasting will not likely be required—and a single seam will be extracted.
Thus selecting a moderately high efficiency of 366,252.45 cu m/cu m
(280,000 yards/yard) of bucket capacity, 53.51 cu m (70 cubic yards) of
machine capacity will be needed.
Figure 124 illustrates the relatively large component of capital costs
attributable to the overburden stripping equipment, and thus suggests the
rationale for focusing on the strip ratio as a primary factor in ultimately
affecting the selling price of coal. There is reason to believe that machine
costs may not exhibit the straight line function with size as is illustrated;
however, since an operator may have the option of selecting multiple pieces
of smaller capacity to meet production needs, the straight line assumption
should remain with the accuracy of this analysis.
The three other lines on the graph show typical capital outlay for
physical investment (the proven reserves, initial equipment, and plant needs),
the start-up mine investment (including interest during construction, prepro-
duction development, and working capital), and the total investment through-
out the mine operating life (including equipment replacements).
Referring to the example, the start-up and total mine investment may
approach $48,000,000 and $73,000,000, respectively.
Using the total capital investment from the preceding graph, the
straight line depreciation and amortization charges per ton of coal produced
can be determined in Figure 125. All mines are presumed to have <± 30-year
operating life, although a similar chart could be constructed if this
simplification is invalid. The number derived from Figure 125 represents
the noncash portion of operating costs and will be summarized with others
to arrive at the minimum acceptable selling price. Since the data for the
mine is outside the limits of the graph in the example for this 1,696,720
tonne (1,670,000 ton/year) operation, per ton depreciation and amortization
charges can be calculated as:
= $,43/toime (.. . ?1,46/ton)
30 x 1,696,7;
This calculation also illustrates the simplicity of the use of the model,
324
-------
OJ
N)
Ui
10
20
30 40 50 60 70 80
Overburden Removed Annually (Cu Yds x 10 )
Figure 12'j. Overburden removal, versus machine capacity.
-------
0
0 20 40 60 80 100 120 140 160 180 200
Machine Requirements (Bucket Capacity in Cu Yds)
Figure 124. Capital requirements for a surface coal mine.
326
-------
1.60
1.50
1.40
I 1.30
h-
V»
JS 1-20
lj
0>
J 1.10
o
9)
O
CO
1.00
•g 0.90
o
CQ
S 0-80
CD
<-> 0.70
c
o
^-
I 0.60
o
E
-o 0.50
o
c
I °-4°
o
-------
Figure 126 is a display of the range and average productivity of surface
mining operations in the United States, measured in tons/man-day. The range
in production is by county, rather than individual mine, so presumably a
range chart based on the latter category would show even greater variation.
The relatively low productivity of the entire United States of 36.57
tonne/man-day (36 tons/man-day) stems from the large proportion of coal
coming from the Appalachian and Interior Regions. It is reasonable to assume
in the example, however, that at a large new well-engineered mine in Illinois
the labor productivity will be 30.48 tonne/man-day (30 tons/man-day).
Figure 127 relates the labor cost per ton of coal produced to the
average labor rate and the productivity of the mine personnel. Mine
personnel in this analysis includes supervisory staff in addition to all
hourly workers.
Assuming a labor rate of $60/day and a productivity of 30.48 tonne/man-
day (30 tons/man-day), for- the example, labor costs are determined to be
$2.26/tonne ($2.3/ton). It is quite possible to vary both productivity and
labor rate, and arrive at different labor costs.
Assuming a productivity of 30.48 tonne/man-day (30 tons/man-day) and
approximately 250 working days/year, the minimum number of men not including
provisions for absenteeism, vacations and turnover is calculated below.
No. of men Annual Production
required (Productivity)(No. of working days/year)
1,696,720 000 . ,1,670,000
OUS x 250 223 emPloyees or < 30 x 250 223
Figure 128 provides the relationship between labor costs per ton of
coal produced and direct operating costs for the operation. As defined
here, direct operating costs include not only labor charges but also
supplies, power, fringe benefits, welfare payments, insurances, royalties,
property and payroll taxes, and so forth. Because of the catch-all nature
of this grouping, and because of the initial premise that new, highly
productive mines are to be considered for analysis, the proportion of labor
costs to direct operating charges is predicated to be somewhat lower than
historical records usually indicate.
The data for the mine is outside the limits of the graph. Therefore,
with labor charges estimated at 30% of total direct charges, the direct
operating costs are calculated to be $7.54 per tonne ($7.67 per ton) of coal
(i.e., $2.3/0.3). Here again, different percent values can be used for
labor charges as a component of the total cost.
328
-------
O
170
160
150
140
130
120
110
100
M
80
70
60
50
40
30
20
10
RANGE IN
PRODUCTION
BY COUNTY
• AVERAGE PRODUCTIVITY
FOR THE STATE
f!
•RANGE DATA UNAVAILABLE
••ONLY A SINGLE OPERATION IN THE STATE
< *' I i 1 i -f $
_llilg
APPALACHIAN
MIDWEST
ROCKY MOUNTAIN
Figure 126. Productivities for surface coal mine operations
in the United States, 1973.
329
-------
0 10 20 30 40 50 60 70 80 90 100 110 120
Average Labor Rates (S Day)
Figure 127. Graph for the calculation of labor costs per ton.
