IC 9502

INFORMATION CIRCULAR/2008 

Department of Health and Human Services

Centers for Disease Control and Prevention

National Institute for Occupational Safety and Health

Guidelines for the Prediction 

and Control of Methane 

Emissions on Longwalls 
 

Information Circular 9502 

Guidelines for the Prediction and Control of 

Methane Emissions on Longwalls 

By Steven J. Schatzel, Ph.D., C. Özgen Karacan, Ph.D., Robert B. Krog, 

Gabriel S. Esterhuizen, Ph.D., and Gerrit V. R. Goodman, 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 

March 2008 
This document is in the public domain and may be freely copied 

or reprinted. 

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DHHS (NIOSH) Publication No. 2008–114 

March 2008 

SAFER • HEALTHIER • PEOPLETM 
 

 

 

CONTENTS
 

Page 

Executive summary..........................................................................................................................1 

Acknowledgments............................................................................................................................4 

Background......................................................................................................................................4 

1. Reservoir modeling for predicting methane emissions in development headings (entries) .......6 

Overview.............................................................................................................................6 

 Effects of mining rate and degasification boreholes production on methane inflow .........8 

 Effects of coalbed parameters on methane inflow............................................................11 

 Coalbed permeability and permeability anisotropy ..................................................11 

 Langmuir volume, Langmuir pressure, and coalbed pressure ..................................13 

 Sorption time constant ..............................................................................................15 

Coalbed thickness .....................................................................................................17 

Initial water saturation of the coalbed.......................................................................18 

 Summary...........................................................................................................................19 

2. Controlling longwall face methane and development mining emissions: predicted improve- 

 ments using in-seam boreholes .............................................................................................21 

Overview...........................................................................................................................21 

 Effects of methane drainage on longwall face emissions .................................................25 

 Effectiveness of shielding the development entries from methane migration and 

 emissions ............................................................................................................................26 

 Summary...........................................................................................................................28 

3. Characterizing and forecasting longwall face methane emission rates for longer longwall faces..29 

Overview 29 

 Longwall face emission monitoring and prediction method.............................................29 

 Characterizing longwall face emissions and extrapolating emissions to longer faces .....31 

 Summary...........................................................................................................................37 

4. Predicting methane emissions from longer longwall faces by analysis of emission 

 contributors ................................................................................................................................38 

Overview...........................................................................................................................38 

Methodology.....................................................................................................................38 

 Assumptions..............................................................................................................39 

 Construction of formulas ..........................................................................................39 

Results...............................................................................................................................40 

 Technique for predicting methane emissions on longer longwall faces...........................43 

 Summary...........................................................................................................................47 

5. Development of numerical models to investigate permeability changes, distributions, and 

gas emissions around a longwall panel.................................................................................48 

Overview...........................................................................................................................48 

 Model development ..........................................................................................................48 

 Rock permeability calculations.................................................................................49 

 Distribution of gob permeability in a mined panel ...................................................51 

 Summary...........................................................................................................................52 

6. Methane emission control during mining of longwall panels using gob gas ventholes ...........53 

Overview...........................................................................................................................53 

Background 53 

 Effect of completion parameters of gob gas ventholes on their performance ..................54 

 Summary...........................................................................................................................59 

.......................................................................................................................

...........................................................................................................................
 

CONTENTS—Continued 

Page 

7. The application of gob gas ventholes to control methane in wider longwall panels and gobs....60 

Overview...........................................................................................................................60 

 Effects of increased longwall panel width on gob methane emissions.............................60 

 Comparison of ideally completed and operated gob gas boreholes with those operating 

in the field .....................................................................................................................61 

 Summary...........................................................................................................................65 

8. Induced fracturing and coalbed gas migration in longwall panel overburden: the NIOSH 

 borehole monitoring experiment...........................................................................................66 

Overview...........................................................................................................................66 

 Description of the borehole monitoring experiment on an active longwall panel............66 

 Methodology and instrumentation ....................................................................................68 

Analysis and interpretation of the field data .....................................................................70 

 Surface measurements of borehole gas.............................................................................73 

 Summary...........................................................................................................................77 

Practical guidelines for controlling longwall coalbed methane.....................................................78 

References 80 

ILLUSTRATIONS 

1. The modeled longwall panel showing entry sections where mining and ventilation data 

 were gathered during development mining..........................................................................7 

2. Methane inflow rate predictions for different development lengths and various premining 

degasification durations .......................................................................................................9 

3. Effect of mining rate and shielding well proximity when mining 610-m and 3,660-m

 (2,000-ft and 12,000-ft) long entries..................................................................................10 

4. Reservoir model predictions of methane inflow rate changes with permeability, permeability anisotropy, and entry length ...................................................................................12 

5. Effects of coalbed pressure on predicted methane inflow rates into development entries 

 using reservoir simulations ................................................................................................14 

6. Effects of Langmuir pressure (PL) and Langmuir volume (VL) on simulated methane 

inflow rates 15 

7.	 Influence of sorption time constant of the coal on reservoir-simulated methane inflow rates 

 during mining.....................................................................................................................16 

8. Effect of coalbed thickness on predicted methane inflow rates into entries of various 

 lengths using reservoir simulation models.........................................................................17 

9. Horizontal methane drainage borehole patterns modeled for degasification of the 

longwall panel....................................................................................................................22 

10. Cumulative methane production and possible reductions in longwall face emission rates 

 for different premining methane drainage time intervals and different horizontal 

borehole patterns................................................................................................................24 

11. Performance of different horizontal methane drainage borehole patterns shown in 

 figure 9 during different phases of degasification, after 3 and 12 months of 

premining degasification 25 

......................................................................................................................................

.........................................................................................................................

....................................................................................................
CONTENTS—Continued 

Page 

12. Sample layout showing additional horizontal, in-seam boreholes on either side of the 

gate roads (A+) in addition to pattern A ............................................................................27 

13. Longwall face emission study area .........................................................................................30 

14. Pittsburgh coalbed longwall face emission study site with instrument locations and 

 face segment identities.......................................................................................................32 

15. Ventilation configuration at the study site ..............................................................................32 

16. Actual portions of the study panel monitored for face methane emissions and associated 

 methane drainage borehole locations.................................................................................33 

17. Average methane emission rates for each day of study determined for each face segment......35 

18. Methane emission prediction curves for the Pittsburgh coalbed ............................................37 

19. Calculated individual methane contributors and total calculated methane emissions for 

study day 1 .........................................................................................................................42 

20. Calculated methane emissions for two full cuts without delays for a 488-m (1,600-ft)

 longwall face......................................................................................................................45 

21. Calculated influence of variable shearer transit speed pass for a 488-m (1,600-ft) 

longwall face......................................................................................................................46 

22. Three-dimensional representation of grid model of the study area that shows the 

 Waynesburg and Pittsburgh coalbeds ................................................................................54 

23. Simulated cumulative methane production from ventholes with varying lengths of 

slotted casing......................................................................................................................56 

24. Simulated methane concentrations in the gas production stream from ventholes completed 

 to varying depths above the Pittsburgh coalbed ................................................................57 

25. Actual methane concentrations measured in the gas production stream from two GGVs 

 completed to different depths above the Pittsburgh coalbed on the same longwall panel ...58 

26. Study mine for wider longwall faces: grid model of a new mining district in the 

 Pittsburgh coalbed..............................................................................................................60 

27. Schematic representation of GGV configurations for the simulation of methane control 

options on wide longwall panels........................................................................................62 

28. Comparison of the performance of actual GGVs at the study site with the optimal GGVs

 for the 442-m (1,450-ft) wide panel...................................................................................62 

29. Stratigraphy and downhole configurations for NIOSH BME.................................................68 

30. Wellhead arrangement for monitoring boreholes ...................................................................69 

31. Water head change as a function of time in boreholes monitoring Sewickley and shale- 

 limestone intervals .............................................................................................................70 

32. A: BH–1 instantaneous change in the permeabilities of monitored horizons during mining; 

 B: BH–1 average permeabilities at certain intervals of water head drop during mining .....72 

33. A: BH–2 instantaneous change in the permeabilities of monitored horizons during mining; 

 B: BH–2 average permeabilities 74 

34. Methane concentration and static shut-in pressures measured in the boreholes during 

 progress of mining .............................................................................................................76 

........................................................................................
CONTENTS—Continued 

TABLES 

Page 

1. Development mining parameters and their range of values used for modeling purposes ........

..............................................................................................

8 

2. Coalbed reservoir properties and their range of values ............................................................

................................................................

..........................................................

8 

3. Linear models for predicting methane inflow rates into development entries of various 

 lengths using key coalbed reservoir parameters 19 

4. Simulated reductions in face emission rates after different phases of methane drainage 

 using horizontal boreholes from the longwall panel 23 

5. Predicted methane emissions into the ventilation system from the surrounding coalbed 

(within the panel and virgin coal) ......................................................................................

...............

......................................................................................................................

28 

6. Face segment emission rate data shown for combined averaged H–T, T–H passes 34 

7. Daily methane emission contributor averages and percentages using daily shield 20 and 

145 constants 41 

 8. Methane contribution percentages from longwall emission contributors during gas delays 

 compared to daily averages for a 315-m (1,030-ft) longwall face.....................................

....................................................................................................................

............................................................................................

