IC 9486 

Information Circular/2006 

Handbook for Methane Control 

in Mining 

Department of Health and Human Services

Centers for Disease Control and Prevention

National Institute for Occupational Safety and Health

TM
Information Circular 9486 

Handbook for Methane Control in Mining 

By Fred N. Kissell, Ph.D. 

DEPARTMENT OF HEALTH AND HUMAN SERVICES 

Centers for Disease Control and Prevention 

National Institute for Occupational Safety and Health 

Pittsburgh Research Laboratory 

Pittsburgh, PA 

June 2006 
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DHHS (NIOSH) Publication No. 2006–127 
CONTENTS
 Page

About this handbook ......................................................................... 1 
 

Acknowledgments ........................................................................... 1 
 

Chapter 1.—Facts about methane that are important to mine safety, by F. N. Kissell ........................ 3 
 

Chapter 2.—Sampling for methane in mines and tunnels, by F. N. Kissell . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 27 
 

Chapter 3.—Methane control at continuous miner sections, by F. N. Kissell, C. D. Taylor, and

 G. V. R. Goodman . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 37 
 

Chapter 4.—Preventing methane ignitions at longwall faces, by F. N. Kissell and A. B. Cecala . . . . . . . . . . . . . . 55 
 

Chapter 5.—Bleeder systems in underground coal mines, by J. E. Urosek, W. J. Francart, and D. A. Beiter . . . . . 63 
 

Chapter 6.—Coal seam degasification, by P. C. Thakur . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 77 
 

Chapter 7.—Managing excess gas emissions associated with coal mine geologic features, by J. P. Ulery . . . . . . . 97 
 

Chapter 8.—Forecasting gas emissions for coal mine safety applications, by C. O. Karacan and

 W. P. Diamond . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 113 
 

Chapter 9.—Control of methane during coal mine shaft excavation and filling, by F. N. Kissell . . . . . . . . . . . . . 127 
 

Chapter 10.—Methane control in highwall mining, by J. C. Volkwein and F. N. Kissell . . . . . . . . . . . . . . . . . . . 135 
 

Chapter 11.—Control of methane in coal silos, by F. N. Kissell . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 141 
 

Chapter 12.—Explosion hazards of coal dust in the presence of methane, by K. L. Cashdollar and M. J. Sapko . 147
 

Chapter 13.—Methane control in metal/nonmetal mines, by H. J. Head and F. N. Kissell . . . . . . . . . . . . . . . . . . 151 
 

Chapter 14.—Preventing methane gas explosions during tunnel construction, by F. N. Kissell . . . . . . . . . . . . . . 169 
 

Index . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 181 
 

ILLUSTRATIONS

 1–1. Depiction of methane being diluted into a moving air stream ..................................... 4

 1–2. Methane explosibility diagram ............................................................ 5

 1–3. Effect of elevated pressure on methane explosibility limits ...................................... 6

 1–4. Estimated methane content of coal versus depth and rank . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 10

 1–5. Original gas content of mined coal versus mine emission . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 11

 1–6. Methane layering with roof, side, and floor sources . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 14

 1–7. Recirculation in a heading . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15

 1–8. Dust scrubber recirculation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 15

 1–9. Ignitions per year per gate road . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 18 
 

1–10. Relative frequency of occurrence of major and minor gas and dust explosions . . . . . . . . . . . . . . . . . . . . . . 20

 2–1. Depiction of methane being diluted into a moving air stream . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 29

 2–2. Methane profile map from a simulated continuous miner face . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 32

 2–3. Recorder chart from a machine-mounted methane monitor . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 33

 3–1. Recorder chart from a machine-mounted methane monitor . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 37

 3–2. Peak average methane emissions for several U.S. coalbeds . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 38

 3–3. Ventilation setback distance . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 39

 3–4. Methane dilution capacity for exhaust line curtain at setbacks of 10 ft and 20 ft . . . . . . . . . . . . . . . . . . . . . 39

 3–5. Spray fan system on a continuous miner . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 40

 3–6. Methane dilution capacity of the spray fan system . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 41

 3–7. Dust scrubber used in conjunction with blowing ventilation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42

 3–8. Methane dilution capacity from MSHA scrubber tests . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 42

 3–9. Test configuration used by Thimons et al. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 43 
 

3–10. Dust scrubber used with exhaust ventilation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 45 
 

3–11. Graph showing that performance of a spray fan depends on water pressure . . . . . . . . . . . . . . . . . . . . . . . . . 47 
 

3–12. Effect of bit wear on frictional ignition . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 49

 4–1. Increased methane liberation during coal bumps . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 56

 4–2. Modified shearer-clearer system . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 56

 4–3. Methane concentration at shearer during tail-to-head pass . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 57

 4–4. Plan view of walkway and gob curtains . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 57

 4–5. Walkway curtain at support No. 3 forces air over the shearer . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 58 
 
ILLUSTRATIONS—Continued 

Page

 4–6. Upwind-pointing venturi spray creates eddy zone and methane buildup . . . . . . . . . . . . . . . . . . . . . . . . . . . 58

 4–7. L-shaped wing curtain, which can cause methane buildup . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 59

 4–8. Anti-ignition back spray . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 60

 4–9. Best methane monitor location . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 61

 5–1. Simplified illustration of a wrap-around bleeder system for a longwall panel . . . . . . . . . . . . . . . . . . . . . . . 64

 5–2. Simplified illustration of a flow-through system with bleeder entries . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 65

 6–1. Gas content of coal versus gas pressure for some U.S. coals . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 80

 6–2. Generalized optimum point for methane drainage . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 81

 6–3. Simplified illustration of a typical longwall panel layout in a U.S. coal mine . . . . . . . . . . . . . . . . . . . . . . . 82

 6–4. Limits of the gas emission space . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 82

 6–5. Longwall panel methane drainage . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 83

 6–6. Premining methane drainage from surface . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 84

 6–7. Methane drainage by the packed cavity method . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 86

 6–8. Partially filled longwall gob . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 87

 6–9. Methane drainage with cross-measure boreholes . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 87 
 

6–10. Methane drainage by the superjacent method . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 89 
 

6–11. Simplified illustration of methane drainage by vertical gob wells . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 89 
 

6–12. Underground methane pipe installation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 91 
 

6–13. Surface installation for horizontal drainage boreholes . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 92

 7–1. Methane drainage scenarios for paleochannels . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 99

 7–2. Methane drainage of a normal fault from the “footwall” side . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 102

 7–3. Methane drainage of a normal fault from the “hanging” wall side . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 102

 7–4. Methane drainage of a reverse fault from the “hanging” wall side . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 103

 7–5. Methane drainage of a reverse fault from the “footwall” side ................................... 103

 7–6. Typical clay vein acting as a barrier to gas migration . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 104

 7–7. Methane drainage of a clay vein gas barrier . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 105

 7–8. Hypothetical gas cell formed by clay veins . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 106

 8–1. Methane content as a function of depth and coal rank . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 115

 8–2. Gas content versus depth and coal rank, Black Warrior basin, Alabama . . . . . . . . . . . . . . . . . . . . . . . . . . 115

 8–3. PFG/FGK method to predict gas emissions in a previously disturbed zone . . . . . . . . . . . . . . . . . . . . . . . . 120

 8–4. Gas pressure method: residual gas pressure lines are dependent on thickness of the mined coalbed . . . . . 121

 9–1. Mine shaft under construction . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 127

 9–2. Ventilation airflow is restricted by work deck . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 129

 9–3. Water ring in shaft wall . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 130

 9–4. Raise bore shaft with second hole for ventilation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 131 
 

10–1. Explosibility diagram for methane-air-inert gas mixtures . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 136 
 

10–2. Schematic of inert gas system for auger mining . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 137 
 

11–1. Desorption test apparatus . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 142 
 

11–2. Gas emitted from conveyor belt grab samples . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 142 
 

11–3. Methane concentration gradient above the coal in coal silo . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 143 
 

12–1. Flammable limits for mixtures of methane and coal dust . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 149 
 

13–1. Relation between quantitative composition and explosibility of mixtures of methane and air . . . . . . . . . . 154 
 

13–2. Examples of methane emission as reported in South African metal/nonmetal mines . . . . . . . . . . . . . . . . . 162 
 

14–1. Low concentrations of methane gas . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 170 
 

14–2. Exhausting system of face ventilation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 173 
 

14–3. Blowing system of face ventilation . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 173 
 

14–4. TBM ventilation system with low airflow zone where methane dilution suffers . . . . . . . . . . . . . . . . . . . . 174 
 

14–5. TBM ventilation system with second duct exhausting air from the face . . . . . . . . . . . . . . . . . . . . . . . . . . . 175 
 

14–6. Auxiliary system with adequate overlap . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 175 
 

14–7. Auxiliary system with no overlap and poor methane dilution . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 175 
 

14–8. Loss of efficiency from mismatched airflow directions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 176 
 
TABLES

 6–1. Major coal mine explosions in the United States, 1970–present . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 78

 6–2. Gassiness of coal seams . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 79

 6–3. Methane capture ratios for postmining methane drainage techniques . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 88 
 

14–1. Initiating events for tunnel methane explosions . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . 180 
 

UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT 

cc/g cubic centimeter per gram 

cfd cubic foot per day 

cfm cubic foot per minute 

cm centimeter 

cm3 cubic centimeter 

ft foot 

ft2 square foot 

ft3 cubic foot 

ft3

/st cubic foot per short ton 

ft/min foot per minute 

ft3

/min cubic foot per minute 

g gram 

gpm gallon per minute 

hr hour 

in inch 

in3 cubic inch 

kg/m3 kilogram per cubic meter 

kg/sec kilogram per second 

kV kilovolt 

L liter 

L/kg liter per kilogram 

lb pound 

lb/ft3 pound per cubic foot 

m meter 

m3 cubic meter 

m/sec meter per second 

m3

/sec cubic meter per second 

mD millidarcy 

min minute 

mm millimeter 

oz/ft3 ounce per cubic foot 

ppm part per million 

psi pound per square inch 

psig pound per square inch gauge 

rpm revolution per minute 

scfm standard cubic foot per minute 

sec second 

vol % volume percent 

°C degree Celsius 

°F degree Fahrenheit 
 

 

HANDBOOK FOR METHANE CONTROL IN MINING 

Fred N. Kissell, Ph.D.,1 Editor 

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and 

Health, Pittsburgh, PA (retired). 

ABOUT THIS HANDBOOK 

This handbook describes effective methods for the control of methane gas in mines and tunnels. 

It assumes the reader is familiar with mining. The first chapter covers facts about methane 

important to mine safety, such as the explosibility of gas mixtures. The second chapter covers 

methane sampling, which is crucial because many methane explosions have been attributed to 

sampling deficiencies. 

Subsequent chapters describe methane control methods for different kinds of mines and mining 

equipment, primarily for U.S. coal mines. These coal mine chapters include continuous miners 

and longwalls, including bleeders. Coal seam degasification is covered extensively. Other coal 

mine chapters deal with methane emission forecasting and predicting the excess gas from 

troublesome geologic features like faults. Additional coal chapters contain methane controls for 

shaft sinking and shaft filling, for surface highwall mines, and for coal storage silos. 

Major coal mine explosion disasters have always involved the combustion of coal dust, 

originally triggered by methane. Thus, a chapter is included on making coal dust inert so it 

cannot explode. Methane is surprisingly common in metal and nonmetal mines around the 

world, as well as in many tunnels as they are excavated. Accordingly, a chapter is included on 

metal and nonmetal mines and another on tunnels. 

Proper ventilation plays the major role in keeping mines free of hazardous methane 

accumulations. The ventilation discussed in this handbook, except for the chapter on bleeder 

systems, deals only with so-called face ventilation, i.e., ventilation of the immediate working 

face area, not ventilation of the mine as a whole. The omission of whole-mine ventilation was 

necessary to keep this handbook to a reasonable size and because a huge amount of excellent 

information is available on the subject.2 

 2

See the following: 

 Hartman HL, Mutmansky JM, Ramani RV, Wang YJ [1997]. Mine ventilation and air conditioning. 3rd ed. 

New York: John Wiley & Sons, Inc. 

 McPherson MJ [1993]. Subsurface ventilation and environmental engineering. Kluwer Academic Publishers. 

 Tien JC [1999]. Practical mine ventilation engineering. Chicago, IL: Intertec Publications. 

ACKNOWLEDGMENTS 

Much credit is due to Joe Schall, the Giles Writer-in-Residence at The Pennsylvania State 

University, for providing valuable help in editing this handbook. 
 

3 

CHAPTER 1.—FACTS ABOUT METHANE 

THAT ARE IMPORTANT TO MINE SAFETY 

By Fred N. Kissell, Ph.D.1 

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

2

In This Chapter 

9 The explosibility of methane gas mixtures 

9 Effect of pressure and temperature on explosibility 

9 Less common sources of methane ignitions 

9 The amount of methane stored in coal 

9 Forecasting the methane emission rate 

9 Layering of methane at the mine roof 

9 When the recirculation of mine air is hazardous 

9 The importance of higher air velocity in preventing methane explosions 

and 

9 Mine explosions, barometric pressure, and the seasonal trend in explosions 

Dealing with methane in mines and tunnels requires knowledge of the circumstances under 

which dangerous accumulations of methane are likely to occur. This knowledge involves the 

properties of the gas itself, an awareness of where these accumulations are likely to occur, and 

facts on how methane mixes safely into the mine air. 

The other chapters in this handbook address the handling of methane under a variety of specific 

circumstances, such as at continuous miner faces or coal storage silos. This chapter addresses 

some broad concepts that serve as a foundation for the suggestions provided in other chapters. 

THE EXPLOSIBILITY OF METHANE GAS MIXTURES 

Methane entering a mine or tunnel often enters as a localized source at high concentration. 

Figure 1–1 depicts a cloud of methane being diluted into a moving air stream. In this illustration, 

methane enters the mine from a crack in the roof. As the methane emerges from the crack, 

it progressively mixes with the ventilation air and is diluted. In the event that this progressive 

dilution reduces the concentration from 100% to 1%,2

Concentration percentage values refer to percent by volume.
 

as shown in Figure 1–1, the methane 

passes through a concentration range of 15% to 5%, known as the explosive range. In the 

explosive range, the mixture may be ignited. Above 15%, called the upper explosive limit 

(UEL), methane-air mixtures are not explosive, but will become explosive when mixed with 

more air. 
4 

 Figure 1–1.—Depiction of methane being diluted into a moving air stream. 

 

 

 

 

Below 5%, called the lower explosive limit (LEL), methane-air mixtures cannot ignite.3 

3

Sometimes the UEL and LEL are referred to as the upper and lower flammable limits (UFL and LFL).

Because methane always passes 

through an explosive range during 

dilution, an effective mine ventilation system will ensure that this 

passage through the explosive range 

is as rapid as possible and that the 

volume of gas mixture in or above 

the explosive range is minimized. 

Even though methane-air mixtures under 5% are not 

explosive, worldwide experience with methane in 

mines has indicated that a considerable margin of 

safety must be provided. 

Addition of inert gases. An inert gas such as nitrogen or carbon dioxide cannot chemically 

react with methane. As a result, inert gases can be added to an explosive methane-air mixture to 

make it nonexplosive. 

Explosibility diagrams are available to find how much inert gas is necessary. For example, 

Zabetakis et al. [1959] have provided a helpful explosibility diagram that shows whether a 

methane-air mixture is explosive after an inert gas such as nitrogen or carbon dioxide is added 

(Figure 1–2). This diagram shows that methane-air-inert gas mixtures fall into one of three 

categories: (A) explosive, (B) explosive when mixed with air, or (C) nonexplosive, depending 

on the percentage of methane and the percentage of “effective inert.” Effective inert is 

calculated from the percentage of “excess nitrogen”4

 and the percentage of carbon dioxide in the 

mixture. 

4

The percentage of excess nitrogen is the percentage of nitrogen in the sample minus the percentage of “normal 

nitrogen.” Normal nitrogen is calculated from the ratio of nitrogen to oxygen normally found in air—a factor of 3.8. 

To calculate the effective inert, suppose, for example, that inert gas is added to a methane-air mixture and that a gas 

analysis shows that the final mixture has 6.6% oxygen, 4% carbon dioxide, 4.3% methane, and 85.1% nitrogen. The 

effective inert is then determined in three steps. First, in this example, the oxygen percentage is 6.6%, so the 

percentage of normal nitrogen is 3.8 times 6.6%, or 25.1%. Second, since the percentage of excess nitrogen is the 

percentage of nitrogen in the sample minus the percent of normal nitrogen, the excess nitrogen is 85.1% minus 

25.1%, or 60%. Third, according to the equation shown in Figure 1–2, since the carbon dioxide in the sample is 

4%, the effective inert is now 60%, plus 1.5 times (4%), or 66%. This gives the “composition point” shown in 

Figure 1–2. (Carbon dioxide has been found to be 50% more effective than nitrogen in inerting, so a multiplying 

factor of 1.5 is used). 
Figure 1–2 shows a “composition point” with 4.3% methane 

and 66% “effective inert.” The 

arrows indicate how the composition point is shifted by the 

addition of more methane, 

more air, or more inert gas. 

For example, adding more air 

shifts the composition point in 

the direction of 100% air (0% 

methane, 0% effective inert), 

whereas adding more nitrogen 

shifts the composition point in 

the direction of 100% effective 

inert (0% methane, 0% air). 

 Figure 1–2.—Methane explosibility diagram (from Zabetakis et al. 

[1959]). 

Addition of other flammable 

gases to air. Mine gas mixtures can contain flammable 

gases other than methane, 

principally ethane, hydrogen, 

and carbon monoxide. The 

explosive limits of these mixtures in air are calculated using Le Chatelier’s law [1891]. This law specifies that if one gas 

mixture at its lower explosive limit is added to another gas mixture also at its lower explosive 

limit, then the combination of the mixtures will be at the lower explosive limit of the combination. Mathematically, 

100 L = , P 1 1 2 2 / L + P / L + ⋅⋅⋅⋅ P X X / L

where P1 + P2 + ⋅⋅⋅⋅ PX = 100 . Here, we have gas mixtures of gas #1, gas #2, and up through 

gas #X. L is the lower explosive limit of the mixture, P is the proportion of each gas in the 

mixture, and L1, L2, and LX are the lower explosive limits in air for each combustible gas 

separately [Jones 1929]. 

Combinations of both flammable and inert gases. For combinations of both flammable and 

inert gases in air, explosive limits can be obtained through diagrams provided by Zabetakis et al. 

[1959]. More explosibility diagrams are available from other sources, and Holding [1992] has 

reviewed the features of each of them. 

EFFECTS OF PRESSURE AND TEMPERATURE ON EXPLOSIBILITY 

5 

Effect of pressure on explosibility limits. According to Kuchta [1985], the flammability limits 

of hydrocarbon vapor-air mixtures (such as methane-air mixtures) vary only slightly with 
 

6 

reduced pressure, except at very low 

pressures, such as below 1⁄4 atmosphere. At elevated pressures, the 

lower limits of hydrocarbon-air 

mixtures generally decrease slightly, 

but the upper limits increase greatly. 

Figure 1–3 shows the variation in 

methane lower explosive limit and 

upper explosive limit with elevated 

pressure. 

Figure 1–3.—Effect of elevated pressure on methane explosibility limits. 

Effect of temperature on explosibility limits. The effect of 

temperature on the explosibility 

limits of methane is modest. For 

example, the LEL of methane-air 

mixtures at -100 °C is 5.6% methane, 

and at +100 °C it is 4.8% methane. 

The UEL of methane-air mixtures 

at +100 °C is 16.3% methane 

[Zabetakis 1965]. 

LESS COMMON SOURCES OF METHANE IGNITIONS 

There are many well-known methane ignition sources in mines, ranging from frictional ignitions 

caused by cutting bits (Chapter 3) to open flames, explosives, and electrical sparking. However, 

there are other less recognized ignition sources. A review of these is worthwhile here. 

Hot solids. The temperature at which a hot solid can ignite methane is quite high. Coward and 

Ramsay [1965] report that the minimum ignition temperature in a closed vessel is about 675 °C, 

but when the hot surface is exposed to convection currents, the minimum temperature is higher. 

For example, ignition from a hot steel bar requires 990 °C. Kuchta [1985] found that ignitions 

by any heated surface depend on the dimensions of the surface. He reports methane ignition 

temperatures ranging from 630 to 1,220 °C. 

However, when an ignitable dust is present on the hot surface, this dust is more readily ignited 

than methane. The burning dust can then ignite the methane. Kim [1977b] reported laboratory 

studies in which the spontaneous ignition temperature of coal dust layers was as low as 160 °C. 

As a result, Mine Safety and Health Administration (MSHA) regulations require that the surface 
temperature of permissible electrical equipment and diesel equipment5

 in coal mines6 not exceed 

150 °C. 

 

7 

5

The MSHA 150 °C requirement applies to diesel equipment intended for use in areas of the coal mine where
 

permissible electrical equipment is required. Some state regulations require a surface temperature maximum of 

150 °C for all diesel equipment in coal mines. 

6

In gassy metal/nonmetal mines, the diesel surface temperature limit is 204 °C (400 °F).

8

These ignitions were not always in coal mines. For example, an explosion in a Detroit water intake tunnel on
 

December 11, 1971, killed 22 workers. The ignition was attributed to sparks caused by dropping a 23-in-diam
 

drill bit a distance of 16 ft onto the concrete tunnel floor [Detroit Water and Sewerage Department 2005]. 

Thermite sparking from light metal alloys. When light metal alloys strike rusty steel, the 

resulting sparks can ignite methane. This so-called thermite sparking can appear in different 

ways. Thomas [1941] showed that striking aluminum-painted rusty iron with a tool could ignite 

methane. Margerson et al. [1953] readily ignited methane by dropping a piece of magnesium 

alloy onto a rusty steel plate. Findings such as these have inhibited the use of light metal alloys 

in mines. 

Today, sparking from light metals is minimized 

by using less incendive alloys. For example, 

MSHA7

 requires that aluminum fan blades 

contain no more than 0.5% magnesium. 

7

Per 30 CFR 57. Code of Federal Regulations. See CFR in references.
 

Adiabatic compression. McPherson [1995] has proposed that adiabatic compression of 

methane-air-coal dust mixtures by falling roof can be responsible for some methane ignitions in 

coal gobs. A theoretical model indicates that the temperatures attained are adequate to ignite 

such mixtures if the roof fall is extensive in plan area, but not necessarily of large thickness. In a 

later laboratory study by Lin et al. [1997], an experimental apparatus was built to simulate the 

adiabatic compression that might result from roof falls. This apparatus, which dropped a 1,320­

lb weight, ignited the methane and dust when they were in the proper concentration range. 

Sliding (or impact) friction between blocks of rock or between rock and steel. Sliding friction between falling blocks of sandstone or pyrites, or between hard rock and steel, can produce 

incendive streaks that ignite methane [Powell and Billinge 1975]. 

According to Coward and Ramsey [1965], methane ignitions from rock falling onto rock were 

reported as early as 1886. Laboratory experiments confirmed this effect. The higher the quartz 

content, the more likely an ignition; however, the necessary rubbing distance was always greater 

than could be envisioned from a fall underground. Ignitions from rock falling onto steel were 

reported as early as 1908 and seen throughout the 20th century, both underground8

 and in the 

laboratory. Today, the most likely source of steel-rock ignitions are cutter picks on mining 

machines, a topic covered in Chapter 3. 
8 

Static electricity. Protection against discharges of static electricity is a common feature of mine 

regulations. Precautions are required for electrical equipment, for explosives loaded into 

blastholes (30 CFR 57.6602), for nonmetallic rotating parts such as belts (30 CFR 18.26), for 

venturi air movers powered by compressed air, and for similar circumstances where static 

charges are likely to collect. The National Fire Protection Association [NFPA 2000] and many 

Internet sites have more information on how to prevent static electricity. 

Although controlling static electricity in mines is important, it has not been a common source of 

methane explosions in underground mines, possibly due to higher humidity underground. 

Nevertheless, extra precaution should be taken where acetylene is used, since acetylene is much 

more easily ignited by static electricity than methane. 

Lightning. The South African underground coal mining industry has seen many incidents 

related to the passage of lightning storms on the surface. These incidents included electrical 

shocks, visible sparking from mining equipment, premature detonation of explosives, and 

methane explosions. The majority were in shallow mines at depths of 300 ft or less. Precautions 

to prevent these lightning-related incidents included lightning warning techniques, the use of less 

sensitive detonators, modified blasting practices, and improved electrical grounding of mining 

equipment [Geldenhuys et al. 1985]. 

In the United States, lightning has been reported as the explosion source at two mines in 

Alabama [Checca and Zuchelli 1995]. Both mines had been worked since the 1970s and had 

large sections that had been abandoned and permanently sealed. The mines were deeper (500 

and 1,200 ft) than those in South Africa. However, in the investigation following each of the 

explosions, it was found that the lightning strike occurred at a location where there was a convenient conduit for electrical current into a sealed area of the mine. In one instance, it was an old 

capped shaft; in the other, it was a test well with a metal casing that extended from a foot below 

the surface to a foot above the mine roof. On the surface, this test well was located in a fenced 

area that enclosed a methane-pumping unit. 

More recently, Novak and Fisher [2000] conducted computer simulations of lightning propagation through the earth to confirm whether lightning could penetrate a 600-ft-deep mine with 

enough energy to trigger methane explosions. They found that the presence of a steel-cased 

borehole dramatically enhances the possibility of lightning starting an explosion. With a steelcased borehole, the calculated voltage difference between a roof bolt adjacent to the borehole 

and a section of rail on the floor was 15.6 kV. 

THE AMOUNT OF METHANE STORED IN COAL 

Coal is the major source of methane gas in mines. Smaller (but still dangerous) amounts of 

methane are found in oil shale, porous rock, and water. Methane in oil shale has been measured 

by Kissell [1975], Matta et al. [1977], and Schatzel and Cooke [1994]. Methane stored in porous 
9 

rock and water is of most concern in tunneling and has been discussed in some detail by Doyle 

[2001].9 

 

 

 

9

Doyle [2001] also provides a helpful discussion of methane in coal.

Methane in coal. The amount of methane in coal is measured by using the “direct-method” test 

during exploration drilling from the surface, or it is estimated from the properties of the coal and 

the gas pressure or depth of the coalbed. The direct-method test for surface exploration drilling 

was first used by Kissell et al. [1973]. Improvements to the method were made by others 

[Diamond and Levine 1981; Ulery and Hyman 1991; Diamond et al. 2001]. McLennan et al. 

[1995] have written a thorough description of how to conduct a direct-method test and analyze 

the results. 

In the direct method, a drill core of coal is brought to the surface, it is enclosed in an airtight 

container, and the methane emitted from the core is measured. The amount of gas that escaped 

the core as it was being brought to the surface is calculated. Later, the core is crushed and the 

residual gas given off during crushing is measured. Added together, these allow one to estimate 

the amount of gas in the coalbed. 

A considerable amount of direct-method testing has taken place, so it is usually possible to get 

gas data for most U.S. coalbeds. For example, Diamond et al. [1986] have given the results of 

1,500 direct-method tests on coal samples from more than 250 coalbeds in 17 states. 

If direct-method results are not available, the amount of gas in coal may be roughly estimated 

from adsorption data. This estimate requires knowledge of the proximate analysis of the coal, 

assumes a standard moisture and ash content, and uses the hydrostatic head to estimate pressure 

[Kim 1977a]. Figure 1–4 summarizes methane content data for different rank coals at various 

depths using the hydrostatic head assumption.10

10The adsorption of mixed gases (methane and carbon dioxide) on coal has been measured by Lama [1988].

However, because the actual pressure is often 

less than the pressure of the hydrostatic head,11 the methane content values shown in Figure 1–4 

are very much an upper limit. 

11Kim [1977a] has compared the actual pressure to the hydrostatic head pressure for several U.S. mines. The results 

varied from 50% to 100% of the hydrostatic head. For Australian mines, Lama and Bartosiewicz [1982] estimate
 

gas pressures ranging from 50% to 90% of the hydrostatic head. For U.K. mines, Creedy [1991] reports that gas 

pressures are generally less than 20% of the hydrostatic head. 

FORECASTING THE METHANE HAZARD 

Additional hazard calls for additional precaution, so an estimate of the expected methane emission is valuable for both new and existing mines. 

Coal mines. When an active mine is nearby, the most effective way to forecast the methane 

emission rate for a mine under development is to use the emission rate from a nearby mine (or 

section) where similar mining methods are used under similar geological conditions. Corrections 

can be made for those factors that are likely to shift the emission rate. Such factors are 
 

 Figure 1–4.—Estimated methane content of coal 

versus depth and rank. Values shown are an upper limit. 

10 

differences in coalbed depth, differences 

in production rate, and geological anomalies12 such as faults. Some of these corrections are simple, if inexact.13 

12The effect of geological anomalies is discussed in Chapter 7. 

13Sometimes very inexact. For example, the methane emission can be assumed as roughly proportional to depth.
 

However, Diamond and Garcia [1999] compared the methane emission rates of two longwall panels a mile apart. 

The second panel was 37% deeper than the first, but gave 61% higher emissions. The emissions were much higher
 

because the elapsed time between development of the panel and retreat of the longwall face was much less in the 

second panel. Thus, there was less time for the second panel to drain gas into the returns, so when it was mined the 

emission was higher than expected. 

When no other mine is nearby, a very 

rough emission forecast for the entire 

mine may be obtained using the gas 

content of the coal. For example, Saghafi 

et al. [1997] have reported the relationship 

between gas content and mine emission 

for Australian mines (Figure 1–5). The 

amount of methane released from the mine 

exceeded the methane in the mined coal 

by a factor of 4. This differs from the 

results of Kissell et al. [1973], who 

measured a factor of about 7 for some 

U.S. mines. The difference is probably 

due to methane emissions from adjacent 

coalbeds and porous rock. Other associations between mine emission and gas 

content have been made without using 

production data. Grau and LaScola [1984] 

have correlated the mine emission of some U.S. mines in cubic feet per day with the in situ gas 

content in cubic feet per ton.14 

14Grau and LaScola report 1980 mine emission data. Reliable U.S. data after 1980 are not available. As degasification programs became widespread, the degas quantities were retained as confidential information by coal companies. 

Much more on forecasting for coal mines is covered in the 

coal mine forecasting chapter (Chapter 8). Forecasting 

for metal and nonmetal mines is covered in this section. 

Forecasting for tunnels is covered in the tunneling chapter 

(Chapter 14). 
 Figure 1–5.—Original gas content of mined coal versus mine 

emission. (Mines producing less than 106

 tons/yr were omitted.) 

11 

In the last 25 years, sophisticated 

computer models have become 

available for coalbed methane 

forecasting. Most of these are 

driven by the need to estimate 

how much gas can be extracted 

to generate revenue. Chapter 8 

deals with coalbed emission 

forecasting. Also, Creedy [1996] 

has provided a comprehensive 

report on methane prediction for 

coal mines. 

Metal and nonmetal mines. 

Metal and nonmetal mines that 

encounter methane emissions are 

placed by MSHA in a special 

regulatory category that requires 

extra precautions against 

methane explosions. Placement in these special regulatory categories (30 CFR 57, Subpart T) is 

usually triggered by a specific incident, such as measurement of a methane concentration of 

0.25% or more, an ignition of methane in the mine, or an outburst in the mine if it is a salt mine. 

Most of these incidents are probabilistic in nature. For example, as the methane hazard level 

increases, the chance of an ignition goes up, but an ignition (especially a small ignition that can 

serve as a warning) is by no means certain. 

How then does the operator of a metal or nonmetal mine estimate the methane hazard level 

without waiting for such an incident to take place? Thimons et al. [1977] established a simple 

guideline that would enable mine personnel to evaluate the methane hazard. In their research, 

they measured trace methane concentrations in 53 metal and nonmetal mines. They found that 

mines with a return concentration exceeding 70 ppm of methane were inevitably classified as 

gassy.15

 

15In 1977, the MSHA classification system for mines with methane was different from the one in existence today. 

However, the triggers that lead to extra precautions (such as measurement of 0.25% or an ignition in the mine) are 

similar.
 

 Although a measurement of concentration alone is not the complete methane story,16 

a return concentration exceeding 70 ppm should serve as an alert to the presence of gas that has 

not yet shown itself in other ways. 

16See the section in this chapter on the importance of air velocity in preventing methane explosions. 
 

 

12 

LAYERING OF METHANE AT THE MINE ROOF 

The density of methane is roughly half that of air, so methane released at the mine roof may form 

a buoyant layer that does not readily mix into the ventilation air stream. Such layers have been 

the source of many mine explosions,17 so it is important to understand the circumstances that led 

to the formation of methane roof layers and the methods used to dissipate them. 

17For example, the 1993 Middelbult coal mine explosion in Secunda, South Africa, was attributed to a methane 

layer [Davies et al. 2000]. 

Creedy and Phillips [1997] have written a 

thorough summary of methane layering and 

its implications for South African mines. 

Detecting methane layers. Methane layers are largely a result of inadequate ventilation. Raine 

[1960] asserted that a measurement of ventilation velocity is of most practical importance. He 

found that under conditions of “normal firedamp emission,”18 a

18The phrase “normal firedamp emission” was not further defined. However, it is clear that at abnormally high gas 

feeds, higher velocities are required. In a laboratory study, Bakke and Leach [1962] found that 230 ft/min air velocity was required to disperse a layer generated by a release of 12 ft3

/min of methane. 

n air velocity of 100 ft/min measured at the roof was enough to prevent layering.19 

19The 100 ft/min applies only to horizontal entries. Higher velocities are suggested for inclined entries [Bakke and 

Leach 1965].

Most current-day estimates of the necessary 

velocity are close to this value.20 

20For example, McPherson [2002] suggests 0.4 m/sec, or about 80 ft/min. 

An alternative approach to estimating the air velocity required to prevent layering is to use a 

“layering number,” devised by Bakke and Leach [1962]. The layering number is a dimensionless number expressed as— 

U 

L = 

V 

W

37 ⋅ 3 

where L is the layering number, U is the air velocity in feet per minute, V is the methane release 

rate in cubic feet per minute, and W is the entry width in feet. In layering experiments conducted 

by Bakke and Leach, methane was released at a single point at the mine roof, and the air velocity 

necessary to dilute the layer was measured. They found that mixing by turbulence began at 

layering numbers larger than 2, but that a layering number of 5 was necessary for adequate dilution.21 

21At high methane emission rates, the layering number suggests that velocities higher than 100 ft/min are necessary 

to prevent layering. For example, for a methane emission rate of 16 ft/min in a 16-ft-wide entry, the velocity 

required to prevent layering is 185 ft/min. 

Compared to the 100-ft/min criterion, the layering number concept is more difficult to 

apply because the methane release rate V is usually not known. 
 

13 

Aside from inadequate ventilation, there are other circumstances under which methane layers are 

probable. Airways next to gobs are an example. Many of the concerns about layers were sharpened by experience in the 1960s with advancing longwalls in the United Kingdom. At these 

longwalls, frequently traveled gate roads were directly adjacent to fresh longwall gob, where 

broken overburden provided a ready pathway for roof gas emissions. 

Thorough gas monitoring is a key to dealing safely with methane layers. Care in monitoring is 

particularly important if— 

•	 The air velocity measured at the roof level is 100 ft/min or less. 

•	 The airway is next to a gob22

22Five miners were killed in a 1972 methane explosion at the Itmann No. 3 Mine in West Virginia. The explosion 

was in a trolley haulageway that ran adjacent to a longwall gob and was attributed to excessive pressure from the 

adjacent strata [Richmond et al. 1983]. 

 or intersects a geologic anomaly, such as a fault, that can 

serve as a conduit for gas. 

•	 The mine roof (or tunnel crown) is not within easy reach, so measurements at roof level 

are less apt to be carried out regularly. 

•	 The airway has cavities [Titman et al. 1965; Vinson et al. 1978] or roof-level obstructions 

to air movement. 

•	 The airway is inclined more than 5° [Bakke and Leach 1962]. 

Workers who test for methane layers should be aware that the gas concentrations in these layers 

may fall outside of the accurate operating range of catalytic heat of combustion sensors. For 

accurate operation of these sensors, the concentration of methane must be below 8% and the 

concentration of oxygen must be above 10%. Also, when measuring methane concentrations 

above 8%, instruments with catalytic heat of combustion sensors can act in a way that is misleading, responding with a rapid upscale reading followed by a declining or erratic reading23 

[CSA 1984]. Such instrument behavior is a tipoff to the possible presence of high, possibly 

explosive methane concentrations. 

23Some instruments will report this as an out-of-range condition. For more information, consult the operating 

instructions for the instrument. 

When the roof is high and beyond convenient reach, measurements may be made in two ways. 

First, the methane detector can be equipped with a remote “sample-draw” capability. Sampledraw systems use a small pump or a hand-squeezed bulb to pull the sample through an extension 

probe and pass it through the detector. Some methane detectors have an accessory sampling 

pump that attaches to the detector; others have a built-in pump. 

Second, the methane detector can be attached to a cradle at the end of a long handle, which is 

then extended to the roof. This method permits a direct reading without aspiration if the user has 
good eyesight.24 Otherwise, the audible alarm on the detector could be set to engage at a low 

methane level. 

Mitigating methane layers. Methane layers are removed by increasing the ventilation quantity 

and reducing the gas flow by methane drainage. In instances where the source and layer size are 

 

24See the sampling chapter (Chapter 2) and the sections on methane detection in the continuous miner chapter 

(Chapter 3).

limited,25

14 

25For example, the immediate face area in a tunnel boring machine. 

a less satisfactory, but workable method is to use a (well-grounded) compressed airpowered venturi air mover or an auxiliary fan at each methane source to blow air at the source of 

the layer and disperse it [Creedy and Phillips 1997]. In either case, an aggressive sampling program is necessary to ensure safe conditions. 

Keep in mind that methane mixed with 

air cannot unmix to form a layer. 

The rib and floor as sources of 

methane layers. Methane layering 

occurs when methane is released at 

the mine roof. When methane is 

released at the mine floor or rib, this 

gas readily mixes into the ventilation 

air stream, losing its buoyancy. 

Figure 1–6, from Bakke and Leach 

[1962], compares 2-cfm methane 

sources at the roof, rib, and floor of 

the mine entry. Only the roof source 

produced a significant methane layer 

at the 2-cfm rate.26 

26The testing did not rule out the possibility of a layer at higher methane flows. 

 Figure 1–6.—Methane layering with roof, side, and floor 

sources (from Bakke and Leach [1962]). 

WHEN RECIRCULATION OF MINE AIR CAN BE HAZARDOUS 

Recirculation leads to higher methane levels 

only when recirculated air replaces fresh air. 
Recirculation of mine air takes place when some portion of return air is picked up by a fan and 

returned to the intake, potentially raising the contaminant level of the intake air. Concerns about 

whether recirculation is a hazard have persisted for decades. The first theory and experiments on 

the recirculation of mine air were reported by Bakke et al. [1964]. They concluded from a 

material balance27 and from experiments that the concentration of methane leaving any region is 

equal to the flow of methane into the region divided by the flow of fresh air into the region. The 

recirculation hazard is higher only if the amount of fresh air is reduced. 

An example of potentially hazardous recirculation in headings is shown in Figure 1–7. Here, the 

region is a heading designated ABCD. Within the heading is an auxiliary fan moving an air 

quantity Q. The fan inlet is in the wrong location, so the air entering the fan is some portion of 

fresh air nQ and some portion of methane-laden return air (1 - n)Q (where n varies between 0 

 

27The basic material balance equations are: air entering the zone equals air leaving the zone, and methane entering 

the zone equals methane leaving the zone. For more details on the material balances used, see Bakke et al. [1964] 

and Kissell and Bielicki [1975].

and 1). The concentration of methane is then: c = V/nQ .

28 

15 

Figure 1–7.—Recirculation in a heading. 

28Strictly speaking, it is c = V/ (nQ + V ) . However, since nQ>>V, the approximation c = V/nQ is adequate. 

Had the fan inlet been positioned at 

a better location, L1, the proportion of fresh air would be greater, the value of n would be higher, 

and the methane concentration lower. Had the 

fan inlet been positioned 

at location L2, the proportion of fresh air would 

be less, the value of n 

would be lower, and the 

methane concentration 

higher. 

Recirculation caused by 

dust scrubbers on continuous miners was 

studied by Kissell and 

Bielicki [1975] (Figure 

1–8). 

Figure 1–8.—Dust scrubber recirculation. 

In this instance, the 

fresh air entering zone 

ABCD was also designated nQ. The scrubber 

moved air quantity R, of 

which a fraction mR 

recirculated back into the 

zone. A new variable Z 

was necessary to account 

for air leaving the zone 

without passing through 

the scrubber. As before, 
 

 

16 

a material balance indicated that the concentration of methane in the zone depended only on the 

amount of fresh air entering the zone, or nQ, and the amount of methane entering the zone, or V. 

However, this left open the question of what factors determine the value of n. 

During experiments conducted with a full-scale model of a mine working face, Kissell and 

Bielicki found that n depended on whether or not the scrubber was turned on, and if turned on, 

where the scrubber exhaust was directed. Turning on the scrubber raised the value of n, reducing 

the methane concentration in the zone. Directing the scrubber exhaust into the return (in this 

case, behind the exhaust line curtain) was the best exhaust configuration, yielding an nQ fresh air 

value over four times higher and a zone methane concentration less than 1⁄4 when compared to 

the test with the scrubber off.29 

29Subsequent studies have confirmed the need to direct the scrubber exhaust into the return. See Figures 3–7
 

and 3–10. 

Bakke et al. and Kissell and Bielicki were primarily concerned with recirculation at coal mine 

working faces. Many other studies have been conducted on so-called district recirculation, i.e., 

recirculation of air in a major portion of a mine. District recirculation is produced by an underground fan that moves air from a return airway back into an intake airway, thus raising the total 

air quantity in that portion of the mine inby the underground fan. Improved dust control can be a 

result. Cecala et al. [1991] used SF6 tracer gas to study recirculation in a trona mine district that 

contained three operating continuous miner sections. The results were consistent with a methane 

material balance. Pritchard [1995] has discussed his own experience and the worldwide experience with controlled district recirculation. Pritchard concluded that— 

1.	 The initial volume of fresh air to the district should be maintained. 

2.	 The recirculation fan should be placed far enough from the face for the dust to settle out, 

but close enough to the face to minimize stopping leakage. 

3.	 District recirculation systems will increase flow and pressure losses in the mine circuit, 

producing a small drop in main fan flow.30 

30The exact amount will depend on the fan locations. 

4.	 Adequate monitoring and controls must be in place. 

In summary, recirculation will raise the methane concentration only when recirculated air is 

substituted for fresh air. If the amount of fresh air entering a zone is unchanged, the methane 

concentration in the zone will be unchanged. At continuous miner faces, if operation of a 

scrubber creates an airflow pattern that enhances the amount of fresh air entering the face zone, 

then operation of the scrubber will lower the methane concentration (and vice versa). 
 

 

 

17 

THE IMPORTANCE OF HIGHER AIR VELOCITY IN PREVENTING 

METHANE EXPLOSIONS 

Low air velocities can lead to poor mixing between methane 

and air. This poor mixing in turn leads to fluctuations in the 

methane concentration that make an ignition more likely. 

Bakke et al. [1967] first suggested that a measurement of the methane concentration alone is an 

incomplete means of assessing the ignition hazard and that other measurable ventilation quantities might be important.31 A study of methane ignitions in U.K. coal mines found that the 

probability of an ignition is determined by both the methane concentration and the densimetric 

Froude number, a dimensionless quantity related to the gas-mixing process in the presence of 

buoyancy forces. The expression for the Froude number F is— 

2

u F = ∆ρ g A

ρ 

∆ρ where u is the air velocity, is the density difference between air and methane divided by the 

ρ 

density of air, and A is the cross-sectional area of the airway. 

The data available to Bakke et al. resulted from 123 ignitions on faces and gate roads at U.K. 

longwalls during 1958–1965. Examination of the data indicated that the risk of an ignition was 

dependent on more than methane concentration alone and that it was possible to combine concentration and Froude number in one variable of the form c2

/F. 

Figure 1–9 shows the normalized number of ignitions P (ignitions per year per gate road) versus 

c

2

/F for the Bakke et al. data. The best fit to the data was P = 0.004 (c2/F)0.9 . A high correlation 

was obtained, indicating that, absent other sources of mixing, the risk of ignition P does depend 

on the variable c2

/F. 

In most mines, the square root of A does not change much compared to changes in c2

 and u

2

. Also, 

the factor of 0.9 is close to 1.0. It follows that ignition risk varies with the quantity (c/u)

2

. 

This departs from any notion that ignition risk depends on the concentration c alone.32 

31Actually, since ignition risk also depends on human factors, there is no reason to expect that ignition risk depends 

only on concentration. Mines with less gas may also have a less vigilant workforce. However, Bakke et al. only 

sought a correlation with measurable ventilation quantities.

32Subsequent work at longwall shearers in the 1980s failed to confirm this finding [Creedy and Phillips 1997; CEC 

1985], probably because water sprays on the shearer provided enough mixing between methane and air to overcome 

any velocity effect on mixing. 
 

 

 

 

 

Figure 1–9.—Ignitions per year per gate road. 

18 

As an example, assume that the 

methane concentration is 1.0% 

and that the air velocity is 100 

ft/min. If then the air velocity is 

raised to just 120 ft/min, the 

methane concentration becomes 

0.83%. If the ignition risk is 

proportional to (c/u)

2

, this 

modest increase in air velocity 

cuts the ignition risk in half.33 

33Some confirmation of the importance of air velocity in reducing ignition risk was obtained by Bielicki and Kissell 

[1974], who conducted a study of the methane concentration fluctuations produced by incomplete mixing of 

methane and air at a model coal mine working face. Poor mixing was characterized by wider concentration 

fluctuations and resulted from low airflow or a high methane release rate. In other studies of methane-air mixing, 

Kissell et al. [1974] found that good mixing was characterized by normally distributed peaks and poor mixing by 

log-normally distributed peaks. Schroeder and Kissell [1983] found the same effect and suggested that the term 

σ(log c), the standard deviation of the logarithms of the sampled peak concentrations c, be used as an indicator of 

mixing. 

The findings of Bakke et al. 

have important implications for 

using higher air velocity to prevent methane explosions: 

•	 In the absence of other means 

to promote mixing, raising air 

velocity is a highly effective 

way to reduce ignition risk. 

Higher air velocity promotes 

better mixing in addition to 

lowering the average 

concentration. 

•	 Water sprays and auxiliary air 

movers (small fans or 

compressed-air venturis) that 

promote mixing can reduce 

ignition risk. 

•	 At similar methane concentration levels, tunnels or mines with large cross-sectional entries 

and low air velocities have higher risk of ignition than those with small cross-sectional 

entries and higher air velocities. Both the lower velocity and higher area will work together 

to give a lower Froude number. 
MINE EXPLOSIONS, BAROMETRIC PRESSURE, 

AND THE SEASONAL EXPLOSION TREND 

19 

Although mine explosions are far less common than in the past, this deadly hazard to miners has 

not disappeared. Mine operators must always be alert to the circumstances that make a mine 

explosion more likely. The chapter on dust explosions (Chapter 12) outlines what must be done 

to prevent a methane ignition from triggering a dust explosion, which is usually lethal. Two 

other important factors, discussed here, are barometric pressure lows and drier dust in the winter. 

In South Africa, most mine explosions have followed a low 

in the barometric pressure; however, there is no seasonal 

frequency trend. In the United States, mine explosions have 

been more frequent in the winter because the dust is drier. 

Barometric pressure lows and mine explosions. Many researchers have documented an 

inverse relationship between barometric pressure and the amount of methane flowing from a 

mine [Carter and Durst 1955; Stevenson 1968; Füssell and Hudewentz 1974; Eschenburg 1977]. 

A falling barometric pressure causes expansion of the methane that has accumulated in underground cavities and crevices. The methane then flows into the mine, making an ignition more 

likely. 

Fauconnier [1992] has examined the role of barometric pressure changes in South African mine 

explosions. Using barometric data corresponding to 59 methane explosions (26 in coal, 33 in 

gold mines) for the period 1970–1989, he concluded that most of the explosions were associated 

with medium-term (longer than 1 day) downward trends in barometric pressure. He also concluded that explosions occur randomly during the year, in contrast with U.S. coal mines, which 

are known to have a seasonal trend. 