330
-------
6.00
5.50 -
2.00
Labor Costs (S Ton)
3.00
4.00
Figure 128. Relationship of labor costs and direct operating costs.
331
-------
The method of mine financing has an important bearing on total mining
costs as does the schedule of depreciation. Figure 129 shows a range of
typical debt-equity ratios commonly employed by the utility and mining
industries. The lines below the 50/50 line are simply the inverse of those
above and are added essentially to allow full graphical application in
determining mining costs.
The start-up nine investment number derived in Figure 124 provides the
basis for financing as it is assumed that equipment replacements and the
like will be purchased from internally generated funds. This approach
results in a steadily decreasing debt/equity ratio throughout the property
life ^nd would tend to increase ultimate mining costs as the risk-taking
investor will require a somewhat greater return on his exposed funds. A
mitigating circumstance, though, is the fact that the remaining book value
of the mine plant will approach zero over the operating life and so the
equity base, on which the return is calculated, will be shrinking. The
assumption in this analysis, then, is that the per ton financial charges
during the life of the property will approximate those established by the
capital structure at the time of mine start up.
The example shows a 75/25 debt/equity ratio which may be occasioned
through utility ownership and development of the reserves. The debt and
equity components of the initial mine investment are $36,000,000 and
$12,000,000, respectively.
From Figure 130, one can determine the financing charges per ton of
coal depending on a choice of interest rates. Dividing the resulting
numbers from Figure 130 by the annual productive capacity will allow direct
determination of these charges. The calculations are illustrated here
because the values for the example mine fall outside the limits of the
graph.
Capital structure/annual ton is calculated as shown below.
Total capital investment $73 million
Start up investment S48 million
Debt component of start up investment = $36 million
Equity component of start up investment = $12 million
Capital structure (Debt component) _ $36 million _ g?1 ^
per annual ton $1.67 m tons
Capital structure (Equity component) _ $12 million = 5 7 19
per annual ton $1.67 m tons
Financing charges debt at 102 = $21.56 x 0.10= $ 2.16
Return on equity at 18* = $ 7.19 x 0.18= $ 1.29
Income taxes = 0.33 x $1.29 $ 0.43
332
-------
70
60
50
•o
CD
0>
13
40
2 30
o
U
20
10
0 10 20 30 40 50 60 70 80 90 100
Start-up Mine Investment (S x 106)
Figure 129. Capital structure for start-up nine investment.
333
-------
3.00
U)
u>
0123
4 56 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26
Capital Structure Annual Production ($ x 106/Tons x 106)
Figure 130. Graph for the calculation of financing charges and rate of return.
-------
Thus assuming in the example a debt servicing rate of 10% and a return
to equity of 18%, the burden will come to $2.12/tonne and $1.27/tonne
($2.16/ton $1.29/ton), respectively. One additional value can be calculated
from Figure 130, the approximate federal and state income taxes. If the
combined tax rate can be taken as 50%, and if one assumes a depletion
allowance will be limited by the 50% of net income ruling, income taxes
will total 33% of the return to equity figure. As illustrated in the
example, these are $0.42/tonne ($0.43/ton).
Table 61 is the culmination of the graphical exercise wherein the
previously determined charges from Figures 125, 128 and 130 are totalled to
arrive at the minimum acceptable selling price per ton necessary to allow
the mine operator an appropriate return on investment.
Referring once more to the example, the selling price for the particular
operating and deposit characteristics should approximate $12.80/tonne
($13.01/ton).
SUMMARY
There are, of course, other approaches to costing an operation. For
example, one major mining company has identified optimum mine sizes and
mining methods on the basis of geographical and market conditions. It
has economically optimized these selections through analysis of past
practice. When the marketing department notifies engineering that it has
received a contract, the engineering department selects an appropriate
property which is held by the company in the geographical vicinity of the
contract or at least as close to the contract delivery point as is possible
to meet the type of product required. For the contract size specified, e.g.,
a five-million ton per year contract, the engineering department has, on
file, the plans for mining development at the particular geographical
location. Such an approach may be desired by large companies which can
exercise economies of scale by utilizing the same, or closely related,
approaches and equipment throughout their operations and are willing to
sacrifice some potentially better approaches. For the average coal mining
company, the sizes and types of deposits will vary over a large scale.
Therefore, the design plans and economic analyses to determine the optimum,
ultimate mine design will proceed in the more traditional fashion. As
already indicated and referenced, computer-oriented cost analysis models
are available for costing total surface mine operations or even individual
unit operations.
335
-------
TABLE 61. CALCULATION OF SELLING PRICE
Depreciation cost/ton
Direct operating cost/ton
Debt financing cost/ton
Return on investment
Taxes @33% ROI
(from Figure 125 or calculation) $ 1.46
(from Figure 128 or calculation) $ 7.67
(from Figure 130 or calculation) $ 2.16
(from Figure 130 or calculation) $ 1.29
(from calculation) $ 0.43
Minimum acceptable selling price
Start-up mine investment
Total mine investment
Mine capacity
Productivity
Personnel requirements:
labor and supervisory
(not including pro-
visions for absenteeism,
vacations, and turnover)
(from Figure 124)
(from Figure 124)
(assumption)
(assumption)
(calculation)
$13.01/ton
$48,000,000
$73,000,000
1,670,000 tons
30 tons/man-day
223 employees
336
-------
SECTION 11
REFERENCES
Adams, L. M. , Capp, J. P., Eisentrout, E., Reclamation of Acidic Coal Mine
Spoil with Fly Ash, U.S. Department of the Interior, Bureau of Mines
Report of Investigation Number 7504, April 1971.