41 

9. Calculated rates for methane emission contributors on idealized passes on longer 

longwall faces 44 

10. Predicted effect of casing diameter on cumulative methane and total gas production for 

 910 days of simulated mining 55 

11. Cumulative predicted methane production difference from gob gas ventholes on a 

442-m (1,450-ft) wide panel 64 
 

 

 

ACRONYMS AND ABBREVIATIONS USED IN THIS REPORT 

BH borehole 

BME borehole monitoring experiment 

 CH4 methane 

FLAC Fast Lagrangian Analysis of Continua 

GEM Generalized Equation-of-State Model 

GGV gob gas venthole 

H–T head-to-tail 

NIOSH National Institute for Occupational Safety and Health 

PRL Pittsburgh Research Laboratory (NIOSH) 

PVC polyvinyl chloride 

 T–H tail-to-head

 USBM U.S. Bureau of Mines 
UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT 

cc/g cubic centimeter per gram

 cfm cubic foot per minute 

 cm/s centimeter per second 

 cps counts per second 

 ft foot

 ft3 

 cubic foot 

ft/day foot per day 

 ft/min foot per minute 

g/cc gram per cubic centimeter 

 in inch

 in/s inch per second 

kPa kilopascal 

m meter

 m/day meter per day 

 m/min meter per minute

 m/s meter per second 

m3 cubic meter 

 m3

/day cubic meter per day 

m3/min cubic meter per minute 

m3

/s cubic meter per second 

md millidarcy 

 min minute

 MMscf million standard cubic feet 

 Pa pascal

 psi pound-force per square inch 

psia pound-force per square inch absolute 

scf standard cubic feet 

scf/day standard cubic feet per day 

scf/ton standard cubic feet per ton 

 sec second 

 

 

 

 

 

 

 

 

 

 
 

 

GUIDELINES FOR THE PREDICTION AND CONTROL OF 

METHANE EMISSIONS ON LONGWALLS 

By Steven J. Schatzel, Ph.D.,1 C. Özgen Karacan, Ph.D.,2 Robert B. Krog,3
 

Gabriel S. Esterhuizen, Ph.D.,2

 and Gerrit V. R. Goodman, Ph.D.4
 

1

Lead research geologist. 

2

Senior service fellow. 

3

Associate service fellow. 

4

Team leader, Ventilation and Explosion Prevention Group, Disaster Prevention and Response Branch. 

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

EXECUTIVE SUMMARY 

Although longwall mining productivity can far exceed that of room-and-pillar mining, 

the total methane emissions per extracted volume associated with longwall sections are generally 

higher than those for continuous miner or pillar removal sections. Increased face advance rates, 

increased productivities, increased panel sizes, and more extensive gate road developments have 

challenged existing designs for controlling methane on longwalls. 

Methane control research by the National Institute for Occupational Safety and Health 

(NIOSH) recently examined a number of practices designed to maintain concentrations in mine 

air within statutory limits and consistently below the lower explosive limit. These included a 

reservoir modeling approach to predict methane inflows in gate road entries. The outputs suggested that emission rates in the gate roads decreased with the use of shielding boreholes and 

increased degasification time. Also, mining perpendicular to the face cleats liberated more gas 

into the mine workings, emissions were almost a linear function of Langmuir pressure and 

volume, and emissions were inversely related to sorption time constant. 

Subsequent simulations predicted changes in methane drainage using in-seam boreholes. 

The results showed that longer degasification times resulted in lower face emission rates. 

Premining degasification produced more methane than that produced during panel extraction, 

a fact attributed to the already decreased methane content of the coalbed. This work concluded 

that longer premining degasification periods would be more advantageous to the operator. 

The industry trend toward increasing longwall face width can produce increased methane 

emissions from the face due to a higher volume of cut coal on the face conveyor. Two methods 

were used to estimate face methane levels on longer faces. In the first method, segmented 

methane data were extrapolated for greater face widths. The second method estimated emissions 

contributions from the shearer, face conveyor, panel belt, longwall face, and ribs, and summed 

these for wider panels. 

Permeability changes in the gob behind the longwall shields were estimated using a 

NIOSH-developed numerical model. The results showed that permeability was highest near the 

edges and corners of the gob. 
Gob gas ventholes (GGVs) represent a very effective means for controlling methane gas 

on longwall faces. NIOSH work showed that (1) larger-diameter GGVs produced more methane, 

but at the expense of increased dilution due to the presence of mine ventilation air in the boreholes; (2) GGVs should not be sunk into the caved zone of a longwall panel since more ventilation airflow will be sent to the boreholes; and (3) increased slotted casing lengths improved 

GGV production. 

NIOSH conducted an extensive borehole monitoring experiment (BME) to assess the 

impacts of longwall mining on development of the coalbed reservoir. Three boreholes were 

drilled to different stratigraphic horizons in advance of undermining by a longwall and instrumented to measure pre- and postmining in situ permeabilities and gas pressures. The data 

showed that mining-induced disturbances occurred 20–46 m (80–150 ft) ahead of the retreating 

longwall face, causing a corresponding increase in formation permeability. 

In this report, several practical guidelines are recommended for controlling longwall 

coalbed methane. All predictions are based on determinations made for the Pittsburgh Coalbed 

in southwestern Pennsylvania. 

•	 It is recommended to use shielding degasification boreholes to decrease emission 

rates by at least 25% for development entries. Drill these boreholes as close as 

practically possible (~27 m (90 ft)) to the entries and operate them for at least 

6 months to achieve a 25%–50% decrease in emission rates (Section 1). 

•	 Equations for predicting methane emissions rates into gate road developments are 

presented in Table 3 (Section 1). These relationships assume a supercritical longwall 

panel developed in the Pittsburgh Coalbed. 

•	 For a longwall panel width greater than 305 m (1,000 ft), a trilateral borehole 

configuration is recommended for effective draining of methane. Short across-panel 

holes at close spacing can also be effective in degassing a panel. The required number 

of holes will depend on seam anisotropic reservoir conditions, but at least 12 holes for 

a 3,400-m (11,000-ft) long panel are recommended to achieve the same degasification 

as the trilateral configuration (Section 2). 

•	 If less than 12 months are available for premining gas drainage, it is recommended 

that degasification be continued until the borehole is approached by mining. This 

approach maximizes the quantity of removed methane and reduces methane emission 

rates (Section 2). 

•	 To avoid shearer coal production delays, it is recommended that continuous GGV 

production be assured while GGVs are within about 500 ft of the working face. In 

many mines, the quantity of coalbed methane removed by a GGV is potentially 75% 

of the volume of gas emissions on the longwall face. A similar finding was observed 

by prior NIOSH research. If an operating GGV ceases producing gas, the gas that was 

being removed will enter the ventilation system (Section 3). 

•	 Assuming a well-caved gob, increasing the longwall face length by X% will increase 

the rate of methane emissions by, at most, two-thirds of X% (Section 4). 

2
 
•	 The length of the slotted casing section of a GGV will strongly influence its level of 

gas production. To effectively design the slotted casing section of a GGV, it is 

recommended to: 

○	 Review the local geology to identify the location of gas-bearing units; and 

○	 Set the top of the slotted section at the highest gas-producing stratigraphic horizon 

(Section 6). 

•	 Completing a GGV into the caved zone is counterproductive and increases the 

likelihood of intermittent production from increased-width, supercritical panels. 

Therefore, the completion depth of GGVs should be at least 14 m (45 ft) above the 

top of coal for longwall panels, particularly in the Northern Appalachian Basin 

(Sections 6 and 7). 

•	 Emphasize continuous GGV production since it will potentially produce 40%–50% 

more coalbed gas than GGVs operating intermittently (Section 7). 

•	 Increasing the longwall panel width increases the quantity of methane present 

because of the increased fractured reservoir volume. However, this increase does not 

enhance the performance of GGVs. As panel width increases, the effectiveness of 

GGVs completed near the tailgate margin will not extend as close to the headgate 

side. Drilling GGVs on the headgate side or near the panel centerline can produce 

coalbed gas from this portion of the panel (Sections 5 and 7). 

•	 In favorable cases, GGVs drilled on the headgate side can be effective. Completion 

depths must isolate the borehole from communication with the ventilation network. 

These findings are based on supercritical panel designs (Section 7). 

•	 Mining-induced fracturing was observed to occur 24–46 m (80–150 ft) ahead of 

the mine face. Boreholes and exhausters should be installed before this occurs 

(Section 8). 

•	 Data were reviewed for GGV configurations completed from 7 to 32 m (24 to 106 ft) 

above the mined coalbed for supercritical panels in the Northern Appalachian Basin. 

It is recommended that GGVs be completed toward the top of this interval and be 

designed to include the Sewickley Coalbed. Permeability increases following 

undermining were dramatic, with increases of about 100–500 times the premining 

values and instantaneous increases of up to about 1,000 times these values. These 

measurements did not differ significantly despite differences in borehole configurations. Fracture permeability pathways remain high to the mined coalbed toward the 

top of the described interval, yet the likelihood of drawing ventilation air into the 

borehole is minimized (Section 8). 

3
 
ACKNOWLEDGMENTS 

The authors thank Thomas P. Mucho, W. P. Diamond, and Fred Garcia, all retired from 

NIOSH, for their significant contributions to the completion of the technical work throughout 

this research project. The authors acknowledge the valuable technical and editorial input 

received from field reviewers Pramod C. Thakur, Ph.D., CONSOL Energy, Inc.; Jerry C. Tien, 

Ph.D., University of Missouri-Rolla; and Thomas McNider, Jim Walter Resources, Inc. The 

authors also acknowledge Robert J. Tuchman, Technical Writer-Editor, Centers for Disease 

Control and Prevention, and Joe Schall, the Giles Writer-in-Residence at The Pennsylvania State 

University, for their substantial contributions to improving the clarity and quality of this report. 

In addition, the authors thank the cooperating mine operators and staff who made these research 

efforts possible and who have requested anonymity. 

BACKGROUND 

In today’s competitive energy market, U.S. deep coal mine producers often choose 

longwall over room-and-pillar mining because of the high productivities and improved safety 

benefits that can be achieved. The longwall mining method requires an adequate and suitable 

coal reserve where the coal is of good quality and where geologic conditions are appropriate. 

Because of the nature of longwall mining, total methane emissions associated with longwall 

mining are generally higher for a specific volume of mined coal than for continuous miner or 

pillar removal sections. During longwall mining, a large block of coal is extracted and the 

immediate roof, three to six times the thickness of the mined coalbed, extensively fractures and 

falls into the mine void [Singh and Kendorski 1981; Palchik 2003]. The stress relief in the 

surrounding strata that results from this caving creates large horizontal fractures along bedding 

planes and vertical fractures in the strata overlying and underlying the caved zone. These 

mining-induced fractures provide extensive pathways for gas migration from the surrounding 

coalbeds and other gas-bearing strata into the mining environment. The fractured zone is present 

in the roof and floor can vary up to 100 times the height of the mined coalbed of the overburden 

[Curl 1978; Thakur 1981] depending on the mechanical properties of the rock layers, thickness 

of the overburden, and size of the panel. 

The methane that originates and accumulates in the gob above the mined-out panel is the 

main source of methane emissions during longwall mining. Prior monitoring studies directed at 

longwall face emissions indicate that only a small portion of the overall methane emission and 

gas production is emitted at the mine face [Tauziede et al. 1997; Diamond and Garcia 1999]. 