The seasonal trend in U.S. coal mine explosions. Historically, U.S. coal mine explosions have 

been more frequent in the winter than the summer [Boyer 1964]. Although barometric pressure 

might be a cause because changes in barometric pressure are more abrupt and intense in the 

winter months, it is also true that mines are drier in the winter because of the low moisture content of the air [Williams 1914; Pappas et al. 2002]. This means that coal dust is drier and more 

easily dispersed and ignited during the cold months. 

According to a study by Kissell et al. [1973], the second factor—drier coal dust—is the most 

influential in making winter explosions in U.S. coal mines more frequent. In this study, coal 

mine accident reports from 1911 to 1970 were examined to see whether winter explosions were 

more likely to occur in regions of the mine more susceptible to barometric pressure fluctuations 

(e.g., gobs). No such tendency was found. Next, based on the accident reports, explosions were 

divided into five different categories: 
 

 

 

20 

•	 All major explosions (where five or 

more miners were killed). Most major 

explosions involved both gas and dust. 

•	 Major dust explosions. Dust explosions are those where the accident 

investigators concluded that dust was 

directly ignited, without gas participating as an intermediate stage. 

Typically, these took place in mines 

known to be relatively free of 

methane and where the dust was 

ignited by a blown-out shot. 

•	 Minor dust explosions (fewer than 

five miners killed). 

•	 Minor gas explosions. Accident 

investigators concluded that dust was 

not involved. 

•	 Explosions in anthracite mines. These 

were known to be “gas only” because 

anthracite dust is not explosive under 

the conditions prevailing in mining. 

Figure 1–10 shows the relative frequency of each of these types of 

explosions for the period 1911 to 1970. 

When all major explosions are considered, the higher frequency in winter months is clearly evident. However, this trend is far more pronounced for the dust explosions. No trend favoring the 

winter months is evident in the anthracite mine explosions or those categorized as “gas only.” 

This provides strong evidence that it is dust, not gas, that accounts for the seasonal trend in 

U.S. coal mine explosions.34 

34The results of Kissell et al. do not contradict the conclusions by Fauconnier. The seasonal trend for coal mine 

explosions in the United States might be due to colder winter months or the fact that a different timeframe was 

examined. Also, Fauconnier combined explosions at both coal and gold mines. 

Figure 1–10.—Relative frequency of occurrence of major 

and minor gas and dust explosions. 

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CHAPTER 2.—SAMPLING FOR METHANE IN MINES AND TUNNELS 

 

27 

By Fred N. Kissell, Ph.D.1 

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

In This Chapter 

9 Instruments available to measure methane in mines and tunnels 

9 Using a portable detector in both accessible and restricted spaces 

9 Machine-mounted monitors: placement and response time 

9 Calibration of catalytic detectors for different gases 

and 

9 Misinterpreting warning signs 

This chapter gives guidelines for methane measurement in 

mines and tunnels. The emphasis is on the measurement 

procedure and the interpretation of the measurement rather 

than on the instrument itself. 

The failure to properly sample for methane is a major contributing factor to methane explosion 

risk. Sampling errors are most likely to occur at mines or tunnels where the presence of methane 

is not suspected or during nonroutine tasks at mines or tunnels known to have gas. 

More specific information on methane sampling at continuous miner sections is in Chapter 3. 

Chapters 4 and 5 discuss sampling at longwall sections, and Chapter 14 discusses sampling at 

tunnels. 

METHANE DETECTORS FOR MINING 

Many models of gas detectors are available to measure methane concentrations, as well as most 

of the other contaminant gases found in mines and tunnels. An example is the iTX Multi-Gas 

Monitor, a portable gas detector available from Industrial Scientific Corp., Oakdale, PA. This 

handheld instrument measures several gases simultaneously. The cost (2004) ranges from 

$1,300 to $2,200, depending on the number of gases measured. Similar instruments are available 

from other manufacturers. 

Most methane detectors used in mining use a catalytic heat of combustion sensor to detect 

methane and other combustible gases. These have been proven through many years of reliable 

operation. For detection of methane, proper operation of catalytic heat of combustion sensors 
 

28 

requires both a methane concentration below 8% and an oxygen content above 10%2— 

requirements that are usually satisfied in mining applications.3 

 

2

The percentages specified in this chapter are percentages by volume. 3

When the methane concentration is over 8% or when the oxygen concentration is under 10%, the sensor response 

declines. As a result, in these circumstances the methane concentration indicated by the instrument will be less, 

possibly much less, than the true methane concentration. For more specifics on operating conditions, check the 

documentation that accompanies the instrument. 

Some methane detectors measure the methane concentration by using infrared absorption as an 

operating method. These detectors (infrared analyzers) can measure accurately without oxygen 

and in a concentration range up to 100 vol % of methane. However, water vapor and dust can 

cause operating difficulties. In some mines [Kim 1973], the methane may be accompanied by 

ethane, which can produce an exaggerated infrared detector response. 

A list of approved gas detectors and gas monitors 

for U.S. mines is available from MSHA’s Approval 

and Certification Center, Triadelphia, WV. 

Based on how they are certified by the Mine Safety and Health Administration (MSHA), 

methane detectors used in mining fall into two categories: portable methane detectors and 

machine-mounted methane monitors. Portable detectors are designed to be hand-carried, 

so measurements can be made at any location. They are approved by MSHA under 30 CFR4 22. 

4

Code of Federal Regulations. See CFR in references. 

Among other requirements, “indicating detectors” must give indications of gas at 0.25% methane 

and must have an accuracy of at least 20% over most of the applicable range. 

Machine-mounted methane monitors are mounted on certain types of mining machinery and 

operate continuously. These monitors are certified under 30 CFR 27, which has different 

requirements than the Part 22 used for portable detectors.5 

5

Thus the distinction between “detectors” and “monitors.” 

The Part 27 requirements include 

a design that prevents the mining equipment from operating unless the methane monitoring 

system is functioning, a warning device that activates when the methane concentration is above 

1.0%–1.5%, and a means to shut off power6

6

Either electrical or diesel power. 

to the equipment when the methane concentration is 

2.0% and above.7 

7

For rapid-excavation machines in tunnels, the Occupational Safety and Health Administration (OSHA) requires 

electrical power to be shut off at 1.0% and above. See Chapter 14. 

MSHA certification requirements for Part 27 monitors are 

different from the certification requirements of Part 22 

detectors. Because of this, Part 27 monitors cannot be used 

for tasks requiring the use of Part 22 detectors (such as the 

20-min gas check task). 
USING PORTABLE METHANE DETECTORS 

In making measurements, consideration must always be given to how methane is released and 

diluted to safe levels. Methane entering a mine or tunnel often enters as a localized source at 

high concentration. An example is shown in Figure 2–1, which depicts a cloud of methane being 

diluted into a moving air stream. As shown in the figure, methane enters through a crack in the 

rock. If no air enters the crack, the methane concentration in the crack can be close to 100%. 

However, as the methane emerges from the crack, it progressively mixes with and is diluted by 

the ventilation air. Suppose this progressive dilution reduces the concentration from 100% to 

1%, as shown in Figure 2–1. In this case, the instrument reading depends highly on the location 

of the measurement—a critical concern if one intends to use the reading to assess whether a 

29 

Taking a gas reading with a portable methane detector is a simple matter. Where to measure and 

how to interpret the reading is not simple. For this reason, the many key points on gas measuring 

in both “accessible” and “restricted” spaces must be addressed. 

Methane measurements in accessible spaces. Accessible spaces are those that can be readily 

entered by a person making a methane measurement with a handheld portable methane detector. 

In accessible spaces, most methane measurements should be made as follows: 

•	 Close to the methane source, where higher concentrations are more likely to be 

encountered. 

•	 Close to the mine roof, where higher concentrations are more likely to be encountered. 

•	 In regions where the dilution of methane is impaired, i.e., those that are poorly ventilated 

and those where air movement is blocked by equipment. 

•	 While cutting is underway, because the methane release rate is higher as coal or rock is 

broken and the mining machinery advances. 

Fortunately, most places in mines where methane is to be 

measured are relatively accessible, i.e., the person making the 

measurement can easily reach the chosen location. The issue 

is how to choose the best location for measurement. 

hazard exists. This problem 

is handled by requiring 

methane measurements at a 

distance of not less than 

12 in from the roof, face, 

ribs, and floor. If there is 

enough gas coming from the 

crack (or other source) to 

exceed statutory limits at a 

12-in distance, then a hazardous condition exists. 

Figure 2–1.—Depiction of methane being diluted into a moving air stream. 
30 

In some cases, measurements must be taken at distances less than 12 in. For example, methane 

is lighter than air, so methane emerging from the roof can form a high-concentration layer along 

the roof of the mine (or crown of the tunnel). The thickness of such layers can be less than 12 in. 

Methane layers8

8

For more information on methane layers, see Chapters 1 and 11. 

are more likely to form if the ventilation air velocity measured at the roof is 

100 ft/min or less [Raine 1960], if the roof has cavities [Vinson et al. 1978], or if the roof has 

drillholes that serve as emission sources. Therefore, the degree of hazard resulting from a highconcentration layer of gas must be assessed from measurements of the size of the layer, as well 

as the location and size of the source. This is why gas readings must be done by a qualified, 

competent person, as prescribed by MSHA and OSHA regulations. 

Methane measurements in restricted spaces. For the purpose of this discussion, a restricted9 

9

This chapter uses the term “restricted space” because it is a more appropriate distinction for mining applications. 

The more common “confined space” term is not used because the whole mine can be regarded as a confined space. 

Further, none of these special terms do a good job of describing measurements at coal mine gobs, a specialized topic 

not considered in this chapter. 

space is one that cannot be readily entered to make a methane measurement. Examples of 

restricted spaces are a mine shaft that has been capped or a mine entry that has been closed to 

travel because of hazardous roof. The lack of accessibility often restricts both the ventilation air 

and the opportunity to make a convenient methane measurement.10 

10For example, a restricted-space methane explosion occurred in January 2003, killing three workers during a shaftsinking operation at a coal mine near Cameron, WV. A 7-ft-high “water ring” space had been excavated back into
 

the strata behind the shaft wall. These water rings facilitate water drainage from the exterior side of the shaft lining. 

Prior to pouring the concrete shaft lining, the water ring space was sealed off with corrugated sheets of steel to prevent it from being filled with concrete. After the concrete had set, the three miners were killed while in the process 

of opening an access door with an acetylene torch. An investigation by the West Virginia Office of Miner’s Health, 

Safety and Training concluded that “an adequate methane test was not made” [Mills 2003]. 

When making methane measurements in restricted spaces, simply making a measurement at the 

entrance of the restricted space is not adequate. The measurements must be made deep within 

and at the top of the restricted space, and also at every location within the restricted space where 

an ignition source (such as sparks from a torch) may be present. 

Restricted-space measurements can be made in two ways. First, the methane detector can be 

equipped with a remote “sample draw” capability. These use a small pump or hand-squeezed 

bulb to pull the sample through an extension probe and pass it through the detector. Some 

methane detectors have an accessory sampling pump that attaches to the detector; others have a 

built-in pump. 

Second, the methane detector can be attached to a cradle at the end of a long handle, which is 

then extended into the restricted space. This permits a direct reading without aspiration, 

provided that the instrument has a large LED readout that can be read from a distance.11 

11See the sections on methane detection in the continuous miner chapter (Chapter 3). 

Remember that gas detectors only sample the air 

that passes through the instrument. 

 
31 

Out-of-range gas concentrations in restricted spaces.12 

 

12This also applies to methane layers. 

Because some restricted spaces have 

little ventilation, the gas concentrations in these spaces may fall outside of the accurate operating 

range of catalytic heat of combustion sensors. For accurate operation of these sensors, the concentration of methane must be below 8% and the concentration of oxygen must be above 10%. 

When measuring methane concentrations above 8%, instruments with catalytic heat of combustion sensors can act in a way that is misleading, responding with a rapid upscale reading followed by a declining or erratic reading13 [CSA 1984]. 

13Some instruments will report this as an out-of-range condition. For more information, consult the operating 

instructions for the instrument. 

Such instrument behavior should be a 

tipoff that very high, potentially explosive methane levels may be present. 

Restricted spaces may also lack the 10% oxygen level necessary to ensure the proper operation 

of catalytic methane detectors. For example, the gas in exploration boreholes often contains little 

to no oxygen. In such circumstances, if the instrument being used has a second sensor to 

measure the oxygen level, an oxygen concentration less than the required 10% will be indicated, 

thereby alerting the user that the methane reading may be incorrect. However, even if the 

oxygen concentration is less than 10%, valid methane measurements are possible with other 

kinds of methane detectors. One approach to sampling low-oxygen atmospheres is to use a 

methane detector that operates by infrared absorption. Another approach is to use a catalytic 

methane detector that provides dilution sampling. The term “dilution sampling” refers to adding 

a controlled quantity of ambient air to the sample in order to raise the oxygen content of the 

sample.14

14Methane gas mixtures with 10% oxygen or less are not combustible, but may become so when mixed with more 

air. See Chapter 10 on using inert gas to prevent highwall methane explosions. 

 For example, if 1 L of sample gas is added to 1 L of ambient air, the oxygen level of 

the mixture will be adequate to operate a catalytic methane detector, and the true concentration 

of methane may be obtained by multiplying the detector methane reading by a factor of two. 

The bump test. It is a good idea to perform a quick “bump test” on every portable methane 

detector to ensure that it is working properly. Before every shift, briefly expose the portable 

detector to a known concentration of methane gas high enough to set off the methane alarm. 

Note the reading to ensure that it is correct. A bump test is not a calibration, but a quick way to 

ensure that the most important functions of the instrument are intact.15 

15A similar test is described by the Canadian Standards Association [CSA 1984]. 

USING MACHINE-MOUNTED METHANE MONITORS 

The disadvantage of portable handheld detectors is that a peak emission can be missed because 

readings at the appropriate locations are only taken at infrequent intervals. By contrast, machinemounted monitors operate continuously and can identify emission peaks and automatically shut 

off electrical equipment when the methane level is excessive. 

Machine-mounted methane monitors are usually mounted on mining and tunnel-boring 

machines. They are designed to have their readout display separated from the sensing head so 
 

 

32 

that the readout is visible to the machine operator and the sensing head is placed in a location 

where methane is most likely to accumulate. 

The usefulness of machine-mounted monitors depends on 

three critical factors: placement of the sensing head in a 

location where methane accumulates, the response time of 

the monitor, and whether or not the sensor head is covered 

by a heavy layer of dust or debris. 

Placement of the sensing head where methane accumulates. Proper placement of monitor 

sensing heads is crucial to the reliable detection of methane levels. Figure 2–2 shows a typical 

methane profile map measured from experiments at a full-scale simulated continuous miner face 

[Wallhagen 1977]. A striking feature of such profile maps is the steep gradient in the methane 

concentration along the length of the machine. Thus, a distance of a foot or two forward or 

backward in the location of the sensing head will greatly change the indicated methane level. 

In the instance depicted, the sensing head should be as far forward as possible to measure higher 

methane levels.16

16A typical sensing head location on continuous miners is on the side of the cutter head boom, a foot or two behind 

the cutter head. 

 Inevitably, some tradeoffs are involved in picking the location, for a sensor 

head located too far forward will quickly become damaged or clogged with dust. 

 Figure 2–2.—Methane profile map from a simulated continuous miner face. 

Response time of the sensor head. It is important that methane monitors have a short response 

time because the methane concentration can 

change quickly. With a short response time, the 

indicated concentration does not lag too far 

behind the true concentration. Figure 2–3 shows 

a recorder chart from a machine-mounted monitor 

at a coal mine working face [Kissell et al. 1974]. 

The peaks correspond to the cutting cycle of the 

mining machine, with the methane concentration 

spiking as the machine cuts into the coal. 

A methane monitor with a short response time 

will follow the spikes, giving warnings at the 

appropriate time. Taylor et al. [2004] reported on 

the response time of methane monitors in a test 

chamber designed to simulate mining conditions, 

while Taylor et al. [2002] reported on the 

response time using calibration caps supplied with 

the instrument. 
 Figure 2–3.—Recorder chart from a machine-mounted methane 

monitor. 

 

 

33 

Dust-clogged sensor heads. 

Machine-mounted methane 

monitors are usually in locations where dust quickly 

accumulates on the sensor 

heads, so some care must be 

directed toward preventing the 

sensor heads from getting 

clogged by heavy dust accumulations or covered with 

debris. Whether a sensor head 

is clogged can be assessed 

during calibration by noting the response of the monitor. All of the monitor manufacturers provide calibration caps that cover the sensing head and allow it to be flooded with calibration gas. 

When the sensor head is flooded with calibration gas, the instrument reading should be at least 

90% of the calibration gas concentration and the response time should be similar to that obtained 

with a clean head.17 

 

17Using calibration caps supplied by the manufacturers, Taylor et al. [2002] measured the 90% response time of 

three models of methane monitors. They found that the response time depends on a host of extraneous factors, such 

as the calibration gas flow rate. If a calibration cap is used to assess monitor response time, the best approach is to 

note the response time of a clean monitor head and then look for corresponding changes as conditions change. 

A lower concentration reading indicates a monitor problem, possibly a dustclogged sensor head. For this reason, sensor heads in particularly dirty locations should be 

cleaned and calibrated more frequently. 

CALIBRATION OF CATALYTIC INSTRUMENTS FOR DIFFERENT GASES 

Methane detectors used in mining must be periodically calibrated with a known concentration of 

methane-air mixture that is injected into the instrument from a tank containing the gas mixture. 

However, combustible gases other than methane are sometimes encountered underground, and 

if the gas being sampled is different from the gas used to calibrate the instrument, the error in the 

instrument reading can be considerable. For instance, a tunnel being excavated under a leaking 

gasoline storage tank might contain gasoline vapor. Under these circumstances, a detector calibrated with a methane-air mixture will read low because higher molecular-weight gases (such as 

those in gasoline vapor) diffuse more slowly into the sensor element. As an example, suppose a 

methane-calibrated instrument is carried into a tunnel containing only gasoline vapor in the air, 

a In this circumstance, the 

actual gasoline vapor concentration in air is about 20% of the LEL—twice the indicated reading. 

The opposite effect on instrument error can also take place. If the gas detector is calibrated with 

a higher molecular-weight gas such as pentane, then carried into a tunnel containing only 

methane, and if it reads 10% of the lower explosive limit (LEL), the actual methane concentration is 5% of the LEL, i.e., half the indicated reading. 

nd the instrument reads 10% of the lower explosion limit (LEL).18

18Do not confuse % methane with % of the LEL. The LEL of a mixture of methane and air is 5% methane by 

volume. Thus, 5% methane by volume is said to be equivalent to 100% of the LEL. It follows that a concentration 

of 1% of methane by volume in air is 20% of the LEL. Other flammable gases have different LELs. For example, 

mixtures of propane and air have an LEL of 2.1% propane by volume; therefore 50% of the LEL is 1.05% by 

volume. 
34 

For this reason, when operators are encountering methane, they should calibrate for methane. 

On the other hand, if higher hydrocarbons are being encountered, operators should calibrate with 

a higher hydrocarbon, such as pentane or propane. 

Calibration-sampling correction value tables for a 

variety of combustible gases are readily available 

[Industrial Scientific Corp. 2004]. More information 

on the response of catalytic sensors to different 

gases is available from Firth et al. [1973]. 

MISINTERPRETING WARNING SIGNS 

It is not unusual to misinterpret a gas warning sign, especially in underground workings thought 

to have no gas. A primary reason is that the gas flow varies with the excavation rate. Suppose, 

for example, a tunnel-boring machine (TBM) begins to cut into an area of gassy ground, releasing methane into the ventilation air. The machine-mounted monitor on the TBM senses this gas 

and shuts it down. After spending some time tracking down the source of the shutdown and 

figuring out what to do, a worker begins to hunt for gas with a handheld detector. The worker 

hunting for gas cannot find much because the emission dropped when the TBM stopped. Thus, 

everyone concludes that the monitor on the TBM is not working properly. Given two instruments, one with bad news and the other with good news, the tendency is to believe the good 

news. However, when methane detection and monitoring instruments fail, they rarely give a 

false alarm or a false high reading; in other words, they rarely indicate gas when there is none. 

The usual failure mode is to not register gas that is present. Therefore, when any instrument 

registers gas, it is better to trust the reading and take appropriate precautions. 

Operators must be especially cautious when successive methane readings vary more than they 

normally do. When the airflow is low or when measurements are taken close to the source, the 

methane will not be well mixed into the air. This could lead to a high reading in one area with a 

low reading just a few feet away. This incomplete mixing can indicate that the ventilation air is 

deficient and that even higher concentrations of gas might be found nearby. 

REFERENCES 

CFR. Code of federal regulations. Washington, DC: U.S. Government Printing Office, Office of 

the Federal Register. 

CSA [1984]. Combustible gas detection instruments. Section 5.3.1, CSA standard C22.2 

No. 152–M1984 (product ID 2001201). Mississauga, Ontario: Canadian Standards Association. 

Firth JG, Jones A, Jones TA [1973]. The principles of the detection of flammable atmospheres 

by catalytic devices. Combust Flame 20:303–311. 
35 

Industrial Scientific Corp. [2004]. Gas detection made easy. P/N 1600–0029. Oakdale, PA: 

Industrial Scientific Corp. 

Kim AG [1973]. The composition of coalbed gas. Pittsburgh, PA: U.S. Department of the 

Interior, Bureau of Mines, RI 7762. NTIS No. PB221574. 

Kissell FN, Banfield JL, Dalzell RW, Zabetakis MG [1974]. Peak methane concentrations 

during coal mining: an analysis. Pittsburgh, PA: U.S. Department of the Interior, Bureau of 

Mines, RI 7885. NTIS No. PB233424. 

Mills B [2003]. Fatal shaft explosion report investigation, McElroy Coal Co., McElroy mine, 

Permit No. U–83–33, Central Cambria Drilling Co., contractor No. C–618, January 22, 2003. 

Fairmont, WV: West Virginia Office of Miner’s Health, Safety and Training. 

Raine EJ [1960]. Layering of firedamp in longwall working. Trans Inst Min Eng (U.K.) 

119(10):579–597. 

Taylor CD, Chilton JE, Mal T [2002]. Evaluating performance characteristics of machinemounted methane monitors by measuring response time. In: De Souza E, ed. Proceedings of the 

North American/Ninth U.S. Mine Ventilation Symposium (Kingston, Ontario, Canada). Lisse, 

Netherlands: Balkema, pp. 315–321. 

Taylor CD, Chilton JE, Zimmer JA [2004]. Use of a test box to measure response times for 

machine-mounted monitors. In: Ganguli R, Bandopadhyay S, eds. Mine ventilation: Proceedings of the 10th U.S./North American Mine Ventilation Symposium (Anchorage, AK, May 16– 

19, 2004). Leiden, Netherlands: Balkema, pp. 177–182. 

Vinson RP, Thimons ED, Kissell FN [1978]. Methane accumulations in coal mine roof cavities. 

Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8267. NTIS No. 

PB277919. 

Wallhagen RE [1977]. Development of optimized diffuser and spray fan systems for coal mine 

face ventilation. Waltham, MA: Foster-Miller Associates, Inc. U.S. Bureau of Mines contract 

No. H0230023. NTIS No. PB277987/AS. 
37 

CHAPTER 3.–METHANE CONTROL AT CONTINUOUS MINER SECTIONS 

By Fred N. Kissell, Ph.D.,1

 

1

Research physical scientist (retired). 

 Charles D. Taylor,2

2

Industrial hygienist. 

and Gerrit V. R. Goodman, Ph.D.3 

3

Mining engineer. 

Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, Pittsburgh, PA. 

In This Chapter 

9 Methane emission peaks 

9 Exhaust line curtain or duct 

9 The spray fan system 

9 Dust scrubbers with blowing ventilation 

9 Dust scrubbers with exhaust ventilation 

9 The ventilation of abnormally gassy faces 

9 Methane detection at continuous miner faces 

9 Ventilation and methane detection at bolter faces

 and 

9 Reducing frictional ignitions 

This chapter gives guidelines for preventing methane gas explosions at continuous miner sections in coal mines, both at continuous miners and at roof bolters. The need to control peak 

methane emissions is particularly stressed. Emphasis is also placed on ventilation principles, 

monitoring for gas, and reducing frictional ignitions. 

METHANE EMISSION PEAKS 

Methane emission from the coal at continuous miner faces varies considerably. Plotted on a 

chart, methane emissions consist of a series of peaks and valleys corresponding to the cutting 

cycle of the mining 

machine, with the 

methane concentration 

spiking as the machine 

cuts into the coal (Figure 

3–1) [Kissell et al. 1974]. 

These methane peaks can 

be substantial. For this 

reason, efforts to safely 

dilute the methane must 

Figure 3–1.—Recorder chart from a machine-mounted methane monitor. focus on the level of the 
 

38 

peaks, not the overall 

methane level. Figure 3–2 

shows peak average4 

4

This is the average of the emission peak values. 

 Figure 3–2.—Peak average methane emissions for several U.S. 

coalbeds. 

methane emissions in several 

U.S. coalbeds [Haney et al. 

1983]. Except for one 

coalbed at 1,300 ft depth, 

peak average coalbed emissions range from 4 to 33 cfm, 

with an overall U.S. average 

of 17 cfm. Any face ventilation system must be able to 

safely handle gas flows of 

this magnitude. 

Although the values cited here are peak averages, the value 

for individual peaks can vary widely. For example, a study 

by Smith and Stoltz [1991] has shown a variation5

 of 46% in 

emission peak values. Similarly, there was a variation5

 of 

50% in methane dilution capacity. 

5

Strictly speaking, the coefficient of variation, which is the standard deviation divided by the mean. 

VENTILATION WITH EXHAUST LINE CURTAIN OR DUCT 

Prior to the development of improved face ventilation systems, most coal mine faces were 

ventilated with exhaust line curtain or ventilation duct. For this reason, exhaust line curtain or 

ventilation duct can serve as a baseline against which newer systems can be measured. Federal 

coal mine regulations [30 CFR6

6

Code of Federal Regulations. See CFR in references. 

75.330] mandate that exhaust systems have a maximum setback 

distance of 10 ft, i.e., the distance from the face being mined to the inlet of the line curtain or 

duct is 10 ft or less (Figure 3–3). However, if a mining machine starts a cutting cycle when the 

setback is 10 ft and subsequently advances another 10 ft, then the curtain or duct ends up at 20 ft 

if it has not been moved forward during the cutting cycle. 
 

 

Figure 3–3 [Ruggieri et al. 1985b] shows why the methane dilution capacity of exhaust line 

curtain is not high. Air flowing up the entry shortcuts to the mouth of the line curtain, leaving 

the off-curtain side of the face poorly ventilated and subject to a buildup of methane. To further 

demonstrate, Figure 3–4 shows the methane dilution capacity7

 f

7

Methane dilution capacity is the highest methane flow that the ventilation system can handle without exceeding 

a 1% methane concentration value anywhere in the face area. It is the best measure of how well a ventilation system
 

is working. See Haney et al. [1983]. 

39 

or exhaust line curtain8

8

The performance for exhaust duct is similar. 

at setbacks of 10 ft and 20 ft 

[Kissell and Wallhagen 

1976; Haney et al. 1982; 

Schultz et al. 1993]. 

A notable feature of 

Figure 3–4 is the reduced 

dilution capacity at the 

20-ft setback. This is why 

Mine Safety and Health 

Administration (MSHA) 

regulations specify a 

maximum setback of 

10 ft. 

Figure 3–3.—Ventilation setback distance. With exhaust line curtain, 

the off-curtain side can be poorly ventilated. 

 Figure 3–4.—Methane dilution capacity for exhaust line curtain at setbacks of 10 ft and 20 ft. 

For those mines using 

exhaust curtain and/or 

duct, an extensible system 

can be used to reduce the 

setback and possibly permit a deeper cut before 

place-changing. 

Extensible duct systems 

are fabricated using a 

duct section a few inches 

in diameter smaller than 

the main duct. This 

smaller section is inserted 

into the main duct at the 

inlet end and is slid 

forward as the miner 

advances. A typical 

extensible curtain system 

is fabricated by attaching 

a 20-ft section of brattice 

cloth to 20 ft of 1⁄2-indiam pipe. The pipe is 
40 

hung on J-hook assemblies, which are installed on the last two roof bolt plates next to the rib 

[Muldoon et al. 1982], and the pipe is slid forward as the miner advances. Urosek et al. [1988] 

described 11 extensible line curtain systems used in coal mines. 

Despite the development of extensible curtain systems, the continuing need for better methane 

control has led to improved face ventilation systems using spray fan and scrubber systems. 

VENTILATION WITH THE SPRAY FAN SYSTEM 

The spray fan system is an auxiliary ventilation system that makes use of the air-moving ability 

of water sprays. Moving droplets in the spray drag the surrounding air forward to create a considerable airflow, particularly when several sprays are arranged in a series as fans in a row. To 

install the spray fan system on a continuous miner (Figure 3–5), spray manifolds 1 and 2 are 

placed on the off-curtain side of the miner. These sprays move air forward to ventilate the offcurtain corner of the face (in Figure 3–5, the right corner). Spray manifold 3 moves air from 

right to left underneath the boom. Spray manifold 4 has 11 sprays angled 30° left to sweep air 

from right to left across the face and 1 spray on the right edge angled right to wet dust and clear 

gas from the right end of the cutter head. These manifolds are arranged for a working face that 

has the exhaust curtain on the left side. 

 Figure 3–5.—Spray fan system on a continuous miner. The numbers correspond to the spray manifolds. 

When the exhaust 

curtain is on the right 

side, another spray 

system, a mirror image 

of the one described, 

must be provided. 

In addition to the spray 

manifolds already 

described, additional 

dust suppression spray 

manifolds 5, 6, and 7 are 

directed at the ends of 

the cutting head and into 

the throat. These also 

help to keep the cutter 

head ends and the throat 

clear of gas, and they 

operate whether the leftor right-side curtain is

in use.
 

41 

The methane dilution performance of the spray fan is shown in 

Figure 3–6. Testing in the laboratory and underground by FosterMiller Associates [Ruggieri et al. 

1985b] at 13,000 cfm and at a 

water pressure of 100 and 150 psi 

yielded impressive methane dilution capacities of 115 and 135, 

respectively. Later testing by 

MSHA gave lower values, in part 

because of different design and 

lower water pressures.9

9

The effectiveness of spray fan systems depends on the water pressure. See Figure 3–11. 

Figure 3–6.—Methane dilution capacity of the spray fan system. 

 Still, the 

improvement with spray fans over 

the baseline 10-ft setback can be

substantial. 

An installation guide for the spray fan is available [Ruggieri et al. 1985a]. Close control of the 

water pressure is particularly important. If the water pressure is too low, little air will be moved. 

High water pressure will move more air, but if the pressure is too high, the spray fan can move 

more air than the line curtain, producing an airflow imbalance that raises the dust level at the 

operator cab. 

Good performance from spray fan systems requires that they be 

used according to established spray location and water pressure 

guidelines [Ruggieri et al. 1985a]. Following these guidelines 

ensures that the amount of air moved is adequate. 

DUST SCRUBBERS WITH BLOWING VENTILATION 

Dust scrubbers were first installed on continuous miners in the 1970s. Today, almost all new 

machines come equipped with them. Their popularity is due to improved methane dilution at 

large curtain setbacks, enabling the coal industry to achieve efficiency gains through extended 

cutting. Figure 3–7 shows a dust scrubber used in conjunction with blowing ventilation, the 

most common ventilation configuration. The scrubber collects dusty air at the boom, removes 

the dust, and discharges the clean air at the rear corner of the miner. This section covers the 

methane dilution effectiveness of scrubbers with blowing ventilation and the operating factors 

that impact the methane dilution. 
Figure 3–7.—Dust scrubber used in conjunction with blowing ventilation. 

42 

Methane dilution effectiveness. The first comprehensive underground 

study to measure the 

methane dilution 

effectiveness of a highvolume scrubber used in 

an extended cut with 

blowing ventilation was 

conducted by Haney 

et al. [1983]. With a 

6,700-cfm dust scrubber, they obtained a 

methane dilution 

capacity of 24.5 cfm using blowing line curtain located on the side of the entry opposite the 

scrubber exhaust. There was no deterioration in performance up to the largest line curtain setback tested (35 ft). 

A subsequent test by Halfinger [1984] gave similar results. The 7,000-cfm scrubber provided a 

methane dilution capacity averaging 26 cfm (see footnote 12). The scrubber system performed 

equally well at all line curtain setbacks tested (25, 35, and 50 ft). Halfinger also noted that the 

methane dilution effectiveness was independent of line curtain airflow between 3,500 and 6,000 

cfm, the lowest and highest line curtain airflows tested. 

Subsequent to the Halfinger study, MSHA conducted an extensive series of scrubber tests in 

mines across the United States [Zuchelli et al. 1993; Schultz et al. 1993; Stoltz and Snyder 1991; 

Snyder et al. 1993; Smith 

and Stoltz 1990; Denk 

et al. 1988; Snyder et al. 

1991; Dupree et al. 1993; 

Mott and Chuhta 1991; 

Denk et al. 1989]. 

Methane dilution results 

(Figure 3–8) indicated that 

methane dilution capacity 

was roughly related to 

scrubber quantity. For 

scrubbers over 4,000 cfm 

in coal heights 60 in or 

more, the average methane 

dilution capacity ranged 

from 28 to 44 cfm; for 

scrubbers over 4,000 cfm 

Figure 3–8.—Methane dilution capacity in coal under 60 in, the from MSHA scrubber tests. 
 

 

43 

average methane dilution 

capacity ranged from 16 

to 27 cfm.10 

10Bear in mind that these figures are only averages. On a cut-to-cut basis, peak methane values vary widely. For 

example, Haney et al. [1983] found that the methane dilution capacity had a coefficient of variation of 55%. Similar 

variability has been found by Smith and Stoltz [1991]. 

With regard 

to the effect of line curtain 

setback, the results 

mirrored those of Haney 

and those of Halfinger. 

The methane dilution 

capacity did not decline at 

the largest line curtain 

setbacks measured, 

typically up to 40 ft. 

An important aspect of scrubber system effectiveness is how well the box is being ventilated 

while the slab cut is being made. An investigation by Thimons et al. [1999] (Figure 3–9) 

indicated that with the blowing line curtain at 50 ft and the continuous miner at the start of a 40­

ft slab, the amount of fresh air reaching the face of the box was 400–600 cfm. However, as the 

miner advanced, more air reached the face of the box. When half of the slab was cut—a 20-ft 

advance—the amount of air reaching the face of the box was 2.5 times higher. Bringing the 

curtain forward also helped. For example, moving the curtain from 50 ft to 40 ft increased the 

airflow by a factor of 1.6 to 2.0 depending on test conditions. 

Figure 3–9.—Test configuration used by Thimons et al. [1999]. 

Operating factors that impact methane dilution. Many factors impact the ability of scrubbers 

to dilute methane safely. Some are related to the original design, others to the quality of 

maintenance. 

It is important to have an understanding of the operating 

factors that impact scrubber methane dilution: air quantity, 

water sprays, exhaust direction, clogging, and turning 

crosscuts. 

• Scrubber air quantity. Taylor et al. [1997] conducted studies in a full-scale surface test 

gallery to assess the impact of changing the scrubber air quantity and intake air quantity. On 

average, raising the scrubber air volume from 6,000 to 14,000 cfm produced a modest 23% 

decrease in methane concentration. The greatest decrease in methane concentration was at an 

intake airflow of 6,000 cfm, where raising the scrubber volume from 6,000 to 14,000 reduced 

methane levels by 38%. These results generally mirror those of MSHA’s scrubber tests shown in 

Figure 3–8, indicating improved methane dilution at higher scrubber airflows. 

• Water sprays. The impact of water sprays on scrubber ventilation effectiveness has been studied by Volkwein and Wellman [1989] and Taylor and Zimmer [2001]. Volkwein and Wellman 
 

44 

found that effective scrubber operation depends on the air movement generated by the dustsuppression water sprays. Turning off the spray system doubled the methane level.11 

11This result is not surprising. Wallhagen [1977] found the same effect with a conventional exhaust ventilation 

system with conventional sprays and no scrubber. 

Volkwein and Wellman also tested a directional spray system (similar in concept to the spray fan 

system described above) to help direct the air into the single inlet scrubber they were testing. 

Switching from a conventional spray system to a directional spray system yielded a 23% reduction in the methane level. Taylor and Zimmer saw no benefit from directional sprays because 

they were testing a dual-inlet scrubber system. 

• Exhaust direction. Taylor and Zimmer [2001] conducted tests to assess the impact of changing 

the scrubber exhaust toward or away from the blowing line curtain. As might be expected, directing the exhaust toward the blowing curtain interfered with the air stream from the curtain and 

gave the highest methane levels. By comparison, directing the exhaust straight back lowered 

methane levels at the face by 30%. Directing the exhaust toward the return-side rib lowered 

methane an additional 25%, for a total decrease of 55%. 

When a scrubber is used in conjunction with blowing 

ventilation, it is important that the blowing curtain (or duct) 

be on the side of the entry opposite the scrubber exhaust 

and that the exhaust be directed at the return-side rib. 

• Clogging. Clogging of the flooded-bed filter panel or the scrubber ductwork will seriously 

inhibit the methane dilution capacity of scrubbers. Denk et al. [1988] conducted a study in an 

Alabama mine that measured the methane dilution impact of a clogged scrubber inlet. The 

scrubber being tested had a metal plate that restricted the airflow at one of the two inlets, and the 

methane dilution capacity of the system was 28.5 cfm. When the metal plate was removed, 

subsequent testing showed that the methane dilution capacity had risen to 39.3 cfm, a 38% 

improvement. 

Clogging from coal particulate can be very rapid. For example, 

Campbell and Dupree [1991] noted a scrubber air decrease of 

23% after just one 30-ft cut of coal. Schultz and Fields [1999] 

noted that some scrubbers lose as much as one-third of their 

airflow after just one cut. 

Schultz and Fields [1999] reported on a method used by one mine operator to block large pieces 

of coal from entering the scrubber inlets under the boom. The mine had installed a flap of conveyor belt about 8 in inby each inlet, and the flaps extended downward about 8 in. The flaps 

forced the air to make an extra turn before entering the inlet, blocking the larger particles flying 

from the cutting drum. These flaps worked so well that the scrubber lost only 10% of its airflow 

capacity after an entire shift of operation. 
 

45 

When a dust scrubber clogs, its air quantity declines. Taylor et al. [1995] investigated ways to 

alert the miner operator to a clogging problem. The most effective was to monitor the fan motor 

current, since an air quantity decline resulting from clogging will lower the fan motor current. 

• Turning crosscuts. Using a small-scale model, Tien [1989] assessed the ventilation provided 

by scrubbers whenever crosscuts were being mined. Results showed that keeping the line curtain 

as close as practical to the rear of the miner is essential for controlling both respirable dust and 

methane. 

• Less critical operating factors. Other operating factors, once thought to be important, have 

turned out to be less critical. For example, Taylor et al. [1997] found that changing the line 

curtain airflow in the range between 6,000 and 14,000 cfm does not change the average face 

methane concentration. Also, methane levels do not increase when the line curtain airflow is less 

than the scrubber airflow (6,000-cfm line curtain versus 14,000-cfm scrubber), a situation that 

leads to a high amount of recirculation. Many years earlier, a study by Kissell and Bielicki 

[1975] led to a conclusion that recirculation per se was not harmful, as long as a sufficient 

quantity of fresh air was provided by the line curtain. 

DUST SCRUBBERS WITH EXHAUST VENTILATION 

Dust scrubbers have also been used with exhaust ventilation. However, the major drawback to 

using scrubbers with exhaust ventilation is the need to ventilate the empty headings that have 

been mined out, but not yet bolted. The jet from a blowing curtain can provide some minimal 

ventilation level, but an exhaust curtain may not [Luxner 1969]. 

Haney et al. [1983] conducted the first tests with scrubbers and exhaust curtain, obtaining a 

methane dilution capacity of 33.4 cfm using a 6,700-cfm dust scrubber. A subsequent study 

using a full-scale model [Taylor et al. 1996] gave methane dilution capacities ranging from 22 to 

58 cfm, depending on 

airflow.12

12This study gave average methane concentrations, whereas the methane dilution capacity is normally calculated 

based on the highest measured concentration (see footnote 7). To obtain the methane dilution capacity values stated, 

we assumed that the highest measured concentration would be 30% greater than the average concentration. 

 Both of these 

studies employed an 

exhaust line curtain 

located on the same side 

of the entry, so the air jet 

from the scrubber fed 

directly into the line 

curtain as shown in Figure 

3–10. 

Figure 3–10.—Dust scrubber used with exhaust ventilation. 
 

46 

When the scrubber exhaust is not on the same side of the entry as the exhaust curtain, methane 

dilution suffers. For example, Stoltz et al. [1991] conducted a scrubber ventilation study at a 

mine that had an exhaust curtain on the opposite side of the entry from the scrubber exhaust. 

The measured methane dilution capacity was only 13 cfm. 

Another exhaust line curtain requirement is that the mouth of the curtain be outby the scrubber 

exhaust (Figure 3–10). Jayaraman et al. [1990] conducted a series of tests that included a line 

curtain setback of 10 ft, which was about 10 ft inby the scrubber exhaust. The methane dilution 

was one-half to one-fourth of that obtained with a curtain setback of 30 ft. 

When using a scrubber in conjunction with 

exhaust ventilation, keep the curtain on the 

same side of the entry as the scrubber 

exhaust and keep the mouth of the curtain 

outby the scrubber exhaust (as shown in 

Figure 3–10). 

THE VENTILATION OF ABNORMALLY GASSY FACES 

Some continuous miner faces have abnormally high methane emissions, and it is helpful to 

explore the various alternatives a mine operator might have to safely ventilate such faces. 

Although mine tests have not been conducted, a high-pressure spray fan and a high-volume 

scrubber have each achieved a methane dilution capacity on the order of 100 cfm in laboratory 

tests, provided that line curtain air quantities were large13 and curtain setback distances were 

modest. 

13Methods to increase airflow by decreasing line curtain and check curtain leakage have been described by Muldoon 

et al. [1982]. A high-capacity duct system designed to deliver 45,000–50,000 cfm has been described by Hagood 

[1980]. 

Abnormally gassy faces may be ventilated with 

diffuser fans, high-pressure spray fans, or highvolume scrubbers. 

However, degasification with horizontal or 

vertical boreholes is necessary if the section 

emits over 300 cfm of methane. For more on 

degasification, see Chapter 6. 

Diffuser fan. The diffuser fan is a small fan mounted on the continuous miner that directs an air 

jet at the working face. Used in conjunction with exhaust line curtain, it was the primary means 

of ventilating gassy faces before the development of the spray fan. Wallhagen [1977] developed 

an optimized two-nozzle, 1,750-cfm fan that essentially induced all 9,000 cfm of the line curtain 
47 

air to the face, yielding a 

methane dilution capacity 

of roughly 65 cfm (see 

footnote 12). The ventilation setback (duct was 

used) was 10 ft. Some 

years later, Haney et al. 

[1982] tested a highvolume 5,000-cfm 

diffuser fan. With line 

curtain airflows of 9,000 

cfm and higher, the minimum methane dilution 

was 74 cfm for a 10-ft 

curtain setback and 

67 cfm for a 20-ft curtain 

setback. 

Although effective for diluting methane, diffuser fans were not popular in the past because they 

were noisy and kicked up dust. They might be more acceptable on today’s remote-control 

machines. Wallhagen [1977] gives tips on how to design a diffuser fan system that is matched to 

a line curtain airflow of 15,000–20,000 cfm. 

High-pressure spray fan. It was mentioned earlier that a spray fan tested by Ruggieri et al. 

[1985b] gave methane dilution capacities of 115 and 135 at pressures of 100 and 150 psi, respectively, but that subsequent MSHA tests gave much lower values. To get full performance from a 

spray fan system, the system must be installed and operated according to established guidelines 

[Ruggieri et al. 1985a] and with line curtain airflows of 15,000 cfm or more. High performance 

from spray fan systems also requires that they be operated at high water pressure, as shown in 

Figure 3–11 [Wallhagen 1977]. 

 Figure 3–11.—Graph showing that performance of a spray fan depends 

on water pressure. 

High-volume scrubber. In a full-scale laboratory test facility, Taylor et al. [1996] tested a 

14,000-cfm scrubber in conjunction with a 14,000-cfm blowing line curtain. At a 25-ft curtain 

setback, the methane dilution capacity was 111 cfm (see footnote 12). At a 35-ft setback, the 

methane dilution capacity decreased to 68 cfm. 

When the 14,000-cfm scrubber was used in conjunction with a 14,000-cfm exhausting line curtain, the methane dilution capacity was 58 and 53 cfm at 25- and 35-ft setbacks, respectively. 

METHANE DETECTION AT CONTINUOUS MINER FACES 

Two methods of methane detection are used at continuous miner faces: intermittent sampling 

with portable methane detectors and continuous monitoring with machine-mounted methane 

monitors. These are both required by MSHA regulations [30 CFR 75]. 
 

 

48 

Methane monitors are usually mounted on the side of the cutting boom of the continuous miner. 

The best practice is to select the side that normally sees the highest concentrations.14

14For additional information on methane monitor placement, see Chapter 2 on sampling. 

 For exhaust 

ventilation systems, including spray fan and scrubber systems, this is normally the same side of 

the entry where the exhaust curtain (or duct) is located. For blowing ventilation used with 

scrubber systems, it is normally the opposite side of the entry from the blowing curtain (or duct) 

[Zuchelli et al. 1993; Schultz et al. 1993; Stoltz and Snyder 1991; Snyder et al. 1993; Smith and 

Stoltz 1990; Denk et al. 1988; Snyder et al. 1991; Dupree et al. 1993; Mott and Chuhta 1991; 

Denk et al. 1989]. 

The required intermittent sampling is a gas check every 20 min with a portable methane detector. 

A common practice is to attach the portable methane detector to the end of an extensible pole, 

then to extend the pole out over the continuous miner as far forward as possible. However, this 

is an awkward procedure that requires a long pole, a methane detector with a large readout, and 

good eyesight. Another approach, used at deep-cut faces, is to tram out the miner and attach the 

methane detector to the head using a magnet. The miner is then trammed back in and the 

detector read. 

VENTILATION AND METHANE DETECTION AT BOLTER FACES 

On faces that are being bolted, the line curtain or ventilation duct must always be extended to the 

last row of bolts and moved forward when a new row of bolts is installed. For particularly gassy 

faces, it may be necessary to use an extensible curtain or duct system [Muldoon et al. 1982]. 

With regard to methane detection, it has always been difficult to make a methane concentration 

measurement at the face while, at the same time, remaining safely under bolted roof. Extendedcut mining methods have increased this difficulty because the freshly cut face can extend 40 ft or 

more beyond the last row of bolts. To deal with this problem, MSHA has published a new rule 

[68 Fed. Reg.15 40132 (2003)].

15See Fed. Reg. in references. 

 This new rule, based on the work of Taylor et al. [1999], allows 

methane tests to be made at intervals not exceeding 20 min by sweeping a 16-ft probe inby the 

last permanently supported roof, provided that a methane monitor is also mounted on the roofbolting machine.16 

16See 68 Fed. Reg. 40132 [2003] for requirements on methane testing.
 

The methane monitor must be capable of giving a warning signal at 1.0% 

methane and capable of automatically deenergizing the machine at 2.0% methane, or if the 

monitor is not working properly. 

Typical ignitions at roof bolter faces have been discussed by Urosek and Francart [1999]. 

REDUCING FRICTIONAL IGNITIONS 

Up to this point, the emphasis of this chapter has been solely on ventilation methods and 

monitoring for gas. However, the chance of a methane ignition may be further reduced by 
49 

dealing directly with the ignition source. When a continuous miner cutter bit strikes rock, abrasion from the rock grinds down the rubbing surface of the bit, producing a glowing hot metal 

streak on the rock surface behind the bit. The metal streak is often hot enough to ignite methane, 

causing a so-called frictional ignition. 

At continuous miner faces, there are two approaches to lower the incidence of frictional ignitions. The first approach concerns the bit itself—providing a regular change-out schedule to 

replace worn bits, providing bits with a larger carbide tip to reduce wear, and possibly changing 

the bit attack angle or the type of bit. 