Andrews, A. B., "Air Blast and Ground Vibrations in Open-Pit Mining,"
Mining Congress Journal, Vol. 61, No. 5, 1975.
Ash, R. L., "The Mechanics of Rock Breakage," Parts.I through IV, Pit and
Quarry, Vol. 56,Nos. 2, 3, 4, and 5, 1963.
Ash, R. L., "Cratering and Its Application in Blasting," Unpublished report,
Department of Mining and Civil Engineering, University of Minnesota,
1970.
Atkinson, T. , "Selection of Open-pit Excavating and Loading Equipment,"
Institution of Mining and Metallurgy, Vol. 80, Bulletin No. 776,
Sec. A, 1971.
Bailey, P. A., "Exploration Methods and Equipment," Chapter 2.1, Surface
Mining, Ed. E. P. Pfleider, SME/AIME, New York, 1968.
Bergmann, 0. R., Wu, F. C., and Edl, J. W., "Model Rock Blasting Measures
Effect of Delays and Hole Patterns on Rock Fragmentation," Engineering
and Mining Journal, Vol. 175, No. 6, 1974.
Boddorff, D., "That's the Way Rock Breaks," The Explosives Engineer, Spring,
1974.
Boyce, T. A., "Economic System Analysis of Large Scale Surface Mining,"
Proceedings of the 3rd Symposium on Surface Mining and Reclamation,
Vol. II. NCA/BCR, Washington, D.C., 1975.
Burton, A. K., "Off-Highway Trucks in the Mining Industry," Mining
Engineering, Vol. 27, No. 8, 1975; Vol. 27, No. 10, 1975; Vol. 27,
No. 11, 1975; Vol. 27, No. 12, 1975; Vol. 28, No. 1, 1976.
Caterpillar Tractor Co., Fundamentals of Earthmoving, Peoria, Illinois,
July 1975.
Caterpillar Tractor, Co., Performance Handbook, 6th Edition, Peoria, Illinois,
1976.
337
-------
Cheek, E. E., "Application of a Standard Wheel Excavator in Coal," Mining
Congress Journal, Vol. 52, No. 11, November 1966, pp. 33-35, 38-39.
Civil Engineering Handbook, 4th Edition, McGraw Hill Co., New York, 1962.
Condon, J. L., and Snodgrass, J. J., "Effects of Primer Type and Borehole
Diameter on AN-FO Detonation Velocities," Mining Congress Journal,
Vol. 60, No. 6, 1974.
Cummins, A. B., Chairman of the Editorial Board, SHE Mining Engineering
Handbook. AIME, New York, 1973.
Daley, D. D., "Emphasizing Environment Natural Beauty Considered," Mining
Engineering, Vol. 26, No. 2, 1974.
Dick, R. A., "Explosives and Borehole Loading," SME Mining Engineers
Handbook, SME/AIME, New York, 1973.
Dombrowsky, N. G., "The New Bucket Wheel Excavator and Its Construction
Problems," Bergbautechnik, Vol. 14, 1964, p. 277.
Doyle, F. J., "Variables Affecting Grading Costs for Strip Mine and Refuse
Bank Reclamation," Proceedings of the First Symposium on Mine and
Preparation Plant Refuse Dispoal, National Coal Association, 1974, pp.
212-227.
Drevdahl, E. R. , Profitable Use of Excavation Equipment, Technical Publica-
tions, Desert Laboratories, Inc., Tuscon, Arizona, 1961.
DuPont Co., Blaster's Handbook, Wilmington, Deleware, 1967.
Ellison, R. D., and Thurman, A. G., "Geotechnology: An Integral Part of Mine
Planning," A Chapter in Coal Exploration, Ed. W. G. Muir, Miller Freeman
Publications, Inc., San Francisco, 1976.
EPRI, Coal Mining Cost Models - Surface Mines, Electric Power Research
Institute, EPRI EA-437, Research Project 435-1, 1977.
Felix, G. L., "Development of the 350 Ton Haulage Truck," Mining Congress
Journal, Vol. 62, No. 3, 1976.
Gartner, K., "Bucket Wheel Excavator on Building Sites," Forden und Heben,
Vol. 3, No. 6, November 1965.
Gordon, R. L. , Historical Trends in Coal Utilization and Supply, U.S. Bureau
of Mines Report, Grant No. G0155116, Washington, D.C., 1976.
Grim, E. C., and Hill, R. D., Environmental Protection in Surface Mining of
Coal, Report No. EPA-670/2-74-093, U.S. Environmental Protection Agency,
Cincinnati, Ohio, 1974.
338
-------
Haley, W. A., "Changing Methods and Equipment Use in Appalachian Coal Mining,"
Second Research and Applied Technology Symposium on Mined Land
Reclamation, NCA/BCR, 1974.
Highway Engineering Handbook, 1st Edition, McGraw Hill Co., New York, 1960.