It has been reported that methane contributions from the subsided strata (gob) generally account 

for 80%–94% of the methane present in the ventilation system of an operating longwall [Curl 

1978; Schatzel et al. 1992]. These research findings suggest that typical longwall face emissions 

account for no more than 20% of longwall emissions and as little as 6% of the total emissions at 

one site [Schatzel et al. 1992] and potentially less at very gassy mines. However, this particular 

emissions source can be critical in relation to underground safety. Face areas on longwall mining 

sections present specific challenges for methane control techniques where high production rates 

can cause relatively high methane emission rates, making it difficult to meet statutory limits on 

methane concentrations at face and tailgate sampling locations. Recent input from NIOSH stakeholders suggests that the longwall tailgate T-junction area is crucial for effective ventilation air 

dilution of methane emissions and will be a subject of upcoming research. 

4
 
Available methane control systems have been challenged by recent developments in 

longwall mining. These developments include increased face advance rates leading to increased 

coal production, increased longwall panel sizes resulting in more extensive gate road developments, and the generally deeper and gassier workings of U.S. coal mines. All of these have 

presented additional challenges for longwall face ventilation. Work by Diamond and Garcia 

[1999] reported the predicted impact of face extensions at two longwall operations in the 

Pocahontas Coalbed from mining operations in northwestern Virginia. Increasing face length 

from 229 to 305 m (750 to 1,000 ft) (33%) led to a predicted increase in face methane emissions 

from 8.0 to 8.6 m3

/min (280 to 300 cfm), or an increase of about 7%. At an adjoining operation, 

a similar increase in face width led to a predicted methane emissions increase of about 13%, 

from 14.2 to 16.1 m3

/min (502 to 567 cfm). Variabilities in mine design and methane control 

practices between the two mine sites were the primary causes of the different predicted methane 

emission rates for the 1,000-ft face widths. Schatzel et al. [2006] and Krog et al. [2006] produced 

empirical models of methane emissions for longer longwall faces for a mine operating in the 

Pittsburgh Coalbed of southwestern Pennsylvania. The calculated methane emissions by Krog 

et al. were determined on the basis of 61-m (200-ft) increment increases in face length. Peak 

methane emissions increased from 6.6 to 9.1 m3

/min (234 to 322 cfm), or 38%, as face length 

increased from roughly 305 to 488 m (1,000 to 1,600 ft). This determination was based on 

computed constants associated with specified methane sources (shearer, face conveyor, belt 

conveyor, background emissions from the coal face, and background emissions from the 

adjoining ribs in the intake gate roads) and a zero time delay “idealized” longwall face pass of 

the shearer. Schatzel et al. [2006] predicted methane emissions on longer longwall faces for 

lengths of 366, 427, and 488 m (1,200, 1,400, and 1,600 ft) of 5.4, 6.4, and 7.4 m3/min (191, 225, 

and 263 cfm), increases of 36%, 61%, and 88%, respectively, over the base case. The lower face 

emission rates predicted by Schatzel et al. compared to those of Krog et al. arose from the 

inclusion of production delays and average emission rates in the first study compared to no 

delays and peak emission rates in the second study. 

The increasing coal productivities from longwall operations have led to greater volumes 

of coalbed methane gas entering the underground mine environments from exposed coal surfaces 

and from cut coal on the conveyor belting. Production advances in longwall mining have significantly challenged currently available methane control designs for reducing the potential for 

methane explosions. Such designs require that reservoir pressures and gas contents be reduced to 

limit migration of coalbed gas into the mine atmosphere, and these techniques require sufficient 

lead time to achieve sufficient reductions in methane inflow. Coalbed methane emission rates 

can vary widely among different coalbeds and should be considered in the application of the 

recommendations provided. Even within the same coalbed, geographic changes can produce 

changes to gas storage or emission-related geotechnical parameters, thereby modifying coalbed 

methane emission behavior. 

5
 
1. RESERVOIR MODELING FOR PREDICTING METHANE EMISSIONS 

IN DEVELOPMENT HEADINGS (ENTRIES) 

Overview 

Planning, designing, and optimizing ventilation are important steps to eliminate any 

accumulation of explosive methane-air mixtures in the development and extraction phases of 

longwall coal mining. Insufficient ventilation and unexpectedly high methane liberation rates can 

overwhelm the existing ventilation capacity and result in elevated concentrations of methane gas, 

which can lead to production interruptions and endanger the safety of workers underground. 

Mine operators usually try to supply maximum ventilation air based on the capacity of the 

system and the predicted needs underground. However, ventilation capacity may decrease over 

time due to leakage, especially as the gate road entries get longer. Also, increased mining 

advance rates and changes in geologic conditions can produce substantial increases in methane 

emission rates that can challenge and overwhelm ventilation system designs. Consequently, 

it has become increasingly difficult to keep methane levels under statutory limits by ventilation 

alone, and ventilation requirements may depart significantly from initially planned values. This 

condition can both limit the advance of a particular section and increase the risk for an ignition 

or explosion. Thus, it is important to develop an understanding of key parameters influencing 

methane emission rates and sound techniques to optimize mine ventilation and methane drainage 

effectiveness to ensure a safer work environment. 

One key requirement in meeting this objective is the ability to predict methane emission 

rates. For gate road developments of longwall panels, the prediction of methane emission rates is 

affected by both mining parameters and coalbed reservoir parameters. Although changes made to 

mining parameters, such as development rate and gate road length, can impact methane inflow, 

similar options do not exist when dealing with coalbed reservoir properties, such as coalbed 

thickness and reservoir pressure. Thus, it is important to understand the effects of these parameters on methane inflow rates in order to be able to control methane emissions effectively. 

When it is difficult to keep methane under statutory limits by ventilation alone, one of the 

most effective approaches is to drill horizontal boreholes in the coalbed to drain excessive 

methane before mining starts [Brunner et al. 1997; Noack 1998]. Effective horizontal methane 

drainage well designs were pioneered by the U.S. Bureau of Mines (USBM) and by CONSOL 

Energy in the 1970s [Cervik et al. 1975; Thakur and Poundstone 1980; Prosser et al. 1981; 

Diamond 1994; Thakur 2006]. In-seam horizontal wells drilled into the planned longwall panels 

and ahead of development mining serve two purposes. First, wells reduce the methane content of 

the coalbed in and around the panel area where development mining will be conducted. Second, 

wells shield the developing entries from methane migration from the coalbed surrounding the 

roadways. This approach has been shown to be effective in reducing methane emission rates. 

Thakur [2006] discussed methane drainage methods where the borehole drilling trailed gate road 

development, but was in advance of the longwall face. Prior research that reported horizontal 

well performance comprised novel studies, but generally used empirical relationships [Cervik 

et al. 1975; Thakur and Poundstone 1980; Diamond 1994]. Thus, an engineering-based analytical 

approach was proposed to develop horizontal borehole design improvements to enhance the 

methane control and ventilation system. 

6
 
 Figure 1.—The modeled longwall panel showing entry sections where mining and ventilation data were 

gathered during development mining. 

A reservoir model was configured to calculate methane emission rates during 

development mining [Karacan, forthcoming; Karacan et al. 2007a]. The model included a threeentry tailgate and headgate development layout, which is typical of coal mines operating in 

Pittsburgh Coalbed in the southwestern Pennsylvania section of the Northern Appalachian Basin 

(Figure 1). 

During the mining of headgate and tailgate entries of the study mine, ventilation air 

qualities and methane in flow rates were measured by the operating mining company using 

methanometers and anemometers at the monitoring locations shown in Figure 1. Methane inflow 

rates were quantified based on the measured data. The data were reported as average monthly 

concentrations of methane, airflow rate, raw and clean tonnages of produced coal, and total 

linear distances advanced during mining. The length of each entry section was around 3,350 m 

(11,000 ft) and required 8–9 months to mine. 

The in-mine measurements of methane inflows were matched with the model predictions 

to calibrate the reservoir model. Various models were developed to investigate methane 

emissions during development mining by relating mining parameters (Table 1) and coalbed 

reservoir parameters (Table 2). In the first set of analyses, the effects of mining parameters with 

and without horizontal degasification wellbores were investigated by keeping the coalbed 

reservoir parameters at their base levels (Table 2) while varying the mining parameters (Table 1). 

In the second set of analyses, mining rate was kept constant and coalbed reservoir parameters 

were varied to investigate the effects of these parameters on methane emissions (Table 2). 

The methane concentrations reported in Table 1 represent values for the last open 

crosscut within development gate roads. The cleat permeability and effective porosity data 

shown in Table 2 represent values for the Pittsburgh Coalbed in southwestern Pennsylvania and 

may vary from determinations made for other locations within the coalbed. The values given for 

shielding well distances to development entries were used to show trends in borehole 

effectiveness for reducing emissions in relation to entry proximity. In practice, drilling a 

borehole 5.8 m (19 ft) from a development entry would create a range of operational challenges. 

Mine operators typically drill methane drainage boreholes 15–46 m (50–150) ft from development entries. 

7
 
 

 

Table 1.—Development mining parameters and their range of values 

used for modeling purposes 

Mining parameter Range of values 

Mining height, ft 5.0–7.0 

Entry development length, ft 1,000–12,000 

Mining rate, ft/day 25–175 

Methane concentration in mine air, % 0.5–1.5 

Distance of shielding wells to entries, ft 19–87 

Duration of degasification before mining, days 0–180 

 

Table 2.—Coalbed reservoir properties and their range of values 

Coalbed reservoir parameter Base value Range of values 

Coalbed thickness, ft 6.0 5.0, 6.0, 7.0 

Coalbed reservoir pressure, psi 90 90, 200, 400 

Sorption time, days 20 20, 125, 250 

Permeability anisotropy, Kx/Ky 4 4, 8 

Face cleat permeability, Kx, md 4 4, 20, 40 

Butt cleat permeability, Ky, md 1 0.5, 1.0, 2.5, 5.0, 10.0 

Langmuir volume, scf/ton 392 200, 392, 600 

Langmuir pressure, psi 326 126, 326, 526 

Initial water saturation, Swi, % 60 40, 60, 100 

Effective porosity, % 4 0.5, 2, 4, 5 

Irreducible water saturation, Swir, % 40 10, 30, 40, 58 

Relative permeability to gas (Krg) at Sg = 1 − Swir 0.35 0.35, 0.50, 0.70, 0.90 

Effects of Mining Rate and Degasification Borehole Production 

on Methane Inflow 

In-seam horizontal boreholes are effective in degasifying coal during development 

mining. To simulate the shielding effects of degasification wells on methane inflow and on the 

ventilation air requirements, 7.6-cm (3-in) diameter horizontal boreholes with no wellbore skin 

were modeled and operated with a negative bottom-hole pressure of 1,400 Pa (0.2 psi) for 

premining degasification periods of 0 months (i.e., no premining degasification), 3 months, and 

6 months. Instead of the conventional distance used by industry, the boreholes were placed 5.8– 

27 m (19–90 ft) away from the entries to demonstrate the importance of proximity to effective 

emission reductions. In all cases, the boreholes were drilled ahead of development workings and 

operated as gate road entries were developed. 