The second approach is to mount a water spray behind each bit, aiming the spray toward the 

location on the rock where the hot metal streak is expected. This anti-ignition back spray 

quenches the hot streak, reducing its temperature and the chance of a frictional ignition. 

Bit changes to reduce frictional ignitions. The most important action one can take to reduce 

frictional ignitions is to replace bits regularly, thus avoiding the formation of wear flats on the 

bits. Frictional ignition with a mining bit always involves a worn bit having a wear flat on the tip 

of the bit [Courtney 1990]. A small wear flat forms a small hot spot, which does not lead to an 

ignition, whereas a large wear flat forms a large hot spot that is more likely to cause an ignition. 

Also, mining bits consist of a steel shank with a tungsten carbide tip. The steel is more incendiary than the tungsten carbide tip, so if the tip is worn off and the steel shank exposed, the chance 

of an ignition is much greater. As an example, Figure 3–12 shows the results of a test in which a 

cutter bit was used to cut a sandstone block in the presence of an ignitable methane concentration. With the tungsten carbide 

tip in place, no ignitions were 

obtained even after 200 or more 

cuts. With the steel shank 

exposed, ignitions quickly 

began. With as little as 0.3-cm 

bit wear, fewer than 10 cuts 

were necessary to produce an 

ignition. 

 Figure 3–12.—Effect of bit wear on frictional ignition [Courtney 

1990]. 

Bits that wear more slowly can 

be changed less frequently. Bit 

wear is reduced by using bits 

that have larger carbide tips or 

by using bits that have a highly 

abrasion-resistant polycrystalline 

diamond layer on the rake face 

of the tip. 

Other methods to reduce frictional ignitions are to change the 

attack angle and tip angle of 

conical bits [Courtney 1990] and 
 

50 

to use radial bits instead of conical bits [Phillips 1996]. McNider et al. [1987] reported a 

decrease in frictional ignitions by using bits with larger carbide tips and by changing the bit 

attack angle.17 

17Changing the bit attack angle can also raise the cutting force, requiring a change in drum design. 

Anti-ignition back sprays. Anti-ignition back sprays, an effective method to reduce frictional 

ignitions, are discussed in the longwall chapter (Chapter 4). Bringing water to the cutter head 

on continuous miners has been an engineering challenge. However, in recent years, practical 

(if expensive) water seals for continuous miner heads have been developed. As a result, a few 

“wet-head” continuous miners equipped with anti-ignition back sprays have been installed in 

U.S. coal mines with a history of frictional ignition problems. Phillips [1997] has provided a 

status report on wet-head cutting drums. 

A thorough review of frictional ignitions in mines, 

including metal-to-metal ignitions and those from 

roof falls, is provided by Phillips [1996]. 

REFERENCES 

68 Fed. Reg. 40132 [2003]. Mine Safety and Health Administration, 30 CFR Part 75: improving 

and eliminating regulations, phase 5, miscellaneous technology improvements (methane testing); 

final rule. 

Campbell CD, Dupree WA [1991]. Evaluation of methane dilution capacity of a blowing face 

ventilation system using a water spray fan and dust scrubber. In: Proceedings of the Fifth U.S. 

Mine Ventilation Symposium (Morgantown, WV, June 3–5, 1991). 

CFR. Code of federal regulations. Washington DC: U.S. Government Printing Office, Office of 

the Federal Register. 

Courtney WG [1990]. Frictional ignition with coal mining bits. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 9251. NTIS No. PB91110072. 

Denk JM, Smith GE, Stoltz RT [1989]. Face ventilation investigation: Wabash mine, I.D. No. 

11–00877, AMAX Coal Company, Keensburg, Wabash County, Illinois, August 29, 1989­

September 1, 1989. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh Health Technology Center, Ventilation Division, Investigative Report 

No. P313–V218. 

Denk JM, Younger RK, Mott ML, Chuhta EJ [1988]. Face ventilation investigations: Mary Lee 

No. 1 mine, I.D. No. 01–00515–0, Drummond Company, Incorporated, Goodsprings, Walker 

County, Alabama, October 28-November 4, 1987, and October 18–26, 1988. Pittsburgh, PA: 
51 

U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh Health Technology Center, Ventilation Division, Investigative Report No. P297–V203. 

Dupree WA Jr., Schroeder ME, Baran JN [1993]. Face ventilation investigation: Island Creek 

Coal Company, Ohio No. 11 mine, I.D. No. 15–03178, Morganfield, Union County, Kentucky, 

May 25–27, 1993. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh Safety and Health Technology Center, Ventilation Division, Investigative 

Report No. P380–V283. 

Hagood [1980]. A high-capacity face ventilation system. In: Proceedings of the Second International Mine Ventilation Congress (Reno, NV, November 4–8, 1979). Littleton, CO: Society of 

Mining Engineers. 

Halfinger G Jr. [1984]. Feasibility of a machine-mounted scrubber system for ventilating coal 

mine working faces. Ingersoll-Rand Research, Inc. U.S. Bureau of Mines contract No. 

J0199080. NTIS No. PB85185742. 

Haney RA, Gigliotti SJ, Banfield, JL [1982]. Face ventilation systems performance in lowheight coal seams. In: Proceedings of the First Mine Ventilation Symposium (Tuscaloosa, AL, 

March 29–31, 1982). 

Haney RA, Smith GE, Denk JM [1983]. Evaluation of a face ventilation system using a 

machine-mounted dust collector. In: Proceedings of the Sixth Annual Mining Institute, 

(Tuscaloosa, AL, October 11–12, 1983). 

Jayaraman NI, Volkwein JC, Kissell FN [1990]. Update on continuous miner dust scrubber 

applications. Min Eng 42(3):281–284. 

Kissell FN, Bielicki RJ [1975]. Methane buildup hazards caused by dust scrubber recirculation 

at coal mine working faces: a preliminary estimate. Pittsburgh, PA: U.S. Department of the 

Interior, Bureau of Mines, RI 8015. NTIS No. PB240684. 

Kissell FN, Wallhagen RE [1976]. Some new approaches to improved ventilation of the working face. In: Proceedings of the NCA/BCR Coal Conference (Louisville, KY, October 19–21, 

1976), pp. 325–338. 

Kissell FN, Banfield JL, Dalzell RW, Zabetakis MG [1974]. Peak methane concentrations during coal mining: an analysis. Pittsburgh PA: U.S. Department of the Interior, Bureau of Mines, 

RI 7885. 

Luxner JV [1969]. Face ventilation in underground bituminous coal mines. Washington, DC: 

U.S. Department of the Interior, Bureau of Mines, RI 7223. 

McNider T, Grygiel E, Haynes J [1987]. Reducing frictional ignitions and improving bit life 

through novel pick and drum design. In: Proceedings of the Third Mine Ventilation Symposium 

(University Park, PA, October 12–14, 1987), pp. 119–125. 
52 

Mott ML, Chuhta EJ [1991]. Face ventilation investigation: Monterey No. 1 mine, ID No. 11– 

00726, Monterey Coal Company, Carlinville, Macoupin County, Illinois, January 15–17, 1991. 

Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh 

Health Technology Center, Ventilation Division, Investigative Report No. P350–V254. 

Muldoon T, Bartlett P, Schroeder W [1982]. Improving check curtains, line curtains, and extensible face ventilation systems. Waltham, MA: Foster-Miller Associates. U.S. Bureau of Mines 

contract No. J0100075. NTIS No. PB83216242. 

Phillips HR [1996]. Identify methods to reduce the risks of frictional ignition hazards. Final 

project report, project COL 226. Braamfontein, Republic of South Africa: Safety in Mines 

Research Advisory Committee (SIMRAC) (www.simrac.co.za). 

Phillips HR [1997]. To establish the current status of research, development, and operational 

experience of wet-head cutting drums for the prevention of frictional ignitions. Final project 

report, project COL 426. Braamfontein, Republic of South Africa: Safety in Mines Research 

Advisory Committee (SIMRAC) (www.simrac.co.za). 

Ruggieri SK, Babbitt C, Burnett J [1985a]. Improved spray fan system: installation guide. 

Waltham, MA: Foster-Miller Associates. U.S. Bureau of Mines contract No. J0113010. 

NTIS No. PB86165065. 

Ruggieri SK, Muldoon TL, Babbitt C, Lee E [1985b]. Improved diffuser and spray fan systems 

for ventilation of coal mine working faces. Waltham, MA: Foster-Miller Associates. U.S. 

Bureau of Mines contract No. J0113010. NTIS No. PB86168440. 

Schultz MJ, Fields KG [1999]. Dust control considerations for deep cut mining sections. SME 

preprint 99–163. Littleton, CO: Society for Mining, Metallurgy, and Exploration, Inc., 4 pp. 

Schultz MJ, Beiter DA, Watkins TR, Baran JN [1993]. Face ventilation investigation: Clark 

Elkhorn Coal Company, Ratliff mine No. 110, I.D. No. 15–16121, Rockhouse Creek, Pike 

County, Kentucky, September 21–24, 1993. Pittsburgh, PA: U.S. Department of Labor, Mine 

Safety and Health Administration, Pittsburgh Safety and Health Technology Center, Ventilation 

Division, Investigative Report No. P385–V286. 

Smith GE, Stoltz RT [1990]. Face ventilation investigation: Greenwich Collieries No. 2 mine, 

ID No. 36–02404, Greenwich Collieries Company, a division of Rochester and Pittsburgh Coal 

Company, Cookport, Indiana County, Pennsylvania, August 28–30, 1990. Pittsburgh, PA: U.S. 

Department of Labor, Mine Safety and Health Administration, Pittsburgh Health Technology 

Center, Ventilation Division, Investigative Report No. P342–V246. 

Smith G, Stoltz R [1991]. Deep-cut mining investigation: case study. In: Proceedings of the 

Fifth U.S. Mine Ventilation Symposium (Morgantown, WV, June 3–5, 1991). 
53 

Snyder MP, Schultz MJ, Baran JN [1991]. Face ventilation investigation: Margaret No. 11B, 

I.D. No. 36–08139, Keystone Coal Mining Corporation, Indiana, Indiana County, Pennsylvania, 

August 5–7, 1991. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh Safety and Health Technology Center, Ventilation Division, Investigative 

Report No. P356–V260. 

Snyder MP, Schultz MJ, Zuchelli DR [1993]. Face ventilation investigation: Peabody Coal 

Company, Camp No. 11 mine, ID No. 15–08357, Union County, Kentucky, May 11–13, 1993. 

Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh 

Safety and Health Technology Center, Ventilation Division, Investigative Report No. P382–
 

V285. 

Stoltz RT, Snyder MP [1991]. Face ventilation investigation: Buck Creek mine, ID No. 12–
 

02033, Laswell Coal Company Incorporated, Sullivan, Sullivan County, Indiana, January 8–10, 

1991. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, Pittsburgh Health Technology Center, Ventilation Division, Investigative Report No. P346–V250. 

Stoltz RT, Snyder MP, Baran J [1991]. Face ventilation investigation: Big Bottom mine, ID No. 

15–12618, Rawl Sales and Processing Company, McCarr, Pike County, Kentucky, October 22–
 

29, 1991. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, 

Pittsburgh Safety and Health Technology Center, Ventilation Division, Investigative Report 

No. P358–V262. 

Taylor CD, Zimmer JA [2001]. Effects of water sprays and scrubber exhaust on face methane 

concentrations. In: Proceedings of the Seventh International Mine Ventilation Congress 

(Krakow, Poland, June 17–22, 2001), pp. 465–470. 

Taylor CD, Kovscek PD, Thimons ED [1995]. Monitoring machine-mounted dust scrubber performance. In: Wala AM, ed. Proceedings of the Seventh U.S. Mine Ventilation Symposium
 

(Lexington, KY, June 5–7, 1995). Littleton, CO: Society of Mining Engineers, pp. 419–422. 

Taylor CD, Rider JP, Thimons ED [1996]. Changes in face methane concentrations using highcapacity scrubbers with exhausting and blowing ventilation. SME preprint 96–167. Littleton, 

CO: Society for Mining, Metallurgy, and Exploration, Inc. 

Taylor CD, Rider JP, Thimons ED [1997]. Impact of unbalanced intake and scrubber flows on 

face methane concentrations. In: Ramani RV, ed. Proceedings of the Sixth International Mine 

Ventilation Congress (Pittsburgh, PA, May 17–22, 1997). Littleton, CO: Society for Mining, 

Metallurgy, and Exploration, Inc, pp. 169–172. 

Taylor CD, Thimons ED, Zimmer JA [1999]. Comparison of methane concentrations at a simulated coal mine face during bolting. In: Tien JC, ed. Proceedings of the Eighth U.S. Mine Ventilation Symposium. Rolla, MO: University of Missouri-Rolla Press, pp. 171–178. 

Thimons ED, Taylor CD, Zimmer JA [1999]. Evaluating the ventilation of a 40-foot two-pass 

extended cut. Pittsburgh, PA: U.S. Department of Health and Human Services, Public Health 
54 

Service, Centers for Disease Control and Prevention, National Institute for Occupational Safety 

and Health, DHHS (NIOSH) Publication No. 99–138, RI 9648. 

Tien JC [1989]. Face ventilation during crosscut development in a room and pillar operation. 

In: Proceedings of the Fourth U.S. Mine Ventilation Symposium (Berkeley, CA, June 5–7, 

1989). 

Urosek JE, Francart WJ [1999]. Methane ignitions on roof bolters in underground coal mines. 

In: Tien JC, ed. Proceedings of the Eighth U.S. Mine Ventilation Symposium. Rolla, MO: 

University of Missouri-Rolla Press.
 

Urosek JE, Battistoni CS, Hazuza GG [1988]. Remote brattice face ventilation systems. In: 

Frantz RL, Ramani RV, eds. Respirable dust in the minerals industries: health effects, characterization, and control. University Park, PA: The Pennsylvania State University. 

Volkwein JC, Wellman TS [1989]. Impact of water sprays on scrubber ventilation effectiveness. 

Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 9259. NTIS No. 

PB90108820. 

Wallhagen RE [1977]. Development of optimized diffuser and spray fan systems for coal mine 

face ventilation. Waltham, MA: Foster-Miller Associates. U.S. Bureau of Mines contract No. 

H0230023. NTIS No. PB277987. 

Zuchelli D, Stoltz R, Baran J [1993]. Face ventilation investigation, Peabody Coal Company, 

Martwick mine, ID No. 15–14074, Central City, Muhlenberg County, Kentucky, January 26–28, 

1993. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, 

Pittsburgh Safety and Health Technology Center, Ventilation Division, Investigative Report 

No. P377–V280. 
 

55 

CHAPTER 4.—PREVENTING METHANE IGNITIONS 

AT LONGWALL FACES 

By Fred N. Kissell, Ph.D.1

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

and Andrew B. Cecala2 

2

Mining engineer, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA. 

In This Chapter 

9 Where methane is emitted at longwall faces 

9 Where methane accumulates at longwall faces 

9 Using the modified shearer-clearer to eliminate ventilation eddy zones 

9 Using a walkway curtain to reduce methane buildup during the headgate cutout 

9 Control of frictional ignitions 

and 

9 The best location for the methane monitor 

The methane released along a longwall face represents only 10%–20% of the total methane 

emitted from the entire longwall panel. Nevertheless, in very gassy coal seams, this methane 

released at the face can pose a problem because the shearer is a ready ignition source. 

Preventing methane ignitions at longwall faces requires four actions. The first is to provide 

better ventilation around the shearer to eliminate the ventilation eddy zones at the drums where 

methane builds up. These eddy zones are eliminated by mounting additional water sprays on the 

shearer to direct air into them. The second action to prevent methane ignitions is to install a 

water spray behind each cutter bit and regularly replace worn bits. Water sprays behind each 

cutter bit act to quench the hot metal streak that follows a worn bit when it strikes rock. The 

third is to ensure that no ventilation eddy zones are inadvertently created by poor placement of 

water sprays. The fourth is to ensure that the methane monitor on the shearer is in the best location to detect methane accumulations. 

ADDRESSING METHANE ACCUMULATIONS AT LONGWALL FACES3 

3

Methane accumulations in longwall gobs are addressed in Chapter 5. 

Cecala et al. [1985a, 1989] and Denk and Wirth [1991] studied methane emission and ventilation 

patterns at longwall faces to find where methane accumulations are most likely. Although not 

always the case, the major source of methane at longwall faces is usually the breakage of coal by 

the shearer. Stress-related fracturing of the coal seam at the face, called bumps or bounces, can 

cause the release of additional gas (Figure 4–1). 
 

Figure 4–1.—Increased methane liberation during coal bumps. 

56 

Methane accumulations around the 

shearer body. The 

quality of the ventilation around the 

shearer body impacts 

the accumulations of 

methane. For example, ventilation eddy 

zones around the 

shearer body are 

known to accumulate 

gas because the air 

exchange in and out 

of these zones is 

limited. Studies reported by Ruggieri 

et al. [1983] on a fullscale mockup of a 

longwall shearer face 

showed that the faceside area around both 

cutting drums and 

the entire area between the drums were 

less ventilated than 

other parts of the 

shearer. Further increasing the primary 

airflow or changing the cutting direction had little impact on improving the ventilation of these 

eddy zones. 

Ventilation at the shearer can be improved by using a “modified” shearer-clearer4

4

The original shearer-clearer system was designed to reduce dust. It is an arrangement of water sprays mounted on 

the body of a longwall shearer. The purpose of these sprays is to induce an air current over the shearer body, which 

serves to hold the dust cloud against the face, keeping the dust out of the operator’s walkway. 

system to 

bring more air into these eddy zones. The modified shearer-clearer is shown in Figure 4–2. 

It differs from the original shearer-clearer system by the addition of three water sprays on the 

return-side splitter arm (shown as A in Figure 4–2) and two sprays on the head-side corner of the 

shearer body (shown as B in Figure 4–2). 

Figure 4–2.—Modified shearer-clearer system. 

These sprays move air5

5

Many studies [Ruggieri et al. 1985] have established that water sprays mounted on mining machines can move air 

in the vicinity of the machine and that this air movement can benefit methane control. 

toward the face side of the 

shearer body and toward the return side of each shearer drum—eddy zone regions where 

methane accumulations are likely. These extra sprays raise the shearer water consumption by 

about 20 gpm. According to Cecala and Jayaraman [1994], the modified shearer-clearer system 

lowers methane concentrations at the shearer by 73%. 
 

57 

Cecala and Jayaraman [1994] have provided a detailed design and 

installation manual for the modified shearer-clearer system. It is 

available as a free pdf download from NTIS6 as PB95104873. 

6

National Technical Information Service, Springfield, VA. See www.ntis.gov. 

Methane accumulations 

during the headgate cutout. 

Ventilation at the shearer also 

suffers and the methane concentration rises as the shearer 

makes the headgate cutout 

(Figure 4–3). This rise in 

methane takes place because the 

air flowing down the headgate 

entry does not readily make the 

90° turn as it reaches the longwall face. Thus, a portion can 

divert to flow through the legs 

of the first 8–10 supports. 

Because of this air diversion, 

the amount of air flowing over 

the shearer is not sufficient to 

avoid methane buildup. This 

buildup can be prevented with a 

walkway curtain. 

 Figure 4–3.—Methane concentration at shearer during tail-to-head 

pass. 

The walkway curtain [Cecala 

et al. 1986], used to force more 

air over the shearer body during 

the headgate cutout, is shown in 

Figures 4–4 and 4–5. The curtain should be located across the 

walkway near support No. 3, 

extending from the support legs 

to the spill plate reaching from 

the roof to the floor. To be 

effective, it must be used in 

conjunction with a gob curtain, 

a dust control device already in 

use at most longwalls (Figure 

4–4). The walkway curtain 

raises the air velocity over the 

shearer by 23% at the headgate Figure 4–4.—Plan view of walkway and gob curtains (shearer not 

shown). 
 

 Figure 4–5.—Walkway curtain at support No. 3 forces air over 

the shearer. 

58 

corner and by 43% at support 

No. 4. These higher air velocities 

reduce the methane concentration 

by 60% during the headgate 

cutout. 

During the longwall cutting cycle, 

the walkway curtain need only be 

in position during the headgate 

cutout. Since it is not needed 

for the remaining 95% of the cutting cycle, it should be tied up to 

keep it out of the way. For the 

remaining 95% of the cutting 

cycle, the modified shearer-clearer 

is used. 

The modified shearer-clearer may need to be turned off during 

the headgate cutout. During testing of the combined systems, 

the high air velocity over the shearer created by the walkway 

curtain overpowered the shearer-clearer sprays and forced 

water mist over the tail-side shearer operator [Cecala and 

Jayaraman 1994]. 

Methane accumulations caused by inadvertent eddy zones. Certain ventilation practices lead 

to the creation of inadvertent eddy zones. Figures 4–6 and 4–7 illustrate two ventilation mistakes 

that can cause an accumulation of methane. Figure 4–6 shows an upwind-pointing venturi spray 

mounted on the headgate end of a shearer. The venturi spray creates an airflow opposing the 

main ventilation flow 

direction along the face, 

thus creating an eddy zone 

where methane builds up. 

Another type of eddy zone 

is shown in Figure 4–7, 

which depicts an L-shaped 

wing curtain used by some 

operators to control dust 

during the headgate cutout. Although convenient 

to use, it can allow the 

accumulation of methane. 

 Figure 4–6.—Upwind-pointing venturi spray creates eddy zone and 

methane buildup. 
Figure 4–7.—L-shaped wing curtain, which can cause methane buildup. 

59 

REDUCING FRICTIONAL IGNITIONS 

Aside from improving the ventilation to reduce methane accumulations in eddy zones, the 

chance of a methane ignition can be reduced by directly addressing the ignition source. When a 

shearer cutter bit strikes rock, abrasion from the rock grinds down the rubbing surface of the bit, 

producing a glowing hot metal streak on the rock surface behind the bit. The metal streak is 

often hot enough to ignite methane, causing a so-called frictional ignition. 

At longwalls, there are two methods to lower the incidence of frictional ignitions. The first 

method concerns the bit itself—providing a regular change-out schedule to replace worn bits, 

providing bits with a larger carbide tip to reduce wear, and possibly changing the bit attack angle 

or the type of bit. These topics are covered in the continuous miner chapter (Chapter 3). 

The second method is to mount a water spray behind each bit, aiming the spray toward the 

location on the rock where the hot metal streak is expected. This anti-ignition back-spray 

(Figure 4–8) quenches the hot streak, reducing its temperature and the chance of a frictional 

ignition. 

Cecala et al. [1985b] reported how a U.S. longwall lowered methane frictional ignitions by 

mounting a water spray behind each bit and by slightly lowering the cutting height of the shearer 

to avoid roof rock. Actions taken in the United Kingdom to reduce frictional ignitions on 

shearers have been reported by Browning [1988]. 
 

Figure 4–8.—Anti-ignition back spray. 

60 

When using water sprays to reduce frictional 

ignitions, the proper spray nozzle selection, 

nozzle placement, and operating pressure of 

anti-ignition back sprays are important if the 

full hot-streak quenching potential is to be 

realized [Courtney 1990; British Coal 1988]. 

For example, if the spray density is too low or 

if too much water is wasted in wetting the back 

of the bit, then quenching effectiveness suffers. 

Longwall drums with anti-ignition 

sprays are commercially available. 

A recent, comprehensive review of frictional 

ignitions in mines, including metal-to-metal 

ignitions and those from roof falls, is provided 

by Phillips [1996]. 

THE BEST LOCATION FOR THE METHANE MONITOR 

Normally, it is required to use two methane monitors at longwall faces, one located at the tailgate 

and another on the shearer. Because the shearer is a primary ignition source at most longwalls 

and because the methane concentration at the shearer is generally higher than the concentration 

at the tailgate,7 

7

On some longwalls, the tailgate concentration is higher. 

the shearer is usually the most critical location for the monitor. For example, during the tail-to-head pass shown in Figure 4–3, the methane concentration at the shearer exceeded 

1.0% several times, and during the headgate cutout it approached 2.5%. However, at no time did 

a methane monitor that was located at the tailgate record a concentration over 1.0%. Even if 

enough gas were released at the shearer to exceed 1.0% at the tailgate monitor, there can be a 

considerable delay as the gas cloud travels down the face from the shearer to the tailgate 

monitor. 

Because the shearer is usually the most critical location for the methane monitor, Cecala et al. 

[1993] conducted a study to establish the best location on the shearer. In this study, a full-scale 

laboratory facility was used to simulate a longwall face with a shearer. Methane was released at 

the drums, and the concentration was measured at several locations on the top of the shearer 

body (Figure 4–9). 
61 

Cecala et al. found that 

locations A through C 

on the face side of the 

shearer gave the highest 

methane concentrations 

and were approximately 

the same value. However, the drawback with 

a location on the face 

side of the shearer is that 

the methane monitor is 

prone to damage or to 

being covered with coal. 

Because of this, the best 

choice for a monitor location is usually at the gob-side tailgate-end of the machine, shown as 

location D in Figure 4–9. A monitor at location D is less likely to be damaged, covered with 

coal, or soaked by water sprays. However, the measured methane concentration is 40%–50% 

lower than that measured at face-side locations A through C. 

Figure 4–9.—Best methane monitor location. 

REFERENCES 

British Coal [1988]. Control of frictional ignitions and dust on shearers: notes for guidance. 

British Coal, Headquarters Technical Department, August. 

Browning EJ [1988]. Frictional ignitions. In: Proceedings of the Fourth International Mine 

Ventilation Congress (Brisbane, Queensland, Australia). 

Cecala AB, Jayaraman NI [1994]. Modified shearer-clearer system for dust and methane 

control. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 9404. NTIS No. 

PB95104873. 

Cecala AB, Jankowski RA, Kissell FN [1985a]. Determining face methane-liberation patterns 

during longwall mining. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, 

IC 9052. NTIS No. PB86156445. 

Cecala AB, Konda BW, Klinowski GW [1989]. A comparison of methane flow patterns on 

advancing and retreating longwalls. In: Proceedings of the Fourth U.S. Mine Ventilation Symposium (Berkeley CA, June 5–7, 1989). 

Cecala AB, O’Green J, Watson RW, Jankowski RA [1985b]. Reducing longwall face ignitions. 

World Min Equip Jan:44–46. 

Cecala AB, Rajan SR, Jankowski RA, Kissell FN [1986]. Cut longwall ignitions by lowering 

methane concentrations at shearers. Coal Age 91(8):66–67. 
62 

Cecala AB, Zimmer JA, Thimons ED [1993]. Determination of optimal longwall face methane 

monitoring locations. In: Proceedings of the Sixth U.S. Mine Ventilation Symposium (Salt Lake 

City, UT, June 21–23, 1993). 

Courtney WG [1990]. Frictional ignition with coal mining bits. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 9251. NTIS No. PB91110072. 

Denk JM, Wirth GJ [1991]. Evaluating methane emissions on an extended longwall face. In: 

Proceedings of the Fifth U.S. Mine Ventilation Symposium (Morgantown WV, June 3–5, 1991). 

Phillips HR [1996]. Identify methods to reduce the risks of frictional ignition hazards. Final 

project report, project COL 226. Braamfontein, Republic of South Africa, Safety in Mines 

Research Advisory Committee (SIMRAC) (www.simrac.co.za). 

Ruggieri SK, Muldoon TL, Babbitt C, Lee E [1985]. Improved diffuser and spray fan systems 

for ventilation of coal mine working faces. Waltham, MA: Foster-Miller Associates. U.S. 

Bureau of Mines contract No. J0113010. NTIS No. PB86168440. 

Ruggieri SK, Muldoon TL, Schroeder W, Babbitt C, Rajan S [1983]. Optimizing water sprays 

for dust control on longwall shearer faces. Waltham, MA: Foster-Miller, Inc. U.S. Bureau of 

Mines contract No. J0308019. NTIS No. PB86205408. 
CHAPTER 5.—BLEEDER SYSTEMS IN UNDERGROUND COAL MINES 

 

63 

By John E. Urosek, P.E.,1

1

Chief, Ventilation Division, Pittsburgh Safety and Health Technology Center, Mine Safety and Health Administration, Pittsburgh, PA. 

William J. Francart, P.E.,2

2

Mining engineer, Ventilation Division, Pittsburgh Safety and Health Technology Center, Mine Safety and Health 

Administration, Pittsburgh, PA. 

and Dennis A. Beiter3 

3

Supervisory mining engineer, Ventilation Division, Pittsburgh Safety and Health Technology Center, Mine Safety 

and Health Administration, Triadelphia, WV. 

In This Chapter 

9 Designing bleeder systems 

9 Examining and maintaining bleeder systems 

9 Evaluating bleeder system effectiveness 

INTRODUCTION 

Bleeder systems are that part of the mine ventilation network used to ventilate pillared areas in 

underground coal mines. Pillared areas are those in which pillars have been wholly or partially 

removed, including the areas where coal has been extracted by longwall mining. Bleeder systems protect miners from the hazards associated with methane and other gases, dusts and fumes, 

and oxygen deficiency that may occur in these mined-out areas. Effective bleeder systems control the air passing through the area and continuously dilute and move any methane-air mixtures 

and other gases, dusts, and fumes from the worked-out area away from active workings and into 

a return air course or to the surface of the mine. A bleeder system includes the pillared area 

(including the internal airflow paths), bleeder entries, bleeder connections, and all associated 

ventilation control devices that control the air passing through the pillared area. Bleeder entries 

are special air courses designed and maintained as part of the mine ventilation system. 

This chapter focuses on the design, examination, 

maintenance, and evaluation of bleeder systems 

in underground coal mines. 

The history of coal mine explosions in the United States is a reminder of the importance of adequate ventilation. Some of those disasters were the result of inadequately ventilated pillared 

areas. The importance of developing bleeder systems to ventilate these pillared areas and evaluating the bleeder system’s effectiveness is reflected in present-day federal regulations. For more 

information on bleeder systems, see Tisdale [1996] and Urosek and Francart [2002]. 
 

 

64 

DESIGNING BLEEDER SYSTEMS 

As part of the mine ventilation system, bleeder systems should be addressed in the overall mine 

design. Designing a good bleeder system requires consideration of ground control and pillar 

design, strata characteristics, contaminant liberation, airflow distribution, and the internal workings of the bleeder system. When bleeder systems are incorporated into an existing mine ventilation system, the capacity of that ventilation system should be considered. 

There are two basic design classifications for bleeder systems used in U.S. coal mines: wraparound and flow-through. Although there are many variations, the concept by which they all 

function is the same.4

4

Some unique systems have been developed for ultragassy mines affected by spontaneous combustion. These 

special cases are not be covered in this chapter. 

 A ventilating pressure differential is established from the active workings, 

across the pillared area, to the bleeder and/or return entries. Sufficient connections are established between the pillared area and these entries for the airflow to be induced and distributed 

through the pillared area by the applied ventilating pressure differential. Gases in the pillared 

area are diluted and moved away from the active workings by the airflow induced through the 

pillared area. The applied ventilating pressure is that pressure 

which actually causes the airflow. The primary difference 

between the two designs is the 

means used to maintain the 

pressure differential across the 

pillared area. 

Wrap-around system designs 

rely completely upon ventilation 

controls constructed between 

the pillared area and bleeder 

entries to establish the ventilating pressure differential. 

Numerous ventilation controls 

located close to the pillared area 

are often necessary with this 

design. The additional ground 

pressures from the pillared area 

increase the susceptibility of 

these ventilation controls to 

damage. Figure 5–1 is a simplified illustration of a wraparound system for a longwall 

panel. Flow-through systems 

use solid coal barriers in 

Figure 5–1.—Simplified illustration of a wrap-around bleeder sys- conjunction with ventilation tem for a longwall panel. 
 

65 

controls to maintain a pressure differential across the pillared area. 

Figure 5–2 is a simplified illustration 

of a flow-through system with 

bleeder entries. The coal barriers 

replace many of the ventilation controls and can provide additional protection for the bleeder entries from 

the ground pressures created as a 

result of extracting the coal pillars. 

These simplified illustrations do not 

show all of the ventilation controls 

sometimes necessary to direct the 

airflow, nor do they show the additional multiple entries that may be 

necessary to carry the required quantity of air. 

Figure 5–2.—Simplified illustration of a flow-through system 

with bleeder entries. 

Ground control and pillar design. 

Ground control and pillar design are 

critical to the overall stability of 

bleeder systems. Some factors 

to be considered include mining 

parameters (e.g., mine opening 

dimensions, pillar size, panel size, 

bleeder geometry, and intersection 

dimensions), overburden and 

horizontal stresses, roof control 

systems, and multiple-seam interaction. Irregular shapes in the 

bleeder system that concentrate 

abutment loads generally require 

additional attention. Ground 

pressures developed as a result of mining can be magnified in such areas and may necessitate 

additional support. 

The life expectancy of the bleeder system must be considered during the design process. The 

necessity of maintaining access and protecting primary internal airflow paths within the pillared 

area cannot be overstated. The long-term nature of bleeder entries and their close proximity to 

pillared areas usually require that supplemental roof support be installed in them. The additional 

support should be installed in advance of overburden stresses from pillaring because, typically, 

limited access to bleeder entries makes remedial actions difficult or impossible. 

Roof support requirements are mine-specific. Wood cribbing has long been recognized as an 

appropriate supplemental roof support for many bleeder systems. Technology and innovative 

design have provided alternative materials. Concrete donuts, canned cribs, pumpable supports, 

yielding posts, metal supports, cable bolts, and trusses have been used. 
66 

Caved area characteristics. The characteristics of the caved material and the overlying strata in 

areas where pillars have been extracted are influenced by several factors. These include the 

geology of the roof strata, the degree of extraction, and the distance between supporting pillars. 

Mining method and percentage of pillar recovery greatly affect the strata response to the extraction. Increased extraction generally results in less permeable caved areas. If the main roof subsides, compaction of the caved material occurs and decreases void space and permeability. 

Compaction occurs in the caved areas created by most longwalls. The limited void space and 

decreased permeability of compacted caved areas limits the volume of gases that can accumulate 

in the rubble of the caved area and diminishes their interaction with mine airstreams. 

Contaminant liberation. Estimation of liberation rates for gaseous contaminants, such as 

methane and carbon dioxide, is an important factor in determining ventilation requirements. 

However, this can be difficult to determine prior to mining when no liberation history has been 

developed. Even with historical data available for the engineer to use, other factors must be considered. The source of the contaminants is one factor that can impact the bleeder system design. 

Some mines considered to be very gassy produce significant methane volumes only while cutting 

coal. In other gassy mines, a majority of the methane is liberated when the strata are broken due 

to second mining. Fractures caused by caving in the pillared area can permit methane and other 

contaminants to enter the mine from overlying and underlying mines as well as from the surrounding strata. Contaminants are also liberated from the coal ribs of the bleeder entries. 

Another factor is the relationship between the size of the pillared area and the total volume of 

contaminants liberated. Liberation of contaminants can increase as the pillared area increases in 

size. Although methane is the primary contaminant in many mines, carbon dioxide can also be 

a concern. 

A relationship exists between production rates and methane liberation in most mines. Increased 

production rates result in increased methane liberation, both in the face area and the pillared area. 

Contaminants in the face ventilation airflow directed into the pillared area will also impact the 

bleeder system. Bleeder system design that addresses the capacity to dilute contaminants such 

that production need not be curtailed is prudent and reduces the potential for exceeding the dilution capacity of the system. 

In some gassy mines, vertical degasification boreholes have been used effectively to reduce the 

methane to be diluted by underground bleeder systems. Horizontal degasification boreholes 

have been used to drain methane from the coal seam in advance of mining and can assist in 

reducing ventilation requirements during development as well. 

Oxidation within the pillared area, as well as contaminants liberated from the strata, can contribute to oxygen deficiency. Oxygen deficiency in the bleeder entries and at measurement point 

locations can impede continued access to examination locations needed for evaluating bleeder 

system effectiveness. Consideration must be given to providing sufficient airflow through the 

pillared area and bleeder entries to prevent oxygen deficiency so that the bleeder entries and 

measurement point locations are maintained safe for access. 
67 

Airflow distribution. The bleeder system should be designed such that the pillared area is 

continuously under the influence of mechanical ventilation that will induce the necessary quantities of airflow in the intended direction. The design should preclude air that has ventilated the 

pillared area from flowing toward or by the working section. The bleeder system must continuously control and distribute airflow through the area. 

Airflow distribution and ventilation capacity affect dilution of contaminants. Proper distribution 

of airflow through the bleeder system facilitates dilution of the contaminants and minimizes the 

possibility of accumulations of methane and other contaminant gases. Managed airflow distribution requires openings, including primary internal airflow paths within the pillared area, bleeder 

connectors, and bleeder entries that function for the duration needed. Connections with the pillared area provide inlets and outlets through which the airflow is distributed and enable gathering 

of information used in evaluating the effectiveness of the airflow distribution. 

Airflow is controlled using ventilation controls and mine layout. Standard construction techniques, considering the life of the system and the conditions anticipated, should be adopted for 

ventilation controls in the bleeder system. Some types of controls, such as curtains, are more 

susceptible to damage by the mine environment or inadvertent change by mine personnel. 

Unplanned changes may adversely impact bleeder system effectiveness. Critical controls should 

remain accessible because adjustments are usually necessary throughout the life of a system. 

Removal of unneeded ventilation controls decreases the likelihood that they will cause unintentional restrictions. 

Internal workings of the bleeder system. It is recognized that the characteristics of individual 

pillared areas vary. However, the primary internal airflow paths of many bleeder systems are the 

remaining development entries and crosscuts between and around the caved material in the pillared area and the perimeter of the caved area. Except for the perimeter, airflow across and 

through much of the caved area generally occurs only to a limited degree unless the permeability 

of the caved material is high and/or the applied ventilating pressure is large. With contemporary 

bleeder system designs, inclusion and support of mine entries within the pillared area often are 

necessary to provide internal airflow paths through which air will continually pass to dilute and 

carry the contaminant gases away from the active areas and into bleeder and/or return entries 

or directly to the surface. 

In longwall bleeder systems, the ventilation of the longwall face area cannot be separated from 

the bleeder system. In longwall panels where caving has occurred, the amount of air flowing 

across the face directly impacts the airflow passing through the parallel primary internal airflow 

path immediately behind the shields. Increased longwall face airflow results in better dilution 

and removal of contaminants in this critical portion of the pillared area closest to the face. 

Loss of the primary internal airflow paths that provide a direct conduit from the active workings 

can compromise bleeder system effectiveness and the ability to evaluate the system’s effectiveness. Although some of the contaminant gases may continue to be moved away from the active 

areas, insufficient airflow through the primary internal airflow paths may result in accumulations 

of hazardous gas concentrations in close proximity to active areas, such as locations behind the 

shields in longwall systems or other open areas near the face. A roof fall or air reversal could 
68 

then move this gas into the active area. Because of limited access, it may not be possible to 

determine the extent of the accumulation or how close it is to the active areas. The failure to 

adequately address the internal workings of the bleeder system may result in hazardous conditions for miners. 

Capacity of the ventilation system. The total ventilation resistance of the pillared area of a 

bleeder system depends not only on the permeability of the caved material, but also on the size 

and shape of the pillared area, including the length and integrity of the primary internal airflow 

paths through and around the caved material. As the pillared area increases in size or age, the 

resistance of the primary airflow paths generally increases. Experience has shown that significant changes to the resistance of primary internal airflow paths can even occur during the mining 

of one panel. 

The mine ventilation system must maintain a ventilating pressure able to overcome the resistance 

in the bleeder system and sustain airflow. The importance of a reserve ventilating pressure 

capacity, as evidenced by the regulation of the air splits, was recognized by ventilation engineers 

over 30 years ago. The reserve ventilating capacity provides flexibility to increase airflow when 

needed and to provide additional ventilating pressure differential across the pillared area 

[Kalasky and Krickovic 1973]. 

Recent trends in longwall mining require more emphasis on the ability of the ventilation system 

to provide and maintain greater ventilating pressure differentials across the pillared areas of 

bleeder systems. Individual longwall panels are increasing in length and width, the number of 

development (gate) entries is decreasing, and the number of connected panels in an individual 

bleeder system is increasing. The result is higher resistance airflow paths and extended longevity. Both often result in greater ventilation requirements. Thus, the need for designing 

bleeder systems with sufficient reserve ventilating pressure capacity is vital. 

Ventilating pressure is an important consideration in the design of all bleeder systems. In some 

mines, the main mine fan can provide adequate airflow for the bleeder system. Other mine operators have found that incorporating bleeder shafts and high-pressure bleeder fans into flowthrough bleeder systems with bleeder entries is necessary to meet the required capacity of 

today’s larger systems, especially in mines with high methane liberation rates. But even these 

high-pressure exhaust fans have limitations. Applying ventilating pressure differential in itself is 

not the equivalent of effectively ventilating the pillared area. Airflow is necessary to dilute and 

carry away the contaminants. Accumulations of hazardous gases can still develop if sufficient 

airflow is not continuously diluting and carrying away the contaminants. 

Other factors also impact bleeder system capacity. The configuration of some bleeder systems 

can severely limit ventilating capacity. The resistance of the airflow paths within the bleeder 

system must be considered. Compared to multiple entries, pressure losses in single bleeder 

entries will reduce the ventilating pressure available to move air through the pillared area. Ventilating pillared areas such that the ventilating pressure is applied in opposing directions may have 

adverse impact on both the airflow through the internal workings of the bleeder system and the 

ability to evaluate the system performance. Additional splits of air directed into the bleeder 

entries, for the purpose of ventilating electrical installations and dewatering systems or to 
 

69 

provide greater access, displaces bleeder airflow and may eliminate a bleeder entry that could 

otherwise carry bleeder airflow. Finally, leakage from splits of air located adjacent to bleeder 

airflow can impact the system capacity, including its ability to dilute contaminants from the 

pillared area. 

Water drainage. The design of the bleeder system should include the means to prevent water 

accumulations that might cause obstructions. The mine water drainage infrastructure should be 

installed before pillar mining limits access to an area. Consideration should also be given to 

preventing water accumulations within the pillared area that could impact bleeder system 

effectiveness. A mine layout that considers anticipated mine floor elevations can impact the 

drainage of water in the bleeder system. 

Sealing. Bleeder systems should be designed to provide ventilation and enable evaluation of 

system effectiveness for the anticipated life of the system. Realistic consideration of the size and 

number of panels and the length of time that ventilation can be provided and the bleeder system 

can be evaluated is essential. Prudent mine operators limit the size and age of the bleeder system 

before a decline in system performance or an inability to evaluate the system necessitates sealing. If it is determined that the bleeder system is not effective or it cannot be determined that the 

bleeder system is effective, the worked-out area must be sealed. The ability to seal the pillared 

area ventilated by the bleeder system must be considered in the design process. Federal regulations require that each mining system be designed so that each worked-out area can be sealed. 

The location and the sequence of construction of proposed seals are required to be specified by 

federal regulations. 

EXAMINING AND MAINTAINING BLEEDER SYSTEMS 

Weekly examination requirements for bleeder 

systems are specified in 30 CFR5

 75.364. 

5

Code of Federal Regulations. See CFR in references. 

Examinations provide the means of collecting the information needed to evaluate bleeder system 

effectiveness. 30 CFR 75.364 specifies the minimum requirements for the examination of 

bleeder systems. Weekly examinations are required at all locations where air enters the workedout area (inlets) and in the bleeder system airflow immediately before the air enters a return split 

of air (outlets). Measurements of methane and oxygen concentrations and of air quantity and a 

test to determine if the air is moving in its proper direction are to be made at all the locations. 

During the weekly examinations, at least one entry of each set of bleeder entries is to be traveled 

in its entirety. Measurement point locations at which examinations are to be made are required 

to be specified in the ventilation plan. These locations are not in lieu of traveling the bleeder 

entries, but rather are the locations within the bleeder system, in addition to the inlets and outlets, 

where the examiner will measure the methane and oxygen concentrations and air quantities and 
70 

perform tests to determine whether the air is moving in the proper direction. Knowledge of the 

specific conditions of each mine and an understanding of how the system functions, including 

the internal airflow patterns, are necessary when considering specific measurement point locations. Examinations should also evaluate the condition of ventilation controls critical to the 

proper function of the bleeder system. Provisions exist for an alternative method of evaluation to 

be specified in the ventilation plan, provided it results in proper evaluation of the effectiveness of 

the bleeder system. 

Mine examiners or persons working or traveling in remote areas, especially in bleeder entries, 

should always be on the alert for changing conditions, such as accumulations of methane or 

oxygen-deficient air. Persons should not enter connectors and the pillared area unless they are 

well-informed about the areas to be entered, have sufficient detection instruments, and have discussed their intent with other persons who will then know their whereabouts. 

Some mine operators have increased the frequency of monitoring the air quality and quantity at 

examination locations by installing sensors connected to the central atmospheric monitoring 

system. In this way, the mine operator can continuously monitor and record this important 

information. Sensors for methane, oxygen, carbon monoxide, and air velocity have been 

installed to enhance the evaluation of the ventilation system. 

Maintenance of bleeder systems can directly impact system effectiveness and/or the ability to 

determine the effectiveness of the bleeder system. Failure to provide and maintain adequate control of the ground conditions often results in roof falls and floor heave. Failure to provide and 

maintain a means to control water often results in water accumulations. The most significant 

consequences of roof falls and water accumulations are the potential for reducing or preventing 

airflow and preventing the completion of the examinations that are required to determine the 

system’s effectiveness. Due to access limitations and practical constraints, deteriorated conditions and obstructions may not be possible to remediate. These same types of conditions and 

obstructions may impact the primary internal airflow paths as well, with fewer or no options to 

remediate. Thus, the need to prevent obstructions within a bleeder system through adequate 

preventive measures cannot be overstated. 

The records that federal regulations require to be maintained concerning the weekly examinations include the results of particular tests and measurements and notations for hazardous conditions observed. Other information not required can also be beneficial. Notes, records, or 

communications of pertinent information made by the examiner can be useful in determining 

whether problems are developing in the bleeder system. This information includes changes in 

water levels and their locations, roof conditions, floor heave, pillar and roof deterioration, and 

damage or deterioration of important ventilation controls. 

Roof fall and water accumulations in bleeder entries have contributed to many serious accidents 

involving coal mine bleeder systems. In one incident, an accumulation of water was a major 

reason that a mine examiner did not travel the bleeder entries of a room-and-pillar wrap-around 

bleeder system. The examiner determined that water had begun to accumulate in a corner of the 

bleeder system. A month later, water had risen to a depth considered by the examiner to be too 

hazardous through which to travel. Consequently, the bleeder entries were examined only to the 
71 

edge of the water. Roof falls in the bleeder entry were not addressed. As a result, the effectiveness of the bleeder system was not evaluated. The bleeder system became ineffective and failed 

to properly remove methane from the mined-out area. Methane accumulated in the mined-out 

area just inby the retreat section pillar line. Removal of ventilation controls on the retreat section 

permitted movement of the accumulated methane to the active workings, where it was ignited. 

Ten miners were fatally injured in the ensuing explosion. 

EVALUATING BLEEDER SYSTEMS 

Examiners, inspectors, and engineers should be trained to evaluate bleeder systems and recognize deficiencies. As they inspect or analyze a bleeder system, they must be able to recognize 

hazardous conditions and take appropriate action. They must have knowledge of the three basic 

factors governing the ventilation of mines: ventilating pressure, airflow, and air quality. In addition, they must have an understanding of the particular bleeder system they are examining or 

evaluating, including airflow directions and air quantities, as well as normal methane or other 

critical gas concentrations; the location of potential problems from inadequate roof support or 

water accumulation; and the location of critical ventilation controls. This knowledge provides a 

base of information with which to evaluate the importance of subtle changes from previous 

examinations and to recognize deteriorating conditions that might cause system failure. 

Control and direction. A bleeder system must continuously control and distribute air throughout the system. Established airflow patterns enable collection of information that is important in 

evaluating the bleeder system’s effectiveness. Effective bleeder systems maintain the established 

airflow patterns. At inlets, outlets, and in the bleeder entries, airflow should be sufficient to be 

readily discernible and be in the proper direction. Airflow direction within the pillared area 

should also be in the proper direction. A bleeder system that does not produce discernible airflow through the pillared area is ineffective. A bleeder system in which the airflow direction has 

changed should be scrutinized. The bleeder system may no longer be effective, or the ability to 

evaluate its effectiveness may have been adversely impacted. Additional information may be 

needed to evaluate the bleeder system’s effectiveness. 