Hill, R. D., and Grim, E. C., "Environmental Factors in Surface Mine
Recovery," Symposium on Restoration and Recovery of Damaged Ecosystems,
VPI & SU, Blacksburg, Virginia, March 1975.
Himmel, W., "The Specific Digging Resistance in Dependency on the Shaving Area
and Shaving Form in Different Soils," Freiberger Forschungshefte, A 265,
1961, pp. 1-40.
International Harvester, "An Analysis of Equipment Selection for Heavy
Construction," Payline Division, 1975.
Katell, S., and Hemingway, E. L., "Cost Analysis of Model Mines for Strip
Mining of Coal in the United States," U.S. Bureau of Mines, Information
Circular 8535, 1972.
Katell, S., Hemingway, E. L., and Berkshire, L. H., "Basic Estimated Capital
and Operating Costs for Coal Strip Mines," Revision of 1C 8661, U.S.
Bureau of Mines, Information Circular 8703, 1976.
Kimball, L. Robert (Consulting Engineers), Slope Stability, Volume 1 -
Report and Field Book, Report ARC-71-66-T3 prepared for the Appalachian
Regional Commission, p. 81, undated.
Krause, R. R., "Predicting Mined Land Soil," Chapter in Ecology and
Reclamation of Devastated Land, Vol. 1, Ed., R. J. Hutnik and Grant
Davis, Gordon Breach, New York, 1973.
Learmont, T., "Productivity in Large Stripping Machines," SME/AIME
Transactions, Vol. 258, No. 3, 1975.
Maneval, D. R., "Coal Mining vs. Environment: A Reconciliation in
Pennsylvania," Appalachia, A Journal of the Appalachian Regional
Commission, Vol. 5, No. 2, February-March 1972.
Mani, M. S., "Design Features of Bucket Wheel Excavators," Journal of Mines,
Metals and Fuels, Vol. 14, No. 3, March 1966, pp. 71-79.
Manula, C. B., Albert, E. K., and Burgos, V., A Report on Anthracite Open-pit
Mining, Part V. Engineering and Mine Cost Analysis, Special Research
Report SR-104, Coal Research Section, The Pennsylvania State University,
1976.
Mathematica, Inc., The Design of Surface Mining Systems in Eastern Kentucky,
Department of Natural Resources and Environmental Protection, Division
of Reclamation, April 30, 1973.
339
-------
McGraw Hill Inc., Coal Age, Vol. 75, No. 7, 1970, pp. 173-179.
MESA, Engineering and Design Manual: Coal Refuse Disposal Facilities,
Prepared by E. D'appolonia Consulting Engineer for U.S. Department of
the Interior, Mining Enforcement and Safety Administration.
Meyer, C. F. , Route Surveying, International Textbook Co., Scranton, PA, 1969.
Morgenstern, N. R., and Price, V. E., "The Analysis of the Stability of
General Slip Surfaces, " Geotechnique, Vol. 15, No. 1, pp. 79-93, 1965.
Morrell, R. J., and Unger, H. F., "Drilling Machines, Surface," SME Mining
Engineering Handbook. SME/AIME, New York, 1973.
Murray, C. W., "A Computer Model for Optimizing Surface Mine Haul Road
Profile," M. Eng. Report, Mineral Engineering Management, The
Pennsylvania State University, 1978.
Nicholls, H. R. , Johnson, C. J., and Duvall, W. I., Blasting Vibrations and
Their Effects on Structures, Bulletin 856, U.S. Department of the
Interior, Bureau of Mines, 1971.
Nordness, H. J., "Horizontal Drills for Surface Mines," Proceedings 3rd
Conference on Mine Productivity, The Pennsylvania State University, 1976.
Obert, L. A., and Duvall, W. I., Rock Mechanics and the Design of Structures
in Rock, John Wiley, New York, 1967.
Otte, J. A., and Boehlje, M., "A Model to Analyze the Cost of Strip Mining
and Reclamation," Proceedings of the 3rd Symposium on Surface Mining and
Reclamation, Vol. II, NCA/BCR, Washington, D.C., 1975.
Peurifoy, R. L., Construction Planning, Equipment and Methods, McGraw Hill,
New York, 1970.
Pfleider, E. P., Ed., Surface Mining, American Institute of Mining,
Metallurgical and Petroleum Engineers, Inc., New York, 1968.
Porter, W. E. , "Multiple Seam Strip Mining: A Survey of Economic Feasibility
Model," Unpublished M. Eng. Report, The Pennsylvania State University,
1971.
Price, G. C., Manula, C. B., and Ramani, R. V., Material Handling Research,
The Bucket Wheel Excavator, U.S. Bureau of Mines, Information Circular
8580, 1973.
Pugliese, J. M. , Designing Blast Patterns Using Empirical Formulas, U.S.
Bureau of Mines, Information Circular 8550, Washington, D.C., 1972.
Ramani, R. V., "Computer Simulation of a Bucket Wheel Excavator," Unpublished
M.S. Thesis, The Pennsylvania State University, 1968.