Figure 2 compares methane inflow rates as a function of development length for 

horizontal boreholes located 5.8 m (19 ft) away from the intake and return entries and operated 

for various premining durations. The data represent a 21-m/day (70-ft/day) development rate in a 

2.1-m (7-ft) thick coalbed. Simulations show that the methane emissions into the development 

entries are highest when shielding degasification boreholes are not used before or during mining. 

Emission rates progressively decrease with increases in the duration of degasification. For 

instance, even if the boreholes did not operate before mining and begin to operate when mining 

starts, the inflow rate decreases about 25% compared to the completely unshielded case. The 

8
 
methane inflow rates to the ventilation air are lowest when longer premining degasification times 

are applied. 

Emission rates in development entries decrease as a result of 

degasification. Even if the boreholes are placed just before 

mining, the methane rate in the ventilation air decreases about 

25% compared to unshielded entries. However, larger emission 

decreases are possible. 

 

 

Figure 2.—Methane inflow rate predictions for different development lengths and various premining 

degasification durations. 

The simulated data shown in Figure 3 are based on a 21-m/day (70-ft/day) development 

rate in a 2.1-m (7-ft) thick coalbed after 6 months of premining degasification, except for the 

“No wells present” case, which does not have any shielding. This figure shows that wellbores 

closest to the entries are most effective in reducing methane inflow rates during development 

mining, about 50% at 3,660 m (12,000 ft) compared to no shielding. 

9
 
 Figure 3.—Effect of mining rate and shielding well proximity when mining 610-m and 3,660-m (2,000-ft and 

12,000-ft) long entries. 

Shielding wells located close to gate road entries and operated 

for longer times are the most effective method in reducing the 

methane inflow rate into the development entries. Mine operators 

mining longer entries at higher mining rates will experience the 

greatest emissions reduction when applying this technology. 

Figure 3 also shows that the methane inflow rate increases with the length of the 

development section because of the increase in surface area of the exposed coalbed. This 

increase is strongly related to the mining rate. However, the effect of the mining rate on methane 

inflow is less pronounced for shorter development distances than for longer distances [Karacan 

et al. 2007a]. 

10
 
 

The coalbed methane inflow rate increases with the length of the 

development section. The magnitude of the increase is strongly 

influenced by the mining rate, about 23,500 m3

/day (1 × 106 

scf/day) with wells 25 m (82 ft) from the development heading. 

However, the effect of the mining rate on methane inflow is less 

pronounced for shorter development distances (610 m (2,000 ft)) 

than for longer entries. 

Effects of Coalbed Parameters on Methane Inflow 

The following subsections summarize the impacts of changing coalbed reservoir parameters (Table 2) on methane inflow. For these simulations, a 15-m/day (50-ft/day) section advance 

rate was assumed. 

Coalbed Permeability and Permeability Anisotropy 

Permeability is an intrinsic property of porous formations that relates pressure drop and 

flow rate. Coalbeds are characterized as anisotropic reservoirs with two orthogonal permeability 

components that run in the direction of characteristic fractures within the coalbed. These are 

termed the “face cleats” and “butt cleats.” Face cleats are continuous and can extend for long 

distances. Butt cleats are short and discontinuous, usually terminating at face cleats. Due to the 

nature of these two cleat systems, face cleats are more permeable to gas and water flow. Vertical 

permeability is usually significantly lower than horizontal permeability. Thus, horizontal 

permeability and its direction are more important in coalbed reservoirs. The permeability 

anisotropy can be defined as the ratio of face cleat to butt cleat permeabilities. The presence, 

lack, or direction of permeability has a profound effect on methane inflow into the wellbores 

[Remner et al. 1986], suggesting that any horizontal drilling in coalbeds should account for 

directional permeability. The average permeability for flow in coalbeds can be defined as 

Ka = Kx × Ky (1) 

where Ka is the average permeability, and Kx and Ky are the directional permeabilities of 

the cleats. 

Direction and magnitude of permeability are important for control of methane emissions 

during mining. When mining advances perpendicular to the face cleats, where permeability is 

higher, much more gas is emitted into mine workings than when the advance is parallel to the 

face cleat. In fact, in one area of a mine operating in the Pittsburgh coalbed, the methane emission from the solid ribs was significantly higher than that from the working face as a result of the 

ribs intersecting the face cleats [McCulloch et al. 1975]. 

11
 
 Figure 4.—Reservoir model predictions of methane inflow rate changes with permeability, permeability 

anisotropy, and entry length. 

Figure 4 shows that increasing average permeability, while keeping the anisotropy 

constant, increases methane inflow into the entries. On the other hand, when anisotropy is 

doubled, methane inflow decreases, even though the face cleat permeability remains the same. 

This is due to a lower flow rate from the decreased butt cleat permeability where the butt cleat is 

perpendicular to the entries in this study. Also, Figure 4 shows that the effects of anisotropy or 

changes in butt cleat permeability are more pronounced for longer development distances. In a 

more general sense, the effect of butt cleat permeability observed in this study should be interpreted as the effect of coalbed permeability in the cleat system, which is perpendicular to the 

entries based on orientation of mining. 

The permeability of cleats perpendicular to the mine entries is 

more important in controlling methane inflow rates than the 

maximum face cleat permeability. 

12
 
 

 

Langmuir Volume, Langmuir Pressure, and Coalbed Pressure 

In coalbeds, the major portion of the gas exists in the adsorbed state within the coal 

matrix rather than in a free state within the cleats. As the pressure is lowered due to fluid production from the cleats, gas starts to desorb from the micropore surfaces and diffuse into the 

macropores [Remner et al. 1986]. Adsorption and desorption of methane from coal is described 

by a Langmuir isotherm that defines the relationship of coalbed pressure to the capacity of a 

given coal to hold gas at a constant temperature. More specifically, the Langmuir isotherm for 

coals relates matrix gas content, V(P), to the coalbed cleat pressure, P, according to: 

V × P

V (P) = L

(2) (P + PL ) 

VL is the Langmuir volume or maximum amount of gas that can be adsorbed. Langmuir pressure, 

PL, is a measurement strongly related to the residence time for a gas molecule on a coal surface 

and represents the pressure at which gas storage capacity is half of the maximum storage 

capacity, VL [Young 1998]. Both VL and PL can be determined from gas sorption measurements 

on the coal core samples. The Langmuir equation (Equation 2) provides a necessary boundary 

condition at the matrix-cleat interface [Young 1998]. For nonequilibrium diffusion sorption 

models, it is assumed that the concentration of methane at the surface of the micropore matrix 

blocks is in equilibrium with the free gas pressure in the cleats. This implies that the external 

boundary condition of the micropore equation is the equilibrium sorption isotherm [King et al. 

1986]. The shape of this isotherm is important for manipulating the boundary condition between 

matrix and cleats and for controlling the desorption rate of gas from coal. 

Coalbed methane emission rates are strongly influenced by 

methane content. The gas content of a coal at a range of coalbed 

pressures is controlled by the Langmuir isotherm, relating 

Langmuir volume, VL, and Langmuir pressure, PL. Cleat pressure 

is also important for determining the gas content of the coalbed 

for a given VL and PL. 

It should be noted that initially the actual amount of gas in the coalbed (gas content) may 

not be on the desorption curve defined by the Langmuir isotherm, but may be below this curve. 

In this case, coalbed cleat pressure needs to be reduced further until it reaches a critical desorption pressure where gas starts to desorb and gas release follows the Langmuir isotherm as pressure declines. 

To obtain the data in Figure 5, mining advance was oriented parallel to the face cleats to 

mimic the modeled mine. Both face and butt cleat permeabilities were varied between the minimum and maximum values shown in Table 2. Permeability anisotropies were increased by 

decreasing the butt cleat permeability. 

Figure 5 shows the effect of coalbed pressure on the methane inflow rate during the 

development mining of gate road entries. A higher coalbed gas or hydrostatic pressure results in 

a higher methane inflow rate at constant VL and PL. The dependency of inflow rate on pressure is 

nearly linear for each of the entry lengths studied. Also, the inflow rate is higher when mining 

13
 
 Figure 5.—Effects of coalbed pressure on predicted methane inflow rates into development entries using 

reservoir simulations. 

longer entries, as expected, since longer entries create a larger drainage area into which gas 

can flow. Furthermore, a much larger portion of the coal seam is depressurized sooner, which 

triggers gas desorption from greater distances. A similar phenomenon is observed when 

increasing the length of horizontal coalbed production wells [Ertekin et al. 1988; King et al. 

1986]. 

The effects of VL and PL on simulated methane inflow into the entries during development mining are shown in Figure 6. This figure shows that methane inflow rate is almost a linear 

function of VL. Since VL represents the maximum gas capacity of coal, higher VL ensures the 

availability of more methane and promotes higher methane emissions when pressure and PL are 

kept constant. The variable PL shows an inverse relationship with the methane inflow rate when 

both VL and pressure are constant (Figure 6). The slope of the isotherm, or the change of gas 

capacity with pressure, is largely controlled by PL. As PL increases, then the isotherm will be flat 

compared to an isotherm generated from a coal sample with a lower PL value. This means that 

when mining a coalbed with a higher PL, the change in gas capacity or change in gas desorption 

from the coal will be less compared to a seam with a lower PL for the same pressure drop. Thus, 

the rates of desorption and methane inflow into entries in a higher-PL coalbed decrease as 

development distances increase. 