Air quality. Air quality is another essential consideration in determining bleeder system effectiveness. The primary air quality considerations for most bleeder systems are: keeping methane 

concentrations to no greater than 2% in the bleeder split of air immediately before it joins 

another split of air, diluting methane concentrations elsewhere within the bleeder system, and 

maintaining oxygen and limiting carbon dioxide concentrations in areas where persons work 

or travel. 

In some highly gassy mines, the air currents coming from the pillared area can contain methane 

concentrations higher than 2% and must be diluted by the air moving in the bleeder entry. As a 

practical matter, when methane in the air moving in the bleeder entries approaches 2%, it can no 

longer dilute additional methane, from the pillared area or that liberated into the bleeder entries, 

to the 2% limit. The bleeder system is no longer effective when methane concentrations in the 

bleeder split cannot be reduced to 2% before entering another split of air. 
72 

A major purpose of the bleeder system is to keep methane accumulations away from mining 

activities, including the primary airflow paths that provide a conduit to the active section. 

Accumulations of unusually high methane concentrations in locations other than small pockets, 

such as in a corner, in the interstices of the rubble material, or in a small roof cavity, indicate that 

changes to the bleeder system may be necessary. 

Methane exists in the pillared area and must be diluted by the air currents within the bleeder 

system and not be allowed to accumulate in open areas. Explosive mixtures of methane in open 

entries or crosscuts can constitute imminent danger. Relevant considerations include the location 

and extent of the accumulation, potential ignition sources, the primary internal airflow paths that 

provide a conduit to the active areas, and the potential for explosive methane-air mixtures to 

move to active areas, including the working section. It is imperative for miner safety that the 

portions of the pillared area adjacent to the working section and the primary internal airflow 

paths providing a conduit to the active working section be free of methane to the extent that a 

pillar fall might displace, or air reversals might move, an explosive methane-air mixture to the 

working section. Several accidents have resulted from the ignition of methane accumulations 

that existed within the pillared area near the working section. 

All possible sources of methane need to be considered carefully. For example, some coal seams 

liberate large amounts of methane continuously from virgin coal ribs. In some instances, 

barometric pressure changes may cause higher liberation from the mined-out area. However, 

a bleeder system with adequate pressure differentials and air distribution will not be substantially 

affected by normal barometric changes. Inactive panels in a mined-out area tend to liberate constant amounts of methane and in lesser quantities than active panels. Increases in methane concentration or reduction in airflows from older panels should be investigated to determine if the 

system is still functioning as designed. Factors influencing methane levels include coal production levels, the production day of the week, the gas-bearing characteristics of the strata near 

the active mining, the proximity of the last vertical methane degasification borehole, recent ventilation changes, and changes in barometric pressure. Changes or trends of deteriorating air quality at measurement point locations or other examination locations may indicate an ineffective 

bleeder system. 

Usually, the oxygen concentration in the bleeder system is a problem only from the standpoint of 

the safety of persons making examinations or assigned to work in the bleeder system. However, 

oxygen deficiency found in traveled areas of the bleeder system may indicate insufficient airflow 

in the bleeder system and should be further investigated. In bleeder systems with insufficient airflow, oxygen deficiency may result in the inability to evaluate the effectiveness of the system. 

Oxygen deficiency could also present a hazard to the working section if the bleeder system is 

ineffective and allows a sizable volume of oxygen-deficient air to exist in the pillared area or 

bleeder entries near the working section. Persons inspecting or examining bleeder systems must 

be alert to locations where low levels of oxygen may exist. The oxygen concentration in areas of 

bleeder entries and mined-out areas where persons work or travel must be at least 19.5%. The 

carbon dioxide levels must not exceed 0.5% time-weighted average and 3% short-term exposure 

limit in areas of bleeder entries and mined-out areas where persons work or travel. Oxygen can 

be displaced by methane, especially in high spots or cavities above roof falls, where the buoyancy of methane can cause it to collect. Oxygen is depleted as coal, wood, or other organic 
73 

materials oxidize. In mines that liberate small amounts of methane, low airflows could result in 

oxygen deficiency before elevated methane levels occur. 

The information collected at the examination locations and recorded in mine records should be 

used in evaluating the effectiveness of the bleeder system. Trends can be monitored. However, 

a thorough understanding of the system is needed to fully assess bleeder system performance. In 

complex systems or those with which there are concerns over limited capacity or effectiveness, 

there may be times when an investigation beyond the regular examination locations, such as into 

the primary internal airflow paths, may be appropriate if an investigation can be conducted 

safely. Confirmation of the internal airflow patterns and assessment of the dilution of contaminants may be appropriate to ensure that the examination locations provide the necessary information and can also provide a better understanding of what the trends at specific measurement point 

locations reflect about the internal workings of the bleeder system. 

Situational indicators. As bleeder systems change, some conditions can develop as a result of, 

or in response to, declining system performance. These circumstances often warrant closer scrutiny to determine the impact of the condition on bleeder system effectiveness. The location and 

number of the present measurement point locations should be reassessed as to the adequacy for 

providing sufficient information for evaluation. Routine changes in the bleeder system can also 

cause problems. Situations such as the following should raise the level of interest and caution of 

persons evaluating bleeder system effectiveness. 

•	 Separate mine fans. Directing airflow from the pillared area to separate mine fans may 

diminish the pressure differential established across segments of the pillared area, sometimes resulting in poorly ventilated areas, unventilated areas, or complication of the 

evaluation of the bleeder system effectiveness. Similar problems can result when the 

airflow ventilating the face area is directed to a mine fan separate from the mine fan ventilating the pillared area. Some of the airflow ventilating the active face should always be 

directed into the pillared area to prevent unventilated areas of the pillared area close to 

the active working section. Depending on the configuration of the system and the control 

of the airflow in the system, similar problems can occur even when a single mine fan is 

ventilating the face area and the mined-out area. 

•	 Splits of air directed into the pillared area (other than the split ventilating the active 

working section). Although other splits are often necessary and important to maintain 

ventilation of the pillared area, understanding the impact of introducing other splits on the 

bleeder system is important. Splits of air directed into the pillared area (other than the 

split ventilating the face(s) from the active working section) can decrease ventilating 

pressure across the pillared area and adversely affect airflow away from the active working section. Depending on ventilating pressure differentials, poorly ventilated areas may 

develop within the pillared area. Extreme conditions would be stagnation of the airflow 

entering the pillared area from the face area or the airflow from the mined-out area moving to or toward the active working section. The proposed introduction of other splits of 

air into the pillared area should be evaluated to predict what overall effect the change will 

have on the system. Following implementation of changes, the bleeder system should be 

evaluated through in-mine examinations to determine if the changes produced the 

expected results. 
74 

•	 Inlets to the pillared area near locations where air exits from the pillared area. Inlets to 

the pillared area near locations where air exits from the pillared area can mask the effectiveness of the airflow distribution and complicate or prevent evaluation of the bleeder 

system. This condition should be closely scrutinized and is generally discouraged. 

•	 Inlets to the pillared area from intake air courses. The pressure differential that exists 

across regulators separating intake air courses from pillared areas should be sufficient to 

prevent air from the pillared area from entering the intake air courses due to normal mining activities that affect ventilating pressures in the area. 

•	 Splits of air separate from the bleeder split. Small splits of intake air that are used to 

ventilate electrical installations are permitted to enter the bleeder airflow and do not generally affect the location of the regulatory 2% methane limit, assuming they have been 

determined to be insignificant in their effect on the bleeder split and ventilation of the pillared area. Leakage from separate splits of air located adjacent to bleeder airflow may 

improve the air quality in the bleeder entries while actually decreasing the ventilating 

capacity of the pillared area by reducing the available ventilating pressure and airflow 

quantity. A full consideration of the significance of separate splits of air located adjacent 

to bleeder airflow should include leakage from that split into the bleeder airflow. The 

location of the regulatory 2% methane limit in the bleeder split may be affected. 

•	 Startup and recovery ventilation. The ventilation of the face and pillared area at the 

startup of both longwall and room-and-pillar systems is dynamic. Significant caving in 

the pillared area sometimes does not readily occur when retreat mining begins. Larger 

open areas often exist in the pillared area, and the primary internal airflow paths that will 

exist during mining of the majority of the panel are not fully established. The ability to 

maintain airflow on a longwall face can be impaired. The conditions often warrant close 

scrutiny of the distribution of the airflow through the pillared area until the primary internal airflow paths are established. Additional ventilation controls may be needed during 

this period to ensure adequate distribution of airflow. 

Setup and recovery of longwall face equipment is another critical time when close scrutiny is warranted. Airflow is often being redistributed in the bleeder system by mine 

management as the new longwall face is set up and equipment on the finished longwall 

face is recovered. Since methane liberation at the face decreases after production is 

stopped, longwall recovery faces are also often ventilated with decreased airflow quantities. That decrease in face airflow directly impacts the ventilation of the pillared area 

nearest the recovery area. Quality examinations are needed to ensure that the bleeder 

system is effective and that methane and other hazardous gases do not accumulate. 

•	 Changes in gas concentrations, ventilating pressures, and air quantities. If examiners 

find rising methane concentrations with no changes in air quantities or pressures, then 

it can be concluded that methane liberation is increasing. Rising methane or decreasing 

oxygen levels with decreasing air quantities and ventilating pressures can indicate possible restrictions in the bleeder system. However, relying on just one parameter such as air 

quantity measurements without considering other related information can sometimes 

result in a misdiagnosis of the condition. 
•	 Dilution capacity of bleeder system less than contaminant production. The ventilating 

capacity of a bleeder system is insufficient if it cannot dilute the amount of contaminants 

that can result from normal coal production. It can be difficult to assess the success of 

limiting production as a corrective action when the capacity to produce coal results in the 

production of more contaminants than the bleeder system can adequately dilute. The full 

effect of excessive coal production cannot be readily assessed in a bleeder system. 

Because of the delay between when methane is liberated and when it exits the pillared 

area, methane can accumulate in the pillared area of a system with insufficient capacity 

before it can be detected at examination locations. This can also result in a continuing 

increase of methane concentration at examination locations after mining has ceased. 

Ceasing production after changes are detected does not have the same impact as decreasing the capable production rate because contaminants have accumulated before detection 

is possible. This process can result in repetitious excursions beyond the bleeder system’s 

capacity to safely dilute contaminants. Such repeated excursions indicate an ineffective 

bleeder system. A more appropriate response would be to improve the capacity of the 

bleeder system. Because of inadequate planning, some mine operators have had to resort 

to sinking bleeder shafts, installing bleeder fans, and drilling degasification boreholes as 

remedial measures to improve the capacity of existing bleeder systems. 

75 

A bleeder system should be designed to provide 

adequate ventilation of the pillared area at the 

maximum expected coal production rate. 

•	 Reserve ventilating pressure relative to the applied ventilating pressure. The magnitude 

of the applied ventilating pressure across the pillared area is often not easily determined. 

The applied ventilating pressure cannot usually be determined merely from a measure of 

the pressure differential across ventilation controls. This information can be obtained 

from altimeter surveys or tube and pressure gauge surveys. Knowledge of the magnitude 

of the applied ventilating pressure is useful in assessing the capacity of the reserve ventilating pressure. This ventilating pressure can be compared with the pressure differentials 

across regulators controlling airflow through the pillared area. Some mine operators have 

installed permanent pressure measurement stations at these locations. 

•	 Confusing examination records. Organization of the examination records greatly 

improves the ability to adequately consider the information collected when evaluating 

bleeder system effectiveness. If the examination records seem unorganized or the information specific to the bleeder system is difficult to collectively review, changes that have 

occurred in the bleeder system are not as easily observed or evaluated. 
76 

REFERENCES 

CFR. Code of federal regulations. Washington DC: U.S. Government Printing Office, Office of 

the Federal Register. 

Kalasky JD, Krickovic S [1973]. Ventilation of pillared areas by bleeder entries, bleeder systems, or equivalent means. Trans Soc Min Eng, AIME, December, Vol. 254, pp. 284–291. 

Tisdale J, ed. [1996]. Bleeder and gob ventilation systems. Arlington, VA: U.S. Department of 

Labor, Mine Safety and Health Administration. 

Urosek JE, Francart WJ [2002]. Bleeder systems in underground coal mines. In: De Souza E, 

ed. Proceedings of the North American/Ninth U.S. Mine Ventilation Symposium (Kingston, 

Ontario, Canada, June 8–12, 2002). Lisse, Netherlands: Balkema. 
 

77 

CHAPTER 6.—COAL SEAM DEGASIFICATION 

By Pramod C. Thakur, Ph.D.1 

1

Manager, Coal Seam Degasification, CONSOL Energy, Inc., Morgantown, WV. 

In This Chapter 

9 Origins of coalbed methane 

9 Reservoir properties of coal seams 

9 Thresholds for coal seam degasification 

9 Methane emissions in mines 

9 Methane drainage techniques 

9 How to transport gas safely in mine pipelines 

and 

9 Economics of coal seam degasification 

Recommended publications on coal seam degasification 

are Gas Control in Underground Coal Mining [Creedy et al. 

1997], Coalbed Methane Extraction [Davidson et al. 1995], 

and Methane Control for Underground Coal Mines 

[Diamond 1994]. 

ORIGINS OF COALBED METHANE AND RESERVOIR PROPERTIES 

OF COAL SEAMS 

Origins of coalbed methane. Coal seams form over millions of years by the biochemical decay 

and metamorphic transformation of plant materials. This coalification process produces large 

quantities of byproduct gases, such as methane and carbon dioxide. The amount of these 

byproducts increases with the rank of coal. It is the highest for anthracite, where for every ton of 

coal nearly 1,900 lb of water, 2,420 lb (20,000 ft3

) of carbon dioxide, and 1,186 lb (27,000 ft3) of 

methane are produced [Hargraves 1973]. Most of these gases escape to the atmosphere during 

the coalification process, but a small fraction is retained in the coal. The amount of gas retained 

in the coal depends on a number of factors, such as the rank of coal, the depth of burial, the type 

of rock in the immediate roof and floor, local geologic anomalies, and the tectonic pressures and 

temperatures prevalent at that time. The gases are contained under pressure and mainly adsorbed 

on the surface of the coal matrix, but a small fraction of gases is also present in the fracture network of the coal. Methane is the major component of gases in coal, comprising 80%–90% or 

more of the total gas volume. The balance is made up of ethane, propane, butane, carbon 

dioxide, hydrogen, oxygen, and argon. 
78 

Methane is released into each mine airway from the coal seam as mining proceeds. Large 

volumes of air, sometimes as much as 20 tons of air for each ton of coal mined, is circulated 

constantly to dilute and carry methane away from coal mines. Methane is a colorless, odorless, 

combustible gas that forms an explosive mixture with mine air in the concentration range of 

5%–15% by volume. The maximum concentration of methane in mine air is restricted by law to 

1%–1.25% in all major coal-producing countries. Nevertheless, methane-air explosions are quite 

common even today. Table 6–1 shows a list of major coal mine explosions since 1970 in the 

United States. In these 13 explosions, 167 lives were lost despite coal seam degasification taking 

place in some mines. 

Table 6–1.—Major coal mine explosions in the United States, 

1970–present 

Year Mine and location Deaths 

2006.........

2001.........

1992.........

1989.........

1983.........

1982.........

1981.........

1981.........

1981.........

1980.........

1976.........

1972.........

1970.........

 Sago Mine, Tallmansville, WV ........................... 

 Blue Creek No. 5 Mine, Brookwood, AL ............ 

 No. 3 Mine, Norton, VA ......................................

 William Station No. 9 Mine, Wheatcroft, KY....... 

 McClure No. 1 Mine, McClure, VA .....................

 No. 1 Mine, Craynor, KY ....................................

 No. 21 Mine, Whitwell, TN.................................. 

 No. 11 Mine, Kite, KY.........................................

 Dutch Creek No. 1 Mine, Redstone, CO............ 

 Ferrell No. 17 Mine, Uneeda, WV ......................

 Scotia Mine, Oven Fork, KY............................... 

 Itmann No. 3 Mine, Itmann, WV.........................

 No. 15 and 16 Mines, Hyden, KY....................... 

12 

13 

 8 

10 

 7 

 7 

13 

 8 

15 

 5 

26 

 5 

38 

Coal has been mined throughout the world for hundreds of years, and the history of coal mining 

is replete with mine explosions and consequent loss of lives. Even today, 60 countries around 

the world mine about 5 billion tons of coal annually with more than 10,000 fatalities per year. 

Before 1950, when coal seam degasification was generally unknown and ventilation was the only 

method of methane control, mine explosions in the United States were much more disastrous 

with a very high number of fatalities. To mitigate this problem, in many instances, mine ventilation can be supplemented by coal seam degasification prior to mining and even after mining. 

Reservoir properties of coal seams. Coal seam degasification techniques to be used in a mine 

depend on the reservoir properties of the coal seams being mined. Good methane control planning depends on accurate information on the reservoir properties of the coal seam and the total 

gas emission space created by the mining process. Reservoir properties governing the emission 

of methane from coal seams can be divided into two groups: (1) properties that determine the 

capacity of the seam for total gas production, e.g., adsorbed gas and porosity, and (2) properties 

that determine the rate of gas flow, e.g., permeability, reservoir pressure, and diffusivity of coal. 

The reservoir properties are highly dependent on the depth and rank of the coal seam. The most 

important of these properties is the seam gas content. 
 

79 

Seam gas content. Based on their gas contents, coal seams can be classified as mildly gassy, 

moderately gassy, and highly gassy, as shown in Table 6–2. 

Table 6–2.—Gassiness of coal seams 

Category 1

Depth, ft Gas content of coal, 

ft3

/ton 

Mildly2

 gassy..................... <600 <100 

Moderately gassy ............. 600–1,200 100–300 

Highly gassy ..................... 1,200–3,000 300–700 1

Depth figures are for high-volatile bituminous coals. 

2

The term “mildly” is not intended to imply that such mines are free from 

gas problems. The potential for gas problems involves many factors, 

not just the coal gas content. If an area of a mildly gassy mine were 

inadequately ventilated, it could easily attain an atmosphere in the 

explosive range. It is likely that many of the mines with explosion fatalities shown in Table 6–1 were in mildly gassy coals. 

By definition, seam gas content is the amount of gas contained in a ton of coal. It includes both 

adsorbed gases and gases in the fracture matrix. Formerly, gas content of a coal seam or the 

gassiness of a coal seam was measured by the specific emission of methane from the mine, 

expressed as the volume of methane emitted from the mine per ton of coal produced. Although a 

rough correlation exists between specific emission and actual gas content of coal, it is not very 

reliable nor can it be used effectively for forecasting.2

2

See Figure 1–5 in Chapter 1. 

Today, gas content of a coal seam is best 

measured directly [Diamond and Schatzel 1998]. If the reservoir pressure is known, an indirect 

estimate of gas content can also be obtained by Langmuir’s equation [Langmuir 1918] for monolayer adsorption: 

V = VmBP /(1+ BP) , 

where V is the estimated gas content of coal, Vm is the volume of gas for full saturation of coal, 

B is a characteristic constant of the coal seam, and P is the reservoir pressure. For U.S. coalbeds, 

the reservoir pressure is roughly correlated with the depth of the coal seam [Thakur and Davis 

1977] and is estimated at 0.303 psi/ft, or roughly 70% of the hydrostatic head. 

Since coal seams and gas in coal are formed together, it is a misnomer to call a coal seam 

nongassy. All coal seams are gassy by definition, but they vary in their degree of gassiness, 

i.e., gas content per ton of coal. The depth of a coal seam and its rank are good indicators of its 

gassiness, but direct measurement of gas content is highly recommended. 

Figure 6–1 shows the gas content of coal versus gas pressure for some U.S. coals [Kissell et al. 

1973]. Both the type of coal and the gas pressure are important. 
 Figure 6–1.—Gas content of coal versus gas pressure for some U.S. coals 

[Kissell et al. 1973]. 

 

80 

THRESHOLDS FOR COAL SEAM DEGASIFICATION 

Methane drainage must be performed when 

the ventilation air cannot dilute the methane 

emissions in the mine to a level below the 

statutory limits. 

Generally, it is economically feasible to handle specific methane emissions3

3

Specific methane emission is the total amount of gas released from the mine divided by the total amount of coal 

mined. Specific emission values are much higher than the gas content of the coal because of methane emission from 

coal that is not mined (ribs and pillars) and methane emission from strata adjacent to the mined coalbed. 

from a mine up to 

1,000 ft3

/ton with a well-designed ventilation system. At higher specific emission rates, a stage 

is reached where ventilation cost becomes excessive or it becomes impossible to stay within 
 

81 

statutory methane limits with mine ventilation 

alone. However, with a well-planned methane 

drainage system and a well-designed ventilation system, even highly gassy mines with 

specific methane emissions in excess of 4,000 

ft3

/ton can be safely operated. 

In some mines, there is often a choice regarding 

how much methane should be drained and how 

much should be handled by mine ventilation 

air. Figure 6–2 shows a generalized optimum 

point. The actual optimum point depends on a 

number of factors, including the rate of mining, 

size of longwall panel, specific methane 

emission, and cost of ventilation and methane 

drainage. 

Figure 6–2.—Generalized optimum point for 

methane drainage, shown here at 70%. Advantages of coal seam degasification can be 

summarized as follows: 

1. Reduced methane concentrations in the mine air, leading to improved safety. 

2. Reduced air requirements and corresponding savings in ventilation costs. 

3. Faster advance of development headings and economy in the number of airways. 

4. Improved coal productivity. 

5. Additional revenue from the sale of coal mine methane. 

6.	 Additional uses of degasification boreholes, such as water infusion to control respirable 

dust. 

7. Advance exploration of coal seams to locate geological anomalies in the longwall panel. 

METHANE EMISSIONS IN MINES 

Underground mining is done in two phases: (1) development and (2) pillar extraction. Development work involves the drivage of a network of tunnels (entries) into the coal seam to create a 

large number of pillars or longwall panels to be mined later. This drivage is usually done with a 

continuous mining machine. This machine cuts and loads coal into a shuttle car, which in turn 

hauls and dumps the coal onto a moving belt. The coal travels out of mines on a series of belts 

and is finally brought to the surface via a slope or shaft. Figure 6–3 shows a typical longwall 

panel layout in a U.S. coal mine. 

All methane produced during the development phase of mining is from the coal seam being 

mined. Methane is emitted at the working face as well as in the previously developed areas. All 

emitted methane is mixed with ventilation air, diluted to safe levels, and discharged on the surface. Methane drainage during or prior to development becomes necessary if the development 

headings will experience a high rate of methane emissions. This is called premining methane 

drainage. Horizontal drilling of longwall panels prior to mining also falls into this category. 
 

 

82 

Figure 6–3.—Simplified illustration of a typical longwall panel 

layout in a U.S. coal mine. Ventilation controls are not shown. 

The second phase of underground 

mining involves complete or partial 

extraction of the coal pillars. 

Smaller pillars are extracted by continuous mining machines by splitting them into even smaller pillars. 

Larger panels of coal (up to 1,000 ft 

by 10,000 ft or more) are extracted 

by the longwall method of mining. 

In either case, the mined coal 

produces methane. In addition, 

extracting these pillars or longwall 

panels causes the overlying strata to 

subside4

4

In the United States, the subsided region is called a gob. 

 and the underlying strata to 

heave. The ventilated mine workings constitute a natural pressure 

sink, into which methane flows 

from the entire disturbed area, or 

what is known as the gas emission 

space. Figure 6–4 shows the limits 

of the gas emission space as suggested by four different authors 

[Lidin 1961; Thakur 1981; Winter 

1975; Gunther and Bélin 1967]. 

The gas emission space may extend 

to 270 ft below the coal seam being 

mined and approximately 1,000 ft 

above it. 

Figure 6–4.—Limits of the gas emission space. 

The gob methane emission rate 

mainly depends on the rate of longwall advance, the geology, the size 

of the longwall panel, and the gas 

content and thickness of any coal 

seams in the gas emission space. 
83 

METHANE DRAINAGE TECHNIQUES 

The ultimate goal of coal seam degasification 

should be to reduce the gas content of the 

coal seam below 100 ft3

/ton prior to mining 

and capture at least 50%, preferably 75%, 

of the postmining emission. 

Various methane drainage techniques are used to capture the gas from the gob so that the mine 

ventilation air does not have to handle all of it. Depending on the magnitude of the problem, 

methane drainage can be performed prior to mining, known as premining methane drainage. 

Methane drainage can also be performed during mining and after the area is completely 

mined out and sealed. These two stages are generally grouped together as postmining methane 

drainage. 

Premining methane drainage. Techniques for premining drainage can be broadly classified 

into four categories: 

1. Horizontal in-seam boreholes 

2. In-mine vertical or inclined (cross-measure) boreholes in the roof and floor 

3. Vertical wells that have been hydraulically fractured (so-called frac wells) 

4. Short-radius horizontal boreholes drilled from surface 

1. Horizontal in-seam boreholes: Early work in premining methane drainage was done with 

short horizontal in-seam boreholes [Spindler and Poundstone 1960]. Figure 6–5 shows the two 

most commonly used variations of degasification with in-seam horizontal boreholes. Success of 

the technique is predicated on good coalbed permeability (≥5 mD). The horizontal drilling technique and its application to degas coal seams are well-documented in published literature 

[Thakur and Davis 1977; Thakur and Poundstone 1980; Thakur et al. 1988]. In highly permeable coal seams, e.g., the Pittsburgh Seam of the Appalachian Basin, nearly 50% of the in situ 

gas can be removed by this technique prior to mining. The major drawback of this technique is 

that only about 6 months to a year—the time between development and longwall extraction— 

is available for degasification. 

Figure 6–5.—Longwall panel methane drainage. 
84 

2. In-mine inclined or vertical boreholes: Short vertical or long inclined boreholes have been 

drilled from an existing mine (or roadways expressively driven for this purpose) to intersect 

other coal seams in the gas emission space, allowing for the seams to be degassed prior to mining. Again, success depends on high permeability. A far better way to degas these coal seams 

lying in close proximity to each other is to use vertical frac wells. 

3. Vertical frac wells: Vertical frac wells are ideally suited to highly gassy, deep, lowpermeability coal seams where it takes several years prior to mining to adequately degas the coal. 

These wells are drilled from the surface on a grid pattern over the entire property or only on 

longwall panels to intersect the coal seam to be mined in the future. 

Vertical wells drilled into the coal seam seldom produce measurable amounts of gas without 

hydraulic stimulation. High-pressure water (or other fluids) with sand are pumped into the coal 

seam to create fractures (Figure 6–6). The fluid (water) is then pumped out, but the sand 

remains, keeping the fractures open for gas to escape to the well bore. Under ideal conditions, 

if the vertical frac wells are drilled more than 5–10 years in advance of mining, 60%–70% of the 

methane in the coal seam can be removed prior to mining. 

Figure 6–6.—Premining methane drainage from surface. 

Vertical frac wells have been very successful in the Appalachian and San Juan Coal Basins of the 

United States. They have also been attempted in the United Kingdom, Germany, Poland, China, 

and Australia, but met with only limited success. Major reasons for the lack of success abroad 

are (1) cost and (2) lack of sufficient permeability, which are further explained below. 

1.	 The cost of drilling and hydrofracing a well in Europe and Australia is typically three 

times the cost in the United States. The cost of permitting and site preparation is also 

higher. In many countries, the drilling and hydrofracing equipment are not conveniently 

available. 
 

 

85 

2.	 Lower permeability (<1 mD) of many European, Asian, and Australian coal seams contributes to the limited success of frac wells. Even well-designed and well-executed frac 

jobs in the Bowen Basin of Australia were ineffective. A solution to this problem may lie 

in “gas flooding,” i.e., injection of an inert gas such as nitrogen or carbon dioxide to drive 

methane out [Puri and Yee 1990]. Increased methane production is, however, obtained 

with an increase in the inert gas content of the produced gas. This may affect the marketing of produced gas adversely. 

4. Short-radius horizontal boreholes: In coal seams with high permeability, methane drainage 

can be performed with boreholes drilled vertically from the surface and then turned through a 

short radius to intersect the coal seam horizontally. The horizontal extension can be up to 3,000 

ft. Methane then flows from the coal seam under its own pressure, as shown in Figure 6–6. The 

technique is well-proven in oil fields, but it has found a very limited application in coal mines for 

two reasons: 

•	 Cost: A short-radius borehole drilled vertically to a depth of 1,000 ft and horizontally 

extended to 3,000 ft may cost up to $500,000. 

•	 Water accumulation in the horizontal borehole: As can be seen in Figure 6–6, any water 

accumulation in the horizontal leg of the borehole will seriously inhibit gas production. 

A solution may lie in deepening the vertical leg below the coal seam being drilled and 

installing a dewatering pump in it, as is commonly done for vertical frac wells. 

Of the above four techniques, vertical frac wells have been the most effective option for premining degasification of most coal seams. Vertical frac wells also allow access to all coal seams 

in the gas emission space for predrainage. Such access becomes necessary in highly gassy mines 

in order to achieve high productivity. The only possible exception is for shallower, very 

permeable coal seams where in-mine drilling is sufficient and more economical. In shallow 

formations, the fracture system created by hydrofracing is like a horizontal pancake and is not 

very productive because the fracture system does not extend far enough from the borehole.5 

5

The ideal fracture is vertical, entirely within the coal, just a few inches wide, and extends upwards of 1,000 ft from
 

the borehole. 

oof and floor are also necessary to contain the fracture system within the coal seam. 

Recently, short-radius horizontal boreholes drilled from the surface have been used to recover 

methane from permeable coalbeds. In the future, carbon dioxide flooding may be used. 

Postmining methane drainage. Techniques for postmining drainage can be broadly classified 

into four categories: 

1. The packed cavity method and its variants 

2. The cross-measure borehole method 

3. The superjacent method 

4. The vertical gob well method 

Strong6

 r

6

A roof with rock compressive strength over 10,000 psi. 
86 

1. Packed cavity method and its variants: This technique is used mainly in Russian coal mines. 

Early methods of methane control consisted of simply isolating the worked-out area in the mine 

using packed walls, partial or complete stowing, and plastic sheets or massive stoppings. A network of pipeline that passed through these isolation barriers was laid in the gob, and methane 

was drained using vacuum pumps. Lidin [1961] reviewed several variants of this technique. 

Figures 6–7 and 6–8 show typical layouts for caving and partially stowed longwall gobs. 

Figure 6–7.—Methane drainage by the packed cavity method. 

Methane capture ratios achieved in practice are shown in Table 6–3. The ratios generally seem 

to improve in going from caving (20%–40%) to fully stowed longwall gobs (60%–80%). In 

Figure 6–7, the gate roads are protected by a packed wall against the gob. Pipelines are laid 

through the packed wall to reach nearly the center line of the gob, then manifolded to a largerdiameter pipe in the gate road. In Figure 6–8, the partially stowed longwall gob, cavities are purposely left between alternate packs. The overlying strata in the cavity area crack and provide a 

channel for gas to flow into these packed cavities. Pipelines are laid to connect the cavity with 

methane drainage mains. Methane extraction is usually done under suction. 

2. Cross-measure borehole method: This is by far the most popular method of methane control 

on European longwall faces. Figure 6–9 shows a typical layout for a retreating longwall face. 

Boreholes 2–4 inches in diameter and about 80 ft apart are drilled from the top gate to a depth of 

60–500 ft. 
87 

Figure 6–8.—Partially filled longwall gob. 

Figure 6–9.—Methane drainage with cross-measure boreholes. 

The angle of these boreholes with 

respect to horizon varies from 20° 

to 50°, while the axis of the borehole is inclined to the longwall 

axis at 15° to 30°. At least one 

hole in the roof is drilled at each 

site, but several boreholes in the 

roof and floor can be drilled at 

varying inclinations depending on 

the degree of gassiness. These 

holes are then manifolded to a 

larger pipeline system, and gas is 

withdrawn using a vacuum pump. 

Vacuum pressures vary from 4 to 

120 in w.g. 

The amount of methane captured 

by the drainage system, expressed 

as a percentage of total methane 

emission in the section, varies 

from 30% to 70%. Some typical 

data for U.K. and U.S. mines are 

given by Kimmins [1971] and 

Thakur et al. [1983], respectively, 

and are shown in Table 6–3. 

The cross-measure borehole 

method is generally more successful for advancing longwall 

panels than for retreat faces. The 

flow from individual boreholes is 

typically 20 ft3

/min, but can occasionally reach 100 ft3

/min for 

deeper holes. Sealing of the casing at the collar of the borehole is 

very important and is usually 

done with quick-setting cement. 

Sometimes a liner (a pipe of 

smaller diameter than the borehole) is inserted in the borehole 

and sealed at the collar to preserve the production from the 

borehole even when it is sheared 

by rock movements. 
Table 6–3.—Methane capture ratios for postmining methane drainage techniques 

Methane drainage technique and 

methane capture ratios Remarks 

Packed cavity method (after Lidin [1961]):

 20%–40%..............................................

 30%–50%..............................................

 60%–80%..............................................

Cross-measure boreholes (after

 Kimmins [1971]):

 59%–70%..............................................

 

Superjacent method: 

 50% .......................................................

Vertical gob wells: 

 30%–80%..............................................

 Caving longwalls.

 Partially stowed longwalls.

 Fully stowed gobs.
 

 Highly gassy mines with specific emissions 3,000–6,000 

ft3

 per ton.
 

 For multiple coal seams in the gas emission space. 

 The methane capture efficiency depends on the number of 

 gob wells per longwall panel and production techniques.
 

88 

3. Superjacent method: This method was used mainly for retreating longwall faces in highly 

gassy seams in French mines. Figure 6–10 shows a typical layout. A roadway is driven 70–120 

ft above the longwall face, preferably in an unworkable coal seam. The roadway is sealed, and 

vacuum pressures up to 120 in w.g. are applied. To improve the flow of gas, inclined boreholes 

in the roof and floor are drilled to intersect with other gassy coalbeds. If the mining scheme proceeds from the top to the bottom seams in a basin, the entries in a working mine can be used to 

drain coal seams at lower levels. Methane flow from these entries is high, averaging 700–1,000 

ft3

/min for highly gassy seams. Nearly 50% of total emissions have been captured using this 

method. 

4. Vertical gob well method: This technique, most commonly used in longwall mining in the 

United States, is relatively new and differs from European systems in several ways. U.S. coal 

seams are generally thin, shallow, and relatively more permeable. Typically, only one seam is 

mined in a given area and retreat longwall mining is the only method being practiced at present. 

Methane emission rates from gobs in various coal basins vary depending on the geological conditions, but deep-seated longwall gobs (e.g., those in the Pocahontas No. 3 Seam in Virginia and 

the Mary Lee Seam in Alabama) produce methane in the range of 1,800–18,000 ft3

/min. Multiple entries (typically four) are driven to develop longwall panels so that necessary air quantities 

can be delivered to the longwall faces via the mine ventilation system. In many cases, however, 

some type of additional methane control becomes necessary. 

The most popular method of methane control is to drill vertical boreholes above the longwall 

prior to mining, as shown in Figure 6–11. Depending on the length of the longwall panel (typically 10,000 ft) and the rate of mining, 3 to 30 vertical gob degas boreholes are needed. The first 

hole is usually within 150–500 ft of the start line of the longwall face. The borehole is drilled to 

within 30–90 ft from the top of the coal. The casing is cemented through the fresh water zones 

near the surface, and a slotted liner is provided over the lower open section to prevent closing of 

the hole by caving. These boreholes are completed prior to mining. Usually, no measurable 

methane production is realized until the longwall face mines past the borehole. 
Figure 6–10.—Methane drainage by the superjacent method. 

 Figure 6–11.—Simplified illustration of meth ane drainage by vertical gob wells. Ventilation controls are not shown. 

89 
 

90 

Early experiences with this method of gob degasification have been described by Moore et al. 

[1976] for the Lower Kittanning Seam, by Moore and Zabetakis [1972] for the Pocahontas Seam, 

and by Davis and Krickovic [1973] and Mazza and Mlinar [1977] for the Pittsburgh Seam. 

Many gob degasification boreholes produce naturally when the longwall face intersects them, 

but vacuum pumps are often added to further improve the flow and, in some cases, to prevent the 

reversal of flow. The capture ratios vary from 30% to 80% depending on the number and size of 

gob wells per panel and the size of vacuum pumps. 

A summary of methane capture ratios for the abovementioned postmining methane drainage 

techniques is presented in Table 6–3. Although each technique offers high capture efficiency in 

some cases, it is the author’s experience that vertical gob wells, if properly designed, offer the 

most universal application with consistently high capture ratios. In addition, this technique is a 

natural outgrowth of the premining degasification technique using vertical frac wells. These frac 

wells can be converted easily into postmining gob wells with minimal additional expense. 

HOW TO TRANSPORT GAS IN UNDERGROUND MINES 

In-mine horizontal drilling and cross-measure boreholes drilled to degas longwall gobs produce 

large volumes of gas. This gas must be conducted out of the mine without being allowed to mix 

with the mine ventilation air. The U.S. coal industry, working with the Mine Safety and Health 

Administration (MSHA), has developed general7

7

The specifics will vary from mine to mine. 

 guidelines for installing and operating underground methane pipelines, as follows: 

1. Underground methane pipeline will be made of well-designed plastic or steel, as detailed in 

Figure 6–12. 

a.	 All underground steel pipelines will be 31⁄2- to 81⁄2-in O.D. schedule 40 pipes joined 

together with threaded couplings. These pipes will be made up tightly using a good 

grade of thread lubricant. Mill collars will be broken out, doped, and remade. 

A flange connection will be used every 10 joints (approximately 210 ft apart) so that 

a section of the pipeline can be removed without cutting the line if one or more joints 

need to be replaced later. 

b.	 All underground plastic line will be 3- to 6-in high-density polyethylene pipe. Plastic 

flange adapters will be fusion bonded to the pipe ends in fresh air. Steel flange 

backup rings installed prior to fusion bonding will be used to connect plastic to plastic 

and plastic to steel. 

2. The entire length of pipeline between the bottom of the venthole and the well head will be 

pressure tested to 1.25 times the shut-in pressure of the borehole or 90 psi, whichever is greater. 

3. Pipeline will be generally laid in the return airway and will not be buried. Whenever the 

pipeline must cross a fresh air entry, it will be conducted through a steel line. 
Figure 6–12.—Underground methane pipe installation. 

91 

4. No hoses will be used in the system, except while a hole is being drilled. Stress-relieving 

flexible tubing will be used at critical points, such as the head-to-pipeline connection. This will 

be stainless steel tubing with a triple wire braid cover. 

5. The steel pipeline will be firmly supported, with no unsupported span greater than 2 ft. 

6. A gas water separator will be installed at the bottom of the vertical venthole to remove condensation that falls back down the casing. Other separators will be installed on the holes or 

on the pipeline if water production from coal warrants. All separators will preferably be commercially made. Water drains will be provided on the line wherever necessary. 

7. If steel pipeline is used, a potential survey will be made and cathodic protection provided 

where needed. 

8. Automatic shut-in valves will be installed at each well head. These will be held open by 

nitrogen or air under pressure contained in a fragile plastic pilot line running parallel to and 

secured on top of the pipeline. Any roof fall or fires serious enough to damage the pipeline will 

damage the pilot line first and close the boreholes immediately. 

9. The pipeline system will be inspected weekly by a competent person familiar with system 

operation. 

10. If the quantity of mine ventilation air flowing over the pipeline is such that a complete 

rupture of the pipeline and consequent discharge of methane in mine air will raise its concentration above the legal limits, a methane monitor will be used. 
92 

11. At the surface installation (Figure 6–13), a commercially made flame arrestor will be 

installed within 10 ft of the top of the vent stack. A check valve shall be used to guard against 

reversal of flow. The check valve can be manually defeated if it is desired to purge the pipeline 

for repairs. An orifice meter may be installed if needed. All surface installations will be periodically inspected to ensure satisfactory performance. 

12. All boreholes drilled for degasification will be accurately surveyed either during or after 

drilling is completed using commercially available borehole surveying tools. These boreholes 

will be accurately plotted on mine maps to prevent any inadvertent mining through them. 

13. Should an occasion arise in the future when cutting into an abandoned, unplugged borehole 

will be necessary, a detailed mining plan will be submitted to MSHA. 

14. A compressor will be required at the surface if beneficial use is made of the gas at a future 

date. Plans for installation will be discussed with MSHA at the appropriate time. 

Figure 6–13.—Surface installation for horizontal drainage boreholes. 
 

93 

ECONOMICS OF COAL SEAM DEGASIFICATION 

The economics of coal seam degasification depend on— 

1. The gas contents of the coal seam mined and the 

other coal seams contained in the gas emission space. 

2. The fraction of the mine methane emissions captured. 

3. Infrastructure costs and the market price of the 

processed gas. 

In general, unless the specific methane emission from the mine (cubic feet of methane per ton 

of mined coal) is high (above 3,000 ft3

/ton), it may not be profitable to process the gas for 

marketing. The cost of compressing and processing the coal mine methane and a complete 

economic analysis to reflect rates of return on the investment is beyond the scope of this chapter. 

A rough estimate of costs associated with coal seam degasification can be derived, however, 

as shown here. 

For all underground longwall mining, a generalized scheme of degasification depending on the 

gassiness of the coal seam has been proposed by Thakur and Zachwieja [2001]. The following 

assumptions were made: 

1.	 The longwall panel is 1,000 ft wide and 10,000 ft long. 

2.	 The coal seam has an average thickness of 6 ft. 

3.	 The coal block to be degassed is 1,300 ft by 10,000 ft, assuming that the width of chain 

pillars is 300 ft. 

4.	 The cost of contract drilling for the in-mine horizontal drilling is $50/ft. 

5.	 The cost of a gob well is $50,000–$200,000, depending on the depth of the mine and 

 the size of the borehole. 

6.	 The cost of hydrofracing a well is $250,000. 

If the total cost of in-mine drilling, including all of the underground pipeline costs, all vertical 

frac wells, and all other gob wells, is added and then divided by the tons of coal in the longwall 

block, the result is the cost of coal seam degasification per ton of coal. 

• Estimated cost for mildly gassy coal seams less than 100 ft3

/ton (see Table 6–2): 

Premining degasification: For coal seams with gas contents less than 100 ft3/ton,8

8

Remember that this figure represents the gas content of the coal, not the specific mine emission. 

 there is generally no need for premining degasification. 

Postmining degasification: Two gob wells are recommended for the longwall panel. The first 

gob well should be installed within 1,000 ft of the setup entry and the second one in the middle 

of the panel. 

The total cost is $100,000, or $0.03/ton. 
 

94 

• Estimated cost for moderately gassy coal seams, 100–300 ft3/ton (see Table 6–2): 

Premining degasification: The longwall panel should be drilled horizontally at 1,000-ft intervals, 

and development boreholes should be drilled to degas development sections. Total in-mine drilling footage for a typical panel may total 25,000 ft. 

Postmining degasification: In moderately gassy coal seams, a proposed longwall panel may need 

five to six gob wells. The diameter and size of exhaust fans will depend on local conditions. 

The total cost is approximately $1.55 million, or $0.50/ton. 

• Estimated cost for highly gassy coal seams over 300 ft3

/ton (see Table 6–2): 

Premining degasification: Highly gassy coal seams must be drained several years ahead of mining with vertical frac wells (wells that have been hydraulically fractured). These frac wells can 

be placed at about a 20-acre spacing. Frac wells drilled about 5 years ahead of mining can drain 

nearly 50% of the in situ gas prior to mining, but this may not be sufficient. Additional degasification with in-mine horizontal drilling can raise the gas drained to nearly 70%. Horizontal 

boreholes are drilled 200–300 ft apart to a depth of 900 ft. Assuming a 200-ft interval, nearly 

45,000 ft of horizontal drilling and about 15 vertical frac wells may be needed to properly degas 

the panel. 

Postmining degasification: Because of very high gas emissions from the gob, the first gob well 

must be installed within 50–100 ft from the setup entry. Subsequent gob wells may be drilled at 

a 6- to 15-acre spacing, depending on the rate of mining and the gas emission per acre of gob. 

In Virginia and Alabama, which have some highly gassy coal seams, gob wells are generally 

9–12 inches in diameter. Powerful exhaust fans capable of a suction of 5–10 inches of mercury 

are needed to capture up to 70% of gob gas emissions. 

The total cost of degasifying a longwall panel in a highly gassy coal seam is approximately 

$11 million, or $3.52/ton. Coal seam degasification is needed for mine safety and high productivity, but in highly gassy mines it becomes quite expensive. In these mines, the processing and 

marketing of coal mine methane becomes necessary to defray the cost. 

REFERENCES 

Creedy DP, Saghafi A, Lama R [1997]. Gas control in underground coal mining. IEACR/91. 

London: IEA Coal Research. 

Davidson RM, Sloss LL, Clarke LB [1995]. Coalbed methane extraction. IEACR/76. London: 

IEA Coal Research. 

Davis JG, Krickovic S [1973]. Gob degasification research: a case history. In: Methane control 

in eastern U.S. coal mines. Proceedings of the Symposium of the Bureau of Mines/Industry 

Technology Transfer Seminar, Morgantown, WV, May 30–31, 1973. Washington, DC: U.S. 

Department of the Interior, Bureau of Mines, IC 8621, pp. 62–72. 
95 

Diamond WP [1994]. Methane control for underground coal mines. Pittsburgh, PA: U.S. 

Department of the Interior, Bureau of Mines, IC 9395. NTIS No. PB94189289. 

Diamond WP, Schatzel SJ [1998]. Measuring the gas content of coal: a review. Int J Coal Geol 

35:311–331. 

Gunther J, Bélin J [1967]. Forecasting methane emission at faces in flat seams. In: Proceedings 

of the 12th International Conference of Mine Safety Research Establishments (Dortmund, Germany, September 11–15, 1967). Vol. 3. Paper No. 45. Detmold, Germany: Hermann Bösmann 

GmbH. 

Hargraves AJ [1973]. Planning and operation of gaseous mines. CIM Bull, March. 

Kimmins EJ [1971]. Firedamp drainage in the northwestern area. Colliery Guardian, Annual 

Review, pp. 39–45. 

Kissell FN, McCulloch CM, Elder CH [1973]. The direct method of determining methane content of coalbeds for ventilation design. Pittsburgh, PA: U.S. Department of the Interior, Bureau 

of Mines, RI 7767. 

Langmuir I [1918]. The adsorption of gases on plane surfaces of glass, mica, and platinum. 

J Am Chem Soc. 

Lidin GD [1961]. Control of methane in coal mines. Washington, DC: Israel Program for Scientific Translations. 

Mazza RL, Mlinar MP [1977]. Reducing methane in coal mine gob areas with vertical 

boreholes. Continental Oil Co. U.S. Bureau of Mines contract No. H0322851. NTIS No. 

PB272768. 

Moore TD Jr., Zabetakis MG [1972]. Effect of a surface borehole on longwall gob degasification (Pocahontas No. 3 coalbed). Pittsburgh, PA: U.S. Department of the Interior, Bureau of 

Mines, RI 7657. 

Moore TD Jr., Deul M, Kissell FN [1976]. Longwall gob degasification with surface ventilation 

boreholes above the lower Kittanning coalbed. Pittsburgh, PA: U.S. Department of the Interior, 

Bureau of Mines, RI 8195. 

Puri R, Yee D [1990]. Enhanced coalbed methane recovery. In: Proceedings of the 65th Annual 

Technical Conference and Exhibition of the Society of Petroleum Engineers (New Orleans LA, 

September 23–26, 1990). Richardson, TX: Society of Petroleum Engineers, pp. 193–202. 

Spindler GR, Poundstone WN [1960]. Experimental work in the degasification of the Pittsburgh 

coal seam by horizontal and vertical Drilling. Preprint No. 60F106. Littleton, CO: American 

Institute of Mining, Metallurgical, and Petroleum Engineers. 
96 

Thakur PC [1981]. Methane control for longwall gobs. In: Ramani RV, ed. Longwall-shortwall 

mining: state of the art. Littleton CO: American Institute of Mining, Metallurgical, and 

Petroleum Engineers, pp. 81–86. 

Thakur PC, Davis JG [1977]. How to plan for methane control in underground coal mines. 

Min Eng Oct:41–45. 

Thakur, PC, Poundstone WN [1980]. Horizontal drilling technology for advance degasification. 

Min Eng Jun:676–680. 

Thakur PC, Zachwieja J [2001]. Methane control and ventilation for 1,000-ft wide longwall 

faces. In: Proceedings of Longwall USA, International Exhibition and Conference (Pittsburgh, 

PA, June 13–15, 2001), pp. 167–180. 