340
-------
Ramani, R. V., and Grim, E. C., "Surface Mining-A Review of Practices and
Progress in Land Disturbance Control," Chapter 14 in the book
Reclamation of Drastically Disturbed Lands, Ed., F. W. Schaller and
P. Sutton, American Society of Agronomy, Madison, Wisconsin, 1978.
Ramani, R. V., Murray, C. W., and Manula, C. B., Application of a Total
System Surface Mine Simulator to Coal Stripping, Volume IV, Surface
Mine Cost Model, National Technical Information Service (NTIS),
Publication No. 294634/AS, 1979.
Rasper, L., and Ritter, H., "Opening Up of the Lignite Mine of the Neyveli
Lignite Corporation and Experience with Bucket Wheel Excavators in Hard
Overburden," Braunkohle, Warme und Energie, Vol. 13, No. 10, October
1961, pp. 390-400.
Rasper, L., "Present Stage of the Development of Bucket Wheel Excavators in
the German Federal Republic," Braunkohle, Warme und Energie, Vol. 16,
No. 11, November 1964, pp. 457-471.
Rumfelt, H., "Computer Methods for Estimating Proper Machine Mass for
Stripping Overburden," Mining Engineering, Vol. 13, No. 5, 1961.
Rumfelt, H., "Application and Performance of Wheel Excavators," Mining
Congress Journal, Vol. 47, No. 6, June 1961, pp. 46-49.
Saperstein, L. W., and Secor, E. S., "Improved Reclamation Potential with the
Block Method of Contour Stripping," Proceedings of the Research and
Applied Technology Symposium on Mined-Land Reclamation, Pittsburgh,
Pennsylvania, Bituminous Coal Research, Inc., NCA/BCR, March 1973.
Scott, R. F., Principles of Soil Mechanics, Addison-Wesley, Reading,
Massachusetts, 1963.
Secor, E. S., "An Economic Comparison Between the Block and Conventional
Methods of Contour Strip Mining," Unpublished M. Eng. report in
Mineral Engineering Management, The Pennsylvania State University,
85 pp., 1977.
Skelly and Loy Engineering Consultants, Economic Engineering Analysis of U.S.
Surface Coal Mines and Effective Land Reclamation, U.S. Bureau of Mines
Report, Contract No. S0241049, Washington, B.C., 1975.
Starr, C., "Energy and Power," Scientific American, Vol. 225, No. 3,
September 1971, pp. 37-49.
Stefanko, R., Ramani, R. V., and Ferko, M. R., An Analysis of Strip Mining
Methods and Equipment Selection, Research and Development Report No. 61,
Interim Report No. 7, Prepared by Coal Research Section, The Pennsylvania
State University, for Office of Coal Research, U.S. Department of the
Interior, Washington, D.C., May 1973.
341
-------
Terzaghi, K. , and R. B. Peck, Soil Mechanics in Engineering Practice, 2nd
Edition, John Wiley, New York, 1967.
Thomas, L. J., Introduction to Mining, Hicks and Sons, Sydney, Australia,
1973.
Toth, G. W., and Stinnett, L. A., "Techniques of Rapid Preliminary Estimates
of Coal Mining Costs," paper presented at the 1976 Mining Convention,
American Mining Congress, Denver, Colorado, September 26-29, 1976.
U.S. Department of the Interior, Surface Mining and Our Environment, A Special
Report to the Nation, 1967.
Wang, F. D., et al., Computer Program for Pit Slope Stability Analysis by
the Finite Element Stress Analysis and Limiting Equilibrium Methods,
U.S. Bureau of Mines, Report of Investigations 7685, p. 53, 1972.
Wendertz, R. L., "Coal Loading Shovels," (a review of factors to consider in
loading equipment selection), Mining Congress Journal, Vol. 56, No. 4,
1970.
342
-------
SECTION 12
BIBLIOGRAPHY
Adams, L. M., Capp, J. P., Gillmore, D. W., "Coal Mine Spoil and Refuse Bank
Reclamation with Powerplant Fly Ash," Proceedings of the Third Mineral
Waste Utilization Symposium, Chicago, Illinois, March 1972.
Alder, L., and Nauman, H. E., Analyzing Excavation and Material Handling
Equipment, VPI, Blacksburg, Virginia, February 1970.
Aiken, G. E., "Bucket Wheel Excavators: How to Choose the Right One for
the Job," Mining Engineering, Vol. 18, No. 1, January 1966, pp. 71-79.
AIME, "Drilling Selection Requires Value Judgement," Mining Engineering,
Vol. 19, No. 10, 1967.
AIME, "Excavation and Loading A Job for the Giants," Mining Engineering,
Vol. 19, No. 10, 1967.
Allen, N., "Experimental Multiple Seam Mining and Reclamation on Steep
Mountain Slopes," Proceedings of the Research and Applied Technology
Symposium on Mined-Land Reclamation, Pittsburgh, Pennsylvania, NCA/BCR,
March 1973.
American Iron and Steel Institute, Handbook of Steel Drainage and Highway
Construction Products, 2nd Edition, New York, 1971.
Archibald, J. C., "Application of Horizontal Drilling as a Replacement of
Vertical Drilling Techniques," Joy Manufacturing Company, Pittsburgh,
Pennsylvania, 1976.
Ash, R. L., "The Design of Blasting Rounds," Surface Mining, Ed. E. P.
Pfleider, SME/AIME, New York, 1968.