14
 
Figure 6.—Effects of Langmuir pressure (PL) and Langmuir volume (VL) on simulated methane inflow rates. 

Methane emission rates are an almost linear function of pressure 

and the distance to be mined. Simulations show that a higher 

Langmuir volume VL ensures the availability of more methane 

and promotes a higher methane drainage rate when pressure 

and Langmuir pressure, PL, are kept constant. The inverse is true 

with a higher Langmuir pressure. 

15
 

Sorption Time Constant 

Once methane desorbs from the micropore surfaces of a coal matrix, it flows through the 

micropores in response to a methane concentration gradient defined by a combination of 

Knudsen, bulk, and surface diffusion processes [Smith and Williams 1984; Kolesar and Ertekin 

1986]. In coalbed modeling, unsteady-state diffusion effects can be quantified by determining a 

sorption time, τ (days), which is related to cleat spacing, sf, and the diffusion coefficient, D: 

s

2 

τ ~ f (3) D 
 Figure 7.—Influence of sorption time constant of the coal on reservoir-simulated methane inflow rates 

during mining. 

 In effect, τ is the time constant that regulates the rate at which gas is released from the 

micropores into the macropore system. For small values of τ, or larger values of D, the diffusion 

and sorption process is faster, and a higher cumulative gas production and a higher production 

rate peak are observed at degasification boreholes [Remner et al. 1986; Spencer et al. 1987]. 

Later in the life of the coalbed methane reservoir, sorption time, together with desorption, may 

also act as an internal pressure maintenance mechanism. 

The simulated effects of different sorption times are shown in Figure 7, which demonstrate that small sorption times cause higher methane inflow rates, as observed in gas production 

from coalbeds using horizontal boreholes [Remner et al. 1986]. It is also evident that the change 

in inflow rates between short and long sorption times is more significant as entry lengths 

increase. This relationship may be related to the regulation of the rate of gas release from 

micropores to cleats and the gas pressure distribution in the reservoir during active mining. 

When the entries being mined are relatively short, the pressure disturbance within the coalbed is 

more limited and the area of methane drainage is diminished. When mining longer entries, the 

inflow rates are higher because the pressure disturbance affects a larger area for a longer time 

and a larger surface area is exposed. A short sorption time, τ, results in high gas flow rates from 

the areas surrounding the entries and from the areas where pressure transients can propagate. 

In the case of longer sorption times, the flow rate will be less due to slow diffusion from the coal 

matrices into the cleats. Also, the pressure transients will travel farther since there will not be 

enough pressure maintenance in the coalbed. 

16
 

Small sorption times cause higher methane inflow rates into gate 

road entries. The effect of the sorption time constant on methane 

inflow rates into mine workings is more significant for the mining 

of longer entries than for shorter ones. 

Coalbed Thickness 

Figure 8 shows that the methane inflow rates increase almost linearly with the coalbed 

thickness. The figure also shows that, as expected, mining longer entries in thicker coalbeds 

results in higher inflow rates. This is due to a combination of the two factors suggested by 

Ertekin et al. [1988] for fracture-stimulated horizontal boreholes: the reservoir volume is larger 

and the surface area for gas flow is increased in fracture-stimulated horizontal boreholes. Ertekin 

et al. reported that methane production from horizontal boreholes increases as the coalbed 

thickness increases even if the borehole surface area is the same because of larger volumes of 

reserves encountered. 

During development mining in thicker coalbeds, larger volumes of methane reserves are 

encountered. Also, the increased thickness of the coalbed results in an increased area of flow due 

to the presence of the natural fracture network. Thus, when mining longer entries, increased 

magnitudes of reserves that are encountered and larger surface areas created combine for increasing inflow rates even further. 

Figure 8.—Effect of coalbed thickness on predicted methane inflow rates into entries of various lengths 

using reservoir simulation models. 

17
 
Initial Water Saturation of the Coalbed 

Coalbeds are almost always partially or completely saturated with water. However, since 

porosity and permeability of coal matrices are extremely low, almost all of the water resides in 

fractures and cleats in the coalbed. Thus, any water saturation for coalbeds represents the fraction 

of cleat porosity that is occupied by water. The definition of saturation is based on the total pore 

volume of cleats. A fraction of the pore volume may be occupied by immobile water, where 

water saturation cannot be reduced further. Thus, water saturation in cleats may vary between a 

minimum value (immobile water saturation) and a maximum of 100% saturation. 

Simulation work by Karacan [forthcoming] showed that reducing irreducible water 

saturation to 10% for a development distance of 610 m (2,000 ft) reduced the methane inflow 

rate. Increasing initial water saturation, while keeping other parameters constant at their base 

values, had a more significant effect on methane inflow rates. When mining 610- and 1,830-m 

(2,000- and 6,000-ft) entries, increasing initial water saturation decreased methane inflow rate 

due to decreased initial gas saturations and gas relative permeabilities. A similar decrease was 

observed in the longer entries when initial water saturation was changed from 40% to 80%. 

It is critical to the successful reduction of underground emission 

rates to remove water from a coalbed in the process of producing coalbed methane from a borehole. The initial water saturation 

of a coalbed has a more significant effect on methane inflow 

rates in the mining of shorter entries (about 1,830 m (6,000 ft) 

or less). 

When mining 3,050-m and 3,660-m (10,000-ft and 12,000-ft) long gate roads, gas inflow 

rate increased when water saturation was increased from 80% to 100%. This may be related to 

capturing negative decline and the peak rates of methane from the coalbed during the dewatering 

phase. When the coalbed is partially saturated with water, both water and methane are produced 

concurrently, and inflow of methane from cleats into the entries may show a declining trend with 

time. When mining begins in a coalbed that is 100% saturated with water, water will initially 

flow into the entries rapidly and trigger a high gas inflow rate as in the “negative decline” period 

of production wells, thus increasing average inflow rate into the entries. Shorter entry developments may not experience the effect of this period, and inflow rate may decrease due to water 

saturation. However, mining longer entries gives sufficient time and longer coalbed exposure for 

the high gas inflow rate experienced during negative decline to impact average methane inflow 

[Karacan, forthcoming]. 

Karacan [forthcoming] used multiple regression to assess the relationships between 

methane inflow rates and various coalbed reservoir parameters (Table 3). Methane inflow rates 

were calculated for 610-to 3,660-m (2,000- to 12,000-ft) gate road lengths assuming a 15-m/day 

(50-ft/day) section advance rate. The units of coalbed parameters are presented in Table 1. 

These analyses revealed that sorption time constant, butt-cleat permeability, and Langmuir 

pressure were found to be the most influential on methane inflow rates for all gate road lengths. 

For shorter gate road lengths, initial water saturation was also found to impact methane inflow 

rates [Karacan, forthcoming]. 

18
 
 

Table 3.—Linear models for predicting methane inflow rates into development entries 

of various lengths using key coalbed reservoir parameters 

Entry 

length (ft) Methane inflow rate (scf/day) = R2

 2,000 −9025.0349 + (18676.8392 × Thick) + (1367.1790 × 

Pres) + (−581.1950 × S.Time) + (13100.6770 × Ky) + 

(121.6918 × L.Vol.) + (−197.1251 × L.Pres.) + 

(−1280.3404 × W.Sat.) 

0.9209

 6,000 −193021.37 + (49697.1335 × Thick) + (2734.4987 × 

Pres) + (−997.3667 × S.Time) + (19103.2963 × Ky) + 

(298.3296 × L.Vol.) + (−544.3628 × L.Pres.) 

0.9011 

10,000 −256242.06 + (73838.2649 × Thick) + (3764.7566 × 

Pres) + (−1256.0067 × S.Time) + (21052.8924 × Ky) + 

(457.1441 × L.Vol.) + (−867.7456 × L.Pres.) 

0.8896 

12,000 −283185.59 + (85055.0763 × Thick) + (4219.7519 × 

Pres) + (−1357.3077 × S.Time) + (21403.8583 × Ky) + 

(532.4142 × L.Vol.) + (−1023.3603 × L.Pres.) 

0.8844 

Thick = Thickness; Pres = Pressure; S.Time = Sorption time constant; Ky = Butt 

cleat permeability in this study, but can be interpreted as the cleat permeability 

perpendicular to entries; L.Vol. = Langmuir volume; L.Pres. = Langmuir pressure; 

W.Sat. = Water saturation. 

NOTE.—Regression was performed using U.S. customary units of measure. 

Summary 

•	 The coalbed methane inflow rate increases with the length of the development section 

and is strongly influenced by the mining rate. 

•	 Shielding degasification boreholes decrease the methane inflow rate in the ventilation air 

compared to unshielded entries, even if operational only during the mining cycle. 

•	 Shielding wells located close to gate road entries and operated for longer times are the 

most effective method application of this methane drainage technology. 

•	 Coalbed methane emission rates are strongly influenced by methane content. The gas 

content of a coal at a range of coalbed pressures is controlled by the Langmuir isotherm. 

•	 Methane emission rates are an almost linear function of pressure and the distance to be 

mined. 

•	 The permeability of cleats perpendicular to the mine entries is more important in controlling methane inflow rates than the maximum face cleat permeability. 

•	 The effect of the sorption time constant on methane inflow rates into mine workings is 

more significant for longer entries than for shorter ones. 

19
 
•	 The initial water saturation of a coalbed has a more significant effect on methane inflow 

rates in the mining of shorter entries. 

•	 The coalbed parameters that most strongly influence methane emission rates for short 

entries under 610 m (2,000 ft) are, in decreasing importance: 

1.	 Reservoir pressure 

2.	 Sorption time constant 

3.	 Permeability 

4.	 Langmuir pressure 

5.	 Langmuir volume 

6.	 Thickness 

7.	 Water saturation 

•	 For long entries (3,660 m (12,000 ft)), the most important parameters are: 

1.	 Reservoir pressure 

2.	 Sorption time constant 

3.	 Langmuir pressure 

4.	 Thickness 

5.	 Permeability 

6.	 Langmuir volume 

20
 
2. CONTROLLING LONGWALL FACE METHANE 

AND DEVELOPMENT MINING EMISSIONS: 

PREDICTED IMPROVEMENTS USING IN-SEAM BOREHOLES
 

Overview 

This section focuses on the optimization of methane drainage technology aimed at 

reducing and controlling longwall face emission rates and concentrations. Reservoir modeling 

software has been used to assess methane drainage methods, which are all based on the application of in-seam boreholes that are rotary or directionally drilled. The boreholes are drilled ahead 

of mining, and coalbed methane production is terminated prior to being mined through by the 

longwall face. The modeling results provide a comparison for the production and emissions 

reductions that can be expected from different borehole configurations over a range of gas 

production durations. 