Thakur PC, Cervik J, Lauer SD [1983]. Methane drainage with cross-measure boreholes on a 

retreat longwall face. Preprint No. 83–398. Littleton, CO: American Institute of Mining, Metallurgical, and Petroleum Engineers. 

Thakur PC, Christopher DA, Bockhorst R W [1988]. Horizontal drilling technology for coal 

seam methane recovery. In: Gillies ADS, ed. Proceedings of the Fourth International Mine 

Ventilation Congress (Brisbane, Queensland, Australia, July 3–6, 1988), pp. 201–207. 

Winter K [1975]. Extent of gas emission zones influenced by extraction. In: Proceedings of the 

International Conference on Coal Mine Safety Research, Washington DC), pp. V3.1–V3.17. 
CHAPTER 7.—MANAGING EXCESS GAS EMISSIONS ASSOCIATED 

WITH COAL MINE GEOLOGIC FEATURES 

 

97 

By James P. Ulery1 

1

Research geologist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA. 

In This Chapter 

9 Geologic features associated with anomalous methane emissions 

9 Gas outbursts and blowers 

9 Methane drainage strategies for mitigating anomalous methane emissions 

and 

9 General considerations for a methane drainage program 

This chapter summarizes how certain geologic features may be associated with unexpected 

increases in gas emissions during coal mining. These unexpected emissions have the potential to 

create explosive conditions in the underground workplace. Also discussed are the generally used 

practices to alleviate potential hazards caused by gas emissions associated with these geologic 

features. 

INTRODUCTION 

Unforeseen mine gas emissions in quantities sufficient to create hazardous conditions have been 

attributed to sources outside the mined coalbed since the first documentation of methane explosions in coal mines [Payman and Statham 1930]. Geologic features such as faults have long 

been recognized as conduits for gas flow from strata adjacent to mined coalbeds [Moss 1927; 

Payman and Statham 1930]. Other features such as sandstone paleochannels, clay veins, and 

localized folding have also been recognized for their impact on gas emissions into mine workings [Darton 1915; Price and Headlee 1943; McCulloch et al. 1975; Ulery and Molinda 1984]. 

The fact that strata adjacent to mined coalbeds can emit large quantities of methane gas into mine 

workings is not surprising from a theoretical perspective. Many researchers have recognized that 

during the burial and diagenesis of the organic matter forming today’s minable coalbeds, similar 

dispersed organic matter in adjacent strata has produced methane in quantities far exceeding the 

storage capacity of the coal and surrounding rock [Juntgen and Klein 1975]. It is not surprising 

then that large quantities of methane can remain trapped in these strata. A potential hazard 

occurs when mining of a nearby coalbed causes pressure differentials and mining-induced fractures conducive to gas flow into the mine workings from these strata. This gas flow may be 

facilitated or temporarily impeded by the presence of geologic structures or anomalies. 
98 

Anomalous, unanticipated methane emissions 

are often related to the interception of geologic 

features, such as paleochannels, faults, and 

clay veins, by mine workings. Furthermore, 

these features can also contribute methane 

emissions before being intercepted by mining. 

For the purpose of discussion in this chapter, gas emissions associated with geologic features are 

divided into two categories. The first includes subtle emission events that are often associated 

with various geologic features or anomalies. These emissions are often not easily detected without instrumentation, but may lead to hazardous accumulations of methane if not remedied. The 

second category includes large-scale, easily recognizable emission events such as blowers or 

outbursts that potentially have immediate and often catastrophic consequences. Documented 

methods to recognize and remedy both types of hazards have been established worldwide and are 

discussed here. 

This chapter summarizes current technologies for recognizing and remediating gas emission 

hazards associated with various geologic features. Although emphasis is placed on recognized 

hazards in U.S. coal mines, hazards not well-documented in these mines, such as gas outbursts, 

are also discussed due to the potential for such occurrences in the future as mines extract deeper 

and gassier seams. 

INCREASED GAS EMISSIONS ASSOCIATED WITH GEOLOGIC FEATURES 

In the United States, gas emission events associated with geologic features constitute a fairly 

common hazard in coal mining. These events are neither as obvious nor as immediate as outbursts or blowers, which are discussed later. However, they can pose significant risks. These 

emission events are often difficult to detect without instrumentation and underground surveys. 

Detailed mapping of geologic features can 

assist in predicting potential emission hazards 

and designing methane drainage systems to 

prevent them. 

This section will consider techniques to detect and remediate anomalous gas emissions associated with geologic features such as sandstone channels/lenses, adjacent source beds, clay veins, 

joints, fractures, and small-scale faulting. Also included is a discussion of igneous intrusions and 

their potential impact on gas storage and emissions. 

Sandstone channels/lenses. Sandstone paleochannel deposits or other lenticular sandstone 

deposits adjacent to mined coalbeds historically have been documented as gas reservoirs [Darton 

1915; Price and Headlee 1943]. The gas may have been generated in situ from organic matter 
 

99 

deposited with the sand, or it may have migrated to the paleochannel from subjacent coalbeds or 

other organic-rich rock strata. These gas-bearing sandstones generally have a greater permeability than other rock strata in coal mining areas. Once a gas flow pathway is established to the 

mine workings, via either the relaxing of naturally occurring joints, clay veins, faults, or mininginduced fractures, then gas emissions from these sandstones may be quite pronounced and often 

produces hazardous conditions. 

Methane emissions from sandstone paleochannels is a well-known problem in the Pittsburgh 

Coalbed in Pennsylvania and West Virginia [McCulloch et al. 1975]. At a mine in northern 

West Virginia, gas was documented to be migrating from a sandstone channel to mine workings 

through clay vein-related fractures [Ulery and Molinda 1984]. The sandstone paleochannel 

deposit above the Pittsburgh Coalbed was a significant gas reservoir that was still flowing 

370,000 cfd of gas into a methane drainage system 2 years after it was installed. 

Delineating problematic sandstone bodies is best accomplished by an exploratory core drilling 

program of sufficiently close spacing to accurately delineate the extent and trend of the sand unit 

[Houseknecht 1982]. Also, evaluating the gas content/flow potential of sandstone units is necessary to determine if emissions will be of sufficient volume to pose a potential hazard. This can 

be accomplished by laboratory testing of core samples for reservoir properties such as porosity 

and permeability, well testing, or direct gas measurements underground in the vicinity of the 

feature [Diamond 1994]. 

When sandstone bodies, either in floor or roof strata, are encountered that pose a gas emission 

hazard, remediation may be accomplished by vertical surface methane drainage boreholes 

(Figure 7–1, borehole A). However, in-mine methane drainage boreholes may be the most economically viable solution. In Figure 7–1, boreholes B, C, and D show vertical, in-mine crossmeasure, and horizontal degasification borehole configurations, respectively, for methane 

drainage of a gas-bearing paleochannel deposit in the roof strata. The same configurations of 

surface or in-mine boreholes would also be applicable to gas-bearing sandstone deposits in 

floor strata. 

Figure 7–1.—Methane drainage scenarios for paleochannels. 
100 

Vertical methane drainage boreholes drilled from the surface usually require both dewatering and 

hydraulic fracturing for maximum effectiveness. Horizontal, cross-measure, or vertical underground degasification boreholes may also need to remove water to effectively produce gas. 

These methane drainage systems are fully explained and referenced by Diamond [1994]. 

Gas-bearing paleochannel deposits or 

adjacent coalbeds may require surface or 

cross-measure methane drainage boreholes. 

Coalbeds and other adjacent gas source beds. Historical mining experience and research have 

shown that coalbeds adjacent to the mined seam can contribute significant quantities of methane 

gas into active workings [Finfinger and Cervik 1979; Ayruni 1984; Diamond et al. 1992]. 

Degasification of these coalbeds may be accomplished through vertical or directional boreholes 

drilled from the surface. Occasionally, in-mine vertical, cross-measure, or directional boreholes 

are used to reduce potentially explosive accumulations of gas [Finfinger and Cervik 1979; 

Ayruni 1984; Diamond 1994]. 

In addition to nearby coalbeds, other adjacent strata may also be significant gas reservoirs and 

contribute unexpected emissions into mine workings. Shales and siltstones rich in organic matter 

often contain significant quantities of methane gas [Darton 1915; Johnson and Flores 1998]. 

Although these formations may have a large gas storage capacity, permeability is usually very 

low. Thus, these rocks may not release gas until mining-induced fractures increase their permeability and provide a pathway for gas migration to the mine workings. In studying the gas content of U.K. coalbeds, Creedy [1988] concluded that although the gas contents and porosities of 

these rocks are low, a well-developed joint or fracture system could facilitate gas release from 

these strata. Therefore, it may be assumed that if certain organic-rich rocks adjacent to mined 

coalbeds have sufficient gas content, migration via joints or fractures into the workings could 

generate hazardous explosive conditions. Methane drainage methods similar to those discussed 

previously for adjacent coalbeds are usually appropriate for adjacent noncoal gas-bearing strata 

as well. 

Large- and small-scale structural faulting. For this discussion, large-scale faults are loosely 

defined as having tectonically activated and structurally mappable features, with lengths greater 

than 500 m (1,640 ft) and vertical movement of at least 10–20 m (33–66 ft). Small-scale faults 

are distinguished from large-scale faults by their limited extent both horizontally and vertically. 

Faults may have a profound effect on gas emissions into mine workings and may also be associated with outbursts and blowers. Usually, the presence of large-scale faulting is known from 

regional geologic mapping and/or exploration boreholes. In certain cases, these large faults may 

act as barriers to gas flow, especially if they contain impermeable fault gouge or the displacement causes impermeable rock above or below the mined coalbed to abut against it [Diamond 

1982]. In these situations, large volumes of gas can be trapped behind the fault at pressures 

higher than that of the mine atmosphere. If mine development proceeds through the fault by 

ramping upward or downward into the displaced, high-pressure coalbed gas reservoir, the potential for sudden, excess gas emissions must be addressed. 
101 

Large-scale faults may also act as conduits for gas flow or blowers into the mine workings from 

gas-enriched strata above or below the mined coalbed. This is especially likely due to stress 

redistributions as mining approaches a large-scale fault. In Germany, Thielemann et al. [2001] 

showed that in nonmined regions, normal faults regularly act as gas conduits for surface emissions into the atmosphere from deep (60–870 m (197–2,854 ft)) formations such as coalbeds. 

Thielemann et al. further demonstrated that distinctly higher surface gas emission rates occurred 

from normal faults in mined areas, presumably caused by the increased permeability of the fault 

and associated strata in response to mining. Therefore, it would seem likely that such faults 

could easily become pathways for gas emissions into mine workings from adjacent source beds. 

In the United States, Clayton et al. [1993] noted similar findings in the Black Warrior Basin. 

Methane drainage from potential gas problem areas associated with large-scale faulting may best 

be accomplished through surface boreholes if the faulted areas in question are well mapped. 

In lieu of this, unexpected problems associated with large-scale faulting may be alleviated by 

underground cross-measure boreholes designed to penetrate the fault zone and/or the gas 

source bed. 

Small-scale faults have limited lateral extent and are often vertically confined to one or two 

strata layers. In coal mining districts, small-scale faults are often, but not exclusively, related to 

differential sediment compaction phenomena. Examples of these faults are illustrated by 

Iannacchione et al. [1981]. 

Little documentation is available on the effects of small-scale faults on gas emissions. However, 

based on descriptions by Iannacchione et al. [1981], it would seem reasonable to conclude that 

these types of faults, if they have any effect at all, could possibly act as more limited barriers to 

gas migration compared to the previously discussed large-scale faults. Prediction of these smallscale features can be difficult, even with detailed underground mapping. The coalbed near these 

features often displays abnormal thickening, undulations, or pulverizing, which may indicate that 

these types of faults are being approached. However, if small-scale faults are encountered and 

found to adversely influence gas emissions, degasification through short horizontal or crossmeasure-type boreholes ahead of the working face is feasible if these faults can be mapped 

and/or anticipated. 

Whether large- or small-scale, displacement faults are generally of two basic types: normal or 

reverse. Small-scale normal faults are often associated with differential compaction phenomena 

near sandstone channels. Large-scale normal faulting is often associated with regional uplifts 

and/or deep plutonic activity. Reverse faults, on the other hand, are often associated with 

mountain-building tectonics and regional compressional forces. Low-angle reverse faults are 

termed “thrust faults” and are often very large-scale regional features. Because normal faults are 

usually associated with tensional forces and reverse faults with compressional forces, they would 

seem most likely to act as gas conduits and barriers, respectively. However, there is no conclusive documentation to that effect. 

Geologic mapping of faults is needed to determine 

the most efficient gas drainage system. 
 

 

102 

Figure 7–2 shows a mine entry approaching a normal fault on the “footwall” side. The remaining coal reserve ahead of (and below, due to the fault) the approaching entry often poses an 

emission hazard, i.e., although the normal fault may potentially be a gas conduit, until mining 

redistributes stresses, it is not generally an open conduit for gas flow as is the normal coal cleat 

system. Therefore, when the entry is ramped downward to mine the remaining reserve, hazardous conditions may occur when the gas trapped behind the fault is suddenly released into the 

mine entry. Optimum degasification of such potential hazards is best accomplished via vertical 

methane drainage boreholes drilled from the surface (Figure 7–2, borehole B). Alternatively, 

directional methane drainage boreholes from the mine entry (Figure 7–2, borehole A) could 

be used. 

Figure 7–2.—Methane drainage of a normal fault from the 

“footwall” side. 

If a normal fault is encountered from the lower “hanging” wall side (Figure 7–3), vertical 

methane drainage boreholes drilled from the surface (Figure 7–3, borehole A) are probably the 

only viable method to remediate the hazard due to the geometry of this condition. 

Figure 7–3.—Methane drainage of a normal fault from the 

“hanging” wall side. 

Reverse faults tend to form by 

compressional forces and therefore 

may often act as barriers to flow, 

causing gas buildup behind them. 

Figure 7–4 shows a mine entry 

approaching a reverse fault on the 

“hanging” wall side. In most 

cases, but especially if the faulting 

is large-scale, vertical methane 

drainage boreholes drilled from the 

surface (Figure 7–4, borehole C) 

are probably the most viable 

option to alleviate potential emission hazards. Other options 

include vertical in-mine boreholes 

(Figure 7–4, borehole A) or 

directional in-mine boreholes 

(Figure 7–4, borehole B). Reverse 

faults, where mining approaches 

from the “footwall” side, are optimally addressed with in-mine 

cross-measure-type boreholes 

(Figure 7–5, borehole A) or 

vertical methane drainage boreholes drilled from the surface 

(Figure 7–5, borehole B). 
 

 

103 

Figure 7–4.—Methane drainage of a reverse fault from the 

“hanging” wall side. 

Figure 7–5.—Methane drainage of a reverse fault from the 

“footwall” side. 

Joints, cleats, and fractures. 

Joints, cleats, and fractures are 

ubiquitous features in most coal 

measure rocks and are related to 

the confining stress fields acting 

upon those strata during burial, 

diagenesis, and uplift. Generally, these features follow a 

systematic pattern. Joints are 

closely spaced with even walls, 

whereas fractures are more 

widely spaced with irregular 

walls [Nickelsen and Hough 

1967]. Joints usually occur in 

orthogonal pairs (at an approximately 90° orientation to each 

other), and in coalbeds, these 

joint sets are referred to as 

“cleats.” 

The main joints and cleats of any 

given set are generally more 

continuous and are the dominant 

migration pathways for gas 

[McCulloch et al. 1974]. They 

are referred to as “systematic 

joints” and “face cleats,” respectively. The corresponding joints 

and cleats at 90° to the main 

features are referred to as 

“nonsystematic joints” and “butt 

cleats,” respectively. These 

joints and cleats generally terminate against the systematic joints 

and face cleats, making them 

notably less continuous. When 

coalbeds are mined, the redistribution of stresses allows cleats to expand and facilitates gas 

migration through the coal to the face. Similarly, the stress redistribution opens joints and 

fractures in roof and floor strata, facilitating gas migration from adjacent strata. 

The ubiquitous nature of joints and the unpredictable spacing of fractures make prediction and 

remediation of abnormal gas emissions related to these features difficult. It is important for 

operators to realize that although joints and fractures may contain free gas at the face, their real 

hazard potential is as a conduit for unexpected gas flows to the mine workings from within the 

coalbed and/or other source beds adjacent to the mined coalbed. These types of conditions are 

most often recognized when the continuous miner or longwall shearer is deenergized due to the 

machine-mounted methane sensors reading concentrations above the allowable limits. The most 
104 

obvious solution to this problem would be to increase the face ventilation, if not already at the 

practical limit. 

If past experience indicates that excess emissions 

may be encountered in a developing section, the 

mine operator should consider an underground 

horizontal borehole methane drainage system or 

vertical methane drainage boreholes drilled from 

the surface to remove gas prior to mining. Mine 

operators should always be aware of regional joint 

and fracture trends to facilitate prediction and 

remediation of abnormal emissions associated with 

joints and fractures. 

Clay veins. Clay veins or clastic dikes (Figure 7–6) are sedimentary intrusions, usually from 

overlying strata, that invade the coal in a vertical or near-vertical orientation. Their appearance 

in cross-section is not unlike an igneous dike, hence their name. Clay veins tend to be systematic 

in occurrence and are often related to differential compaction/diagenetic processes, but can also 

be influenced by tectonic processes. 

Figure 7–6.—Typical clay vein acting as a barrier to gas migration. 

These features, their characteristics, and modes of formation have been widely documented in 

U.S. coal mines [Chase and Ulery 1987]. Clay veins, which are generally composed of very 

fine-grain sediment or clay, 

are virtually impermeable 

barriers to gas migration in 

coalbeds. Therefore, they 

tend to have a “damming” 

effect on gas flow when 

approached in a developing 

section. When a continuous 

miner or longwall shearer 

penetrates a clay vein with 

gas trapped behind it, the 

abnormally high emissions 

may cause production interruptions due to methane 

concentrations above the 

legal limit, and in a worstcase scenario, an explosive 

methane/air mixture could 

ignite. 
105 

Clay veins have a well-documented history of causing unexpected high gas emissions in the 

Pittsburgh Coalbed during mining [McCulloch et al. 1975]. Prosser et al. [1981] measured 

increases in gas flow from 47,000 to 80,000 cfd when a horizontal in-mine methane drainage 

borehole penetrated a clay vein approximately 800 ft from the face. For a different horizontal 

borehole in the same study, the gas flow increased from 144,000 to 214,000 cfd when a clay vein 

was penetrated approximately 2,175 ft from the face. High gas flow rates (>80,000 cfd) from 

low-angle cross-measure boreholes penetrating clay veins near a gas-bearing sandstone above 

the Pittsburgh Coalbed have been documented in northern West Virginia [Ulery and Molinda 

1984]. 

In-mine horizontal boreholes can effectively 

drain gas trapped behind clay veins in advance 

of mining. 

Only experience at a given mine can indicate to the operator if clay veins have the potential for 

gas emission problems. If methane emission problems are encountered, then underground mapping of clay veins is needed to predict where they will occur in developing sections. Because 

clay veins can frequently extend hundreds of feet along a given trend [Chase and Ulery 1987], 

predicting their occurrence in a developing section will allow the operator to anticipate and/or 

alleviate the potential gas problem. 

If the clay vein network in a specific mine has a history of gas emission problems, then horizontal methane drainage boreholes drilled ahead of the face and penetrating the clay vein 

are probably the most economical and timely method to remediate the potential problem 

(Figure 7–7, borehole A). Theoretically, if a clay vein network is well-mapped and a large, 

isolated “cell” is delineated (Figure 7–8), then the gas could be drained through a surface borehole (Figure 7–7, borehole B). Generally, however, this would be a cost-prohibitive course of 

action, necessitated only in extreme cases or where drilling costs would be low due to shallow 

depths. 

Figure 7–7.—Methane drainage of a clay vein gas barrier. 
Figure 7–8.—Hypothetical gas cell formed by clay veins (adapted from Chase and Ulery [1987]). 

106 

Igneous intrusions. Igneous intrusions into coalbeds and coal measure rocks are not frequent, 

but not uncommon either. Igneous intrusions into coal measure strata generally will be either 

discordant features such as dikes, which cut across bedding planes, or concordant features such 

as sills, which are injected parallel to bedding. Massive discordant features such as plutons are 

rare in coal measures. 

Because igneous intrusions involve magmatic rock injected at elevated temperatures, they cause 

an alteration and increase in the thermal maturity (rank) of nearby coalbeds and organic matter in 

rocks [Dutcher et al. 1966]. The degree of coal alteration caused by an igneous intrusion 

depends on many factors, including the intrusion’s temperature, thickness, distance from the coal 

seam, and cooling rate. Such thermal alteration of organic matter is accompanied by methane 

gas generation. Therefore, for a given coalbed, localized areas affected by igneous intrusions 

can be expected to have higher gas contents than normal [Gurba and Weber 2001]. Larger igneous intrusions such as sills may also be responsible for potentially high carbon dioxide concentrations in some coal basins [Clayton 1998]. 

Discordant igneous dikes, which cut across the coalbed, may not only be expected to increase the 

in situ gas content due to thermal alteration, they can also act as a barrier or “dam” to gas 
107 

migration and present hazards similar to clastic dikes when mined through. Predicting the location and orientation of igneous dikes in developing sections is best accomplished by detailed 

underground mapping in adjacent developed sections. Remediating potential gas emission hazards associated with igneous dikes, as with clastic dikes, is best accomplished by horizontal boreholes drilled from the face to penetrate the dike. 

Concordant igneous features such as sills usually cover a far greater area than dikes and can elevate the thermal maturity and gas content of a coalbed over a similarly large area [Gurba and 

Weber 2001]. However, the greater extent of these features is more conducive to prediction and 

mapping through conventional exploratory core drilling programs. If associated gas contents and 

emissions are expected to present a potential hazard, premining methane drainage through 

vertical methane drainage boreholes drilled from the surface is the optimal method to alleviate 

the hazard. 

GAS OUTBURSTS AND BLOWERS 

 

Although not generally considered to be hazards in domestic mines at present, both outbursts and 

blowers historically have occurred in certain U.S. mining districts [Darton 1915; Campoli et al. 

1985]. The two features are mainly distinguished by their duration of occurrence. Outbursts are 

sudden, often violent expulsions of large quantities of gas, usually methane, and are generally 

associated with the ejection of great quantities of coal or other rock material. Blowers, on the 

other hand, historically have been viewed as the release of large quantities of gas, but over an 

extended time period of months or even years. Also, blowers are not associated with the expulsion of coal or rock material. A subset of blowers is methane bleeders, which also continually 

emit gas, but at lower rates and generally for shorter timeframes. 

Although not typically associated with U.S. coal mines, gas outbursts occur regularly in certain 

mining districts worldwide. Typically, the mines in these districts are in a coalbed with high inplace gas contents, coupled with steeply dipping and/or very deep workings. As shallower, more 

easily extracted coal reserves are depleted in the United States and as mining progresses to 

deeper, more structurally complex and gassier coalbeds, the potential for gas outbursts will likely 

increase. Campoli et al. [1985] delineate more than a dozen U.S. coalbeds with outburst potential based on internationally recognized criteria. In fact, gas outbursts have been documented 

throughout history in U.S. mines with similar conditions. 

Historical examples of U.S. outbursts are mentioned by Darton [1915] as occurring in Pennsylvania. Two of these occurred in anthracite mines in steeply dipping coalbeds. Another took 

place in western Pennsylvania near Connellsville, where Darton noted that 100,000 ft3

 of fresh 

air per minute for 3 days was required to reduce the methane concentration in the mine air to safe 

levels. Little additional documentation is presented, and it is not known if rock material was also 

ejected with the gas. Darton also summarized extensive European documentation of gas outbursts and concluded that these phenomena were usually related to crushed coal zones associated 

with folds, buckles, and faults. 
108 

Lama and Bodziony [1998] compiled a comprehensive overview of outbursts worldwide and 

their causative factors and prevention. They conclude that the following factors contribute to 

outbursts: (1) gas content, (2) gas pressure, (3) permeability, (4) sorption/desorption characteristics, (5) stress conditions, (6) coal strength, and (7) geologic factors (often related to tectonic 

activity). Other modern research on these phenomena has demonstrated two major indicators of 

outburst potential in coal mines. The first indicator is the coal lithotype. Beamish and Crosdale 

[1998] demonstrated that coals with high vitrain and/or inertodetrinite lithotypes were more 

likely to retain the large quantities of gas needed to produce outbursts. A second indicator, documented by Cao et al. [2001], is the association of outbursts with tectonically altered, faulted 

coals. Cao et al. noted that outbursts in China seem to be associated with tectonic activity that 

has produced regional thrust and reverse faulting. Such faulting often manifests itself in coalbeds because of their brittle nature compared to the surrounding strata. The coal adjacent to such 

faults is often severely crushed and pulverized, resulting in significant local changes in the gas 

storage and migration characteristics of the coal. 

Blowers, like outbursts, are not normally associated with coal mining in the United States, but 

historically they have been noted in the United States and in other mining districts abroad. 

Darton [1915] summarized documented blower occurrences worldwide and noted the occurrence 

of blowerlike features in the Pennsylvania anthracite district. 

Detection and remediation of outbursts and blowers. Based on past observations, outbursts 

and blowers are often associated with tectonically disturbed and faulted strata where gassy coals 

are mined at considerable depth. Thus, mine planners who are aware of such conditions should 

give some thought to the possibility that they will be extracting coal under conditions that have 

produced outbursts and blowers in other mining districts. 

If large-scale faulting is known and adequately mapped in future development areas, a detailed 

core-drilling program, coupled with gas content testing of core samples and in situ gas pressure 

measurements, can detect potential outburst-prone areas. Beamish and Crosdale [1998] recommend, as do Lama and Bodziony [1998], the use of any one of several published gas emission 

indices as an indicator of proneness to outbursting. 

Since outbursts often occur in “nests” or clusters, when such conditions are encountered or 

anticipated, remediation may be achieved by using gas drainage techniques such as vertical 

methane drainage boreholes drilled from the surface, horizontal boreholes drilled underground 

ahead of the face, or (if outburst-prone strata are in the roof rock) cross-measure boreholes 

[Diamond 1994]. Typically, these boreholes will penetrate the fault system that has altered the 

coal structure and allowed large quantities of gas to accumulate at great pressure behind it. The 

boreholes are used to drain gas from the outburst-prone area and to relieve the gas overpressure 

that drives outbursts. Lama and Bodziony [1998] stress that vertical surface boreholes may be 

preferred over holes drilled in the coalbed because of the difficulty of maintaining borehole 

integrity during horizontal drilling in the outburst-prone strata. The difficulty of maintaining 

borehole integrity is due to the crushed nature of the coal in these areas. Beamish and Crosdale 

[1998] also recommend water infusion to reduce outburst hazards. 
109 

Outbursts are often driven not only by gas pressure, but also by inherent, concentrated stress 

fields in the rock mass. Hyman [1987] summarizes several methods to reduce in situ strata gas 

pressures to prevent outbursts. These techniques include the use of modified mine opening 

geometries, shot firing, and water infusion, all of which have been successfully used abroad to 

abate outbursts. 

Blowers most often emanate from underlying strata. The recommended remediation technique is 

the use of cross-measure holes angled downward from the mine heading to intercept the fissure 

or fault, which acts as a gas conduit. The borehole(s) may then be used to drain gas away from 

the blower outlet in the mine workings. 

GENERAL REMEDIATION CONSIDERATIONS 

When unanticipated gas emissions cause repeated 

production interruptions, mine operators must 

understand that they have a gas problem. 

In order to determine the most appropriate course of action to remediate a gas emission problem, 

the mine operator must make a thorough evaluation of the cause, extent, and severity of the problem. The cause of the problem may be as simple as underestimating the original gas content of 

the coalbed, or it may be more complex and involve gas sources outside the mined coalbed. 

Determining the extent of the problem may only entail additional gas content testing [Diamond 

and Schatzel 1998] of the mined coalbed through exploratory boreholes, or it may require extensive underground methane monitoring surveys or gas monitoring instrumentation. Also, detailed 

underground mapping of geologic features may be needed to delineate and predict gas emission 

trends. Remediation may require only an increase in ventilation airflow to the face, or it may 

require an extensive mapping and drilling program to delineate and alleviate the problem. 

Often when unusually high methane emissions are unexpectedly encountered during mining, 

operators must make quick decisions about how to address the problem. A prudent operator 

should weigh several pertinent factors before embarking on any course of action. 

If increased ventilation capacity alone cannot alleviate the problem, operators, especially smaller 

ones, do not always have the available expertise, human resources, or equipment to evaluate the 

problem and implement either a surface or in-mine borehole methane drainage program and will 

need to rely on outside consultants. Methane drainage systems drilled from the surface generally 

require fewer boreholes, but need good geologic control to effectively hit the gas-bearing zone. 

These boreholes generally require dewatering, hydraulic fracturing, and sufficient time to be 

optimally effective. Methane drainage systems drilled from the surface have associated issues 

with the procurement of appropriate and accessible drilling sites and environmental concerns for 

water disposal and site reclamation. 
 

110 

In-mine methane drainage systems generally require more boreholes and may also require 

dewatering. However, they can be drilled relatively quickly and require less time to be optimally 

effective. Additionally, in-mine methane drainage systems will require an underground gasgathering system to transport gas from the boreholes to the surface, usually via one or more 

vertical boreholes drilled for that purpose. In-mine systems may also have accessibility constraints due to poor roof or floor conditions or other mining-related safety issues in the area 

where holes need to be drilled. 

It should be noted that there may be regulatory requirements that need to be addressed when 

methane drainage systems are associated with mining operations. In the United States, a recent 

bulletin issued by the Mine Safety and Health Administration (MSHA) states that “MSHA has 

determined that [coalbed methane] wells are subject to the ventilation plan and mapping requirements that apply to methane degas holes” [McKinney 2005]. If coalbed gas of sufficient quality 

and quantity is produced by the methane drainage system, the gas has the potential to be sold for 

commercial use, which helps defray the costs of methane drainage. 

After a methane drainage system is put into place, it can only be effective as long as it is operating properly. Operators must consider who will operate and maintain the system once it is 

installed. If installed in-house, personnel may need to be permanently assigned to the project. 

If outside contractors are used for the installation, will they be retained for long-term operation 

and maintenance, or will mine personnel need training to operate and maintain the system once 

the contractor leaves? 

The economics of any methane drainage system under consideration involves weighing the pros 

and cons of all of the factors discussed above. The final remediation plan will hopefully be one 

that, under the site-specific circumstances, will create a safer underground workplace for the 

miners while minimizing capital investment and human resources. 

REFERENCES 

Ayruni AT [1984]. Theory and practice of mine gas control at deep levels. Rockville, MD: 

Terraspace, Inc. 

Beamish BB, Crosdale PJ [1998]. Instantaneous outbursts in underground coal mines: 

an overview and association with coal type. Int J Coal Geol 35:27–55. 

Campoli AA, Trevits MA, Molinda GM [1985]. Coal and gas outbursts: prediction and prevention. Coal Min 22(12):42–44, 47. 

Cao Y, He D, Glick DC [2001]. Coal and gas outbursts in footwalls of reverse faults. Int J Coal 

Geol 48(1–2):47–63. 

Chase FE, Ulery JP [1987]. Clay veins: their occurrence, characteristics, and support. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 9060. NTIS No. PB87204517. 
111 

Clayton JL [1998]. Geochemistry of coalbed gas: a review. Int J Coal Geol 35(1–4):159–173. 

Clayton JL, Leventhal JS, Rice DD, Pashin JC, Mosher B, Czepiel P [1993]. Atmospheric 

methane flux from coals: preliminary investigation of coal mines and geologic structures in the 

Black Warrior basin, Alabama. In: Howell DG, ed. The future of energy gases. USGS Professional Paper 1570. Reston, VA: U.S. Department of the Interior, U.S. Geological Survey, 

pp. 471–492. 

Creedy DP [1988]. Geological controls on the formation and distribution of gas in British coal 

measure strata. Int J Coal Geol 10(1):1–31. 

Darton NH [1915]. Occurrence of explosive gases in coal mines. Washington, DC: U.S. Department of the Interior, Bureau of Mines, Bulletin 72. 

Diamond WP [1982]. Site-specific and regional geologic considerations for coalbed gas drainage. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 8898. NTIS No. 

PB83157685. 

Diamond WP [1994]. Methane control for underground coal mines. Pittsburgh, PA: U.S. 

Department of the Interior, Bureau of Mines, IC 9395. NTIS No. PB94189289. 

Diamond WP, Schatzel SJ [1998]. Measuring the gas content of coal: a review. Int J Coal Geol 

35(1–4):311–331. 

Diamond WP, Ulery JP, Kravits SJ [1992]. Determining the source of longwall gob gas: lower 

Kittanning coalbed, Cambria County, PA. Pittsburgh, PA: U.S. Department of the Interior, 

Bureau of Mines, RI 9430. 

Dutcher RR, Campell DL, Thornton CP [1966]. Coal metamorphism and igneous intrusives in 

Colorado. In: Coal Science American Chemical Society, Advances in Chemistry Series, vol. 55, 

pp. 708–723. 

Finfinger GL, Cervik J [1979]. Drainage of methane from the overlying Pocahontas No. 4 coalbed from workings in the Pocahontas No. 3 coalbed. Pittsburgh, PA, U.S. Department of the 

Interior, RI 8359. NTIS No. PB296882. 

Gurba LW, Weber CR [2001]. Effects of igneous intrusions on coalbed methane potential, 

Gunnedah basin, Australia. Int J Coal Geol 46(2–4):113–131. 

Houseknecht DW [1982]. Probability of encountering coalbed discontinuities during vertical 

and horizontal borehole drilling. Washington, DC: U.S. Department of the Interior, Bureau of 

Mines, RI 8665. NTIS No. PB82232091. 

Hyman DM [1987]. A review of the mechanisms of gas outbursts in coal. Pittsburgh, PA: 

U.S. Department of the Interior, Bureau of Mines, IC 9155. NTIS No. PB88231717. 
112 

Iannacchione AT, Ulery JP, Hyman DM, Chase FE [1981]. Geologic factors in predicting coal 

mine roof-rock stability in the upper Kittanning coalbed, Somerset County, PA. Pittsburgh, PA: 

U.S. Department of the Interior, Bureau of Mines, RI 8575. NTIS No. PB82145376. 

Johnson RC, Flores RM [1998]. Developmental geology of coalbed methane from shallow to 

deep in Rocky Mountain basins and in Cook Inlet-Matanuska basin, Alaska, U.S.A. and Canada. 

Int J Coal Geol 35(1–4):241–282. 

Juntgen H, Klein J [1975]. Formation of natural gas from coaly sediments (in German). Erdöl 

und Kohle Erdgas Petrochemie 28(2):65–73. 

Lama RD, Bodziony J [1998]. Management of outburst in underground coal mines. Int J Coal 

Geol 35(1–4):83–115. 

McCulloch CM, Deul M, Jeran PW [1974]. Cleat in bituminous coalbeds. Pittsburgh, PA: U.S. 

Department of the Interior, Bureau of Mines, RI 7910. NTIS No. PB233420. 

McCulloch CM, Diamond WP, Bench BM, Deul M [1975]. Selected geologic factors affecting 

mining of the Pittsburgh coalbed. Pittsburgh, PA: U.S. Department of the Interior, Bureau of 

Mines, RI 8093. NTIS No. PB249851. 

McKinney R [2005]. Coalbed methane wells. Program information bulletin No. P05–10. 

May 10, 2005. Arlington, VA: U.S. Department of Labor, Mine Safety and Health Administration. [http://www.msha.gov/regs/complian/PIB/2005/pib05-10.asp]. 

Moss KN [1927]. Gases, dust, and heat in mines. London: Charles Griffen and Co., Ltd. 

Nickelsen RP, Hough VD [1967]. Jointing in the Appalachian plateau of Pennsylvania. Geol 

Soc Am Bull 78:609–629. 

Payman W, Statham ICF [1930]. Mine atmospheres. London: Methun & Co. Ltd. 

Price PH, Headlee AJW [1943]. Natural coal gas in West Virginia. AAPG Bull 27(4):529–537. 

Prosser LJ Jr., Finfinger GL, Cervik J [1981]. Methane draining study using an underground 

pipeline: Marianna mine 58. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, 

RI 8577. NTIS No. PB82142860. 

Thielemann T, Krooss BM, Littke R, Welte DH [2001]. Does coal mining induce methane emissions through the lithosphere/atmosphere boundary in the Ruhr Basin, Germany? J Geochem
 

Explor 74(1–3):219–231. 

Ulery JP, Molinda GM [1984]. Influence of overlying strata on methane emissions in a northern 

West Virginia coal mine. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, 

RI 8879. NTIS No. PB84235282. 
 

113 

CHAPTER 8.—FORECASTING GAS EMISSIONS FOR 

COAL MINE SAFETY APPLICATIONS
 

By C. Ozgen Karacan, Ph.D.1

1

Senior service fellow. 

2

and William P. Diamond2 

Supervisory physical scientist. 

Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, Pittsburgh, PA. 

In This Chapter 

9 Measuring the gas content of coal 

9 Predicting gas emissions based on geologic and coal reservoir property data 

9 Determining the gas storage capacity of coalbeds and other gas-bearing strata 

9 Methane drainage borehole monitoring to forecast the remaining gas-in-place and the 

 influence on mine emissions 

9 Forecasting gas emissions during mining as a function of mining parameters 

and 

9 Gas emission prediction based on numerical simulation 

This chapter provides guidelines for determining the gas content of coalbeds, estimating the gasin-place, and predicting gas flow and emissions before and during coal mining operations. The 

techniques are discussed briefly in the following sections. However, detailed information on the 

techniques is provided in the cited references. 

INTRODUCTION 

Coalbed methane, if not properly controlled in the underground mine environment, is a safety 

concern due to the potential risk for an explosion. This is a particular problem during longwall 

mining, where the high rate and volume of coal extraction can result in the release of large 

amounts of methane from the mined coalbed and other adjacent gas-bearing strata. The 

variability and potential hazards of these sometimes unexpectedly high gas flows provide the 

impetus to develop methods to predict methane emissions into the underground workplace. 

A forecast of the volume of gas that might be released during coal mining is helpful for 

designing ventilation systems and for implementing optimum methane drainage strategies to 

help mitigate expected gas emission problems. 

A complete assessment of the need for methane drainage prior to mine development generally 

requires both an empirical and a theoretical approach. If there are active mines in the general 

area with similar geologic conditions and coal characteristics, a review of gas problems in those 

mines provides an initial insight into the level of gas emissions to be expected at a new location. 

In addition, relatively simple methods exist to determine the in situ gas content (volume of gas 

per unit weight of coal) of the coalbeds in a particular mining area, as well as the gas-in-place 

(volume of gas in the coalbed(s) within a defined geographic area). 
114 

More sophisticated reservoir engineering methods are also available not only to estimate the gasin-place, but also to simulate gas flow patterns in the mining horizon, as well as in the surrounding strata. With a reservoir modeling approach, gas flow to various configurations of methane 

drainage boreholes can be investigated to optimize the interception and extraction of coalbed 

methane before it can enter the mine ventilation system. 

Although potentially providing valuable insights about gas flow and methane drainage in the 

mining environment, the site-specific input data required for reservoir modeling is not routinely 

available at many, if not most, mine sites. For this reason, it is recommended that if the reservoir 

modeling approach is anticipated due to high in situ gas contents, then the necessary geologic, 

engineering, and reservoir data should be obtained early so that methane drainage options can be 

evaluated before methane emission problems become acute. 

Forecasting gas emissions requires knowledge 

of the relationships among gas storage in coal 

(and adjacent strata), the factors affecting gas 

emissions, and the techniques used to predict 

emissions. 

METHANE CONTENT OF COAL 

The gas content of coal can be measured or estimated using various techniques. These techniques usually fall into two categories: (1) direct methods that actually measure the volume of 

gas released from a coal sample (preferably wire line core) sealed in a desorption canister, and 

(2) indirect methods based on empirical correlations or laboratory-derived gas storage capacity 

data from sorption isotherms. An extensive review of direct techniques for gas content measurement for coal was published by Diamond and Schatzel [1998]. One of the most commonly 

used methods to determine the gas content of coal is the U.S. Bureau of Mines direct method 

[Diamond and Levine 1981; Diamond et al. 1986]. Properly conducted direct-method testing of 

coal cores provides relatively accurate estimates of in-place gas contents for most mine planning 

purposes while allowing for resource evaluation at a reasonably low cost. A modified-directmethod procedure [Ulery and Hyman 1991] provides an increased level of accuracy, but at a 

higher level of instrumentation sophistication, procedural complexity, and cost. 

Direct-method testing of coal cores provides 

sufficient estimates of in-place gas contents 

for most mine planning purposes. Greater 

accuracy can be obtained by using the 

modified direct method. 
 

 

 

115 

In the absence of extensive measured gas contents in an area of 

interest, an alternative approach to 

obtain the gas content is to use the 

relations proposed by Kim [1977] 

based on gas content determinations 

from adsorption analysis on different coals of various ranks and depths 

(Figure 8–1). This approach can be 

considered in parts of a basin where 

coal samples are initially not available for direct gas content testing. 

However, it is important to note that 

these are only estimated values and 

should be confirmed with subsequent direct gas content testing 

within the actual area of interest. 

Figure 8–1.—Methane content as a function of depth and 

coal rank (modified from Kim [1977]). 

For estimating in-place gas contents 

for a specific area, regional gas content data on individual coal samples 

can be used along with data on 

coal rank and/or depth to construct 

curves such as those in Figure 8–2. 

Such curves are generated for a 

particular coalbed or closely associated group of coalbeds and can be 

used to estimate gas content values 

only if the rank or depth are known 

for the coalbed of interest. As an 

example, the graph in Figure 8–2 

presents such curves for the Black 

Warrior Basin in Alabama [McFall 

et al. 1986]. Coal lithotype 

characteristics also affect the 

methane content of coal. For 

instance, significantly higher 

methane capacity was observed for 

bright bands (850 ft3

/ton) versus 

dull bands (570 ft3/ton) in the same 

coalbed during an evaluation of 

compositional effects on coals from western Canada [Lamberson and Bustin 1993]. Total gas 

content varied with the amount of vitrinite and liptinite, which usually offer high methane 

storage capacity, whereas no obvious relationship was observed with the inertinite content. 

Some studies report increases in gas yield with fusain content, which tends to allow rapid 

desorption of methane [Creedy 1986]. Despite these examples, while the general influence of 

Figure 8–2.—Gas content versus depth and coal rank, 

Black Warrior Basin, Alabama (McFall et al. [1986]). 
116 

coal lithotype on gas content is of interest, it cannot be sufficiently quantified for use as a 

predictive gas content method. 

Estimated gas contents should only be used 

for preliminary assessments. They are not a 

substitute for site-specific gas content 

determinations. 

GAS-IN-PLACE CALCULATION 

One of the key steps in forecasting gas emissions during and after mining is to calculate the 

volume of gas-in-place that will potentially migrate to the underground mining environment. 

During mining, these emissions are primarily from the mined coalbed, whereas postmining emissions include not only the mined coalbed (ribs and pillars), but also gas-bearing strata above 

(gob) and below the mined coalbed. 

The simplest method for calculating the gas-in-place for coalbeds is based on commonly available geologic mapping data for the mine site and the site-specific gas content data [Diamond 

1982], as follows: 

GIPc = (ρ × h × A)GC, (1) 

where GIPc = coal gas-in-place, ft3; 

ρ = coal density, tons/acre ft; 

h = coal thickness, ft; 

A = area, acres; 

and GC = gas content (volume-to-mass ratio), ft3

 gas/ton of coal. 

Depending on the variability of measured gas contents within the area of interest, multiple gasin-place calculations for individual zones (based on gas content versus depth and/or coal rank 

data as shown in Figure 8–2) may be necessary to obtain the best total gas-in-place value. For 

gas-bearing strata other than coal (organic shales, etc.) where the gas is primarily stored by 

adsorption (as in coal) or for low matrix permeability rocks such as siltstones where the stored 

gas cannot readily escape from a core sample before it is sealed in a desorption canister, the gasin-place estimate can be calculated as follows [Diamond et al. 1992]: 

GIPr = (h × A)GC, (2) 

where GIPr = rock gas-in-place, ft3

; 

h = rock strata unit thickness, ft; 

A = area, ft2; 

and GC = gas content (volume-to-volume ratio), ft3

 gas/ft3 rock. 
 

 

117 

For high matrix permeability rock units, like some sandstones, where the direct-method-type gas 

content determinations are not appropriate, traditional reservoir engineering methods for estimating the volume of gas-in-place are more appropriate. These methods may include the use of well 

logs, laboratory reservoir property core testing, and well/production testing. 

There are various approaches to determine an 

area’s in-place gas volume. The preference 

usually depends on data availability, degree of 

sophistication required in the analysis, and 

the technical background of the personnel 

conducting the analysis. 

Reservoir-analysis-based approaches may also be used to determine the gas-in-place for coalbeds 

in a specific geographic area. These approaches relate the volume of gas in the reservoir at reservoir conditions to the volume at STP2

 conditions and use the differences in remaining gas volumes as the reservoir pressure is depleting. Although reservoir-based approaches can be an 

essential part of gas-in-place calculations, particularly for mines that are considering marketing 

the produced methane, these approaches require significantly more data (at a relatively high cost) 

than is usually available for most mine safety applications. Two of the reservoir-analysis-based 

methods for calculating gas-in-place are the volumetric calculation and the material balance 

calculation. 

The volumetric method of calculation (Equation 3) is very similar to Equation 1 above. In addition to estimating the gas-in-place (free gas, if any, and adsorbed gas) for the coal in an area 

based on the direct-method gas content data, this method also calculates the amount of gas in the 

cleats and fractures by taking into account the water saturation (Swfi), cleat/fracture porosity (φf), 

and gas formation volume factor (Bgi).3

2

Standard temperature and pressure. 

3

Refer to McLennan et al. [1995] for more information on the volumetric method of calculation and for further 

definition of the terms in Equation 3.
 

 ⎛ 43560φ f (1 − Swfi ) ⎞ Gi = Ah ⎜

⎜

Bgi

+1.359Cgiρ c (1− f a − f m )⎟

⎟ (3) ⎝ ⎠

Another reservoir engineering method of estimating gas-in-place is the use of material balance 

calculations derived for coalbeds from conventional material balance equations. Terms are 

added to the equations to account for desorption mechanisms. However, this is an iterative technique and requires more data and calculation complexity than the previous methods. Details on 

the use of material balance calculations for coalbed methane applications were published by 

King [1993]. 
118 

FORECASTING REMAINING GAS-IN-PLACE FROM 

PRODUCING METHANE DRAINAGE BOREHOLES
 

IN THE AREA OF INTEREST 

(PRODUCTION DECLINE ANALYSIS)
 

Analysis of production decline trends for premining methane drainage boreholes in an area of 

interest, when combined with the original gas-in-place estimate, can provide a reasonably accurate estimate of the volume of gas remaining in the coalbed and available for flow to the mining 

environment. This method is widely accepted in the natural gas industry due to its ease of application and requires only the gas production histories from existing wells. However, in contrast 

to conventional gas reservoirs, it usually takes a long time (potentially a year or longer) for a 

coalbed methane borehole to exhibit a production decline. Thus, there may be a long delay before 

the data can be analyzed. Also, borehole spacing, formation permeability, desorption properties of 

coal, and production problems not related to the reservoir can affect the production profile. 

Production decline analysis can be used for forecasting the future methane flow and emission 

potential from a coalbed by analyzing the time-resolved production trends of methane drainage 

boreholes. However, these boreholes are generally completed only at the mined coalbed interval, 

which makes analysis for mine safety applications more complex and difficult. During longwall 

mining, gas emissions in the relaxed strata are usually a combination of different sources of 

migrating gas (from coalbeds plus other gas-bearing strata in the overburden and underburden) 

due to horizontal and vertical fracturing of the surrounding strata. Therefore, it is difficult to 

compare the estimated gas flow and emission rates from decline analysis of methane drainage 

boreholes completed in the mined coalbed with the actual gas emissions observed during mining. 

The result is that the forecasts of gas emissions based on decline curve analysis of commercial 

coalbed methane wells (or vertical degasification wells), completed at a single interval (the 

mined coalbed), will likely underestimate the volume of gas that will be released from the mined 

coalbed and surrounding strata into the mine environment. 