Atkinson, T., and Pryor, R. , "Selection of Open-Pit Excavating and Loading
Equipment discussion on," Institution of Mining and Metallurgy, Vol. 80,
Bulletin No. 777, Sec. A, 1971.
Battelle Columbus Laboratories, Energy from Coal; Guidelines for the Prep-
aration of Environmental Impact Statement, NTIS Publication PB-242-960,
April 1975.
Bellum, D. P., "Open-Pit Mining," Mining Engineering, Vol. 28, No. 2, 1976.
343
-------
Bengston, G. W. , Mays, D. A., and Allen, J. C., "Revegetation of Coal Spoil
in Northeastern Alabama: Effects of Timing of Seeding and Fertilization
on Establishment of Pine-Grass Mixtures," Proceedings of the Research and
Applied Technology Symposium on Mined-Land Reclamation, Pittsburgh,
Pennsylvania, NCA/BCR, March 1973.
Burgos, V. T., "A Study on the Feasibility of Open-Pit Anthracite Mining,"
Unpublished M. Eng. Report, The Pennsylvania State University, 1975.
Caruccio, F. T., "An Evaluation of Factors Affecting Acid Mine Drainage
Production and the Ground Water Interactions in Selected Areas of Western
Pennsylvania," Second Symposium on Coal Mine Drainage Research,
Pittsburgh, Pennsylvania, NCA/BCR, May 1968.
Day, D. A., Construction Equipment Guide, A. Wiley & Sons, New York, 1974.
Davis, G., Strip-Mine Reclamation in Appalachia (Review Draft), U.S. Depart-
ment of Agriculture, Forest Service, Northeastern Forest Experiment
Station, July 1971.
Dawson, V. E. , "Observations Concerning On Site Brake Testing of Large Mining
Trucks in British Columbia," SAE Paper 75060, Society of Automotive
Engineers, Warrendale, Pennsylvania, 1975.
Dotson, G. K. , Constraints to Spreading Sewage Sludge on Croplands^ U.S.
Environmental Protection Agency, Cincinnati, Ohio, 1973.
Duggan, C., Scanlon, D. H., and Bean, S. D., "Evaluation of Municipal Compost
for Strip Mine Reclamation," Compost Science; Journal of Waste Recycling,
Vol. 14, No. 3, May-June 1973.
Environmental Protection Agency, Guidelines for Erosion and Sediment Control
Planning and Implementation, Environmental Protection Technology Series,
EPA-R2-72-015 August 1972, [Project 15030 FMZ. Prepared by Department of
Water Resources, State of Maryland, and Hittman Associates, Inc.,
Columbia, Maryland'].
Enyedy, G. , Jr., "AACE's Cost of Major Equipment (COME) Project in the
Mineral Industry," AIME Council of Economic Proceedings, 1971.
Fites, D. V., "Tractor Scrapers Break New Ground," Mining Engineering, Vol. 21,
No. 5, May 1969, pp. 69-70.
Ford, W. W., "Slope Reduction Methods," Unpublished Report, Division of
Reclamation, Department of Natural Resources, Frankfort, Kentucky, 1968.
Ford, W. W., "Slope Reduction Method (Parallel Fill)," Unpublished Report,
Division of Reclamation, Department of Natural Resources, Frankfort,
Kentucky, 1970.
344
-------
Gartner, E., "Trends of Development in Mining and Haulage Equipment of the
Rhenish Brown Coal Open Cut Mines," Braunkohle Warme und Energie,
Vol. 7, Nos. 11 and 12, June 1955, pp. 226-241.
Grant, A. F., and Lang, A. L., Reclaiming Illinois Strip Coal Land with
Legumes and Grasses, Bulletin 628, University of Illinois, Agriculture
Experiment Station, Urbana, Illinois, 1958.
Griffith, F. E., Magnuson, M. S., and Kimball, R. F., Demonstration and
Evaluation of Five Methods of Secondary Backfilling of Strip-Mine Areas,
U.S. Department of the Interior, Bureau of Mines, Report of
Investigations No. 6772, 1966.
Grube, W. E., Jr., Smith, R. M., Singh, R. and Sobek, A. A., "Characterization
of Coal Overburden Materials and Proceedings," Research and Applied
Technology Symposium on Mined-Land Reclamation, Pittsburgh, Pennsylvania,
NCA/BCR, March 1973.
Gwinn, T. A., "An Improved Environment Through Intelligent Mined-Land
Reclamation," Paper presented at the fall meeting of the Institute of
Mining, Metallurgical, and Petroleum Engineers, Inc., Salt Lake City,
Utah, September 17 to 19, 1969, pp. 28-30.
Hartman, H. L., Principles of Drilling, Chapter 6.1, Surface Mining,
Ed. E. P. Pfleider, SME/AIME, New York, 1968.
Heine, W. N., Gukert, W. E., "A New Method of Surface Coal Mining in Steep
Terrain," Proceedings of the Research and Applied Technology Symposium
on Mined-Land Reclamation, Pittsburgh, Pennsylvania, NCA/BCR, March 1973.
Hewes, L. I., and Oblesby, C. H., Highway Engineering, John Wiley and Sons,
New York, 1960.
Hinesly, T. D., Braids, 0. C., and Molina, J. E., Agricultural Benefits and
Environmental Changes Resulting from the Use of Digested Sewage Sludge
on Field Crops, U.S. Environmental Protection Agency (SW-30d),
Cincinnati, Ohio, 1971.