As mines progress into deeper and gassier coalbeds, or as longwall panel size increases, 

ventilation and GGVs together may not be sufficient to maintain methane levels within statutory 

limits. To decrease the explosion risk associated with methane emissions under these circumstances, in-seam horizontal methane drainage is often used to reduce the gas content of the 

coalbed prior to mining. Degasification of the coalbed prior to mining is effective in reducing 

face emission rates and volumes (emissions from the freshly exposed face of the coalbed, broken 

coal pieces during mining, and the coal load on the conveyor belts). Potentially, if enough time 

elapses after gate road development and prior to longwall mining, gas emissions into the 

development entries from the panel may help reduce future gas emissions during panel extraction 

[Diamond and Garcia 1999]. However, there may not always be enough time for the coalbed to 

drain before the longwall starts production. This is especially true since longwall productivity 

has increased considerably in recent years because of advances in mining methods and 

machinery. Aul and Ray [1991] reported that longwall productivity increased by 200%–400% 

between 1983 and 1990; the accompanying methane emissions increased by as much as 200%– 

300% in several mines operating at depths ranging from 370 to 730 m (1,200 to 2,400 ft) in the 

Pocahontas No. 3 Coalbed in Virginia. 

To investigate these issues, a reservoir model was developed and calibrated for a 

longwall mining area in the Pittsburgh Coalbed in Greene County, PA. The study area at the 

mine was located in a new mining district where panels were initially 381 m (1,250 ft) wide and 

were increased to 442 m (1,450 ft) starting with the third panel. The study site is described in 

greater detail by Karacan et al. [2007a]. 

A methane drainage program using in-seam horizontal boreholes in the Pittsburgh 

Coalbed is used at the study site to shield the gate roads during development and to reduce the 

in-place methane content of the outlined longwall panel. As the gate roads advance, two to three 

sets of two horizontal boreholes are generally drilled from what will eventually be the tailgate 

entries of each panel. The number of sets depends generally on the length of the panels and the 

length of the individual boreholes. The holes are drilled toward what will be the startup end of 

each panel, with one hole paralleling the tailgate side of the panel. The individual horizontal 

boreholes are connected to a common underground pipeline for transmission to the surface 

[Karacan et al. 2007a]. 

Five different horizontal, in-seam, methane drainage borehole patterns were evaluated in 

this study (Figure 9, A through E). The gate road entries shown represent a three-entry system 

21
 
 

with intervening coal pillars. The boreholes in patterns C and D were spaced equally. In all of the 

patterns, each borehole was modeled as a 7.6-cm (3-in) diameter, unstimulated well drilled from 

the tailgate entries into the Pittsburgh Coalbed. The total lengths of the simulated boreholes were 

5,610 m, 8,480 m, 1,160 m, 2,320 m, and 1,840 m (18,400 ft, 27,800 ft, 3,800 ft, 7,600 ft, and 

6,050 ft) for patterns A, B, C, D, and E, respectively. The boreholes operated at bottom-hole 

pressures equal to atmospheric pressure to represent the absence of any exhausters used to aid 

gas flow. To evaluate gas production potential during mining, the 3- and 12-month premining 

methane drainage models for each borehole pattern were extended by 268 days to include the 

production of gas during extraction of the longwall panel [Karacan et al. 2007a]. 

Figure 9.—Horizontal methane drainage borehole patterns modeled for degasification of the longwall panel 

(not to scale). 

The horizontal borehole production data were evaluated to estimate the potential average 

reductions in face emission rates as a result of premining methane drainage. The calculated 

average face emission reductions were also determined. For this calculation, cumulative methane 

production was divided by the duration (268 days) of mining for the study panel. The assumption 

was that either the majority (75%) or the total (100%) of the methane produced would be 

released as face emissions during mining if it was not removed prior to mining. This assumption 

is based on the work of Noack [1998], who proposed that, in the absence of empirical data, 

the degree of gas emission (percentage of gas-in-place) could be assumed to be 100% in 

stratigraphic zones within 20 to −11 m (66 to –36 ft) of the top and bottom of the mined coalbed, 

and 75% in the mined coalbed. The calculated average face emission rates that can be expected if 

methane drainage is not utilized are 27 m3

/min (940 cfm) for the 100% gas emission basis and 

20 m3/min (710 cfm) for the 75% emission basis. These predictions are based on the calculated 

total in-place methane content of 10 × 106

 m3 (360 MMscf) in the outlined panel and 268 days of 

longwall mining. 

Using horizontal borehole pattern B in the Pittsburgh Coalbed reduces the average face 

emission rates by as much as 10 m3/min (360 cfm) for the 100% emission basis and by 

22
 
 

7.6 m3/min (270 cfm) for the 75% basis after 12 months of premining degasification. The 

methane emission reduction values for pattern A are as much as 6.8 m3

/min (240 cfm) for the 

100% basis and 5.1 m3/min (180 cfm) for the 75% basis. The possible face emission reduction 

values calculated for D are 6.2 and 4.5 m3

/min (220 and 160 cfm) for the 100% and 75% basis 

respectively, after 12 months of degasification. Lesser amounts of face emission reductions are 

calculated for patterns C and E. Figure 10 compares cumulative methane productions, different 

horizontal borehole patterns, and the reductions in face emission rates for the 100% degree of gas 

emission basis. 

The overall reduction in average face emission rates due to methane drainage can be 

evaluated by analyzing the panel extraction phase and combining it with premining methane 

drainage performance. After 12 months of premining degasification, the additional reductions in 

face emissions with continued degasification during mining are 0.9 m3

/min and 1.2 m3/min 

(32 and 42 cfm) for the 100% basis with patterns A and B, respectively. The additional face 

emission reduction that can be achieved with patterns D and E due to continued degasification is 

similar to that of A (0.8 and 0.7 m3

/min (28 and 25 cfm), respectively), and the reduction is less 

with C after 12 months of premining degasification. 

The data in Table 4 show that after only 3 months of premining degasification, continued 

degasification during panel extraction becomes more important. For instance, after 3 months of 

premining degasification, continued methane drainage during mining using borehole pattern A 

reduces face emission rates by 1.8 m3/min (64 cfm) for the 100% degree of emission basis and 

1.4 m3

/min (49 cfm) for the 75% basis. These reductions are more than 50% of what can be 

achieved in addition to 3 months of premining degasification. Similar impacts are achieved with 

other borehole patterns as well as a result of continued degasification during panel extraction 

after a short (3-month) premining degasification period. 

After 3 months of premining degasification, additional drainage during mining using 

borehole pattern B reduces face emission rates by 2.7 m3/min (95 cfm) for the 100% emission 

basis and 2.1 m3/min (74 cfm) for the 75% basis. These data (Table 4) show that additional 

degasification after a 3-month premining production period with pattern B is almost as effective 

as pattern A. 

Table 4.CSimulated reductions in face emission rates after different phases of methane drainage 

using horizontal boreholes from the longwall panel 

Premining 

degasification, 

months 

Wellbore pattern1 

A B C D E 

100% 

emission, 

cfm 

75% 

emission, 

cfm 

100% 

emission, 

cfm 

75% 

emission, 

cfm 

100% 

emission, 

cfm 

75% 

emission, 

cfm 

100% 

emission, 

cfm 

75% 

emission, 

cfm 

100% 

emission, 

cfm 

75% 

emission, 

cfm

Premining 3 100 78 160 120 49 35 92 67 71 53 

degasification 12 240 180 360 270 110 85 220 160 180 130 

Degasification 3 64 49 95 74 25 18 57 42 49 35 

during mining 12 32 21 42 32 14 11 28 21 25 18 

Total emissions 

reduction 

3 164 130 255 190 74 53 149 110 120 92 

12 272 200 402 310 124 95 248 180 205 150 1

From Figure 9. 

23 
 Figure 10.—Cumulative methane production and possible reductions in longwall face emission rates for 

different premining methane drainage time intervals and different horizontal borehole patterns (A through E). 

The total reductions in average face emissions after 12 months of methane drainage 

before and during mining are significant, particularly with horizontal borehole patterns A, B, 

and D. The total reduction that can be achieved with pattern B is between 8.8 and 12 m3/min 

(310 and 410 cfm). This reduction can decrease the average maximum emission rate that may 

be expected from the Pittsburgh Coalbed from 20–27 m3/min (710–940 cfm) to less than 

11–15 m3/min (400–540 cfm). 

Longer degasification times result in lower face emission rates. 

For the Pittsburgh Coalbed, the most effective borehole pattern 

is the trilateral configuration (pattern B), followed by the duallateral arrangement (pattern A). High emission reductions can 

also be achieved or exceeded by a large number of short, crosspanel boreholes drilled in a closely spaced, linear manner. 

During simulation of longwall degasification, GGVs were put into production as the 

longwall face advanced to their location [Karacan et al. 2007a]. Figure 11 shows methane 

production from the simulated horizontal methane drainage borehole patterns shown in Figure 9 

before and during panel extraction. The results show that methane production during panel 

extraction is substantially less than that of the premining phase. Additionally, as the premining 

degasification time increases from 3 to 12 months, the amount of methane produced by the 

horizontal boreholes during panel extraction decreases even more. The reduced methane 

production is probably due to a combination of two factors. First, the methane content of the 

coalbed is reduced during the premining methane drainage phase, leaving less methane available 

for production at a slower rate during mining, especially after a longer premining degasification. 

Second, the methane production from each horizontal borehole is progressively terminated as the 

longwall face reaches its location [Karacan et al. 2007a]. The simulation results shown in 

24
 
 

 

Figure 11 indicate that the highest cumulative methane production is achieved with long horizontal borehole patterns A and B. However, as the number of shorter, cross-panel, horizontal 

boreholes perpendicular to the higher-permeability face cleats increases, as in pattern D, their 

cumulative production approaches that of pattern A. 

The period of premining degasification should be extended for 

as long as possible. Methane production during panel extraction 

is substantially less than that of the premining phase due to the 

decreased methane content of the coalbed and the successive 

termination of the boreholes during mining. 