In a case where one can be sure that there is no gas source other than the mined coalbed, the 

decline curve analysis technique may be applicable for estimating the remaining gas-in-place 

for that coalbed that might still migrate to the ventilation system during future mining activities. 

In order to be able to use decline curve techniques for this situation, all or most of the criteria 

below need to be met for a high degree of confidence in production forecasts: 

• Decreasing gas and water rates 

• A stable slope in gas rates for at least 6 months 

• A length of producing well life greater than 2 years 

• Bounded wells and well spacing 

It is also recommended that decline-based gas-in-place and future methane flow/emission potential projections be compared against projections from volumetric or other available analytical 

techniques [Hanby 1991]. 

Alternatively, type-curve matching techniques are a reservoir engineering tool in which the 

solutions of complex equations for various situations are represented in graphical form. The 
techniques rely on matching the actual gas production data plots (prepared with the same set of 

units and graphical form as in the type-curves) to the theoretical curves. These techniques can 

also be used to analyze production decline curves to predict remaining gas-in-place and future 

emissions for mine safety assessments [Chen and Teufel 2000]. In addition, if the gas production is exclusively from the coalbed to be mined, type-curve analysis can provide other reservoir 

information that needs to be determined for emission forecasting studies (e.g., modeling). 

In summary, type-curves can contribute to: 

• Stimulation (fracturing) effectiveness diagnosis 

• Recovery efficiency (recovery factor based on initial gas-in-place) 

• Estimation of reservoir flow properties (permeability, flow capacity, etc.) 

• Reservoir storage properties 

• Future prediction of production 

119 

Both production decline analysis and type-curve matching 

techniques can be used to analyze methane drainage 

boreholes for future production rates and thus predict the 

remaining gas-in-place at the time of mining. However, 

gas emissions from different gas sources may be 

commingled, especially during longwall mining. 

Therefore, one can expect higher emission rates during 

mining compared to what is predicted by the analysis. 

PREDICTING GAS EMISSIONS DURING MINING 

The main sources for gas (generally predominantly methane) that can be released into the underground mine workings are the mined and adjacent coalbeds and other surrounding gas-bearing 

strata [Mucho et al. 2000; Diamond et al. 1992]. Mining activities disturb the existing stress 

equilibrium in the rock mass and create changes to the structural integrity of the affected strata. 

The mining processes can thus create sudden and unstable gas problems, which may increase the 

risk of an explosion in the underground workplace. Gas flow from these sources is initiated and 

maintained by differential pressures between the source (higher pressure) and the mine workings 

(lower pressure). The flow paths are both the naturally occurring rock joints, faults, and coal 

cleat, as well as mining-induced fractures in the surrounding strata. 

It is generally observed that the amount of gas released during the mining process is greater than 

that contained in the actual volume of coal mined at the face [Kissell et al. 1973]. This apparent 

discrepancy is due to the continual emission of gas from the coal that is left in place as ribs and 

pillars, as well as the migration of gas from the surrounding strata, including the longwall gob 

[Mucho et al. 2000]. Methane emission rates change over time in the life cycle of a mine 

because of the interaction of variable geotechnical, mine design, and operational factors. The 

following mathematical formula by Lunarzewski [1998] addresses this phenomenon and calculates the quantity of gas released into a mine during various stages in the life of a mine: 
 ⎛

⎜

⎜ 1 

⎞

⎟

⎟

m m +1 ⎛

⎜

⎜

⎝

⎞

⎟

⎟

⎠

+
1 −
⎛

⎜

⎜

⎝

y

0 

∑C⎞

⎟

⎟

⎠

y

0 

Q( y) =
 ∑ g

CA

+
 ,
 (4) 

⎠

C
 

⎝

 

120

where Q(y) is the average methane emission (cubic meters of methane) in a year “Y” of the mine’s 

existence, CA is the coal output in 1 year only (tons), C is the total coal output for the life of the 

mine up to year “Y”, and g and m are coefficients dependent upon geological and mining 

conditions. 

The highest gas emissions can be expected as the coal is extracted and the floor and roof strata 

are relaxed. The instantaneous volume of gas released from all potential sources when 1 ton of coal 

is extracted can be calculated, and practical experience has shown that gas emissions are related to 

daily and weekly coal production levels and to the time factor, as follows [Lunarzewski 1998]: 

Q = a CP + b , (5) 

where Q is the total methane emission rate expressed in liters of methane per second, CP is the 

daily coal production rate in tons, and a and b are empirical constants related to weekly coal 

production levels and number of working days per week [Lunarzewski 1998]. 

Another empirical method to predict the total gas emissions from 

longwall mining is the use of 

degree-of-gas emission curves. 

Figure 8–3 is an example of a 

degree-of-gas emission curve for 

previously disturbed roof and 

floor strata in a slightly to moderately dipping coalbed [Noack 

1998]. In such a condition, the 

prediction can be made assuming 

that the emitted methane proportion is not a function of the initial 

gas content, but rather a function 

of the geometric location with 

respect to the longwall face 

[Noack 1998]. For practical 

purposes, the upper boundary of 

the zone from which gas can be 

released is assumed to be at 

+541 ft (+165 m), whereas the 

lower boundary is at -194 ft 

(-59 m). In the absence of gas 

emission measurements, a mean 

degree of gas emission of 75% of 

the gas content in the mined Figure 8–3.—PFG/FGK method to predict gas emissions in a 

previously disturbed zone [Noack 1998]. PFG = degree of gas 

emission curve for the roof; FGK = degree of gas emission curve 

for the floor. 
121 

coalbed is assumed, as is the case for Figure 8–3. Above the coalbed from 0 to 66 ft (20 m) and 

below the coalbed from 0 to -36 ft (-11 m), the degree of gas emission is assumed to be 100%. 

Because these curves are empirical correlations or standard assumed degrees of emissions, there 

may be considerable variations when they are applied to other locations. As always, it should be 

remembered that the best information for prediction is the measured data and the derived empirical correlations at a specific site of interest. 

On the other hand, if the roof and floor have not been fractured before, the prediction can be 

based on gas pressure, and thus a remaining gas content. In this case, the proportion of gas 

emitted depends on the gas pressure (gas content) and the location of the strata. The gas emission prediction for such a situation can be based on the remaining gas profiles, as shown in 

Figure 8–4. There are three zones designated in the roof and two in the floor, which are characterized by varying the remaining gas gradients. 

 Figure 8–4.—Gas pressure method: residual gas pressure lines are 

dependent on thickness of the mined coalbed [Noack 1998]. 

Based on Figure 8–4, the residual gas pressures are first determined layer by layer in accordance 

with the mean normal distance of a gas-bearing layer from the mined coal seam. The residual 

gas pressures are converted to 

remaining gas contents using 

the Langmuir isotherm. The 

difference between the remaining and initial gas contents 

represents the emitted portion 

of the adsorbed gas, which is 

the required value [Noack 

1998]. Free gas is then added 

to this value. 

The gas pressure method has 

the advantage of not defining 

upper and lower zones strictly 

compared to the prediction 

based on the degree of gas 

emission. Also, this method 

takes into account both the 

adsorbed gas and the free gas 

in the surrounding strata. 

There is another method based 

on using zones of emission, 

reviewed extensively by Curl 

[1978]. This model describes 

methane emissions in terms of 

the geometry of the zone of 

emissions, the size of the zone 

of emissions, and the degree of 
122 

emissions. The geometry and size of the zone of emissions simply refer to the shape and extent 

of the zone. The degree of emissions refers to the percentage of desorbable gas that is released 

into the mine workings at a given location near the coalbed being mined. In this model, underburden emission zones and the degree of emissions are generally more limited in extent. The 

lateral extent of the zone of emissions is generally limited to the dimensions of the panel. In 

these models, sandstone units within the emission zone are ascribed 10% of the gas contained in 

a nearby coalbed of equal thickness, whereas shale is assigned 1% of the gas contained in coalbeds of equal thickness. 

Schatzel et al. [1992] used such a model to predict methane emissions from longwall panels. 

They reported that this approach performed well for longwall panels in Cambria County, PA, 

but poorly in the Central Appalachian Basin of southwestern Virginia. This suggests that 

although the simplistic predictive techniques and empirical methods may offer quick calculation 

advantages, in general they are not sufficiently reliable for making emission estimates given the 

complex interplay of the geotechnical and mining variables involved. Thus, the use of numerical 

models to simulate the physics of both the failure mechanics of rock strata and the fluid flow in 

porous media is more appropriate for obtaining reliable emission estimates, for flexibility in 

adapting the models to different situations, and for optimizing methane drainage systems and 

mine designs accordingly. 

Simple calculations and empirical models are usually sitespecific and are very limited in their capabilities to estimate 

methane emissions. Realistic numerical simulations offer 

flexibility, confidence in estimates, and guidance for 

optimizing methane drainage systems and mine designs. 

GAS PREDICTION TECHNIQUES BASED ON NUMERICAL SIMULATION 

Reservoir simulation is the process of integrating geology, petrophysics, reservoir engineering, 

and production operations to more effectively develop and produce hydrocarbon resources. 

Numerical reservoir simulations can also be useful in mine safety applications. In fact, reservoir 

simulations are currently the only analytical method that can be used to establish the complex 

relationships between coalbed methane reservoir properties, methane drainage, and mining 

operations in a reliable and cost-effective manner. Numerical simulation is also the only practical method to describe how reservoir properties affect both gas and water flow and can address 

the intricate mechanisms of gas desorption and diffusion in coal due to either methane drainage 

and/or mining of the coalbed reservoir. 

Reservoir simulators can be used to perform a variety of analyses. The primary applications 

relative to coalbed methane/mining are: 

• Determining the volume of gas-in-place 

•	 Developing optimum methane drainage systems to reduce the flow of gas into 

 underground mine workings 

•	 Predicting the methane emission consequences of changing mining methods and 

practices 
123 

• Identifying and diagnosing production problems in operating methane drainage systems 

• Predicting gas recovery from methane drainage systems associated with underground mines 

In general, three different types of coalbed methane reservoir simulators are available: gas sorption and diffusion simulators, compositional simulators, and black oil simulators. The compositional simulators with coalbed methane options that can handle the sorption and diffusion 

processes are widely used and are more appropriate for coalbed methane applications due to their 

capability for simulating different gas mixtures. 

Reservoir simulators for coalbed methane applications are also classified based on their treatment of the gas sorption process. More than 50 coalbed methane reservoir simulators are 

described in the literature [King and Ertekin 1989a,b; 1991], which are classified as equilibrium 

sorption (pressure-dependent) and nonequilibrium sorption (time- and pressure-dependent) 

simulators. The basic difference between these two classifications is that when using equilibrium simulators, it is implicitly assumed that as the pressure declines, the gas immediately enters 

the fracture system. This oversimplification gives optimistic gas flow rates in some cases. Nonequilibrium models, which take the sorption time into account and include modifications to the 

conventional dual-porosity models, are more realistic. The primary modifications required to 

enhance the simulation capability of the dual-porosity models are to account for methane storage 

by adsorption on the matrix-coal surface and control of gas transport through the coal matrix by 

diffusion until the gas reaches the fracture network, where conventional Darcy flow mechanics 

are the controlling transport factor. 

The most realistic simulations of gas flows in coalbeds are 

provided by compositional, nonequilibrium, dual-porosity 

reservoir models. These models account for sorption time, 

methane storage by adsorption, and gas transport by 

diffusion through the coal matrix to the fracture network. 

Although numerical reservoir simulation techniques offer more reliable emission predictions and 

guidance for optimum methane drainage system designs, building objective-oriented models 

requires more time and effort for gathering site-specific data, careful analysis of field data, and 

detailed planning. 

The basic steps of performing a gas flow/production study using a reservoir simulator are as 

follows [Saulsberry et al. 1996]: 

• State the study objectives 

• Select a reservoir simulator 

• Collect and evaluate all geologic and engineering data 

• Construct a geologic model for reservoir 

• Design the simulation grid 

• Digitize the maps 

• Install engineering data into the model 

• Define the well operating constraints 

• Perform simulations 
 

124 

Reservoir models require a substantial amount of site-specific data to provide reliable simulations of gas flow and production from boreholes/wells. Commonly, all of the reservoir property 

data required to conduct a simulation are not available or are unknown. This is particularly true 

in the mining industry where reservoir modeling is relatively new and coalbed reservoir property 

base data acquisition is not part of the routine site evaluations. However, as more mines are considering the commercial production of coalbed methane, the value of obtaining coalbed reservoir 

data is becoming more widely recognized. It is an accepted reservoir engineering practice to use 

measured gas production data from boreholes and wells in “history matching” exercises to estimate some of the unknown reservoir properties. For mining-related applications, gas production 

data from both vertical and horizontal methane drainage boreholes and gob gas ventholes can be 

used for history matching. Because of the potential for gas production variabilities due to nonreservoir-related reasons (such as mechanical problems with pumps, etc.), using multiwell data 

sets for history matching is usually more dependable than single-well simulations, and they provide a better representation of the reservoir. 

Although reservoir simulators are very successful in the representation of the multiphase flow 

and time-dependent gas diffusion processes in coalbeds, they do not readily model the dynamics 

of the mining process on the coalbed reservoir and surrounding strata. The progressive advance 

of the mine face and associated removal of the coalbed reservoir is a key dynamic that must be 

accounted for in mining-related simulations and can be accomplished with “frequent restart” files 

representing periodic updates of the reservoir geometry consistent with expansion of the mine. 

Another aspect of longwall mining that cannot be predicted by conventional reservoir simulators 

is the geomechanical response in the surrounding strata, causing permeability changes that 

influence the drainage of gob gas. This problem can be manually overcome by computing the 

changes in rock properties (in particular, permeability) with a geomechanical program such as 

FLAC4

4

Fast Lagrangian Analysis of Continua. 

[Itasca Consulting Group 2000] and representing those reservoir property changes in the 

appropriate reservoir simulation steps. Karacan et al. [2005] discuss how these dynamic miningrelated processes can be addressed in a longwall mining simulation to optimize gob gas venthole 

methane drainage. 

Other alternatives to reservoir simulation to predict gas emissions due to mining are “Roofgas” 

and “Floorgas” programs specifically designed for mining applications. They produce graphical 

representations of strata relaxation and gas flow using boundary-element and bed separation 

techniques to calculate the strata response and the rate of gas release [Lunarzewski 1998]. 

Numerical analysis and modeling techniques are the most 

powerful tools available to simulate gas flows and emissions 

in the mining environment. However, the successful 

application of these methods is highly dependent on the 

availability of valid, site-specific, reservoir property data. 
125 

REFERENCES 

Chen H–Y, Teufel LW [2000]. A new rate-time type curve for analysis of tight-gas linear and 

radial flows. SPE paper 63094. Richardson, TX: Society of Petroleum Engineers. 

Creedy DP [1986]. Methods for the evaluation of seam gas content from measurements on coal 

samples. Min Sci Technol 3(2):141–160. 

Curl SJ [1978]. Methane prediction in coal mines. London: IEA Coal Research. 

Diamond WP [1982]. Site-specific and regional geologic considerations for coalbed gas drainage. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 8898. NTIS No. 

PB83157685. 

Diamond WP, Levine JR [1981]. Direct method determination of the gas content of coal: procedures and results. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8515. 

NTIS No. PB81196735. 

Diamond WP, Schatzel SJ [1998]. Measuring the gas content of coal: a review. Int J Coal Geol 

35(1–4):311–331. 

Diamond WP, LaScola JC, Hyman DM [1986]. Results of direct-method determination of the 

gas content of U.S. coalbeds. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines 

IC 9067. NTIS No. PB86205325. 

Diamond WP, Ulery JP, Kravits SJ [1992]. Determining the source of longwall gob gas: lower 

Kittanning coalbed, Cambria County, PA. Pittsburgh, PA: U.S. Department of the Interior, 

Bureau of Mines, RI 9430. 

Hanby KP [1991]. The use of production profiles for coalbed methane valuation. Paper 9117. 

In: Proceedings of the International Coalbed Methane Symposium (Tuscaloosa, AL, May 13–16, 

1991). Tuscaloosa, AL: University of Alabama. 

Itasca Consulting Group [2000]. Fast Lagrangian analysis of continua. 2nd ed. Minneapolis, 

MN: Itasca Consulting Group, Inc. 

Karacan CO, Diamond WP, Esterhuizen GS, Schatzel SJ [2005]. Numerical analysis of the 

impact of longwall panel width on methane emissions and performance of gob gas ventholes. 

In: Proceedings of the International Coalbed Methane Symposium (Tuscaloosa, AL, May 18–19, 

2005). Tuscaloosa, AL: University of Alabama, pp. 1–28. 

Kim AG [1977]. Estimating methane content of bituminous coalbeds from adsorption data. 

Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8245. NTIS No. 

PB271218. 

King GR [1993]. Material balance techniques for coal seam and Devonian shale gas reservoirs 

with limited water influx. SPE Reservoir Eng Feb:67–72. 
126 

King GR, Ertekin T [1989a]. A survey of mathematical models related to methane production 

from coal seams. Part 1: Empirical and equilibrium sorption models. In: Proceedings of the 

International Coalbed Methane Symposium (Tuscaloosa, AL, April 17–21, 1989). Tuscaloosa, 

AL: University of Alabama, pp. 125–138. 

King GR, Ertekin T [1989b]. A survey of mathematical models related to methane production 

from coal seams. Part 2: Nonequilibrium sorption models. In: Proceedings of the International 

Coalbed Methane Symposium (Tuscaloosa, AL, April 17–21, 1989). Tuscaloosa, AL: 

University of Alabama, pp. 139–155. 

King GR, Ertekin T [1991]. State-of-the-art modeling for unconventional gas recovery. SPE 

Formation Eval Mar:63–71. 

Kissell FN, McCulloch CM, Elder CH [1973]. The direct method of determining methane content of coalbeds for ventilation design. Pittsburgh, PA: U.S. Department of the Interior, Bureau 

of Mines, RI 7767. NTIS No. PB221628. 

Lamberson MN, Bustin RM [1993]. Coalbed methane characteristics of Gates Formation coals, 

northeastern British Columbia: effect of maceral composition. AAPG Bull 77(12):2062–2076. 

Lunarzewski LW [1998]. Gas emission prediction and recovery in underground coal mines. 

Int J Coal Geol 35(1–4):117–145. 

McFall KS, Wicks DE, Kuuskraa VA [1986]. A geologic assessment of natural gas from coal 

seams in the Warrior basin, Alabama. Gas Research Institute, GRI Topical Report 86/0272. 

McLennan JD, Shafer PS, Pratt TJ [1995]. A guide to determining coalbed gas content. GRI– 

94/0396. Chicago, IL: Gas Research Institute. 

Mucho TP, Diamond WP, Garcia F, Byars JD, Cario SL [2000]. Implications of recent NIOSH 

tracer gas studies on bleeder and gob gas ventilation design. SME preprint 00–8. Littleton, CO: 

Society for Mining, Metallurgy, and Exploration, Inc. 

Noack K [1998]. Control of gas emissions in underground coal mines. Int J Coal Geol 35(1–4): 

57–82. 

Saulsberry JL, Shafer PS, Schraufnagel RA, eds. [1996]. A guide to coalbed methane reservoir 

engineering. GRI–94/0397. Chicago, IL: Gas Research Institute. 

Schatzel SJ, Garcia F, McCall FE [1992]. Methane sources and emissions on two longwall 

panels of a Virginia coal mine. In: Proceedings of the Ninth Annual International Pittsburgh 

Coal Conference (Pittsburgh, PA, October 12–16, 1992), pp. 991–998. 

Ulery JP, Hyman DM [1991]. The modified direct method of gas content determination: 

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127 

CHAPTER 9.—CONTROL OF METHANE DURING COAL MINE 

SHAFT EXCAVATION AND FILLING 

By Fred N. Kissell, Ph.D.1  

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

In This Chapter 

9 Ventilation and methane sampling guidelines for conventional shafts 

9 Dealing with restricted spaces where methane can accumulate 

9 Ventilation and methane sampling at raise bore drills 

and 

9 Filling shafts at closed coal mines 

It is not unusual to encounter methane gas during shaft excavation or shaft-filling operations. 

Shafts into coalbeds usually produce the most methane, so the information in this chapter applies 

primarily to shafts at coal mines. Nevertheless, much of the information is relevant for noncoal 

mines that have methane in the mine or in the overlying strata. 

METHANE IN SHAFTS EXCAVATED BY CONVENTIONAL MEANS 

(DRILL, BLAST, MUCK) 

Ventilation. Shafts excavated by conventional 

means are ventilated with a fan on the surface 

connected to ductwork that extends down into 

the shaft (Figure 9–1). Shafts into coalbeds 

should have ventilation systems designed to 

handle the higher gas flows to be expected as 

the shaft excavation passes through any overlying gas-containing strata and nears the coalbed to be mined. Following are some simple 

ventilation guidelines for the minimum amount 

of air required and the selection of the ventilation duct: 

Figure 9–1.—Mine shaft under construction. 

•	 Provide enough air to meet the minimum OSHA tunnel requirement of 

30 ft/min air velocity in the open shaft, 

as specified in 29 CFR2

2

Code of Federal Regulations. See CFR in references.
 

1926.800(k)(3). 

For example, a 20-ft-diam shaft having 
 

 

128 

an area of 314 ft2 would require 9,420 cfm at the inby, or bottom, end of the ventilation 

duct. 

•	 Assume at least 35% leakage between the fan and the inby end of the ventilation duct. 

Using the 9,420-cfm figure above, the fan would have to deliver at least 14,500 cfm. 

•	 Size the ventilation duct diameter to achieve a duct velocity of 3,000 ft/min or less. 

Using the 14,500-cfm fan quantity above, the duct would have an area of 4.8 ft2 or more 

and a diameter of at least 2.5 ft. 

•	 Locate the inby end of a duct exhausting air from the shaft within 10 ft of the point of 

deepest penetration. If exhaust ventilation is used, auxiliary ventilation of the space 

below the work deck may be necessary. 

•	 Locate the inby end of a duct blowing air into the shaft within 15 duct diameters of the 

point of deepest penetration. For example, using the 2.5-ft duct diameter figure above, 

the maximum distance between the end of the duct and the point of deepest penetration 

would be 37.5 ft. 

•	 Limit fan shutdown times when workers are in the shaft to a maximum of 15 min. 

Monitor the methane level in the air during fan shutdowns. 

Gas sampling. Preshift and on-shift examinations for methane are governed by federal coal 

mine regulations at 30 CFR 77.1901. This regulation requires that methane be measured by a 

certified person within 90 min before each shift, at least once during the shift, and both before 

and after blasting. Other examinations for methane must be made immediately before and 

periodically during any welding or cutting in the shaft, per 30 CFR 77.1916(c). In all instances, 

work must not continue when the air contains 1.0 vol % or more of methane. 

Particular attention should be paid to sampling the gas level in restricted spaces (as described in 

the later section on water rings) and in any portion of the shaft where the free movement of air is 

restricted. Figure 9–2 illustrates how the free movement of ventilation air is inhibited by the 

presence of a work deck, even if the deck is fabricated from metal grating. In such circumstances, methane measurements must be made with an extensible probe positioned to draw air 

samples immediately below the work deck and at regular intervals all the way down to the muck 

pile. Measurements are necessary at regular intervals because assumptions cannot be made as to 

where the methane is likely to accumulate.3 

3

Cook [1998] describes a shaft explosion in a South African mine where the workers assumed that any methane 

would layer just below the work deck, making any measurements further down unnecessary. They failed to 

recognize that methane released from the muck at the bottom is very unlikely to layer and that once the methane is 

mixed into the air, it cannot unmix to form a layer. Because of this, measurements must be made at regular intervals 

over the entire distance between the work deck and the muck pile. For more information on layering, see Chapters 1 

and 11. 
 Figure 9–2.—Ventilation airflow is restricted by work 

deck. Therefore, methane must be monitored at regular 

distances between the work deck and the muck pile. 

 

129 

Welding and cutting. To the extent 

possible, welding and cutting should be 

done on the surface. When welding and 

cutting in the shaft cannot be avoided, 

gas check procedures must be carefully 

followed, and the number of people in 

the shaft must be held to the minimum 

required to conduct the work and check 

for gas—usually two individuals. 

Water rings. A water ring cross-section 

is shown in Figure 9–3. Water rings are 

circular spaces excavated from the rock 

in a shaft wall. They are used to provide 

water drainage from the exterior side of 

the concrete shaft lining and are often 

located to drain water from an adjacent 

aquifer. Methane gas may be drained 

along with the water, resulting in additional hazard. 

Water ring spaces in shaft walls are an example of 

a so-called restricted space, where the ventilation 

and methane sampling require extra effort.4 

4

Sampling for methane in restricted spaces is described in more detail in the sampling chapter (Chapter 2). 

Failure to monitor restricted spaces such as water rings can have disastrous consequences. For 

example, in January 2003, three miners were killed by a methane explosion during a shaftsinking operation in West Virginia [Mills 2003]. A water ring space had been excavated back 

into the strata from the shaft wall. Before pouring the concrete shaft lining, the water ring space 

was sealed off with corrugated sheets of steel to prevent it from being filled with concrete. After 

the concrete had set, the three workers were killed while in the process of opening an access door 

with an acetylene torch. The torch ignited methane that had accumulated in the water ring space 

after it was sealed off. 
 

 

130 

Figure 9–3.—Water ring in shaft wall (adapted from 

Oakes et al. [2003]. 

In the case of the abovementioned West 

Virginia mine, any one of the following 

measures would have reduced the chance 

of an explosion: 

•	 Thoroughly ventilating the water 

ring space with compressed air. 

•	 Carefully checking for methane 

using a methane detector with a 

probe inserted into the water ring 

space. 

•	 Opening the access door without 

using an acetylene torch. 

To avoid methane problems when dealing 

with water rings, the appropriate action is 

to incorporate all three measures into 

standard operating practice. 

METHANE IN SHAFTS EXCAVATED WITH RAISE BORE DRILLS 

Like conventional shafts, raise bore shafts at coal mines must be well-ventilated and the methane 

level must be monitored frequently. Maksimovic [1981] reported on some methods to ventilate 

raise bore shafts. These involve moving air through the drill stem or the annular space between 

the drill stem and the wall of the pilot hole. A more recent approach is to drill a second hole for 

ventilation and methane sampling (Figure 9–4). 

The methods reported by Maksimovic involve the use of air compressors to inject air into the 

shaft in quantities ranging from 600 to 3,000 cfm, and vacuum pumps to withdraw air from the 

shaft in quantities ranging from 600 to 1,000 cfm. These are used separately or in combination 

as follows: 

•	 Alternating the use of the air compressor and vacuum pump. Compressed air is injected 

down the drill stem5

 for a period of 2–4 hr, and then a vacuum pump is used for a short 

interval to bring an air sample through the drill stem to the surface for a methane 

measurement. 

5

At the same time, air can also be forced down through the annulus. 
 

131 

•	 Simultaneous use of the air 

compressor and vacuum pump. 

Compressed air is continuously 

injected down the drill stem. 

At the same time, for methane 

sampling, a vacuum pump is 

used to draw air up through the 

annulus between the drill stem 

and pilot hole wall. This 

method allows for continuous 

sampling of methane, but 

requires the installation of a seal 

(a stuffing box) around the drill 

stem at the surface to prevent 

surface air from entering the 

annulus and contaminating the 

sample. Any leaks in this seal 

will cause the indicated methane 

concentration to be lower than it 

actually is. 

•	 Use of the vacuum pump only. 

A vacuum pump is used to draw 

air up through the drill stem. 

This approach may be suitable 

when the methane level is very 

low, and it allows for continuous sampling. However, 

dust builds up inside the drill 

stem and accumulates on the 

joints when sections are 

removed. 

Figure 9–4.—Raise bore shaft with second hole for 

ventilation. 

In recent years, an alternative technique to ventilate raise bore shafts has been to drill a second 

hole for ventilation and methane sampling. This second hole is drilled close to the pilot hole and 

within the perimeter of the hole to be reamed by the raise drill (Figure 9–4). For gassier mines or 

those that are likely to have gas in the overburden, this is a better method than using the drill 

stem or the annulus. Larger quantities of air can be drawn to the surface using a vacuum pump, 

and the methane concentration can be monitored continuously. Surface leaks are also less of a 

problem. Finally, the diameter of the hole can be matched to the air quantity requirements. 

A disadvantage of using a second hole for ventilation and sampling is that it can only be used for 

reamed hole diameters of about 8 ft or more. It also requires careful drilling to ensure that the 

holes do not wander too far apart. 
 

 

 

132 

FILLING SHAFTS AT CLOSED MINES 

Filling a shaft at a closed coal mine can be hazardous because 

of methane accumulations in the shaft or at the surface. This 

gas may be ignited by rock dumped into the shaft or by cutting 

torches used to dismantle surface structures such as fan 

housings. The key to maintaining safe conditions is adequate 

methane and barometric pressure monitoring. 

Shaft filling at U.S. coal mines. Under 30 CFR 75.1711–1, the Mine Safety and Health 

Administration (MSHA) requires that shafts be filled with incombustible material or covered 

with a 6-in-thick concrete cap that is equipped with a 2-in vent pipe extending upward 15 ft or 

more.6

6

Under MSHA regulations, shafts must be either filled or capped. Other federal or state agencies may require that 

shafts be filled.
 

 In addition, precautions to deal with methane are necessary during the shaft-filling 

operation. 

Denk et al. [1987] discussed the methods used to monitor methane and the precautions taken to 

ensure worker safety during a shaft filling operation at a U.S. coal mine. At this mine, an explosion occurred as rock was being initially dumped into the 16-ft-diam, 953-ft-deep shaft. Following the explosion, MSHA measured the shaft methane concentration by extending a sampling 

tube down the shaft to the bottom and pumping air samples through the tube to the surface.7 

7

An alternative method, using newer technology, is to lower a data-logging methane detector into the shaft. 

The methane concentration ranged from 2.2% to over 12%.8 

8

MSHA used an infrared analyzer to measure the methane concentration. Infrared analyzers are accurate at high 

methane concentrations and/or low oxygen levels. Bear in mind that methane detection instruments that use heat of
 

combustion sensors are not accurate at methane concentrations above 8% or oxygen concentrations below 10%. 
 

See the sampling chapter (Chapter 2) for more information on the distinction between infrared analyzers and heat of 

combustion sensors. 

At this mine, the most cost-effective way of dealing with this gas was to pump compressed air 

into the shaft to dilute it. Air from a compressor rated at 750 scfm and 100 psi was delivered to 

the shaft bottom through a 2-in PVC9

9

Polyvinyl chloride. 

natural gas line secured with a hemp rope. A copper 

ground wire was attached along the entire length of the gas line to guard against explosion due to 

static electricity. The gas line was to be pulled up as the shaft was filled. After 4 hr of operation 

with this system, methane at the 835-ft level in the 16-ft-diam shaft was reduced from 9.5% 

to 1.4%. 

Subsequently, MSHA specified that while work was being conducted in the shaft area, the 

methane concentration at the bottom of the shaft was to be maintained at 2.0% or less and elsewhere in the shaft at 3.0% or less. Alternating 1-hr periods with the air compressor turned on 

and off was enough to keep the methane concentration within these limits as the shaft was being 

filled. 
 

 

133 

After the shaft was filled, mine gases continued to leak to the surface. Explosive concentrations 

of methane were measured at the surface, so air-powered tools were used to dismantle the fans. 

In addition, elevated carbon dioxide levels and oxygen deficiencies persisted in the fan housing 

and the structures surrounding the shaft collar. Nevertheless, the project was completed without 

further incident. 

Shaft filling in Germany. Hinderfeld [1995] reported on the methane safety precautions taken 

during shaft filling in Germany, where more than 100 shafts had been filled in the previous 

10 years. When shafts are not under the influence of a fan, the flow of methane is controlled by 

the barometric pressure. High flows of methane have been observed when the barometric pressure falls and the gas-laden mine air expands and fills the shaft from below. In this situation, it is 

important to monitor the barometric pressure. 

The preferred approach to shaft filling at each mine is to ensure that the shaft being filled is 

downcast and ventilated with a fan in another shaft. The last two shafts at the mine are then 

closed off simultaneously and filled as quickly as possible. 

During filling, the methane concentration is continuously measured at a point 50 m (164 ft) 

below the surface. If a threshold value—established for that particular shaft—is reached, then 

filling work stops and the shaft is probed to the bottom. 

When high methane concentrations are recorded, and if these are not diluted during a subsequent 

barometric pressure increase, the shaft atmosphere is inerted by adding enough carbon dioxide or 

nitrogen to reduce the oxygen content below 10%.10 

10When a shaft is under the influence of barometric pressure, adding carbon dioxide or nitrogen can be a better 

approach than diluting methane with compressed air. When compressed air is used, a spike in the methane flow due 

to a falling barometer may move the shaft atmosphere into the explosive range. When the shaft is inerted, a spike in 

the methane flow may drive the shaft atmosphere at the bottom further into the inert range. 

REFERENCES 

CFR. Code of federal regulations. Washington DC: U.S. Government Printing Office, Office of 

the Federal Register. 

Cook AP [1998]. The occurrence, emission, and ignition of combustible strata gases in 

Witwatersrand gold mines and Bushveld platinum mines, and means of ameliorating related 

ignition and explosion hazards. Final project report, project GAP 504. Braamfontein, Republic 

of South Africa, Safety in Mines Research Advisory Committee (SIMRAC) (www.simrac.co.za). 

Denk JM, Francart WJ, Baran JN [1987]. Monitoring mine gases during shaft filling operations. 

In: Proceedings of the Third Mine Ventilation Symposium (University Park, PA, October 12–14, 

1987). 
134 

Hinderfeld G [1995]. Ventilation-technical and gas management measures within closing down 

of coal mines. In: Proceedings of the 26th International Conference of Safety in Mines Research 

Institutes (Katowice, Poland, September 4–8, 1995). Vol. 2. Katowice, Poland: Central Mining 

Institute, pp. 201–217. 

Maksimovic SD [1981]. Control of methane by ventilation of shafts during raise drilling. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 8847. NTIS No. PB81234502. 

Mills B [2003]. Fatal shaft explosion report investigation, McElroy Coal Co., McElroy mine, 

Permit No. U–83–33, Central Cambria Drilling Co., contractor No. C–618, January 22, 2003. 

Fairmont, WV: West Virginia Office of Miner’s Health, Safety and Training. 

Oakes JK, Tortorea JS, Stoltz RT, Stephan CR, Brown VF, Penigar RA, et al. [2003]. Report of 

investigation, fatal shaft sinking explosion, January 22, 2003. McElroy mine, McElroy Coal 

Company (subsidiary of CONSOL Energy Incorporated), Cameron, Marshall County, West 

Virginia, ID No. 46–01437. Arlington, VA: U.S. Department of Labor, Mine Safety and Health 

Administration. 
 CHAPTER 10.—METHANE CONTROL IN HIGHWALL MINING 

 

 

135

By Jon C. Volkwein1

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA. 

and Fred N. Kissell, Ph.D.2 

2

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

In This Chapter 

9 How inert gas works to prevent methane explosions 

9 How inert gas is generated and delivered at highwall mines 

9 Volume and quality requirements for inert gas at highwall mines 

9 How an inert gas system is operated 

and 

9 Precautions to take during mining to prevent methane explosions 

This chapter discusses a method, originally developed by Volkwein and Ulery [1993], to prevent 

methane explosions during highwall mining. In highwall mining, a horizontal auger or a mining 

machine enters the coal seam from a surface mine pit at the bottom of a highwall, and the coal is 

mined out from a series of parallel holes. Explosions can be prevented by injecting inert gas into 

each hole as it is mined. 

Coal near the surface has lost most of its methane gas over time. However, in recent years, 

surface mining has been used for deeper reserves of coal. This trend toward mining deeper 

reserves has increased the chance of encountering methane, and methane explosions at highwall 

mining operations have resulted in injuries. 

HOW INERT GAS WORKS TO PREVENT METHANE EXPLOSIONS 

A methane explosion requires the presence of sufficient amounts of both methane and oxygen, 

as well as an ignition source. If the methane cannot be reduced and the ignition source cannot be 

eliminated, then explosions may be prevented by adding an inert gas, which contains little to no 

oxygen, to the mixture [FWQA 1970]. Just how much inert gas must be added depends on the 

mining rate, as well as the composition of the inert gas. 

An explosibility diagram can be used to show whether a methane mixture is explosive after inert 

gas is added [Zabetakis et al. 1959] (Figure 10–1). This diagram indicates that gas mixtures 

fall into one of three range categories—explosive, explosive when mixed with air, and 

nonexplosive—depending on the percentage of methane and percentage of “effective inert.” 

Effective inert is calculated from the percentage of “excess nitrogen”3 and percentage of carbon 

dioxide in the mixture. 

3

The percentage of excess nitrogen is the percentage of nitrogen in the sample minus the percentage of “normal 

nitrogen.” Normal nitrogen is calculated from the ratio of nitrogen to oxygen normally found in air—a factor of 3.8.
 
 

136 

To calculate the effective 

inert, suppose, for example, 

that inert gas is added to a 

methane-air mixture and 

that a gas analysis shows 

that the final mixture has 

10% oxygen, 6% carbon 

dioxide, 6% methane, and 

78% nitrogen. The effective inert is then determined in three steps. First, 

in this example, the oxygen 

percentage is 10%, so the 

percentage of normal nitrogen is 3.8 times 10%, or 

38%. Second, since the 

percentage of excess nitrogen is the percentage of nitrogen in the sample minus the percent of 

normal nitrogen, the excess nitrogen is 78% minus 38%, or 40%. Third, according to the equation shown in Figure 10–1, since the carbon dioxide in the sample is 6%, the effective inert is 

now 40% + (1.5 × 6%), or 49%.4 

4

Carbon dioxide has been found to be 50% more effective than nitrogen in inerting, so a multiplying factor of 1.5 

is used. 

Figure 10–1.—Explosibility diagram for methane-air-inert gas mixtures. 

The point representing a gas mixture containing 6% methane and 49% effective inert is shown in 

Figure 10–1, placing this mixture in the “nonexplosive” range. To minimize the explosion hazard during highwall mining, the objective is to add enough inert gas to keep the final mixture 

well out of the explosive range. 

A handy rule of thumb is that the oxygen content of the mixture must be reliably maintained 

below 12%. Nitrogen-oxygen-methane mixtures with 12% oxygen fall along the dotted line in 

Figure 10–1. Mixtures with less than 12% oxygen fall to the right of this line and are either 

“nonexplosive” or “explosive when mixed with air.” 

The explosibility of mixtures with more than 12% oxygen must be evaluated in the context of 

Figure 10–1. For example, a mixture of 15% methane, 15% oxygen, and 70% nitrogen has 13% 

effective inert, so it falls in the “explosive when mixed with air” range. 

THE INERT GAS SYSTEM 

Preventing explosions on highwall mining machines using inert gas requires a source of inert gas 

and a method to keep the hole completely filled with inert gas as it is mined. As Figure 10–1 

indicates, if an inert gas completely displaces all of the air in the hole, then any gas source having an effective inert concentration of 34% or greater (or an oxygen concentration of 12% or 

less) will prevent methane from igniting. To ensure that all of the air in the hole is displaced, the 
volume of inert gas delivered to the hole must equal or exceed the volume of coal extracted 

from the hole. 

Auger mining. Logical sources of inert gas for auger mining are the auger machine’s diesel 

engine and an auxiliary gasoline engine. The inert gas system that was originally developed by 

Volkwein and Ulery [1993] is shown in Figure 10–2. 

Figure 10–2.—Schematic of inert gas system for auger mining. 

The Volkwein and Ulery system was evaluated at an auger mining operation at a surface mine in 

Kentucky. During initial testing of their system, Volkwein and Ulery found that the oxygen content of the diesel engine powering the auger (a Cummins turbocharged 270) was not always low 

enough to prevent explosions. The diesel engine exhaust oxygen concentrations range from 8% 

at full load to 17% at no load (typical for diesel engines). 

It is known that gasoline engine exhaust has consistently lower oxygen concentrations than 

diesel engine exhaust, so a 305-in3

137 

gasoline engine was used as the primary source of the inert 

gas. This gasoline engine was mounted on the roof of the auger drill, and a new catalytic converter was installed on its exhaust manifold. The catalytic converter burned excess hydrocarbons 

and carbon monoxide, further lowering the oxygen content of the exhaust. This engine was 

operated at 3,600–3,900 rpm. Although the gasoline engine alone produced oxygen concentrations in the 1%–4% range, the exhaust volume was insufficient to keep the hole completely filled 

with inert gas. To make up for this lack of volume, a portion of the diesel engine exhaust was 

added to the gasoline engine exhaust. 

As shown in Figure 10–2, the diesel engine exhaust was conducted from the engine muffler 

(no catalytic converter was used on the diesel) to the roof of the auger drill through a flexible 

5-in steel pipe. This flexibility allowed the engine carriage to travel freely as the auger drill 

moved forward. The volume of diesel exhaust entering the auger hole was controlled by a 

critical orifice restriction, selected to allow equal proportions of diesel and gasoline engine 
138 

exhaust. The remaining diesel exhaust was vented at the roof of the auger drill. The combined 

exhaust flows were routed through a section of 5-in-diam flexible pipe that connected to a vertical descending pipe that brought the exhaust to the level of the hole. A 7-ft length of 3-in stub 

pipe extended the exhaust gas discharge about 5 ft into the hole. 

With this system, inert gas mixture was released at the collar of the hole, reaching the cutting 

head by simple displacement of the extracted coal. The inert gas followed the auger head into 

the hole. As angering proceeded, the hole remained in an inert condition provided that a sufficient volume of inert gas always flowed out of the hole to keep out the surrounding air. 

Ripper-head miners. Adapting inert gas technology from auger miners to ripper-head miners 

involves few changes. Ripper-head systems are in use in Australia with the addition of an automated stub pipe that discharges the inert gas farther into the hole. The inert gas has to be 

discharged farther into the hole because the hole is wider. With a short stub pipe, the wider 

openings may allow equipment movement and external wind to dilute the inert gas before it can 

displace the removed coal. 

An inert gas system designed and used by a mine operator 

does not have to be identical to the one designed by 

Volkwein and Ulery. However, any inert gas generation 

system must deliver an adequate quantity of gas with a 

sufficiently low oxygen concentration. 

SYSTEM REQUIREMENTS AND OPERATION 

Inert gas quantity. In order for inert conditions to be maintained at the cutting head of the 

auger string, the volume of inert gas produced must exceed the volume of coal removed and the 

inert gas must be supplied continuously as augering proceeds. This keeps the surrounding air out 

of the hole. 

During testing of the system designed by Volkwein and Ulery [1993], time studies of coal 

removal showed that auger sections 17 through 27 required an average of 102 sec per cycle. 

Of that cycle time, approximately 20 sec was required for retraction of the kelly bar, leaving 

82 sec for coal removal. The fastest cycle time recorded was 70 sec for coal removal. Each 

added auger section removed a coal cylinder 3.25 ft in diameter by 6 ft long, or 49.7 ft3

 of coal. 

The average coal removal rate was calculated to be 35.0 ft3/min, with a maximum removal rate 

of 42.0 ft3

/min. At greater hole depth (auger numbers 55 through 60), the average removal rate 

was calculated to be 27.0 ft3/min, with a maximum rate of 30.3 ft3/min. Smaller diameter augers 

or slower penetration speeds will decrease this volume, and vice versa. 

Since the engine gas cools and water vapor condenses inside the auger hole, the amount of inert 

gas actually available is the cooled gas, not the hot gas. When the volume of hot gas is measured, a large correction factor must be used to determine the available inert gas volume. 

A correction factor of 0.125 was obtained during the testing, so 0.125 was multiplied by the hot 
 

139 

exhaust gas volume to yield the available inert gas volume.5

5

This seems like a surprisingly large reduction in volume, but much of it is due to water vapor condensation. 

 Hot gas velocities can be measured 

with a pitot tube installed in the gas delivery pipe. The readings must be corrected for air density 

using a value of 0.0415 lb/ft3 to reflect the elevated temperature of about 500 °F. 

For the system tested, the minimum cooled gas volume found during testing with the combined 

engine exhaust was 50 ft3/min. The maximum rate of coal removal was 42 ft3/min. This calculates to a 16% excess volume of inert gas for the worst-case conditions—minimum gas volume 

and maximum coal removal. 

Oxygen concentration. If the oxygen concentration can be maintained at 12% or less, measurement of the oxygen concentration alone is sufficient to indicate the inert condition of the gas. 

These measurements could be made with a handheld oxygen detector or an in-line continuous 

oxygen detector. 

During testing [Volkwein and Ulery 1993], a level of 12% oxygen was maintained along with 

6% carbon dioxide. Since combustion engines always produce carbon dioxide in addition to 

lowering the oxygen level, the presence of carbon dioxide will provide a safety factor if the 

oxygen is 12% or less. 

Placement of the stub pipe and purging the starter hole. For inert gas to be continuously 

maintained at the front of the auger hole, the region just inside the collar of the hole must be 

continuously provided with inert gas. However, when the head and lead guide augers are starting the drilling, there is no room to insert the stub pipe. The ideal time to extend the stub pipe 

into the collar of the hole is after a smaller-diameter auger is attached and the hole is just deep 

enough to make room for the stub pipe (see Figure 10–2). Then the auger is stopped and the stub 

pipe is installed. After the stub pipe is installed, the auger is not rotated until the starter hole is 

purged with inert gas. 

When placing a stub pipe, be certain that it extends at least 

5 ft into the hole. The jet from a shorter stub pipe might 

entrain outside air. 

Purging of the starter hole is necessary because of the air drawn in by the head and lead guide 

augers. The time required to purge the starter hole depends on the volume of the hole and the 

gas flow rate. During testing in the Volkwein and Ulery study, the empty hole volume was 

216 ft3

 (3.25 ft diameter by 26 ft deep) with about one-half of this volume occupied by the auger 

steel and cut coal, leaving 108 ft3

. At an inert gas flow rate of 56 ft3/min, one complete air 

change occurred in less than 2 min. During testing, engines were run for about 4 min to fill the 

starter hole with inert gas before augering proceeded. 

When insertion of the inert gas stub pipe is delayed, deeper starter holes require much longer 

times to become inert. For example, during testing when the hole was augered 44 ft before 

inserting the stub pipe, it took about 12 min to reach inert conditions. By contrast, a 26-ft hole 
140 

required only 4 min. As a result, timely placement of the stub pipe is important so that the starter 

hole does not become too deep before inert gas is added. 

PRECAUTIONS TO TAKE DURING MINING 

An inert gas system will not prevent explosions if it is not operating properly. During mining, 

the operator of the system must ensure that the concentration of engine exhaust gases stays at or 

below 12% oxygen, that the workplace is free of carbon monoxide, and that there is a steady 

movement of gas from the hole. 

The oxygen level in the engine exhaust gases is easily measured 

with a real-time oxygen indicator that has a readout and provides 

a warning if the level rises above 12%. 

Because the exhaust gas from engines contains carbon monoxide, personal exposure to carbon 

monoxide should be monitored. During the testing by Volkwein and Ulery [1993], the highest 

personal exposure to carbon monoxide measured with the passive dosimeters was less than 

20 ppm. This occurred near the auger pin puller. The conveyor side helper had only a trace of 

exposure, and the machine operator had no detectable exposure. The combination of dilution 

and distance from the collar of the hole accounted for the observed low personal exposure to 

carbon monoxide. 

A steady movement of inert gas from the hole will keep out the surrounding air. This requires 

the operator to observe the direction of movement of dust or smoke during periods of maximum 

auger penetration. This direction of movement should always be out of the hole. 

Auger removal. The inert gas system should be left on during auger removal so that the dilution 

of gas in the hole takes place with inert gas rather than with air. However, auger removal is 

usually rapid––about 25 sec per cycle, with 15 sec of that cycle required for pin pulling and 

stacking of the auger. During the testing by Volkwein and Ulery [1993], rapid removal of the 

auger steel volume slightly exceeded the inert gas volume, but this excess was not considered 

to be significant. 

REFERENCES 

FWQA [1970]. Feasibility study on mining coal in an oxygen-free atmosphere. Washington, 

DC: U.S. Department of the Interior, Federal Water Quality Administration. NTIS No. 

PB197446. 

Volkwein JC, Ulery JP [1993]. A method to eliminate explosion hazards in auger highwall 

mining. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 9462. 

Zabetakis MG, Stahl RW, Watson HA [1959]. Determining the explosibility of mine atmospheres. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, IC 7901. 
 