Hinesly, T. D., Jones, R. L., and Sovewitz, B., "Use of Waste Treatment Plant
Solids for Mined Land Reclamation," Mining Congress Journal, September
1972.
Hollingsworth, J. A., "History of Development of Strip Mining Machines,"
Bucyrus-Erie Company Publication, pp. 1-16.
Holzman, R. S., Executive Tax Management, Thomas V. Cromwell Co., New York,
1974.
Howarth, G. R., "How Will Major New Mines be Financed in the Future," SME
Preprint No. 77-K-125, Paper presented at the AIME-Annual Meeting,
Atlanta, 1977.
345
-------
Hubbert, M. K., "The Energy Resources of the Earth,-" Scientific American,
Vol. 225, No. 3, September 1971, pp. 61-70.
Jarpa, S. G. , "Capital Investment and Operating Cost Estimation in Open-Pit
Mining," Application of Computer Methods in the Mineral Industry, Ed.
R. V. Ramani, Society of Mining Engineers of AIME, Salt Lake City, 1977.
Jelen, F. C., Editor, Cost and Optimization Engineering, McGraw Hill Inc.,
New York, 1970.
Kentucky Department for Natural Resources and Environmental Protection, A.
Manual of Kentucky Reclamation, Frankfort, Kentucky, May 1973.
Killebrew, C. E., "Tractor Shovels, Tractor Dozers, Tractor Scrapers," Surface
Mining, Pfleider, Eugene P. (Ed.), The American Institute of Mining,
Metallurgical, and Petroleum Engineers, Inc., New York, 1968, pp. 463-
477.
Krumrey, A., "Krupp Bucket Wheel Excavators with Stacker Boom," Krupp
Technical Review, Vol. 23, No. 1, 1965.
Levene, H. D.,"An Unusual Coal Mine/Power Plant Complex" Coal Mining and
Processing, October, 1971.
Limstrom, G. A. and Dextschman, G. H., "Reclaiming Illinois Strip Coal Lands
by Forest Planting," University of Illinois Bulletin, 547, 1951, p. 50.
Linden, G. , "The Bucket Wheel Excavator and the Versatility in Practice,"
Krupp Technical Publication. Also published in Deutsche Hebeword
Fordertechnik, Nos. 3 and 4, July and October 1957.
Loftis, J. B., "The Evaluation of Horizontal Drilling In Open-Pit Mining,"
SME/AIME Preprint No. 75-AO-319, 1975.
Loper, F. G. , "Simulation Techniques Pay Their Way as Management Tools,"
Mining Engineering, Vol. 19, No. 1, 1967.
Lyon, T. L. , Buckman, H. 0., and Brady, N. C., The Nature and Properties
of Soil, 5th Ed., MacMillian Company, New York, 1952.
Mackellar, A. J., "Cost Engineering Who Needs It?" SME Preprint 75-B-23,
Paper presented at the AIME Annual Meeting, New York, 1975.
Manula, C. B. , Albert, E. K. , and Burgos, V., A Report on Anthracite Open-Pit
Mining Part II. Engineering and Mine Cost Analysis, Special Research
Report SR-104, Coal Research Section, The Pennsylvania State University,
1976.
Manula, C. B. , Albert, E. K. , and Ramani, R. V., "Application of a Total
Systems Simulator to Coal Stripping, Volume II: OPMHS User's Manual,"
U.S. Bureau of Mines, 1977.
346
-------
Manula, C. B., "OPMHS - Open-Pit Materials Handling Simulator," Xlth APCOM
Symposium, University of Arizona, 1973.
Maraboli, L. 0., "Computer Simulation and Cost Analysis of Materials Handling
in Chuquicamata, Chile," Unpublished M. Eng. Report, The Pennsylvania
State University, 1973.
Marcus, J. J., "The Buckingham Method - An Aid in Equipment Selection,"
Mining Engineering, Vol. 17, No. 9, September 1965, pp. 92-94.
McClay, J., "A Dynamic Costing Model for Mining Systems," Unpublished M.S.
Thesis, The Pennsylvania State University, 1974.
McFadden, D. H., "Pit Planning and Design-Coal Mines," Surface Mining,
Pfleider, Eugene P. (Ed.), The Institute of Mining, Metallurgical, and
Petroleum Engineers, Inc., New York, 1968, pp. 211-216.
McGraw Hill Inc., Coal Age, Vol. 75, No. 7, 1970, pp. 173-179.
Morgan, W., Fleet Production and Cost Program, Caterpillar Tractor Co.,
Peoria, Illinois, 1975.
Nalle, P. B., and Week, L. W., "The Digital Computer Application in Mining
and Process Control," Mining Engineering, Vol. 28, No. 9, 1976.
National Coal Association, Mined-Land Conservation Conference, Vol. 7, No. 4,
Washington, D.C., November-December, 1971.
Nutter, F. B., "Mining with Front-End Loaders," Mining Congress Journal,
Vol. 56, No. 6, June 1970, pp. 60-64.
Olivares, L. V., "Theory, Research and Practical Results of Inclined
Drilling," An Unpublished Report in Mineral Engineering Management, The
Pennsylvania State University, December 1967.