4 

5 

6 

m 3

) ne (x10

-200 

-150 

1 

2 

3

Cumulative Metha 6

-100 

- 50 

0 

A - E 

Before Mining 

A - E 

During Mining 

A - E 

Cumulative 

-0 

Solid Fill Colors: 3-months Pre-mine.Degas. 

Patterned Fill Colors: 12-months Pre-mine.Degas. 

1 x 103 ft 3 

Figure 11.—Performances of different horizontal methane drainage borehole patterns shown in Figure 9 

during different phases of degasification, after 3 and 12 months of premining degasification. 

Effects of Methane Drainage on Longwall Face Emissions 

The overall reduction in average face emission rates due to methane drainage can be 

evaluated by analyzing the panel extraction phase and combining it with premining methane 

drainage performance. The average reductions in longwall face emission rates due to methane 

drainage before and during mining were calculated and are shown in Figure 11 and Table 4. The 

data show that longwall face emissions were reduced by 11 m3

/min (400 cfm) with 12 months of 

premining methane drainage using borehole pattern B. As would be expected, the simulations 

show that 3 months of premining methane drainage is less effective in reducing the total 

emission rate than 12 months of premining drainage. 

25
 
After 12 months of premining degasification, the additional reductions in face emissions 

with continued degasification during mining are 0.9 m3

/min and 1.2 m3/min (32 and 42 cfm) for 

the 100% basis with patterns A and B, respectively. The data in Table 4 show that after only 

3 months of premining degasification, continued degasification during panel extraction becomes 

more important. 

The data presented in Table 4 show that after 3 months of premining degasification, 

continued methane drainage during mining using borehole pattern A reduces face emission rates 

by 1.8 m3/min (64 cfm) for the 100% degree of emission basis and 1.4 m3/min (49 cfm) for the 

75% basis. These reductions are more than 50% of what can be achieved in addition to 3 months 

of premining degasification. Similar impacts are achieved with other borehole patterns as well as 

a result of continued degasification during panel extraction after a short (3-month) premining 

degasification period. On the other hand, the incremental reduction obtained with continued 

degasification during panel extraction in average face emission rates after 12 months of premining degasification using the same borehole pattern is about 15% of what can be achieved by 

12 months premining degasification. Again, similar reductions in average face emissions were 

observed for the other borehole patterns after continued degasification. 

It is important to continue methane drainage during mining to 

maximize reductions in longwall face methane emissions. 

Methane borehole production during panel extraction and its 

resultant effect on face emissions is substantially less than 

during the premining phase. However, if the duration of premining methane drainage is short, the contribution of methane 

produced during panel extraction and the subsequent face emission reduction may be significant. 

Effectiveness of Shielding the Development Entries 

From Methane Migration and Emissions 

In-seam methane drainage boreholes not only degas the longwall panel itself, but can also 

shield the advancing development entries from methane emissions from the surrounding virgin 

coalbed gas reservoir [Diamond 1994]. To study this phenomenon, the impacts of borehole 

pattern on the effectiveness of shielding gate roads from methane migration were evaluated. In 

this analysis, two approaches were used. In the first approach, methane emissions were predicted 

in conjunction with operating in-seam horizontal methane drainage boreholes, configured as 

shown in Figure 9. In the second approach, three horizontal boreholes were placed in the virgin 

coalbed along the gate roads on both sides of the outlined panel, as shown in Figure 12, for 

pattern A+ (applicable only to the first longwall panel in a new mining district) [Karacan et al. 

2007a]. The first approach is intended to analyze the amount of methane migrating into the 

gate roads from the panel area and the surrounding coalbed only in the presence of in-panel 

horizontal methane drainage boreholes. The second approach analyzes the effects of additional 

boreholes on shielding the entries from methane emissions from the surrounding coalbed during 

premining degasification. 

Table 5 gives the predicted methane inflow into the ventilation system from the 

surrounding virgin coalbed at the end of various methane drainage time intervals based on the 

26
 
 

presence of in-panel horizontal methane drainage boreholes (A+ to E+), along with methane 

inflow into the ventilation in the absence of any degasification borehole. These comparative 

results show that borehole patterns A and B, where near-margin horizontal boreholes are drilled 

from the tailgate, are more effective against methane inflow into the gate roads during premining 

degasification than the other patterns because of their extended length along the entries. The 

simulation indicates that by using either pattern A+ or B+, the potential methane inflow into the 

entries can be reduced by as much as 62 × 104

 m3 (22 × 106 ft3) over 12 months, which 

corresponds to a decrease in methane inflow by ~1.1 m3/min (~40 cfm) simply by shielding 

against methane inflow originating from the coalbed within the panel area. As would be 

expected, this suggests that the middle horizontal borehole segments in pattern B+ do not 

contribute much to gate road shielding. 

The results of a borehole drilling study for shielding the entries from methane inflow 

were reported by DuBois et al. [2006]. This research was conducted in the Pittsburgh Coalbed at 

the same mining district modeled in this study. The horizontal boreholes of pattern A were 

drilled to maximize the shielding for both belt and return entries during headgate and tailgate 

development. This approach resulted in a pattern similar to the one designated as A+ in this 

study. The drilling strategy described by DuBois et al. permitted horizontal boreholes to be 

active for 6–24 months prior to any mining. They reported that methane concentration decreased 

by 41%, and methane emission into the entries decreased between 30% and 35%, close to the 

predicted values in this study for patterns A+ and B+. 

The emissions into the gate road entries originate mostly from 

the gas desorbing from the margins of the panel where the 

in-seam, near-margin borehole segments are located and where 

these boreholes are most effective in capturing this gas. 

Figure 12.—Sample layout showing additional horizontal, in-seam boreholes on either side of the 

gate roads (A+) in addition to pattern A (not to scale). 

Additional reductions in methane emissions into the gate roads 

can be achieved by employing horizontal methane drainage 

boreholes in the surrounding virgin coalbed gas reservoir on 

either side of the gate roads. Methane migration into the mine 

ventilation airflow in the gate road entries surrounding the 

outlined longwall panel can be reduced effectively by using dual 

and trilateral horizontal methane drainage holes paralleling the 

gate roads. 

27
 
 

 

 

 

Table 5.—Predicted methane emissions into the ventilation system from the 

surrounding coalbed (within the panel and virgin coal) 

Wellbore pattern 

Methane inflow into mine ventilation system (×106

 ft3) 

for various durations of methane drainage 

3 months 6 months 9 months 12 months 

No degasification 78.8 126 159 186 

A (A+) 74.5 (66.7) 116 (99.6) 144 (119) 164 (132) 

B (B+) 74.5 (66.7) 116 (99.6) 144 (119) 164 (132) 

C (C+) 78.0 (72.0) 124 (108) 156 (132) 181 (150) 

D (D+) 77.3 (69.6) 123 (107) 155 (131) 179 (147) 

E (E+) 78.0 (70.3) 124 (107) 156 (131) 178 (147) 

NOTE.—“+” indicates drainage boreholes placed on both sides of the gate roads. 

Summary 

•	 To reduce longwall face emissions from the Pittsburgh Coalbed, the most effective 

borehole pattern is the trilateral configuration (pattern B), followed by the dual-lateral 

arrangement (pattern A). 

•	 The greatest reduction in methane emissions can be achieved by extending the duration 

of premining degasification for as long as possible. 

•	 Additional reductions in longwall face methane emissions can be realized by continuing 

methane drainage during the mining phase. 

•	 The emissions into the gate road entries originate mostly from gas desorbing from the 

margins of the panel where in-seam, near-margin borehole segments are located. 

•	 Further reductions in methane emissions into the gate roads can be achieved with 

horizontal methane drainage boreholes on either side of the gate roads, although this 

method is only applicable to the first panel in a new set. Data for this study are based on 

the Pittsburgh Coalbed. A more permeable seam may not benefit significantly from the 

presence of boreholes on either side of the gate road. 

28
 
3. CHARACTERIZING AND FORECASTING LONGWALL FACE 

METHANE EMISSION RATES FOR LONGER LONGWALL FACES 

This section discusses one of two NIOSH-developed methods for predicting methane 

emission rates when longwall face lengths increase. The first method by Schatzel et al. [2006] is 

an adaptation of an earlier NIOSH technique that essentially extrapolates emission rates to longer 

faces using field monitoring data. The strength of this methodology is in the computation of 

methane emission rates for each face segment length being mined. A total of eight methane 

emission rates are shown in the case study. Changing methane emission rates over the length of 

the longwall face, as well as emissions rate variations over subsequent days, are clearly visible 

and the causes of these changes are discussed. A second method by Krog et al. [2006] is discussed in Section 4 and is probably better suited to meet the needs of most coal mine operators 

for forecasting methane emission rates to wider longwall faces. Both methods describe results 

using the same data set. The two methods produce complementary views of longwall face 

methane emissions at an operating mine. 

Overview 

Continuous enhancements in longwall mining equipment have significantly improved 

face advance rates. This increasing longwall advance rate has generally outpaced the continuous 

miner gate road advance rates, which had led operators to decrease the relative amount of 

development mining required by increasing longwall panel sizes, in particular face lengths. The 

mining of longer longwall faces has the advantage of less development per ton of coal mined on 

the longwall, but can result in increased methane emissions. Increases in longwall face length 

can create problems such as increased cumulative face methane emissions and increased potential for methane-related production delays [Krog et al. 2006]. This may be further exacerbated by 

the complex airflow movements along the face itself. Although airflow movement along the 

longwall face is generally assumed to be linear, evidence of exchanges between face and gob 

atmospheres have been noted [Balusu et al. 2001; Wendt and Balusu 2001]. Other researchers 

have suggested that an increased level of opportunity for air exchanges between face and gob 

regions result from resistant roof units and associated void spaces behind the longwall face 

[Noack 1998; Balusu et al. 2001; Wendt and Balusu 2001]. 

The questions asked by ventilation engineers are: how much of an increase in methane 

emissions can be expected with the longer longwall face, and how can this increase be mitigated 

to maintain a safe underground workplace? An increase in the ventilation airflow to dilute the 

expected increase in methane emissions might not be possible because many modern longwalls 

are at, or near, their reasonably practical airflow limits. Therefore, the extra methane emissions 

will generally have to be handled by a combination of increased ventilation airflow, methane 

drainage, and production management. 