 

141 

CHAPTER 11.—CONTROL OF METHANE IN COAL SILOS 

By Fred N. Kissell, Ph.D.1 

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

In This Chapter 

9 Measuring the gas emission from the coal 

9 Methane at the top of the silo 

9 Methane at the load-out area 

and 

9 Actions taken after a silo explosion 

Methane accumulations in coal silos have resulted in the occasional silo explosion. These can be 

quite violent and dangerous because coal dust adds to the strength of the blast. However, with 

the appropriate precautionary measures, methane accumulations in silos can be greatly reduced. 

Mine Safety and Health Administration (MSHA) regulations at 30 CFR2

2

Code of Federal Regulations. See CFR in references. 

77.201 require that the 

methane content in the air of any coal silo be maintained below 1.0 vol %.3

3

The presence of coal dust reduces the methane lower explosive limit (LEL) value below 5%, and so the safety 

factor from the specification of a 1% value can be less than 5. For more information, see Chapter 12 on dust
 

explosions. 

Also, MSHA 

requires that methane tests be conducted before any electrical equipment is energized, unless a 

continuous monitor capable of deenergizing the electrical equipment is used.4 

4

Equally important, monitor heads must be placed in locations where methane is likely to accumulate. 

Measuring the gas emission from the coal. The first necessary step in dealing with silo 

methane issues is to measure the gas emission from the coal going into the silo. Such measurements allow one to estimate the silo ventilation needs and permit a comparison with the methane 

controls used at other mines that have similar gas levels. 

The gassiness of the coal can be measured by taking conveyor belt grab samples.5

5

This must be done safely, i.e., the belt must be stopped before the sample is removed. 

Matta et al. 

[1978] measured the gas emission from conveyor belt grab samples using a simple desorption 

test. To conduct the test, they collected several grab samples of coal, weighing a few pounds 

each, from the conveyor belt entering the silo.6

6

Most of these mines had an overall mine emission rate exceeding 1million ft3 per day, placing them in the ranks of
 

the gassiest U.S. mines.
 

 They then sealed the coal into an airtight sample 

container that was equipped with a valve and short hose along with a pressure gauge. Every few 

hours they opened the valve and bled the emitted gas into a water-filled graduated cylinder that 

had been inverted and placed in a pan of water (Figure 11–1). The results are shown in Figure 

11–2. 
 

 

142 

Figure 11–1.—Desorption test apparatus. 

Figure 11–2.—Gas emitted from conveyor belt grab samples. 

LaScola et al. [1981] also 

collected conveyor belt 

grab samples. Since the 

coal in the silos they investigated was usually stored 

in the silo for about 24 hr, 

they used the 24-hr cumulative emission value as a 

comparative index. These 

24-hr emission rates ranged 

from 3.3 to 86 ft3/ton. 

LaScola et al. noted that all 

of the mines with 24-hr 

emission values exceeding 

14 ft3

/ton had open-top 

silos to provide better ventilation at the top. Also, 

most had forced ventilation 

of the reclaiming areas, 

at ventilation rates ranging 

from 5,600 to 20,000 cfm. 

Kolada [1985] reported 

similar gas amounts from 

silo conveyor belt samples 

at Canadian mines.7  

 

7

More information on this study is available from AMCL [1985]. 

Interestingly, grab samples 

from coal entering clean 

coal silos sometimes gave 

emission rates five times 

higher than coal entering 

raw coal silos. This is not 

what one would expect 

since clean coal has been 

out of the mine longer. 

However, in these 

instances, the clean coal had been passed through a fluidized bed dryer to remove moisture and 

its temperature was 40 °C as it was being loaded into the silo. The higher temperature greatly 

increased the methane emission rate.8 

8

The breakage of coal during the cleaning process could also have been a contributing factor to the elevated methane 

emission. Friable coals will fracture into smaller size particles during coal cleaning, and smaller size particles will 

give off methane more quickly [Mikhail and Patching 1980]. 
 

 

143 

Closed-top silos should always be recognized as a potential 

methane problem. Provision must be made for continuous 

mechanical ventilation if the silo has a closed top. 

Methane at the top of the silo. LaScola et al. [1981] also measured the silo gas concentration 

above the stored coal pile at a wide variety of coal mines. Both open-top and closed-top silos 

were visited. Open-top silos allow large air movements above the coal pile, reducing the hazard 

of a methane explosion at the top. However, dust emissions during silo loading can be a problem. Closed-top silos are usually ventilated by openings at the top of the silo. These are typically 1- by 2-ft holes spaced around the perimeter immediately below the concrete roof. Some 

closed-top silos have ventilation fans or dust collectors. 

In the LaScola et al. study, gas measurements of the open space above the coal pile were 

conducted by lowering flexible plastic tubing down the center line of the silo and pumping the 

gas to a methane detector. Measurements were made at 10-ft increments until the coal pile was 

reached. The results are shown in Figure 11–3. Methane concentrations were not excessive, and 

there was no layering9 of methane at the top of the silos. 

9

The lack of layering is not surprising since the coal pile where methane is released is below the silo roof. Studies 

on methane layering in mine entries have usually measured mine roof layers caused by methane released at the roof 

of the mine. When methane is released at the mine rib, the tendency to layer at the roof is much less, and the tendency to layer at the roof is even less so for methane released at the mine floor. Moreover, once methane is mixed 

into the air, it does not unmix to form layers. See the discussion on layering in Chapter 1. 

 Figure 11–3.—Methane concentration gradient above the coal in 

coal silo. 

Kolada [1985] conducted measurements above the coal pile in Canadian coal silos following a 

similar procedure. The methane found was within a few inches of the coal, where the concentration ranged from 0% to 2% methane. When the silo was discharging, the methane concentration 

within a few inches of the 

coal ranged from 0% to 6% 

methane. 

Methane at the load-out 

area. The load-out area at the 

bottom of the silo is always a 

potential location for methane 

accumulations. These 

accumulations may increase if 

the coal has been stored for 

longer than normal periods. 

Methane detectors and 

adequate ventilation must be 

provided. Electrical 

switchgear should be 

minimized in load-out areas, 

especially if the load-out is 
 

 

 

144 

enclosed in a tunnel-like structure. Special attention should be given to railroad load-outs when 

electrical locomotives are used because of the additional ignition source. 

At the bottom of every silo, the methane emission should be 

measured as coal is reclaimed, and mechanical ventilation 

should be provided if there is any likelihood of methane buildup. 

Although methane measurements taken during the reclaiming of coal are valuable, the values 

obtained only reflect the circumstances at the time. These measurements can be supplemented 

by an estimation of ventilation requirements calculated from the gas concentration in the 

coal pile. 

Gas concentration inside the coal pile has been measured directly and also calculated from the 

coal emission measurements. In the study by Kolada [1985], tubing was extended down into 

several silos, where it was buried with coal as the silo was filled. At the same time, a conveyor 

belt grab sample was taken and the emission from the grab sample was measured. At one silo, 

the conveyor belt grab sample emitted 0.013 L/kg in the first 30 min. The coal pile was known 

to have a bulk density of 800 kg/m3

 and 41% void space, so the amount of gas given off by a 

cubic meter of coal pile was 0.013 × 800, or 10.4 L of methane. Next, the concentration of 

methane in the coal pile was calculated to be equal to the volume of methane divided by the volume of air plus methane, or 2.5%.10

10If 41% of the coal pile is void space, the void space in the cubic meter would be 410 L and the concentration of 

methane in the void space would be 10.4/(410 + 10.4), equal to a calculated concentration value of 2.5%. 

The measured value, which Kolada obtained by pumping air 

from a tube buried in the coal pile for 30 min, was about the same as this calculated concentration value. 

Using the above approach at several silos, Kolada obtained methane concentration values as high 

as 35%. However, for any given silo the concentration will depend on both the emission rate and 

the amount of time the coal remains in the silo. 

During the reclaiming of coal, methane gas in the void space will emerge into the coal discharge 

gallery. Kolada has given a sample calculation, assuming a peak coal discharge rate of 1,021 

k in 

the void space.

g/sec, a bulk density of 800 kg/m3

, 41% void space, and a methane concentration of 35%11

11A value of 35% may seem high, but Kolada and Chakravorty [1987] measured methane concentrations as high as 

40% in a silo coal pile within an hour of filling the silo. These concentrations are not in the flammable range, but 

will become so when mixed with air. See the discussion on flammability in Chapter 1. 

 A discharge rate of 1,021 kg/sec corresponds to 1,021/800 = 1.28 m3/sec. The 

methane discharged is then 1.28 × 0.41 × 0.35 = 0.184 m3/sec = 389 cfm. Reducing this flow of 

methane to a 1% concentration will require an airflow of 38,900 cfm. 

Actions taken after a silo explosion in British Columbia. Stokes [1986] reported on the 

actions taken after an explosion at a closed-top silo in British Columbia, Canada. These 

postexplosion actions serve as a good model for mines desiring to prevent a methane explosion 

in a coal silo. 
The coal had been surface mined. Surface-mined coal normally has a very low methane emission; however, the coal had been heated in a dryer12

12Hot coals from the dryer were a probable source of the ignition. Hot coal detection systems can flag this problem
 

and should be installed if the coal is gassy.

to remove moisture just before being loaded 

into the silo. The 24-hr conveyor belt grab sample emission was measured at 14 ft3

/ton. Before 

the explosion, the top was ventilated with a 7,500-cfm wet dust scrubber system that operated 

only during loading. An unworkable13 

13The idea behind a natural ventilation methane stack is to use the density difference between methane and air to
 

produce a chimney effect that ventilates the silo. However, neither the stack height nor the methane concentration at
 

any silo would ever be high enough to make a natural ventilation methane stack function.

natural ventilation methane stack was located on the silo 

roof. There was also a methane detector at the roof of the silo (probably in a location where the 

methane did not accumulate). 

After the explosion, the silo was put back into operation with these new methane control and 

damage prevention measures: 

•	 Continuous ventilation was provided at the top. A 14,000-cfm dust scrubber system14 

14Dust problems precluded the conversion to an open-top silo. 

operated when the silo was loading. When loading stopped, another fan, a 20,000-cfm 

forcing fan, automatically turned on, and this fresh air was deflected downward toward 

the coal surface. 

•	 Other openings at the top were provided to supply fresh air in the event of fan failure. 

•	 A new methane monitoring system that used several sensing heads was installed. Using 

several heads reduced the chance that a methane accumulation would be missed. 

•	 A large portion of the roof was provided with a lightweight sheet metal cover that could 

provide some explosion relief without damage to the main structure of the silo.15 

 

 

145 

15The National Fire Protection Association has a guide for venting of deflagrations [NFPA 1998], i.e., methane and
 

dust explosions. 

These measures provided for continuous ventilation in a quantity matched to the gas level and 

provided for monitoring in locations where methane was likely to accumulate. They considerably reduced the chance of a silo explosion. 

REFERENCES 

AMCL [1985]. Survey of methane accumulation within coal silos. Calgary, Alberta, Canada: 

Associated Mining Consultants Ltd. 

CFR. Code of federal regulations. Washington DC: U.S. Government Printing Office, Office of 

the Federal Register. 
146 

Kolada RJ [1985]. Investigation into methane accumulation in coal storage silos. SME preprint 

85–346. Littleton, CO: Society for Mining, Metallurgy, and Exploration, Inc. 

Kolada RJ, Chakravorty RN [1987]. Controlling the hazard of methane explosions in coal storage facilities. In: Mutmansky JM, ed. Proceedings of the Third Mine Ventilation Symposium 

(University Park, PA, October 12–14, 1987), pp. 334–339. 

LaScola JC, Matta JE, Kissell FN [1981]. Assessing the methane hazard of gassy coals in storage silos. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8525. NTIS No. 

PB81226920. 

Matta JE, LaScola JC, Kissell FN [1978]. Methane emissions from gassy coals in storage silos. 

Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8269. NTIS No. 

PB277912. 

Mikhail MW, Patching TH [1980]. Size degradation of bituminous coal from western Canada. 

CIM Bull May:105–110. 

NFPA [1998]. NFPA 68: Guide for venting of deflagrations. Quincy, MA: National Fire Protection Association. 

Stokes DA [1986]. Fording Coal Ltd. silo explosion. CIM Bull 75(891):56–60. 
 

147 

CHAPTER 12.—EXPLOSION HAZARDS OF COAL DUST 

IN THE PRESENCE OF METHANE 

By Kenneth L. Cashdollar1

1

Research physicist. 

 and Michael J. Sapko2 

2

Research physical scientist. 

Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, Pittsburgh, PA. 

In This Chapter 

9 Methane ignition as initiation source for much larger secondary coal dust explosions 

9 Rock dusting requirements to prevent coal dust explosions 

9 Dangers of hybrid mixtures of methane and coal dust 

Although methane explosions are dangerous, those that involve coal dust are even more so. 

If exploding methane disperses and ignites the coal dust that has accumulated on the mine ribs 

and floor, the burning coal dust immeasurably increases the strength of the explosion. Such 

methane-dust explosions are prevented by inerting the coal dust in a way that prevents the 

exploding methane from igniting it. This chapter discusses the dust hazard and how it is prevented in U.S. coal mines. 

METHANE IGNITION AS INITIATION SOURCE FOR MUCH LARGER 

SECONDARY COAL DUST EXPLOSIONS 

The typical scenario for coal mine explosions starts with the ignition of a flammable methane-air 

atmosphere near the face. The turbulent winds from the primary methane explosion then disperse the coal dust. If there is insufficient rock dust (usually limestone), a secondary coal dust 

explosion then propagates throughout large sections of the mine. These scenarios have been 

studied extensively at the Bruceton Experimental Mine (BEM) and the Lake Lynn Experimental 

Mine (LLEM) of the NIOSH Pittsburgh Research Laboratory. 

The minimum quantity of methane required to initiate a coal dust explosion was studied in 1930 

in the BEM [Rice et al. 1933; Nagy 1981] and then later in the LLEM, whose cross-sectional 

area (130 ft2

) is over twice that of the BEM [Sapko et al. 1987a]. Studies conducted in the BEM 

closely simulated conditions that existed in operating mines in the early 20th century. The later 

tests in the 20-ft-wide entries of the LLEM simulated the geometries of modern mines with 

advanced roof support technology. 

The data from the BEM tests show that 13 ft3

 was the minimum quantity of methane at the face 

that, when ignited, would disperse and ignite coal dust. In the BEM, this amount of methane was 

mixed with air to form a total flammable volume of about 140 ft3

 of a 9% methane-in-air mixture. In the wider entries of the LLEM, about 37 ft3

 of methane was required to disperse pure 

coal dust and start a self-sustained coal dust explosion. This amount of methane was mixed with 
 

148 

air in a 6-ft by 9-ft by 6.5-ft high plastic containment zone to form a total flammable volume of 

about 350 ft3

 of a 10% methane-air mixture. Based on the 54-ft2 cross-section of the BEM and 

the 130-ft2

 cross-section of the LLEM, both of these methane-air volumes would correspond to a 

linear distance of about 2.5 ft from the face. 

Although the entire cross-sections are used for this comparison, the actual methane-air zones in 

both the BEM and LLEM only partially filled the cross-sections. If obstacles were added to the 

methane zones to create turbulence, the methane explosions could be even more effective at dispersing coal dust. It should also be noted that these experimental mine tests were somewhat 

idealized conditions because there was no rock dust mixed with the coal dust, as would normally 

be the case under real mining conditions. In addition, the dust was all on shelves near the roof. 

This arrangement provided the easiest conditions for initiating a dust explosion. In a mine with 

rock dust added to the coal dust and with most of the dust deposited on the floor, more methane 

would be required to disperse and ignite the dust mixture. 

ROCK DUSTING REQUIREMENTS TO PREVENT COAL DUST EXPLOSIONS 

The primary method of preventing coal dust explosions in underground mines is to add sufficient amounts of an incombustible rock 

dust (usually limestone) to the coal dust. Then, even if the coal and 

rock dust mixture is dispersed into the air by a methane explosion, 

a secondary dust explosion will not occur. The rock dust acts as a 

heat sink to cool the explosion temperature below the temperature 

needed for continued propagation. 

30 CFR3

3

Code of Federal Regulations. See CFR in references. 

75.402 to 75.404 requires rock dusting in all underground bituminous coal mines: 

All underground areas of a coal mine, except those areas in which the dust is too wet or 

too high in incombustible content to propagate an explosion, shall be rock dusted to 

within 40 feet of all working faces...Where rock dust is required to be applied, it shall be 

distributed upon the top, floor, and sides of all underground areas of a coal mine and 

maintained in such quantities that the incombustible content of the combined coal dust, 

rock dust, and other dust shall be not less than 65 per centum, but the incombustible content in the return aircourses shall be no less than 80 per centum. 

The higher incombustible content required for returns is based on the finer size of coal dust 

found in returns. The regulations further state that— 

Where methane is present in any ventilating current, the per centum of incombustible 

content of such combined dusts shall be increased 1.0 and 0.4 per centum for each 

0.1 per centum of methane where 65 and 80 per centum, respectively, of incombustibles 

are required. 
 

149 

The above paragraph means that the incombustible content of the dust in intakes must be 

increased from 65% to 75% if the ventilating air contains 1% methane. Similarly, the incombustible content of the dust in returns must be increased from 80% to 84% if the ventilating air 

contains 1% methane. The incombustible content of the dust mixture includes the rock dust, 

the ash content of the coal dust, and the moisture content. These regulations are based on 

research conducted at the BEM [Rice and Greenwald 1929; Rice et al. 1933; Nagy 1981]. 

There is an additional hazard when the rock dust is not well mixed with the coal dust. If there is 

a thin layer of float coal dust (dust that has been carried and deposited by the ventilation air) on 

top of a thick layer of properly rock-dusted floor dust, a weak methane explosion may preferentially lift the top layer of coal dust. The explosion can then continue to propagate through and 

beyond the length of the float coal dust deposit. This has been demonstrated in full-scale experimental mine tests [Nagy et al. 1965; Sapko et al. 1987b]. 

DANGERS OF HYBRID MIXTURES OF METHANE AND COAL DUST 

Chapter 1 discusses the lower flammable limit (LFL)4

4

Also called the lower explosive limit (LEL). 

for methane and some of the ways that 

methane can be ignited. With only methane present, the LFL is 5% methane in air. However, 

when coal dust is added to the methane-air mixture, the LFL of the mixture is reduced. This can 

occur at the mining face, where methane is being liberated and coal dust is being generated. 

The Le Chatelier linear mixing law5

5

See the section entitled “Addition of other flammable gases to air” in Chapter 1.
 

for the LFL of gases is also roughly applicable to a hybrid 

mixture of methane gas and coal dust. Figure 12–1 shows flammable limit data from BEM tests 

[Cashdollar et al. 1987] for coal 

dust dispersed with methane. 

Additional laboratory data [Cashdollar 1996; Cashdollar et al. 

1987] confirm the roughly linear 

relationship for various coal dusts 

mixed with methane. In the 

example in Figure 12–1, the LFL 

of methane alone is 5% and the 

LFL of the coal dust by itself is 

0.10 oz/ft3

. It should be noted that 

this is only an example and that 

the LFL of various coal dusts will 

vary with the particle size and 

volatility. The area below and to 

the left of the dashed line in the 

figure represents nonflammable 

mixtures. The area above and 

to the right of the dashed line Figure 12–1.—Flammable limits for mixtures of methane and 

coal dust. 
150 

represents flammable mixtures. For example, if 3.5% methane is present, then any coal dust 

concentrations above 0.03 oz/ft3

 are flammable when mixed with the methane. If 2% methane is 

present, then any coal dust concentrations above 0.06 oz/ft3

 are flammable. This means that, 

if sufficient coal dust is present, an ignition can still occur at the mine face even if the methane 

concentration is below the 5% LFL. 

REFERENCES 

Cashdollar KL [1996]. Coal dust explosibility. J Loss Prev Process Ind 9(1):65–76. 

Cashdollar KL, Sapko MJ, Weiss ES, Hertzberg M [1987]. Laboratory and mine dust explosion 

research at the Bureau of Mines. In: Industrial dust explosions. ASTM Special Technical 

Publication (STP) 958. West Conshohocken, PA: American Society for Testing and Materials, 

pp. 107–123. 

CFR. Code of federal regulations. Washington DC: U.S. Government Printing Office, Office of 

the Federal Register. 

Nagy J [1981]. The explosion hazard in mining. Arlington, VA: U.S. Department of Labor, 

Mine Safety and Health Administration, Informational Report (IR) 1119. 

Nagy J, Mitchell DW, Kawenski EM [1965]. Float coal hazard in mines: a progress report. 

Washington, DC: U.S. Department of the Interior, Bureau of Mines, RI 6581. 

Rice GS, Greenwald HP [1929]. Coal dust explosibility factors indicated by experimental mine 

investigations, 1911 to 1929. Washington, DC: U.S. Department of the Interior, Bureau of 

Mines, Technical Paper (TP) 464. 

Rice GS, Greenwald HP, Howarth HC [1933]. Explosion tests of Pittsburgh coal dust in the 

experimental mine, 1925 to 1932 inclusive. Washington, DC: U.S. Department of the Interior, 

Bureau of Mines, Bulletin 369. 

Sapko MJ, Hertzberg M, Watson RW, Cashdollar KL [1987a]. Dust explosion size scaling. In: 

Proceedings of the Shenyang International Symposium on Dust Explosions (Shenyang, China, 

September 14–16, 1987). 

Sapko MJ, Weiss ES, Watson RW [1987b]. Explosibility of float coal dust distributed over a 

coal-rock dust substratum. In: Proceedings of the 22nd International Conference of Safety in 

Mines Research Institutes (Beijing, China, November 2–6, 1987). 
CHAPTER 13.—METHANE CONTROL IN METAL/NONMETAL MINES 

 

151 

By H. John Head, P.E., C.Eng.,1

1

Senior principal engineer, Continental Placer, Inc., Wheaton, IL. 

and Fred N. Kissell, Ph.D.2 

2

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

In This Chapter 

9 Gas reports from around the world 

9 Regulations for gassy mines in the United States 

9 Differences between metal/nonmetal mines and coal mines 

9 Monitoring for methane and taking action 

9 Diluting methane with additional ventilation 

9 Eliminating ignition sources 

9 What experienced mine operators say about methane control 

and 

9 Looking for methane when starting a new mine or expanding an existing mine 

This chapter gives guidelines for preventing methane gas explosions during metal and nonmetal 

mine development and subsequent production operations.3 

3

Those mines extracting metallic ores such as copper or nickel are referred to as “metal mines.” Those mines 

extracting nonmetallic minerals such as salt, potash, trona, and limestone are referred to as “nonmetal mines.” 

Emphasis is placed on recognizing 

the differences between coal mines, where the potential for methane hazards is relatively well 

understood, and metal/nonmetal mines, where methane may accumulate unexpectedly. Also, 

interviews with experienced mine operators add much to a complete understanding of what must 

be done to address methane problems in metal/nonmetal mines. 

METHANE GAS IN METAL/NONMETAL MINES 

Gas reports from around the world. The presence of methane gas in metal/nonmetal mines 

around the world is more common than one might imagine [Edwards and Durucan 1991]. 

For example: 

•	 The former Soviet republics have occurrences of methane and hydrogen in apatite, gold, 

and diamond ores, where solid or liquid bitumen occurs in the rock. 

•	 Scandinavian iron ore deposits include methane and other hydrocarbons in boreholes that 

intersect pitch and asphalt within the deposits, and methane and nitrogen in boreholes and 

fissures in arsenic and sulfide ores. 
152 

•	 In Eastern Europe, petroleum and gas has been observed in igneous and metamorphic 

rocks in Yugoslavia, in some copper mines in Hungary, and in mica schists containing 

limestone intrusions in Romania. 

•	 In the United Kingdom, granites in Cornwall and Aberdeen and iron ore deposits in 

Cleveland all report hydrocarbon gases associated with overlying bituminous shales. 

Also, Derbyshire lead mines have reported methane along with bitumen. 

•	 Canadian Shield mines contain methane, other hydrocarbons, and sometimes hydrogen 

and helium [Fritz et al. 1987; Andrews 1987]. These are widespread and occur in almost 

all the mines, particularly where carbonaceous materials are found in the rocks. The 

emissions are usually associated with boreholes [Sherwood et al. 1988] and are relatively 

short-lived and easily dissipated. The Kidd Creek mines have methane pockets associated with sulfide deposits. At some mines, the occurrences of methane and hydrogen 

increase with depth, and the resulting gas mixtures reduce the lower explosive limit to as 

low as 4.5%.4

 

 

4

For more information on the lower explosive limit, see Chapter 1.

The Ontario Ministry of Labour (OML) has approximately eight reports 

per year of combustible gas in an underground working place [OML 1996]. These 

reports are almost always for boreholes, with measured concentrations of 0.1%–10%. 

Gas is very seldom detected in the general body of the mine’s atmosphere, although 

methane ignitions due to cigarette smoking and friction between metal and sandstone 

have been reported to the OML. 

•	 U.S. mines report methane emissions associated with oil shales, salt, trona, potash, limestone, copper, and uranium ores. 

•	 In Australia, hydrocarbon gases are reported from copper mines and from Precambrian 

rocks at Kalgoorlie. The usual type of methane encounter is a diamond drill blower and 

methane is readily dispersed. 

•	 The Republic of South Africa has combustible gases in almost all gold and platinum 

mines, as well as kimberlite pipes. Along with the methane, there can be hydrogen and 

helium. The usual assumption is that the methane is associated with overlying Karoo 

strata, which are coal-bearing [Searra 1990; Eschenburg 1980; Jackson 1957]. The gas is 

transported downward through the rock dissolved in water. 

REGULATIONS FOR GASSY METAL/NONMETAL MINES 

IN THE UNITED STATES 

The United States developed new federal standards for controlling methane hazards in metal/ 

nonmetal mines in 1985. These are contained in 30 CFR5

5

Code of Federal Regulations. See CFR in references. 

 57, Subpart T—Safety Standards for 

Methane in Metal and Nonmetal Mines. Considering that there is such a wide variety of metal/ 

nonmetal mines in the United States, these standards are quite comprehensive and detailed. 
Because of this, any discussion of controlling methane must first begin with a discussion of the 

regulations and their history. 

The impetus for the revision to the standard, which previously had been based on the simple 

observation and measurement of methane in the mine atmosphere, was the Belle Isle Mine 

disaster of 1979 [Plimpton et al. 1979]. The Belle Isle Mine was an underground salt mine in a 

salt dome in southern Louisiana. The salt domes are known for their proximity to petroleum 

production facilities, with oil and gas often found in the sedimentary structures adjacent to the 

sides of the up-thrusting salt domes. The mine had produced gas intermittently for many 

years since it was opened in 1962. There had also been “outbursts”6

 

153 

6

An outburst is a sudden, violent release of solids and high-pressure occluded gases, including methane, in a domal 

salt mine (30 CFR 57.22002).

of salt found after regular 

production blasts. What was not understood at the time was the mechanism for the release of 

huge quantities of methane gas from these outbursts. When postblast crew members went down 

into the mine after a blast at Belle Isle, an ignition source, possibly the diesel pickup truck that 

they were riding in, set off a massive explosion, killing all five members of the crew underground at the time. 

Earlier, a gas explosion at the Cane Creek potash mine in Utah had occurred in 1963 wherein 

18 miners were killed during development operations [Westfield et al. 1963]. Several of these 

miners survived the initial explosion itself, only to die in a barricaded dead-end drift when their 

oxygen supply ran out. 

Significantly, neither Belle Isle Mine nor Cane Creek 

Mine had reached the threshold of 0.25% methane in 

the general atmosphere of the mine (as required by 

the regulations of the time). Therefore, neither was 

considered to be a “gassy mine.” 

There is a wide variety of metal/nonmetal mines, with many different ways in which methane is 

released into the mine atmosphere. To address the numerous mine-specific potential methane 

hazards, the Mine Safety and Health Administration (MSHA) defined various categories of gassy 

mines (30 CFR 57.22003) in the 1985 federal standards, as summarized below.7 

7

A precedent for developing gassy mine standards in this manner that took into account the different hazards associated with differing methane gas occurrences was found in Spanish mining regulations [Lumsden and Talbot 1983]. 

Category I applies to mines that operate within a combustible ore body and either liberate 

methane or have the potential to liberate methane. Within Category I, there are several subcategories, depending on the actual presence of methane gas (at 0.25% or more) or the occurrence of an ignition (Subcategory I–A) or not (Subcategory I–B). Subcategory I–C is intended to 

include the potential hazard from flammable dust. Category I applies mainly to oil shale and 

gilsonite mines. 
 

154 

Category II applies to domal salt 

mines where the history of the mine 

or geological area indicates the 

occurrence of or potential for an 

outburst. As with Category I, there 

are two subcategories, depending on 

the occurrence of an outburst that 

released 0.25% or more of methane 

(Subcategory II–A) or not (Subcategory II–B). 

Category III applies to mines in 

which noncombustible ore is 

extracted and which liberate a 

concentration of methane that is 

explosive, or is capable of forming 

explosive mixtures with air, or has 

the potential to do so based on the 

history of the mine or the geological area in which the mine is located. 

The flammability of the gas is determined by its position on Figure 

13–1, an illustration contained at 

30 CFR 57.22003(a)(3). Category 

III applies mainly to trona mines. 

Figure 13–1.—Relation between quantitative composition 

and explosibility of mixtures of methane and air (from 30 CFR 

57.22003(3)). 

Category IV applies to mines in which noncombustible ore is extracted and which liberate a 

concentration of methane that is not explosive or capable of forming explosive mixtures with air. 

This somewhat unusual concept derives from the fact that New Mexico potash mines have 

methane contained within the clay and shale seams in the strata along with high percentages 

of inert nitrogen. The flammability of this gas mixture is determined by its position on 

Figure 13–1. Category IV applies mainly to potash mines. 

Category V applies mainly to petroleum mines. 

All mines that are not placed in any of the above categories or subcategories are considered to be 

Category VI, or nongassy, mines. 

Each category (or industry sector) has its own set of requirements for monitoring and control 

measures. Categories I, III, and V most closely match coal mining standards, with fully permissible equipment in production settings. Category II recognizes that gas is only likely to be 

liberated in hazardous quantities during drilling, cutting, and blasting, so those activities are 

controlled. Category IV has few limitations, with monitoring of gas being the primary 

requirement. 
 

155 

DEALING WITH METHANE IN METAL/NONMETAL MINES 

Dealing with methane in metal/nonmetal mines requires an 

understanding of five important issues: 

1. The differences between coal mines and metal/nonmetal 

mines, and recognizing why explosions happen even 

with low gas emission rates. 

2. How to monitor for gas and what gas concentrations 

require action. 

3. The importance of continuously diluting methane with 

ventilation air. 

4. The importance of eliminating all ignition sources. 

5. Avoiding outburst hazards. 

Although the regulations discussed in the previous section are important to preventing methane 

explosions in metal/nonmetal mines, they represent only a starting point in achieving a safe 

mine. A broader understanding of five important issues is necessary, which are detailed below. 

1. Why metal/nonmetal mines are different from coal mines. Unlike coal mines, methane 

emission rates in metal/nonmetal mines are not consistent. This irregularity often makes an 

accumulation of methane an unexpected event, and an unexpected event by definition is difficult 

to anticipate.8 

8

In the manufacturing industry, quality inspections have recognized the problem of unexpected events, which is why 

deliberately faulty parts are slipped into an inspection line to keep the inspectors on their toes. 

Methane can be detected in coal mines everywhere and almost all the time; therefore, monitoring becomes a regular pattern of activity. Ventilation controls are rigorously maintained, and large quantities of ventilation air are blown through the mine to sweep the gas away. 

Permissible equipment is used to minimize ignition sources, and workers are constantly on notice 

that coal mines are potentially dangerous places. All of these factors lead to a constant awareness of the potential methane hazard, promoting consistent efforts to reduce the risk of an ignition or explosion. 

By contrast, workers in metal/nonmetal mines may never detect methane gas or only encounter it 

infrequently. It is easy to become complacent, testing for gas in a cursory or offhand fashion, 

or not even bothering to test at all. Also, the same attention to ventilation controls is lacking, 

with series9

 ventilation circuits and recirculation being common practices. 

9

In series ventilation, the return from one working area is used as the intake to the next one downstream. 

See Chapter 14 for a parallel discussion on ventilation 

controls. Poor ventilation allows for a dangerous 

accumulation of gas in a location where little to no 

gas had been previously detected. 
 

156 

2. Monitoring for gas and taking control measures. 

An important part of methane monitoring is knowing 

what control measures to take when gas is detected. 

The actions described here for methane testing are 

only a summary. For complete details, consult the 

Code of Federal Regulations. 

Metal/nonmetal mines with a history of gas emissions are already on notice that they have a 

methane problem. The U.S. regulatory standards have very specific monitoring requirements 

and actions to be taken depending on the category of mine and its methane history. These provide a good, commonsense approach to methane monitoring and control measures. 

Those mines that liberate significant quantities of gas must monitor for it each shift in a preshift 

examination similar to that for coal mines (30 CFR 57.22226 and 57.22228). No testing is mandated for those mines where methane is not expected to be present, but there are action levels and 

prescribed actions when gas is detected at certain levels,10

10Keep in mind that one needs to regularly test for gas only in the mine atmosphere, not in boreholes. Chapter 2 

contains more information on sampling for methane. 

as follows: 

Actions at 0.25% methane: If the mine had never before measured 0.25% or more or never had an 

ignition, then changes must be made to improve ventilation, and MSHA must be notified immediately. 

Actions at 0.5% methane: Ventilation changes are required to reduce methane below 0.5%. 

In the meantime, depending on the category of mine, one or more of the following are necessary: 

electrical power must be deenergized, diesel equipment must be shut off or removed, and/or 

work must stop. 

Actions at 1.0% methane: Ventilation changes are required to reduce the methane. In the 

meantime, depending on the category of mine, one or more of the following are necessary: 

all workers from the affected areas must be withdrawn except those needed to make the ventilation changes, electrical power must be deenergized, and/or diesel equipment must be shut off or 

removed. 

All persons must be withdrawn from the mine if the 1.0% accumulation results from a main fan 

failure or if 1.0% is measured at a main exhaust fan. 

Actions at 2.0% methane: Ventilation changes are required to reduce the methane, and all persons must be withdrawn except those necessary to make the ventilation changes. Depending on 

the category of mine, one or more of the following are necessary: MSHA must be informed, the 

methane must be reduced to below 0.5%, and/or the methane must be reduced to below 1.0%. 

For all of the above scenarios, the mine category also impacts (1) the frequency and location of 

methane testing, (2) the use of atmospheric monitoring systems, (3) the use of methane monitors 

on mining equipment, and (4) whether explosion-proof electrical equipment is used, among other 

factors. 
 

 

157 

In Canada, Ontario has similar requirements in Section 35 of its regulations for mines and 

mining plants under the Ontario Occupational Health and Safety Act [1990],11

11Section 35 applies wherever mining is being carried out and methane is likely to be present. 12More information on controlling with methane layers is presented in Chapter 1. 

 but with a few 

additional features (such as the need to provide written instructions), which makes Section 35 

worthwhile to read. It is reproduced in Appendix A of this chapter. 

Chapter 2 gives more information on sampling 

for methane. 

3. Diluting the gas with more ventilation. Gas only presents a flammability or explosion problem in the explosive range: 5%–15% methane. If it is diluted with sufficient air, then it ceases 

to be an immediate hazard. The emphasis on adequate ventilation to dilute methane is specifically mentioned by knowledgeable mining operators. 

In addition to diluting methane, it is important that the potential for layering of methane gas be 

eliminated.12 Thus, fans must not only add air into the general body of the mine atmosphere, but 

must also stir up the air within a roadway or heading. Gas can collect in cavities in the roof or in 

the end of an inclined ramp or at the top of a raise, and the ventilation must be directed to stir this 

gas up and dilute it into the body of the mine atmosphere. 

In the U.S. standards for gassy metal/nonmetal mines, several sections address airflow 

requirements: 

30 CFR 57.22213 – Air flow (Category III mines).—The quantity of air coursed through 

the last open crosscut in pairs or sets of entries, or through other ventilation openings 

nearest the face, shall be at least 6,000 cubic feet per minute, or 9,000 cubic feet per minute in longwall and continuous miner sections. The quantity of air across each face at a 

work place shall be at least 2,000 cubic feet per minute. 

This standard for Category III gassy mines includes all underground trona mines. These mines 

may experience gas emissions on a regular basis, so the standard provides a useful guideline for 

quantities and for the importance of those quantities to be directed at each workplace. 

Another method of specifying airflow has been adopted in the standard for oil shale mines: 

30 CFR 57.22211 – Air flow (Category I–A mines).—The average air velocity in the last 

open crosscut in pairs or sets of developing entries, or through other ventilation openings 

nearest the face, shall be at least 40 feet per minute. The velocity of air ventilating each 

face at a work place shall be at least 20 feet per minute. 
These mines typically have large openings, e.g., 25 ft high by 50 ft wide, and therefore slow air 

velocities despite relatively large air volumes. In this case, the standard recognizes that actual air 

velocity, rather than the total volume of air flowing in a mine roadway, provides the necessary 

turbulence to remove methane gas layers. A 20 ft/min flow in a 25-ft by 50-ft roadway translates 

to a total air volume of 25,000 ft3

/min.13

 

158 

13Air volume (ft3

/min) = Air velocity (ft/min) × roadway area (ft2). 

 Obviously, a much smaller 6,000 ft3/min is required in 

the trona mines, which typically have smaller roadways, say, 12 ft high by 20 ft wide. In this 

case, the 6,000 ft3

/min translates to a velocity of 25 ft/min. 

A more general standard exists for mines experiencing gas emissions on an irregular or intermittent basis, such as salt and oil reservoir mines, along with mines containing combustible dust: 

30 CFR 57.22212 – Air flow (I–C, II–A, and V–A mines).—Air flow across each working face shall be sufficient to carry away any accumulation of methane, smoke, fumes, 

and dust. 

This is a “performance-oriented” standard that outlines the desired result: the airflow should 

“carry away any accumulation of methane.” The standard does not specify exactly what airflow 

quantities are needed to accomplish the desired result. 

The key to methane control is to dilute and render 

harmless the methane gas in the mine. Good 

ventilation is required to accomplish this. 

4. Eliminating ignition sources. Electric and diesel-powered mining equipment can provide 

the spark to ignite a gas explosion, and if this equipment is to be used in gassy atmospheres, 

it must be approved by MSHA. MSHA-approved diesel equipment is designed so that no external surface gets hot enough to ignite methane. 

MSHA-approved electrical equipment can take two forms. “Intrinsically safe” equipment 

implies that no electrical spark will have enough energy to ignite a gas mixture. An example 

would be portable methane detectors. The electrical circuit in these detectors is designed not to 

provide a spark strong enough to ignite gas. The other form of approved equipment surrounds 

electrical circuits with an explosion-proof box. If gas enters the box and if it is ignited by sparking inside the box, the resulting explosion is contained within the box and cannot propagate into 

the external atmosphere. 

One of the most common forms of removing ignition sources has nothing to do with equipment. 

In Category II–A (mainly domal salt) mines in the United States that are considered to be 

outburst-prone, all blasting is done with the mine evacuated of all personnel. Only when the 

mine is determined to be clear of gas, using remote methane monitoring systems, are workers 

allowed to enter the mine to conduct a preshift examination. 
 

 

159 

The most obvious ignition source has nothing to do 

with equipment. The thought of risking your life for a 

cigarette is dreadful to contemplate. Yet one of the 

recent U.S. mining disasters was almost certainly 

caused by a miner smoking in a coal mine. It takes 

continued vigilance by all miners to make sure that 

accidents like that never happen again. 

Matches and other smoking materials must not be 

carried into mines where methane gas may be present. 

5. Avoiding outburst hazards. To avoid outburst hazards in domal salt mines, continuous 

mining machines used in these mines should be operated only in areas that are known to be relatively gas-free. This is because methane gas trapped within the salt mass can be at high pressure. 

 This pressure represents a source of mechanical energy that could be suddenly released as an 

“outburst.” An outburst is a sudden, violent release of solids and high-pressure occluded gases.14 

14More information on outbursts in domal salt mines is available from Iannacchione et al. [1984], Schatzel and 

Hyman [1984], Molinda [1988], and Grau et al. [1988]. 

The key word here is “occluded.” Typically, the gas is trapped in tiny pockets within the crystal 

structure of salt or voids of an impermeable rock mass. During a change in stress conditions, 

these pockets can link up, resulting in a significant volume of gas—at full lithostatic pressure— 

immediately beneath the surface. If the pressure is sufficient, the thin layer of containment is 

burst, releasing the gas amid a shower of broken rock. The secondary shock wave caused by this 

primary burst of gas sometimes starts a chain reaction, with several million cubic feet of gas and 

several thousand tons of rock being ejected, resulting in voids a hundred feet or more in height 

[Plimpton et al. 1979].15

15Small quantities of methane gas may also be liberated while drilling or undercutting in domal salt mines, thus the 

need for permissible equipment in Category II–A mines. 

 This is the reason for the precautions taken at the Boulby potash mine 

in the United Kingdom, where the excavation is required to stay well beneath potentially gasbearing shale formations [Lumsden and Talbot 1983]. 

Outbursts can be triggered by the stress redistribution that follows blasting or excavation by continuous mining machines [Lumsden and Talbot 1983]. Blasting is performed only with workers 

on the surface; however, continuous mining machines require an operator. This is why continuous mining machines used in domal salt mines should be operated only in areas that are known 

to be relatively gas-free. 

WHAT EXPERIENCED MINE OPERATORS HAVE TO SAY 

ABOUT METHANE CONTROL 

The perspective of experienced mine operators adds much to a complete understanding of what 

must be done to address methane problems in metal/nonmetal mines. In our discussions with 

operators of gassy mines, the one concern expressed by all was the need to be vigilant. Safety 

precautions can always be defeated by careless or foolish actions. Below are summaries of interviews with five such operators across the United States. 
160 

Dave Graham is the safety and health manager of General Chemical’s trona mine in Green 

River, WY. This mine liberates large quantities of methane from the oil shales above and below 

the trona beds. Dave says that everyone knows what to do in “normal” mining operations, where 

continuous-reading methanometers keep track of gas levels. However, his concern is whether 

miners will recognize unusual and infrequent situations. Once they are aware of a hazard, they 

know what to do, but it may not be obvious that a hazard exists. Dave comments that “Miners 

can’t let their guard down. They have to be constantly asking themselves, ‘Will this 

situation create a hazardous buildup of gas?’ ” 

Charlie Young is the plant manager of the Weeks Island Mine in New Iberia, LA, a large domal 

salt mine prone to gas outbursts. They blast with the mine evacuated and must test for gas 

remotely from the surface before sending miners back underground. Charlie concentrates on 

three approaches to methane control: 

•	 Ventilation. The primary ventilation system must be capable of flushing out large quantities of gas if a methane outburst occurs after a remote blast, without the use of auxiliary 

ventilation, because the power to the mine is automatically deenergized by the mine-wide 

methane monitoring system. 

•	 Remote gas monitoring. Remote gas monitoring depends on sensors placed close to the 

face line to detect gas concentrations well below the explosive limit (the sensors are sensitive to .01% methane). 

•	 Maintenance of permissible equipment. Face and bench undercutting and drilling equipment, along with auxiliary face fans, must be approved by MSHA as permissible and 

maintained in approved condition. 

Rick Steenberg is the mine manager of FMC’s trona mine in Green River, WY. The mine is 

classified as a Category III gassy mine, so all production equipment must conform to MSHA 

standards of permissibility. 

Rick is constantly aware of the need to make sure that methane is flushed out of the mine with 

adequate quantities of fresh air. Also, a comprehensive audit system designed to check the 

equipment for permissibility compliance has become institutionalized and has proven effective. 

Jim Lekas owns and operates the ITM Mine in Vernal, UT, which produces gilsonite. Gilsonite 

is solid hydrocarbon resin, a black carbonaceous mass. Gilsonite mines include an additional 

hazard along with methane gas: the high flammability of the carbon-rich dust. 

Jim emphasizes training to be alert for possible gas problems in poorly ventilated working places 

and to make sure that there is enough air to dilute any accumulations before hand-operated metal 

tools that could create sparks are used. 

Ventilation is accomplished by introducing fresh outside air into the upper levels of active mining areas through surface-mounted, forced-air ventilating fans, with the pneumatic conveying 

system exhausting air from the lower levels back to the surface. The fans that provide the intake 
 

161 

air and the motive force for the pneumatic conveyors are all outside of the mine. No electric 

equipment is used in the mine, permissible or otherwise. 

Dick Heinen is the manager of mines for the Intrepid potash mines in Carlsbad, NM. The potash mines have gas within the evaporite strata and are classified as Category IV gassy mines. 

The noncombustible ore is extracted and liberates a concentration of methane that is not explosive or capable of forming explosive mixtures with air based on the history of the potash mines. 

However, gas accumulating in possible bed separations in the roof of the mine can provide additional pressure to cause slabbing in mine intersections. 

Dick insists that pressure relief holes be drilled 20–27 ft deep in the roof of every intersection. 

These serve both as detectors for bed separation and also as gas relief vents. The mine checks 

for methane gas every shift in every panel. The measured gas levels are almost always below the 

detection levels of the instruments. Any gas coming from relief holes is effectively diluted by 

the mine’s ventilation system. 

LOOKING FOR METHANE WHEN OPENING A NEW OR 

EXPANDING AN EXISTING METAL/NONMETAL MINE 

From both safety and economic perspectives, when opening a new or expanding16

16This includes shaft excavation. For more information on controlling methane during shaft excavation, see 

Chapter 9. 

an existing 

metal/nonmetal mine, it is critical to know whether methane will be present. For example, the 

presence of methane will impact ventilation design and permissible equipment purchases, two of 

the many items that will affect the safety and profitability of the mine. Answers can be provided 

by knowledge of the local geology and by gas testing. 

Knowledge of the local geology. The most obvious source of information about the potential 

for gas in any new mine or new section of an existing mine is the geology and history of the area 

and any nearby mines. If mines that are geologically similar to the new mine have gas problems, 

then the new mine will almost certainly share in those problems. 

If no mines exist in the area to allow for comparison, then other geological information should be 

studied. The key geological factors to look for are the presence of coal seams, carbonaceous 

shales, and other strata containing oil or gas production wells. All of these raise the risk of having gas in the mine. Methane originates from the decay of the carbonaceous materials inherent 

in coal seams, oil shales, and other carbon-bearing rocks. Methane is also embedded within the 

deep mantle rocks of the earth. It can be dissolved under pressure within water and other fluids 

and carried with them until the liquid emerges into an underground void. As the void is reached, 

the reduction in pressure releases the gas into the atmosphere. Methane can also remain in its 

gaseous form and migrate, independently of any carrier fluids, over great distances. 
 

162 

Figure 13–2 illustrates most 

of the mechanisms for 

methane transport in South 

African mines. In Figure 

13–2, emission source No. 1 

results from the simple 

decomposition of the 

carbonaceous material in the 

gold ore body (the reef),17 

17A “reef” is a lode or vein, a term commonly used in South Africa to describe the quartzite host rock for the flatlying, gold-bearing formations. 

 Figure 13–2.—Examples of methane emission as reported in South 

African metal/nonmetal mines [Cook 1998]. (Courtesy of SIMRAC.) 

KEY: 

1 - Associated with carbonaceous material in reefs. 

2 - In inclusions in alkaline dykes (although unlikely in dykes 

generally). 

3 - Thermogenic methane where dykes have heated carbon in reefs. 

4 - Coal seams in overlying Karoo sediments, with transportation to 

Witwatersrand strata in solution in water via fault planes. 

5 - General seepage of mantle methane via joints and bedding. 

6 - Strong seepage of methane along major faults, which are often 

 

along dyke contacts. 

7 - Collection of methane in highly jointed areas, e.g., adjacent to 

dykes. 

8 - Seepage of methane from reef carbonaceous material into fault 

systems. 

whereas No. 3 requires heat 

to release the methane gas 

from the carbon in the ore 

body. Nos. 2, 6, and 7 are 

variations on the same theme, 

without regard to the original 

source of the gas. Nos. 4 and 

8 represent a common origin 

for methane gas in metal/ 

nonmetal mines. The 

methane comes from coal or 

other carbonaceous seams 

and is carried into the mine 

via joints and faults. In the 

case of No. 4, the gas is carried in solution in the hot, 

pressurized ground water, but 

No. 8 shows the gas entering 

the mine via direct connections, such as geologic discontinuities or exploration 

drillholes. 