O'Neil, T. J., "Computer Simulation of Materials Handling in Open-Pit Mines,"
Unpublished M.S. Thesis, The Pennsylvania State University, 1966.
O'Neil, T. J., "The Minerals Depletion Allowance: Its Importance in
Nonferous Metal Mining," Mining Engineering, Vol. 26, No. 10, 1974.
Peabody Coal Company, "A Model Mining Company," Coal Age, Vol. 76, No. 10,
1971.
Perez, R., "Application of a Time-Oriented Total Systems Simulator for Open-
Pit Mining," Unpublished M. Eng. Report, The Pennsylvania State
University, 1973.
Peterson, J. R., and Gschwind, J., "Amelioration of Coal Mine Spoils with
Digested Sewage Sludge," Proceedings of the Research and Applied Tech-
nology Symposium on Mined-Land Reclamation, Pittsburgh, Pennsylvania,
NCA/BCR, March 1973.
347
-------
Plass, W. T., and Capp, J. P., "Physical and Chemical Characteristics of
Sulfur Mine Spoil Treated with Fly Ash," Journal of Soil and Water
Conservation, Vol. 29, No. 3, 1974.
Ramani, R. V., "Environmental Concerns and Decisions for the Mining Industry,"
Journal of the Institution of Engineers (India), MM Vol. 54, No. 2,
March 1974.
Ramani, R. V., "Surface Mining," Mining Engineering, Vol. 26, No. 2, 1974.
Ramani, R. V., Iroegbu, M. , and Manula, C. B., "Application of a Total Systems
Simulator to Coal Stripping, Volume III: Equipment Selection Models,"
U.S. Bureau of Mines, 1977.
Riley, C. V., "Design Criteria of Mined Land Reclamation," Paper presented at
the SME Fall Meeting and Exhibit, Birmingham, Alabama, October 1972.
Riley, C. V., "Furrow Grading—Key to Successful Reclamation," Proceedings
of the Research and Applied Technology Symposium on Mined-Land
Reclamation, Pittsburgh, Pennsylvania, NCA/BCR, March 1973.
Rodgers, H. C. C., "Developments in Equipment Used for Overburden Removal and
Coal Winning in the Brown Coal Industry. Excavating Equipment for Brown
Coal, Part I," Proceedings of the Australian Institute of Mining and
Metallurgy, No. 194, June 1960, pp. 127-249.
Rose, B. A., "Sanitary District Puts Sludge to Work in Land Reclamation,"
Water and Sewage Works, A Scranton Gillette Publication, September, 1968.
Sathit, T. , "Principles of Drilling," Mining Engineering Handbook, Cumins and
Given, SME, Mudd Series, New York, 1973.
Sanford, R. L. , "Stochastic Simulation of a Belt Conveyor System," Unpublished
M.S. Thesis, The Pennsylvania State University, 1965.
Schellhorn, H. W., "Some Aspects of High Capacity Production with Bucket Wheel
Excavators," Krupp Technical Publication. Also published in The K. A.
Clark Volume, Research Council of Alberta, Edmonton, October 1963.
Schmidt, R. A. , and Stoneman, W. C., A Study of Surface Coal Mining in West
Virginia, Stanford Research Institute, Menlo Park, California, 1972.
Scott, R. B. , Hill, R. D., and Wilmoth, R. C., Cost of Reclamation and Mine
Drainage Abatement—Elkins Demonstration Project, U.S. Environmental
Protection Agency, Cincinnati, Ohio, 1970.
Skelly and Loy Engineering Consultants, Surface Mine Haul Road Design Study,
U.S. Bureau of Mines Report, Contract No. J0255037, Washington, D.C.
1976.
Society of Mining Engineers of AIME, Mining Engineering Handbook, New York,
1973.
348
-------
Staff, Surface Mining of Coal, Coal Age Guidebook, July 1970.
Stermole, F. J. , Economic Evaluation and Investment Decision Methods,
Investment Eval. Corp., Golden, Colorado, 1974.
Taylor, D. W., Soil Mechanics, John Wiley and Sons, New York, 1960.
Taylor, G. A., Managerial and Engineering Economy, V. Nostrand Reinhold Co.,
New York, 1964.
U.S. Department of Agriculture, Kentucky Guide for Classification, Use, and
Vegetative Treatment of Surface Mine Spoil, Soil Conservation Service,
Lexington, Kentucky, 1971.
Weimer, W. H., "Production and Engineering in Surface Coal Mines," Surface
Mining, Pfleider, E. P. (Ed.), The Institute of Mining, Metallurgical,
and Petroleum Engineers, Inc., New York, 1968, pp. 229-238.
Wimpfen, F. P., "Mine Costs and Control," Section 31, SME Mining Engineer's
Handbook, AIME, New York, 1973.
Woodruff, S. D., Methods of Working Coal and Metal Mines, Planning and
Operations, Vol. 3, Pergamon Press, Inc., New York, 1966.
Wurtz, C. B., "Mining and Conservation," Mining Congress Journal, Vol. 56,
No. 5, May 1970, pp. 40.
Zemp, J. C., "Improved Blasting Techniques Give Wider Applications for Wheel
Loaders," Mining Magazine, Vol. 121, No. 3, September 1969, p. 207.
349
------- |