Methane can be carried in joints, faults, dykes, 

sills, and other geologic discontinuities. The 

more prevalent the discontinuities, the more 

permeable the rock and the greater the potential 

for gas storage and transportation. 
 

163 

Exploration drillholes can be a major conduit for gas migration. Typically, these will be drilled 

during the initial phases of opening up an ore body or new sections of the same ore body.18

 

18In the Republic of South Africa, there has been considerable study of the occurrence of methane gas associated 

with gold mining in the Witwatersrand. The Chamber of Mines published a comprehensive text on mitigating gas 

problems entitled Flammable Gas in Metal Mines: A Guide to Managers to Assist in Combating Flammable Gas in 

Metal Mines [Association of Mine Managers 1989]. This guide contains specific sections on methane occurrence 

and detection, the prevention of flammable gas accumulations, ventilation systems, mining methods, equipment 

modifications, “hot work” permits, and the responsibilities of mine officials with regard to methane control. 

In at 

least one case in the United States, decades-old deep wells that had not been plugged properly 

may have provided a link between coal measures and permeable strata, which in turn funneled 

the gas into a major fault system. Gas subsequently entered the mine via water carried along a 

secondary fault that had branched off from the major fault zone. 

Gas testing. Another primary task when opening up a new mine or expanding an existing mine 

is testing for methane gas at all stages in the exploration and development of the mine. Gas tests 

must be conducted at the collar of exploration drillholes19

19For more on sampling from boreholes, see Chapter 2. 

and at the roof of newly exposed faces. 

If no gas is found during those two stages, it is less likely to be present in the production phase of 

the mine’s life. However, gas at just one drillhole or at one newly exposed face indicates a 

potentially larger gas problem. Immediate measures must be taken to confirm the presence of 

methane by laboratory analysis and to carefully sample all of the other drillholes and newly 

exposed faces that are part of the project.20 

20In addition to methane, laboratory analysis should test for other gases that may be flammable or toxic, such as 

ethane or hydrogen sulfide. 

Methane can sometimes be associated with water in underground mines. Any gas bubbling from 

water coming from a fault zone or from pools of water collecting in the floor of the mine or tunnel should be sampled and analyzed for methane and other gases, such as carbon dioxide and 

hydrogen sulfide. 

For mines where the presence of methane is not definitely established, Thimons et al. [1979] 

established a simple guideline that would enable mine personnel to evaluate the methane hazard. 

In their research, they measured trace methane concentrations in 53 metal/nonmetal mines, finding that mines with a return concentration exceeding 70 ppm of methane were inevitably classified as gassy.21 

21In 1979, the MSHA classification system for metal/nonmetal mines with methane was different from the current 

standard. However, the triggers that lead to extra precautions (such as measurement of 0.25% or an ignition in the 

mine) are similar. 

Although a measurement of concentration alone is not the complete methane 

story, a return concentration exceeding 70 ppm should serve as an alert to the presence of gas 

that has not yet shown itself in other ways. 
164 

REFERENCES 

Andrews JN [1987]. Noble gases in groundwaters from crystalline rocks. In: Saline water and 

gases in crystalline rocks. St. John’s, Newfoundland, Canada: Geological Association of 

Canada, Special Paper 33, pp. 234–244. 

Association of Mine Managers [1989]. Flammable gas in metal mines: a guide to managers to 

assist in combating flammable gas in metal mines. Johannesburg, Republic of South Africa: 

Chamber of Mines of South Africa, Safety and Technical Services. 

CFR. Code of federal regulations. Washington DC: U.S. Government Printing Office, Office of 

the Federal Register. 

Cook AP [1998]. The occurrence, emission, and ignition of combustible strata gases in Witwatersrand gold mines and Bushveld platinum mines, and means of ameliorating related ignition 

and explosion hazards. Final project report, project GAP 504. Braamfontein, Republic of South 

Africa, Safety in Mines Research Advisory Committee (SIMRAC) (www.simrac.co.za). 

Edwards JS, Durucan S [1991]. The origins of methane. Min Sci Technol 12(2):193–204. 

Eschenburg HMW [1980]. Sources and control of methane in gold mines. J Mine Vent Soc 

S Afr, August. 

Fritz P, Frape S K, Miles M [1987]. Methane in the crystalline rocks of the Canadian shield. 

In: Saline water and gases in crystalline rocks. St. John’s, Newfoundland, Canada: Geological 

Association of Canada, Special Paper 33, pp. 211–223. 

Grau RH, Marshall TE, Finfinger GL [1988]. Evaluation of methane liberations from production 

blasts in domal salt mines. In: Proceedings of the Fourth International Mine Ventilation Congress (Brisbane, Queensland, Australia, July 3–6, 1988). Carlton, Victoria, Australia: Australasian Institute of Mining and Metallurgy. 

Iannacchione AT, Grau RH III, Sainato A, Kohler TM, Schatzel SJ [1984]. Assessment of 

methane hazards in an anomalous zone of a Gulf Coast salt dome. Pittsburgh, PA: U.S. Department of the Interior, U.S. Bureau of Mines, RI 8861. NTIS No. PB84214162. 

Jackson KAB [1957]. Methane on St. Helena Gold Mined Limited. Association of Mine Managers of South Africa. 

Lumsden AM, Talbot R [1983]. Develop data for review of metal and nonmetal mine methane 

hazard classification. Golder Associates, Inc. U.S. Bureau of Mines contract No. J010060. 

NTIS No. PB83258624. 

Molinda GM [1988]. Investigation of methane occurrence and outbursts in the Cote Blanche 

domal salt mine, Louisiana. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, 

RI 9186. NTIS No. PB90270083. 
165 

OML [1996]. Health and safety guidelines, draft #4. Sudbury Ontario, Canada: Ontario Ministry of Labour. 

Ontario Occupational Health and Safety Act [1990]. Revised regulations of Ontario (R.R.O.) 

1990, regulation 854, mines and mining plants. Part II, fire protection. Section 35. [http:// 

www.e-laws.gov.on.ca/DBLaws/Regs/English/900854_e.htm#BK3]. Date accessed: February 

2006. 

Plimpton HG, Foster RK, Risbeck JS, Rutherford RP, King RF, Buffington GL, et al. [1979]. 

Final report of mine explosion disaster, Belle Isle mine, Cargill, Inc., Franklin, St. Mary Parish, 

Louisiana, June 8, 1979. Arlington, VA: U.S. Department of Labor, Mine Safety and Health 

Administration. 

Schatzel SJ, Hyman DM [1984]. Methane content of Gulf Coast domal rock salt. Pittsburgh, 

PA: U.S. Department of the Interior, Bureau of Mines, RI 8889. NTIS No. PB85109593. 

Searra JA [1990]. Gas-related problems in the Witwatersrand supergroup [Thesis]. Johannesburg, Republic of South Africa: University of the Witwatersrand. 

Sherwood B, Fritz P, Frape SK, Macko SA, Weise SM, Welhan JA [1988]. Methane occurrences in the Canadian shield. Chem Geol 71(1–3):223–236. 

Thimons ED, Vinson RP, Kissell FN [1979]. Forecasting methane hazards in metal and nonmetal mines. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8392. NTIS 

No. PB80100696. 

Westfield J, Knill LD, Moschetti AC [1963]. Final report of major mine explosion disaster, 

Cane Creek mine, Potash Division, Texas Gulf Sulphur Company, Grand County, Utah, 

August 27, 1963. Washington, DC: U.S. Department of the Interior, Bureau of Mines. 
 

166 

APPENDIX A.—ONTARIO OCCUPATIONAL HEALTH AND SAFETY ACT 

R.R.O.22 1990, REGULATION 854 

MINES AND MINING PLANTS, SECTION 35 

22Revised Regulations of Ontario (Canada). 

35. (1) If a flow of flammable gas is encountered in a mine or in an enclosed building housing a 

diamond drill on the surface and the concentration of the flammable gas is unknown23, 

23Emphasis ours. 

(a) all sources of ignition in the affected area shall be eliminated; 

(b) all electrical equipment in the affected area shall be de-energized; 

(c) the affected area shall be evacuated; 

(d) precautions shall be taken to prevent persons from entering the affected area inadvertently; 

(e) a supervisor shall be notified; 

(f) the affected area shall be tested by a competent person; and 

(g) the affected area shall be designated as a fire hazard area. O. Reg.24 236/99, s. 3. 

24“O. Reg.” stands for Ontario Regulation. 

(2) Subject to subsections (3), (4) and (5), work may resume if the concentration of flammable 

gas is below 1.0 per cent. O. Reg. 236/99, s. 3. 

(3) If the concentration is less than 0.25 per cent and the affected area is tested periodically to 

ensure that the level of concentration is known, no precautions are required. O. Reg. 236/99, 

s. 3. 

(4) If the concentration is 0.25 per cent or greater but not more than 0.5 per cent, all of the 

following precautions shall be taken: 

1. The supervisor shall provide written instructions of any special precautions. 

2. The instructions, if any, shall be communicated to the workers. 

3. The affected area shall be designated as a fire hazard area. 

4. The affected area shall be tested at least once per shift before work begins and, again, 

on release of any further flow of gas. 

5. A flammable gas detector shall remain in the affected area for the purpose of continued 

testing. O. Reg. 236/99, s. 3. 
167 

(5) If the concentration is 0.5 per cent or greater but not more than 1.0 per cent, all of the precautions set out in subsection (4) shall be taken and the electrical equipment, diesel engines, 

tools and other material used in the workplace shall be designed to function safely in a flammable gas atmosphere. O. Reg. 236/99, s. 3. 

(6) If concentrations of flammable gas exceed 1.0 per cent in an area, all of the following precautions shall be taken: 

1. All sources of ignition in the affected area shall be eliminated. 

2. All electrical equipment in the affected area shall be de-energized. 

3. All persons, other than competent persons necessary to measure the concentration of 

flammable gas and to make ventilation changes, shall be removed from the affected area. 

O. Reg. 236/99, s. 3. 

(7) In mines where flammable gas is known to occur, workers who are underground or diamond 

drillers who are on the surface shall be advised of, 

(a) the probability of encountering a flow of the gas; and 

(b) the measures and procedures prescribed in this section. O. Reg. 236/99, s. 3. 

(8) For the purposes of this section, the concentration of flammable gas means the percentage, 

by volume, of flammable gas in the general atmosphere. O. Reg. 236/99, s. 3. 
 

169 

CHAPTER 14.—PREVENTING METHANE GAS EXPLOSIONS 

DURING TUNNEL CONSTRUCTION 

By Fred N. Kissell, Ph.D.1 

1

Research physical scientist, Pittsburgh Research Laboratory, National Institute for Occupational Safety and Health, 

Pittsburgh, PA (retired). 

In This Chapter 

9 Early indicators of a gas problem 

9 How the methane hazard is reduced 

9 Ventilation principles for gassy tunnels 

9 Monitoring for gas 

9 Eliminating ignition sources

 and 

9 The all-important human factors component 

This chapter gives guidelines for preventing methane gas explosions during tunnel construction. 

Emphasis is placed on assessing the hazard potential, on ventilation principles, and on monitoring for gas. 

The chapter also emphasizes the importance of human factors in reducing explosion risk. Ensuring safe conditions is much more than just good engineering design. It also involves the everyday vigilance of those working underground. This does not imply that the engineering design 

can be ignored, only that the job of providing safe conditions has just begun with design. 

EARLY INDICATORS OF A GAS PROBLEM 

For the engineer planning a tunnel project, reliable early indicators of methane are scarce. However, the local geology can often provide some information.2

2

A good source of information on hazardous ground gases is Doyle [2001]. 

 Carbonaceous rocks and tar sands 

are a likely methane source. Gas is also a distinct possibility if it is known to be present elsewhere in the same sequence of geologic formations. Swampy areas, sewerage systems, and 

landfills are also candidates because the decomposition of organic materials produces methane. 

The gas in a tunnel can originate in the strata being excavated, or it can migrate a considerable 

distance from adjacent strata. 

Test borings at the project site can also serve as initial indicators of gas. Methane has no odor, 

but may be emitted along with gases that do. If gas is emitted from the borehole, a sample may 

be collected by inserting a tube into the hole as far as possible and pumping the gas out. The 

gas sample should be collected in a sampling bag or canister for later analysis by a chemical 

laboratory.3 

3

Most handheld methane detectors require the presence of 10% oxygen in the sample to operate properly, and this 

much oxygen is not normally found in borehole samples. For more information on methane detection, see the sampling chapter (Chapter 2). 
170 

Gas flowing from just one test boring indicates a potentially larger gas problem. Immediate 

measures must be taken to confirm the presence of methane by laboratory analysis and to sample 

all of the other boreholes that are part of the project. In addition to testing for methane, laboratory analysis should also test for other gases that may be flammable or toxic, such as ethane or 

hydrogen sulfide. 

INDICATORS OF GAS UNDERGROUND 

In tunneling and hard-rock mining, methane is not 

normally encountered. Therefore, the mistaken 

inclination is to not suspect the presence of gas. 

There have been many methane explosions in places where the existence of gas was never 

suspected or was thought to be minimal. A basic problem is that the commonly used catalytic 

detectors are not very good at detecting very low concentrations of gas. Figure 14–1 illustrates a 

representative situation (the ventilation quantities in the figures are only provided as examples). 

In this figure, the main tunnel fan moves air at a rate of 10,000 cfm and the scavenger fan moves 

air at 5,000 cfm. Methane gas enters the tunnel at the face at a rate of 1 cfm. First, the gas 

concentration is measured in the main fan line. With 1 cfm of methane in 10,000 cfm of air, the 

concentration will be 1/10,000, or 0.01%. Next, the methane concentration in air returning from 

the scavenger fan duct is measured; here, the concentration is 1/5,000 or 0.02%. With most commonly used catalytic detectors, these low percentages will show up as zero. However, even if 

methane were detected, such low concentrations would usually be considered negligible. 

Figure 14–1.—Low concentrations of methane gas. 

Is this level of gas hazardous? Obviously not under the conditions in which the measurements 

were made. However, consider the following scenario. The 5,000-cfm scavenger ventilation 

goes off for 10 min because of an electrical problem. Ten cubic feet of methane then accumulates in the face area. This quickly dilutes to 100 ft3

 of a 10% methane explosive mixture, and 

thus an explosion occurs in a tunnel where no one had initially measured any gas. 

This might sound farfetched, but many workers 

have died under very similar circumstances. Gas 

checks had been made 

when the ventilation system was working well. 

Later, the ventilation 

failed for some reason, 

and lethal quantities of 

gas accumulated, causing 

an explosion.
 

 

 

171 

Ways to confirm the presence of gas underground. Given that low emissions of methane can 

be hazardous, how does one determine if there is a potential methane problem, assuming that 

there were no clues from exploration boreholes? There are three possible ways. 

1. Look for gas in the parts-per-million range. When testing for methane gas, return air samples 

should be collected in a bag or bottle specially designed for gas sampling. A laboratory analysis 

that uses a chromatograph to look for gas in the parts-per-million range is then conducted. 

Applying this method to the scenario in Figure 14–1, the return air sample would have shown 

100 ppm, a definite indicator of gas in low quantities.4

 

4

Low levels of methane gas have been found in a wide variety of hard-rock and noncoal mines. For mines classified 

as gassy by the Mine Safety and Health Administration, Thimons et al. [1979] found that the return air methane 

concentrations were 70 ppm or higher. No similar research on methane has been conducted in tunnels, but there is 

little doubt that a comparable value in a return airflow of 10,000 cfm would show hazardous quantities of gas. 

When sampling, the ambient air on the 

surface should also be measured, as it generally contains a few parts per million of methane. 

2. Hunt for gas when ventilation is temporarily off. It is common for tunnel ventilation systems 

to be down for short periods while fan changes are being made or ductwork extended. Because 

even low gas emissions accumulate to measurable levels quickly, this is an opportune time to 

hunt for gas accumulations with a handheld methane detector. If gas has already been shown to 

exist, this hunt is an important safety measure. 

3. Look for gas in those places where it is most likely to accumulate. Gas emitted at the face 

will accumulate in unventilated corners near the face. Emissions from small cracks or fissures 

near the crown may produce a methane layer there because methane is much lighter than air. 

In operations using a tunnel boring machine (TBM), gas accumulations near the muck discharge 

point are likely. If the tunnel is unlined, any location along the entire length is a potential site for 

a methane layer at the crown. 

As with surface samples, the initial presence of gas underground must be confirmed to a higher 

level of accuracy by laboratory analysis. If a field instrument shows that gas is present, an air 

sample must be collected in a bag or bottle specifically designed for gas sampling. The analysis 

is normally conducted with a carefully calibrated gas chromatograph. To be effective, the hunt 

for gas in tunnels must be conducted at frequent intervals. This is the only way to detect gas in 

isolated pockets. If the presence of gas is suspected but not yet confirmed, the tunnel air should 

be tested for methane with a handheld instrument at least twice per shift. 

PROVIDING ADEQUATE VENTILATION 

Ample dilution to safe levels. Enough ventilation air must be provided to immediately dilute 

the methane gas to safe levels as soon as the gas enters the tunnel. Methane is combustible when 

mixed with air in the range between 5 and 15 vol % of gas. The 5 vol % value is the lower 

explosive limit (LEL). Methane concentrations in air that are below the LEL are not explosive. 

The 15 vol % value is the upper explosive limit. Gas mixtures with concentrations above this 

limit are not explosive, but may become so if mixed with more air. 
 

172 

Simultaneous application of three basic elements 

reduces the methane hazard: 

•	 Adequate ventilation. 

•	 Regular monitoring of air quantities and gas 

concentrations, with automatic equipment shutoff 

at high gas concentrations. 

•	 Elimination of ignition sources, including those 

that are worker-related. 

The simultaneous application of several elements is 

necessary because if one fails, the others continue to 

ensure safety. 

When methane is emitted from the strata, it is usually at high concentration. As it progressively 

mixes with air, the concentration will pass through the explosive range and down below the LEL. 

A good ventilation system will supply enough fresh air to reduce all of the gas to far below the 

LEL as soon as the gas is emitted from the strata. 

In referring to gas concentrations, different government agencies may use different terminology. 

For example, in regulating coal mines, the Mine Safety and Health Administration (MSHA) 

specifies that the concentrations of methane at coal mine working faces remain below 1.0 vol %. 

This is the same as 20% of the LEL. With the LEL of methane in air at 5 vol %, 20% of 5 vol % 

is 1.0 vol %. Specifying a percentage of the LEL is advantageous when mixtures of flammable 

gases are emitted. 

Main ventilation systems. Main ventilation systems carry air from the portal into the TBM 

trailing gear. These are classified as either blowing or exhausting. In blowing systems, fans 

located on the surface and along the ductwork push air through the ductwork into the tunnel. 

In exhaust systems, air in the ductwork flows out of the tunnel. Each system has its advantages. 

Selection of either an exhaust or a blowing main ventilation system will depend on whether a 

face shield and scrubber are used, on whether or not a scavenger system is used, and on the type 

of ductwork used.5 

5

When planning a tunnel ventilation system, make a simple diagram of all ductwork and airflow movement to 

ensure that the ventilation mistakes described in this section are not incorporated into your plans. 

Face ventilation systems. Face ventilation systems carry air from the trailing gear to the face of 

the tunnel where rock is broken and removed. In most instances, the primary source of gas is at 

the face, so it is vital to provide adequate ventilation air all the way to the face, that is, to the 

last foot.6 

6

In some instances it is also necessary to focus attention on the muck discharge point. For example, in earth pressure balance machines most of the methane may be released at the end of the screw conveyor. A nearby fan or compressed air venturi can be used to dilute this gas, but does not relieve the need to provide adequate ventilation all the 

way to the last foot. 

For this reason, the ventilation focus of this chapter is on face ventilation. 
 

 Figure 14–2.—Exhausting system of face ventilation. For clarity, 

the equipment is not shown. 

The tunnel face is usually ventilated with much less air than 

you think. If 20,000 cfm goes down the shaft but only 2,000 cfm 

reaches the face, then as far as methane control is concerned 

the tunnel is being ventilated with only 2,000 cfm. 

There are two categories of face 

ventilation: exhausting (Figure 

14–2) and blowing (Figure 14–3). 

The exhausting system is the less 

efficient in clearing out gas from 

the face. For example, the face 

7

7

FVE is an indicator of the proportion of air reaching the last foot, i.e., a distance of 1 ft from the face. 

 Figure 14–3.—Blowing system of face ventilation. If the blowing 

duct diameter is 2 ft, the maximum distance from the face would be 

20–30 ft. 

173 

ventilation effectiveness (FVE) 

of a 10,000-cfm, 24-in-diam 

exhaust duct located 10 ft from a 

mine face is only about 0.10. In 

other words, the concentration of 

methane measured near the face 

is 10 times higher than the 

concentration in the air passing 

through the duct [Wallhagen 

1977]. If the end of the exhaust 

duct is more than 10 ft from the 

face, the FVE is even less. 

Therefore, the end of an exhaust 

duct must always be 10 ft or less 

from the tunnel face unless other 

means are used to ventilate the 

face, such as venturi air movers 

powered by compressed air. 

Blowing face ventilation (Figure 

14–3) is better for clearing out 

gas than exhaust ventilation 

because the momentum of the air 

in a blowing jet carries it farther. 

However, blowing systems also 

lose effectiveness as the face-toduct distance increases. The duct 

must be kept as close to the face 

as possible, with the end of the 
174 

duct not more than 10–15 duct diameters from the face. There also must be no obstructions that 

would prevent the emerging jet of air from reaching the face. 

Studies of blowing ventilation at coal mine faces show that the FVE for a 10,000-cfm, 24-indiam blowing duct at 20 ft (10 duct diameters) is about 0.40, indicating that the concentration at 

the face is 2.5 times that in the return [Wallhagen 1977]. Thus, although 10,000 ft3 of air 

emerges from the duct per minute, only 4,000 ft3 of air actually reaches the face. If the face 

emits 20 ft3

 of gas per minute, the average concentration in the immediate face area will be 

20/4,000 or 0.5%, rather than 20/10,000 or 0.2%. 

Whether exhausting or blowing ventilation is used, the end of the duct should be kept as close to 

the face as possible. If the face is drilled and blasted, keeping the ductwork in place is particularly difficult. Blast shields can help, but may hinder clearing the face. It is sometimes possible 

to move flexible ductwork forward and back on a trolley wire. Another possibility is inflatable 

cloth ductwork, which is inexpensive and may be considered expendable. Whatever method is 

used, when methane is present the need to keep the ventilation ductwork within the required face 

distance cannot be ignored, regardless of cost or inconvenience. 

Which face ventilation system is best? In principle, blowing ventilation systems provide better 

dilution of methane at the face, but it does not always follow that it is better to use blowing face 

ventilation in a tunnel. For example, Figure 14–4 illustrates the face ventilation of a smalldiameter TBM with an enclosed cutter head. In this example, 5,000 cfm is withdrawn from the 

cutter head enclosure through duct #1. An airflow above 5,000 cfm would be better, but there is 

not space for larger ductwork at the front of the TBM. So, an additional 5,000 cfm is provided 

with a second ventilation duct (duct #2) that extends to the front of the trailing gear and blows air 

toward the face. 

 Figure 14–4.—TBM ventilation system with low airflow zone where 

methane dilution suffers. 

The problem with this system is that it has equal duct airflows moving in opposite directions, 

leading to a stagnant zone of low airflow (see Figure 14–4) where methane may accumulate. 

Also, it only delivers 5,000 cfm to the face, even with the two ducts. To avoid zones of low 

airflow, ventilation designs that move air through ductwork in opposite directions should be 

avoided. To demonstrate, if ventilation duct #2 exhausted air from the face instead of blowing 

toward it (as shown in Figure 14–5), then the front of the tunnel would be ventilated with 

10,000 cfm of air instead of 

5,000 cfm, and there would 

be no zone of low air 

movement. 

In this example, venturi air 

movers powered by compressed air are used for additional air movement in the 

space between the end of 

duct #2 and the cutter head 

enclosure. 
 

 

175 

Figure 14–5.—TBM ventilation system with second duct exhausting
 

air from the face. Venturi air movers provide additional air movement. 

Auxiliary face ventilation 

systems. A common way to 

ventilate the tunnel face is to 

use an auxiliary ventilation 

system (Figure 14–6). 

Auxiliary face ventilation 

systems ventilate the tunnel 

face with a fan and duct that 

are separate from the main 

ventilation system. Auxiliary 

systems are often called 

scavenger fans. 

Figure 14–6.—Auxiliary system with adequate overlap (not to scale). 

A critical feature of auxiliary 

systems is the required 

overlap with the main 

ventilation duct, since 

auxiliary systems that do 

not overlap properly suffer 

huge efficiency losses. 

Figure 14–6 shows a simple 

two-duct auxiliary system 

that is working properly. 

The main duct is on exhaust, 

with the fan on the surface; 

the scavenger, or auxiliary 

fan, is blowing toward the 

face. Note that the inlet of 

the scavenger fan is in the 

fresh air stream of the main 

ventilation duct. Figure 14–7 

shows the same arrangement, 

but with no overlap. The 

inlet of the scavenger fan 

picks up contaminated air 

returning from the face 

rather than fresh air, 

creating recirculation. 

Eddy currents between 

the two inlets provide the 

only air to the face, greatly 

reducing the amount of 

fresh air available to 

dilute methane. To prevent this scenario, the two ducts must overlap by at least twice the 

tunnel diameter (Figure 14–6). 

Figure 14–7.—Auxiliary system with no overlap and poor methane 

dilution. 
176 

Unfortunately, overlap is not something that can be engineered into the system from the start. 

New sections of fan line must be promptly added as the tunnel advances. Adequate overlap is 

maintained only through continued around-the-clock vigilance of the tunnel crew. For this reason, it is a major problem area. 

With auxiliary systems, ventilation efficiency also suffers when airflow directions are not 

coordinated. Figure 14–8 depicts a scenario in which the airflow directions are not coordinated. 

Figure 14–8 is similar to Figure 14–6 except that the main duct is now blowing. The scavenger 

fan inlet is now in the contaminated return air, and contaminants are recirculated back to the 

face. The impact is that the fresh air reaching the face is reduced by up to one-half. Whatever 

the arrangement of ducts and fans, workers must check carefully to be sure that fresh air is not 

being replaced by recirculated air (see footnote 5). 

Figure 14–8.—Loss of efficiency from mismatched airflow directions. 

A scavenger fan with inadequate overlap can recirculate 90% of the air 

returning from the face. In such an instance, a 3,000-cfm scavenger fan 

will deliver only 10% or 300 cfm of fresh air to the face. 

Minimizing leakage. Leakage in both main and face ventilation systems is another source of 

airflow losses. Factors that impact leakage are fan placement, ductwork diameter and length, 

pressure drop, and duct condition. It is not unusual to lose half of the airflow in a long run of 

ductwork. In planning, the largest practical diameter of duct should be used. Damaged ductwork should never be installed, particularly if the ends are buckled. 

Another major source of leakage (and recirculation) is the “trombone” section in the trailing 

gear, where concentric ventilation ducts slide apart and new sections of duct are added as the 

trailing gear moves forward. Recirculation of contaminated air can be particularly high as new 

sections of ductwork are added. This leakage and recirculation may be minimized by locating 

fans on both sides of the trombone section and balancing the fan flows to minimize the pressure 

drop between the air passing through the trombone 

and the outside tunnel air. 

Cumulative ventilation 

inefficiencies. Cumulative 

ventilation inefficiencies 

include leakage in the main 

duct, leakage at the trombone connection, auxiliary 

system problems, and low 

face ventilation effectiveness because the end of the 

ductwork is too far from the 
 

177 

face. While one of these elements alone may not be significant, the cumulative effect of several 

will certainly be. 

Venturi air movers. Compressed air is ineffective as a primary fresh air source because it 

cannot deliver enough air for adequate dilution of gas. However, there are some circumstances 

where compressed air can serve as an adjunct to conventional ventilation, particularly to enhance 

the air velocity over short distances. For example, if an exhaust ventilation system is being used, 

a venturi-type air mover powered by compressed air can provide better dilution of methane at the 

face, provided that the air mover is located in fresh air.8 

8

Venturi air movers must be grounded to prevent the buildup of static electricity. 

A good way to align venturi air movers is shown in Figure 14–5. Here the air movers are placed 

on the opposite side from the exhaust duct so as to generate a U-shaped airflow pattern that feeds 

contaminated air to the duct inlet. Note in Figure 14–5 that venturi #1 is outby the inlet of 

duct #2 and is in the 10,000-cfm fresh air stream produced by both ducts #1 and #2. Also, 

venturi #2 is placed directly forward of venturi #1. 

Venturi systems will recirculate a high proportion of the airflow, and the amount of recirculation 

will grow as the distance and the number of venturis grow. As a result, a venturi system is not 

effective for distances over 25 ft, as indicated in Figure 14–5. 

Checking the ventilation system. To adequately check the ventilation system, a regular program of airflow measurements must be used, with airflows measured at least weekly. Airflow in 

all ducts must be measured, along with the airflow in the center line of the tunnel. Even if there 

are no leaks, it is common for ductwork to be clogged with muck. 

Tunnel workers should always be on the lookout for ventilation 

danger signals. Does ventilation duct always extend throughout 

the tunnel and close to the face? Is there always an adequate 

overlap? Is the ductwork sealed against leaks? Are the fans 

always running? Unless the tunnel has a diameter of 20 ft or more, 

is there obvious air movement everywhere in the tunnel? Is the air 

unusually warm or dusty? 

MONITORING FOR METHANE 

If methane is found, either at boreholes or during tunnel construction, regular monitoring must be 

scheduled. The most likely place to find methane is in the face area of the tunnel.9

9

There may be exceptions. See footnote 6. 

 Gases emitted at the face will collect there in unventilated corners. Emission from feeders or faults near the 

crown may produce a methane layer there, particularly in unlined tunnels. On faces that are 

drilled and blasted, workers must check for methane before blasting. If a TBM is being used, 
 

 

178 

methane can accumulate at the muck discharge and behind the face shield. Finally, workers 

should also check for gas before and during welding and cutting operations. 

In tunnels known to have methane, preshift and midshift gas checks are minimum requirements. 

The frequency of other checks depend on whether continuous detectors are also present, the 

extent of the hazard, and the applicable regulations. The Occupational Safety and Health 

Administration (OSHA) safety and health standards for underground construction [29 CFR10 

10Code of Federal Regulations. See CFR in references. 

1926.800] require continuous monitoring when rapid TBMs are used. Other flammable gas 

requirements from 29 CFR 1926.800 are as follows: 

When an air sample indicates— The necessary action is— 

5% or more of the LEL Increase ventilation, control gas.1 

10% or more of the LEL Suspend hot work such as welding or cutting. 

20% or more of the LEL 2 Cease work, cut power, withdraw employees.

 1A flammable gas concentration of 5% of the LEL or higher (0.25 vol % of methane) indicates an action 

level to take improved safety measures. OSHA requires steps to increase ventilation air or otherwise 

control the gas in such cases. However, it is wise to also implement a better monitoring program and 

training for workers. Any ventilation improvements should generally be permanent, the goal being to 

consistently operate below 5% of the LEL if at all possible. 

 2

Detector warnings and equipment shutdowns triggered by high gas levels indicate an immediate need 

for better ventilation. 

Handheld detectors are used to check for gas in any location. However, a peak emission can be 

missed because readings are taken at infrequent intervals. Fixed-site monitors operate continuously and can identify emission peaks and shut off electrical equipment when the methane level 

is excessive. Fixed-site monitors typically have two or more heads; the important ones are near 

the face and/or the muck discharge point. 

Monitor heads should not be located where they are directly bathed by a stream of fresh air; 

this can prevent gas from reaching the head. Also, regular cleaning of monitor heads is necessary. Dirt-clogged heads can fail to detect methane, so monitor heads should not be located 

where muck spatter or water sprays will make them ineffective. 

As part of a monitor check, use a “shutdown test” to ensure 

that the fixed-site monitor is hard-wired into the tunnel electrical 

system properly. Bathe each monitor head with a gas mixture 

that has more than 1% methane, and check to see that the TBM 

and its auxiliary equipment shut down as they should. 

Do the shutdown test as excavation begins, and then a few 

more times over the course of the project. 
ELIMINATING IGNITION SOURCES 

179 

Electrical equipment in tunnels may or may not be explosion-proof, depending on the level of 

the hazard. The OSHA safety and health standards for underground construction [29 CFR 

1926.800] contain the applicable requirements and definitions. OSHA has two hazard classifications, denoted “potentially gassy” and “gassy.” These are based on the results of air monitoring, on the local geology, on whether there has been a flammable gas ignition, and on whether 

there is a connection to another tunnel that is gassy. For the air monitoring, the classification 

trigger level is 10% of the LEL, and the specific classification depends on the length of time for 

which this gas level or higher is observed. Tunnels so classified must meet additional ventilation, gas monitoring, and equipment requirements. Some states have their own regulations as well. 

It was mentioned earlier that a flammable gas concentration of 5% of the LEL or higher should 

be regarded as an action level to improve safety. Taking action at the 5% level will improve the 

chances that the 10% level will not be reached. 

In the event that a large pocket of gas is encountered, some equipment may still be used. At a 

minimum, this includes fans and telephones. However, such equipment must always be 

explosion-proof. 

Tunnel contractors must bear in mind that providing explosion-proof equipment does not in itself 

eliminate the possibility of a spark source. For instance, sparks generated by cutting tools striking rock often have enough energy to ignite an explosive mixture. Welding or striking a match 

to light a cigarette can have the same effect. 

THE IMPORTANCE OF HUMAN FACTORS AND MULTIPLE PREVENTIVE ACTIONS 

The importance of human factors and multiple preventive actions in reducing methane explosion 

risk was identified in a study by Kissell and Goodman [1991]. Using a fault tree, they examined 

the possible causes of tunnel methane explosions. The intent was to provide a relative ranking of 

the events or combinations of events most likely to contribute to an explosion. 

Human factors. In the Kissell and Goodman study, 15 “initiating events” were identified 

to represent starting conditions that lead to an explosion (Table 14–1). As evidenced in Table 

14–1, most initiating events involve a human factor rather than an engineering specification. 

In other words, safe conditions require the everyday vigilance of those working underground. 

This does not undermine the importance of good engineering design, only that the job of providing safe conditions just begins with design. For example, workers must maintain overlap in 

auxiliary systems as mining advances, regularly check the ventilation quantity and methane 

concentration, and adequately service the methane monitors. Equally important, workers must 

not smoke underground; those who do risk causing an explosion if methane is present. 

Multiple preventive actions. Another conclusion from the fault-tree study was that large 

reductions (over 90%) in the risk of an explosion only result from multiple preventive actions. 

For example, a ventilation upgrade or a methane monitor upgrade by itself offers risk reductions 

under 50%. A risk reduction of 90% or more would typically require both of these, plus additional actions such as a no-smoking rule and more thorough gas checks during welding. 
180 

Table 14–1.—Initiating events for tunnel methane explosions 

(from Kissell and Goodman [1991]) 

Human factors primarily involved:

 1. Ventilation duct setback from face is too great 

2. Use of a scavenger system with inadequate overlap 

3. A fan is turned off 

4. Fan performance is seriously degraded 

5. Ductwork has serious leaks 

6. Ductwork is seriously pinched 

7. Smoking or welding occurs 

8. Methane monitor calibration is off 

9. Equipment used is not explosion-proof operationally 

10. Gas checks are not made before or during welding 

Combination of human factors and engineering specifications:

 1. Methane monitor disabled or not present 

2. No other warnings of excess gas are provided 

Engineering specifications primarily involved:

 1. Ductwork is seriously undersized 

2. Equipment not explosion-proof by design 

Neither engineering or human factors involved:

 1. Cutter pick sparking 

REFERENCES 

CFR. Code of federal regulations. Washington, DC: U.S. Government Printing Office, Office of 

the Federal Register. 

Doyle BR [2001]. Hazardous gases underground: applications to tunnel engineering. New 

York: Marcel Dekker, Inc. 

Kissell FN, Goodman GVR [1991]. Preventing tunnel methane explosions: what’s most 

important. In: Proceedings of the Fifth U.S. Mine Ventilation Symposium (Morgantown, WV, 

June 3–5, 1991). Littleton, CO: Society for Mining, Metallurgy, and Exploration, Inc., pp. 605– 

610. 

Thimons ED, Vinson RP, Kissell FN [1979]. Forecasting methane hazards in metal and 

nonmetal mines. Pittsburgh, PA: U.S. Department of the Interior, Bureau of Mines, RI 8392. 

NTIS No. PB80100696. 

Wallhagen RE [1977]. Development of optimized diffuser and spray fan systems for coal mine 

face ventilation. Waltham, MA: Foster-Miller Associates. U.S. Bureau of Mines contract No. 

H0230023. NTIS No. PB277987. 
181 

INDEX 

A 

Abnormally gassy faces...........................................46 

Air velocity, effect of.........................................12, 17 

Anomalous methane emissions................................98 

Anti-ignition back spray ..........................................59 

Auger mining – See Highwall mining ...................135 

Auxiliary air movers................................................18 

Auxiliary fans ..........................................................18 

B 

Barometric pressure, shaft sealing.........................133 

Bit wear, ignitions....................................................49 

Bleeder splits .....................................................68–69 

Bleeder systems .................................................63–76 

 additional air splits........................................68–69 

 designing ............................................................64 

 evaluating ...........................................................71 

 examining ...........................................................69 

 flow-through .......................................................65 

 maintaining ........................................................69 

 water drainage.....................................................69 

 wrap-around........................................................64 

Blowers..................................................................107 

Blowing ventilation .................................................41 

Bolter faces..............................................................48 

Boreholes, degas 

 cross-measure .....................................................86 

 horizontal............................................................83 

 short-radius .........................................................85 

 vertical ................................................................84 

Bump test, methane detectors..................................31 

C 

Calibration of catalytic detectors, different gases....33 

Catalytic heat of combustion sensors.......................13 

Clastic dikes...........................................................104 

Clay veins ..............................................................104 

Cleats .....................................................................103 

Coal dust and methane mixtures, explosibility......149 

Coal dust explosions..............................................147 

Coal seam degasification 

 economics ...........................................................93 

 thresholds............................................................80 

Coal silos.......................................................141–146 

 closed-top silos .................................................143 

 explosion in British Columbia, Canada ............144 

 load-out areas....................................................143 

 reclaiming areas................................................143 

Crosscuts, ventilation...............................................45 

Cross-measure boreholes.........................................86 

D 

Degasification of coal seams .............................77–96 

 economics of.......................................................93 

 horizontal boreholes ...........................................83 

 methane capture ratios ..................................83–90 

 vertical frac wells................................................84 

Diffuser fans............................................................46 

Dikes......................................................................104 

Dilution capacity of bleeder systems.......................75 

Dilution of methane ..................................................3 

Direct method ..................................................10, 114

District recirculation................................................16 

Domal salt mines ...................................................154 

Drainage of methane................................................83 

Dust collectors ...................................................41, 45

Dust explosions .....................................................147 

Dust scrubber recirculation......................................15 

Dust scrubbers ...................................................41, 45 

E 

Economics of coal seam degasification...................93 

Eddy zones, longwall.........................................56, 58 

Effective inert, in explosibility diagram ....................4 

Emission peaks ........................................................37 

Emission rate .............................................................9 

Excess nitrogen, in explosibility diagram..................4 

Exhaust ventilation ..................................................45 

Explosibility 

 inert gases, effect on .............................................4 

 pressure, effect on.................................................5 

 temperature, effect on ...........................................5 

Explosibility diagram ................................................5 

Explosibility of gas mixtures.....................................3 

Explosions ...............................................................78 

Explosive range .........................................................3 

Extensible curtain systems.......................................39 

F 

Fan spray system .....................................................40 

Faults .............................................................100–104 

 large-scale.........................................................101 

 normal...............................................................102 

 reverse...............................................................103 

 small-scale ........................................................101 

Filling shafts at closed mines.................................132 

Forecasting methane emissions .........................9, 119 

Frac wells.................................................................84 

Fractures ................................................................103 

Frictional ignitions.............................................48, 59 

Froude number.........................................................17 
182 

G 

Gas accumulation, longwall...............................56–58 

Gas blowers ..........................................................107 

Gas content of coal ........................................114, 116 

Gas detector.............................................................27 

Gas emission from coal in silos.............................141 

Gas emission sources, longwall...............................55 

Gas from bumps, longwall.......................................55 

Gas-in-place...........................................................116 

Gas mixtures, explosive range...................................3 

Gas outbursts .........................................................107 

Gas readings ............................................................29 

Gassiness of coal seams...........................................79 

Gilsonite mines......................................................153 

Gob curtain, longwall ..............................................57 

H 

Headgate cutout, longwall .......................................57 

Heat of combustion sensors...............................13, 27

High-volume scrubbers............................................47 

Higher air velocity, effect of..............................12, 17 

Highly gassy coal seams..........................................79 

Highwall mining ...........................................135–140 

 auger removal ...................................................140 

 inert gas quantity ..............................................138 

 oxygen concentration........................................139 

 purging..............................................................139 

 starter hole ........................................................139 

Horizontal boreholes, degasification .......................83 

Human factors .......................................................179 

Hydrofracing, degasification ...................................84 

I 

Igneous intrusions..................................................106 

Ignition ....................................................................59 

Ignition of methane by— 

 adiabatic compression...........................................7 

 hot solids...............................................................6 

 lightning................................................................8 

 sliding friction.......................................................7 

 static electricity.....................................................8 

 thermite sparking ..................................................7 

Ignition sources in metal/nonmetal mines .............158 

Indirect method......................................................114 

Inert gas .............................................................4, 135 

Inert gas generator .................................................137 

Infrared absorption, sampling..................................28 

Inlets to pillared areas..............................................74 

Instruments to measure methane .............................27 

J 

Joints......................................................................103 

L 

Large cross-sectional entries....................................18 

Layering number......................................................12 

Layering of methane................................................12 

Le Chatelier’s law......................................................5 

Lightning ...................................................................8 

Lower explosive limit............................................3, 5 

Lower flammable limit ..........................................3, 5

L-shaped wing curtain, longwalls............................59 

M 

Machine-mounted methane monitors ......................28 

Metal and nonmetal mines, methane in .........151–167 

 actions to take ...................................................156 

 Australia ...........................................................152 

 Canada ..............................................................152 

 difference from coal mines ...............................155 

 estimating the hazard ........................................161 

 Europe...............................................................152 

 ignition sources.................................................158 

 Scandinavia.......................................................151 

 South Africa, Republic of.................................152 

 starting a new mine...........................................161 

 United Kingdom ...............................................152 

Methane 

 explosions..........................................................78 

 origins................................................................77 

Methane and coal dust mixtures, explosibility ......149 

Methane capture ratios, degasification ..............83–90 

Methane content, coal seams ...................................79 

Methane detection..............................................47–48 

Methane detectors

 calibration of.......................................................33 

 dust-clogged heads .............................................33 

 machine-mounted ...............................................28 

 portable.........................................................27–28 

 sensor head response ..........................................32 

 where to take readings ..................................29–30 

Methane dilution by scrubbers...........................42, 45 

Methane dilution capacity of scrubbers...................39 

Methane drainage techniques ..................................83 

Methane emission peaks..........................................37 

Methane ignition......................................................59 

Methane layers.........................................................12 

Methane monitor, longwall......................................60 

Methane pipeline, degasification .............................90 

Methane stored in coal...............................................8 

Mildly gassy coal seams ..........................................79 

Mine emissions......................................................119 

Mixing into air, methane .....................................4, 17 

Modeling techniques, methane reservoirs .............124 

Moderately gassy coal seams...................................79 

Modified shearer-clearer..........................................56 
N 

Natural ventilation, coal silos ................................145 

Numerical simulation, methane reservoirs ............122 

O 

Oil shale mines ......................................................153 

Origins of coalbed methane.....................................77 

Outbursts................................................................107 

Oxygen deficiency...................................................66 

P 

Paleochannels ..........................................................98 

Petroleum mines ....................................................154 

Pipelines in mines....................................................90 

Placement of sensing heads, methane monitors.......32 

Portable methane detectors................................27–28 

Potash mines..........................................................154 

Prediction of gas emissions ...................................119 

Pressure, effect on explosibility.................................5 

Production decline analysis, degas boreholes........118 

R 

Raise bore drills, ventilation of..............................130 

Recirculation, hazards of .........................................14 

 district.................................................................16 

 dust scrubber.......................................................15 

 in headings..........................................................15 

Reserve ventilating capacity, bleeders.....................68 

Reservoir engineering methods .............114, 117–118 

Reservoir modeling........................................122–123 

Reservoir properties, coal seams .............................78 

Restricted spaces, sampling in.................................30 

Rock dusting requirements ....................................148 

S 

Sample draw systems in methane detectors.............13 

Sampling accessible spaces for methane .................29 

Sampling restricted spaces for methane.............30–31 

Sandstone channels..................................................98 

Scrubber clogging, continuous miners.....................44 

Scrubber exhaust direction, continuous miners...44–45
 

Scrubbers, continuous miners............................41, 45 

Sealing pillared areas...............................................69 

Seasonal trend in explosions....................................19 

Sedimentary intrusions ..........................................104 

Separate mine fans, effect on bleeders.....................73 

Setback, brattice, continuous miners .......................39 

Shafts .............................................................127–134 

 barometric pressure...........................................133 

 cutting and welding ..........................................129 

 filling, at closed mines......................................132 

 restricted spaces........................................129–130 

 sampling guidelines ..........................................128 

 ventilation guidelines........................................127 

 water rings ........................................................129 

 work decks........................................................128 

Shearer-clearer, longwalls .......................................56 

Silos – See Coal silos ............................................141 

Specific methane emission ......................................80 

Splits of air, bleeders ...............................................73 

Spray fan system, continuous miners ......................40 

Sprays, longwall ................................................56, 58 

Startup and recovery ventilation, bleeders...............74 

Static electricity .........................................................8 

T 

Temperature, effect on explosibility..........................5 

Thresholds for coal seam degasification..................80 

Trona mines...........................................................154 

Tunnels ..........................................................169–180 

 assessing the hazard..................................169–171 

 auxiliary ventilation..........................................175 

 blowing ventilation ..........................................173 

 checking the ventilation system........................177 

 cumulative ventilation inefficiencies................176 

 dilution of methane to safe levels .....................171 

 early indicators of a gas problem......................169 

 exhaust ventilation............................................173 

 face ventilation effectiveness............................173 

 face ventilation systems....................................172 

 fixed-site methane monitors .............................178 

 gas checks.........................................................177 

 handheld methane detectors..............................178 

 human factors ...................................................179 

 ignition sources.................................................179 

 inadequate overlap in auxiliary systems ...........175 

 lower explosive limit ........................................171 

 main ventilation systems...................................172 

 monitoring for methane ....................................177 

 multiple preventive actions...............................172 

 overlap in auxiliary systems .............................175 

 planning ............................................................169 

 potentially gassy ...............................................179 

 recirculation......................................................175 

 shutdown test ....................................................178 

 stagnant zones...................................................174 

 upper explosive limit ........................................171 

 ventilation efficiency ........................................176 

183 
184 

U 

Underground methane pipeline, degasification .......90 

Unforeseen gas emissions........................................98 

Upper explosive limit ............................................3, 5 

Upper flammable limit...........................................3, 5 

V 

Ventilation eddy zones ......................................56, 58 

Ventilation failure, tunnels ....................................170 

Ventilation setback ..................................................39 

Ventilation systems, tunnels ..................................172 

Venturi air movers, tunnels....................................177 

Vertical frac wells, degasification ...........................84 

W 

Walkway curtain, longwalls ....................................57 

Warning signs during sampling.........................31, 34 

Water sprays, longwalls.....................................56, 58 

Wells, oil or gas at metal/nonmetal mines.............161 
Delivering on the Nation’s Promise: 

Safety and health at work for all people 

through research and prevention 

For information about occupational safety and health topics contact NIOSH at:

1–800–35–N IOSH (1–800–356–4674)

Fax: 513–533–8573

E-mail: pubstaft@ cdc.gov

www.cdc.gov/niosh

S AFE R • H E ALT H IE R • P E OP L ET M

DHHS (NIOSH) Publication No. 2006 127 –
Handbook for Methane Control in Mining • IC 9486