Longwall Face Emission Monitoring and Prediction Method 

A longwall mine operating in the Pittsburgh Coalbed in southwestern Pennsylvania was 

studied. The dimensions of the outlined coal block were about 3,250 m (10,700 ft) long by about 

315 m (1,030 ft) wide (Figure 13). The study began with 1,250 m (4,100 ft) of panel length 

remaining. The longwall bleeder system was ventilated by the use of a centrifugal bleeder fan 

29
 
Intake Air Shaft

Intake/Return 

Shaft Fan

Bleeder 

Shaft Fan

and four vertical gob ventilation boreholes per panel placed at regularly spaced intervals of 

610 m (2,000 ft) to provide additional methane control capacity. Four horizontal in-seam 

methane drainage boreholes were also present adjacent to the gate roads of the study panel 

[Schatzel et al. 2006]. 

Intake Air Shaft 

Intake/Return 

Shaft Fan 

Bleeder 

Shaft Fan 

Figure 13.—Longwall face emission study area. 

The original methodology for this study was first developed by Diamond and Garcia 

[1999] to predict methane emission rates for 305-m (1,000-ft) faces based on face emissions 

monitoring on a 229-m (750-ft) wide face at two adjacent mines operating in the Pocahontas 

No. 3 Coalbed in Virginia. The face was divided into three equal segments of 76.2 m (250 ft). 

Average cumulative methane emissions data for each of the three segments were plotted as a 

function of face length. Curves were fit to the actual emission data and then extrapolated to the 

305-m (1,000-ft) face widths to predict methane emission rates on the longer faces. 

For this study in the Pittsburgh Coalbed, continuously recording methane monitors were 

installed along the longwall face. Methane concentrations were recorded by data loggers. 

Airflow measurements were made at various locations on the face. One shift was monitored on 

each of the 3 days of the study. A production time study consisting of shearer location on the 

face (recorded as shield numbers) and shearer mining direction (head-to-tail (H–T) or tail-tohead (T–H)) was conducted throughout the 3 days of the face emissions monitoring [Schatzel 

et al. 2006; Krog et al. 2006]. In applying this methodology, the following assumptions were 

made to project methane emissions to longer faces: (1) the mine advance rate and the frequency 

of mining delays will occur at about the same rate, (2) face methane emissions are assumed to be 

constant within each segment, (3) all sources of methane emissions change at a constant rate with 

increased face length, and (4) the solutions are site-specific for the Pittsburgh Coalbed, the 

30
 
ventilation system configuration, and the methane drainage systems applied at the study 

mine site. 

To analyze the movement of methane emissions in the longwall ventilation airflow, the 

face was divided into four segments of equal length (Figure 14). Methane emission rates were 

determined for each face segment of each pass of the shearer using the associated methane 

concentration, ventilation airflow, and time study data. Since the principal focus of this study 

was longwall face ventilation and methane emissions, the decision was made to position the 

monitors away from the headgate and tailgate corners to avoid including ventilation air in these 

areas, which did not traverse the face. The methane sensors were installed at shields 20, 80, and 

145 (see “Instrument locations” in Figure 14). On the study panel, belt air was used on the 

longwall face (Figure 15). 

Longwall face methane emissions data can be used as the basis 

for predicting methane emissions on longer longwall faces. An 

accounting of underground activities near the face is necessary 

for data analysis. In the methodology described, methane 

monitors are positioned away from the headgate and tailgate 

corners to address methane in air moving along the face, not 

ventilation air losses at the headgate corner or gob gas leakage. 

Characterizing Longwall Face Emissions and Extrapolating Emissions 

to Longer Faces 

A total of 27 methane-related delays were noted during the study. Methane delays were 

typically of short duration, averaging about 7 min. The belt line made up the great majority of 

the methane measured at the shield 20 methanometer. Consequently, for analytical purposes, 

face emissions were assumed to be zero for the first 20 shields of an H–T pass (up to 34 m 

(114 ft) from the headgate corner) (Figure 15). Methane emissions in this portion of the face 

were quite low, usually measured as between 0.0 and 0.021 m3/s (0 and 44 cfm) [Schatzel et al. 

2006]. 

31
 
 

 

 Tailgate Headgate 

Gob 

S 

Instrument 

locations 

290 

(953) 

157 

(516) 

35 

(114) 

315 

(1032) 

236 

(774) 

157 

(516) 

79 

(258) 

0 

Face segment endpoints, distances from headgate 

corner, m (ft) 

4 3 2 1 (H-T) 

(T-H)1a 2a 3a 4a 

Face segment identities 

Figure 14.—Pittsburgh Coalbed longwall face emission study site with instrument locations and 

face segment identities (not to scale). 

Fresh Air 

Return/Bleeder Air 

Face Air 

Mining Direction 

Mining Direction 

Tailgate 

Headgate 

Gob 

S 

Belt entry 

Figure 15.—Ventilation configuration at the study site (not to scale). 

32
 
 

The sums of gas emissions from the eight face segments for the H–T and T–H passes 

were normalized to match actual gas emission totals measured at the shield 145 methane emissions monitor. The cumulative methane emission rates for the face segments for each of the 

3 days of the study are shown in Table 6. 

Face methane emission rates were consistently higher by about 

41% in the H–T passes than the T–H passes. This difference in 

methane emissions rate was caused by the longwall face 

configuration, which produced a much longer “wedge” cut 

toward the tailgate side of the face; a slower shearer cutting 

rate in the T–H passes; and the pan line direction of movement 

toward the headgate. 

One methane drainage borehole was drilled on each side of the panel. The boreholes were 

oriented parallel to the gate road entries, 30 m (98 ft) from the respective gate roads (Figure 16). 

The holes were drilled toward the advancing face, terminating in the study area. Prior to interception by mining, the holes were filled with water so that coalbed gas could not accumulate in 

the borehole. 

Production delays due to increasing methane concentration were most common when the 

shearer was near the tailgate side of the longwall face of the H–T pass, but began occurring with 

increasing frequency at locations nearer the headgate side of the face as the study progressed 

over the 3 days [Schatzel et al. 2006]. 

Figure 16.—Actual portions of the study panel monitored for face methane emissions and associated 

methane drainage borehole locations. 

33
 
 

 

 

 

Table 6.—Face segment emission rate data shown for combined 

averaged H–T, T–H passes 

Face 

segments 

Day 1 

m3

/s cfm 

Day 2 

m3/s cfm 

Day 3 

m3/s cfm 

1, 4a 

2, 3a 

3, 2a 

4, 1a 

0.003 

0.009 

0.031 

0.020 

7 

18 

66 

42 

0.010 

0.006 

0.018 

0.008 

20 

13 

37 

18 

0.023 

0.029 

0.011 

0.016 

49 

61 

24 

33 

In many mines, the quantity of coalbed methane removed by a 

gob gas venthole is significant, potentially 75% of the volume of 

gas emissions on the longwall face. Pressure differentials are 

also important in defining the flow paths for gas movement. 

A nearby active gob gas venthole is much more influential than a 

distant one. 

The variation in emissions behavior can be seen more clearly in Table 6, which shows the 

average methane emissions from each segment (H–T and T–H combined data) during the 3 days 

of monitoring (Figure 17). Methane emissions produced during passes on day 1 were relatively 

low (from 0.003 to 0.009 m3

/s (7 to 18 cfm)) in the segments nearest to the headgate (1, 2, 3a, 

4a). Emissions increased significantly in segments 3 and 2a (0.031 m3/s (66 cfm)) and diminished somewhat from that level toward the tailgate in segments 4 and 1a. On day 2, methane 

emission rates in the near-headgate segments 1 and 4a were higher than those on day 1, then 

decreased in the next segments, 2 and 3a. Similar to day 1, methane emissions increased in face 

segments 3 and 2a (0.018 m3

/s (37 cfm)), then decreased in the face segments toward the tailgate, i.e., face segments 4 and 1a. On day 3, methane emissions were much higher (0.023 m3

/s 

(49 cfm)) near the headgate face segments 1 and 4a than during the prior 2 days of the study 

(Table 6). The methane emissions rate increased in the next face segments (0.029 m3/s (61 cfm)) 

2 and 3a, but then decreased significantly in face segments 3 and 2a (0.011 m3

/s (24 cfm)), with 

a slight increase (0.023 m3

/s (49 cfm)) in the tailgate segment face segments 4 and 1a (Figure 17) 

[Schatzel et al. 2006]. 

Methane delays were most frequent when the shearer was 

mining near the tailgate. Methane delays on the longwall face 

increased from the first to third day of the study as the longwall 

face reached the approximate maximum distance from the 

nearest operating gob gas venthole prior to interception of the 

next venthole, 2 days after the study was completed. 

34
 
 

Figure 17.—Average methane emission rates for each day of study determined for each face segment. 

On day 2 of the study, the presence of a horizontal borehole on the face was noted near 

the tailgate side of the panel at shield 140. On day 3 of the study, the interception of the horizontal borehole near the headgate at shield 23 was observed. It is likely that coalbed methane 

production from the first borehole encountered near shield 140 on the tailgate side of the face 

had decreased the methane content of the coalbed in the vicinity of the borehole. Consequently, 

after the borehole was intercepted on day 2 of the study, methane emissions dropped toward the 

tailgate side, as shown by the comparison of face segments 4 and 1a on day 1 with those on day 

2 (Figure 17). A similar drop in methane emission rates was not observed in face segments near 

the headgate from day 2 and day 3 when the borehole was intercepted near shield 23 (Figure 16). 

It is not known if methane emissions would have diminished in the region of face segments 1 

and 4a if the study had been continued for another day. 

Methane drainage via horizontal boreholes near the study area 

seems to be an effective method for reducing longwall face 

emissions. The variable methane emission characteristics 

observed on the longwall face were likely produced by emission 

rate reductions associated with methane drainage boreholes. 

Production delays also affected face methane emission rates. These delays, including 

those due to high methane emissions, result in lower calculated pass segment methane emission 

rates because the time to complete the pass segment increases while the longwall face equipment 

is idle. Therefore, the increased number of methane-related production delays on day 3, and to a 

lesser extent on day 2, resulted in lower average methane emission rates on some individual pass 

35
 
segments, particularly on tailgate-side face segments 4 and 1a (Figure 17, Table 6). By day 3, 

methane-related production delays were occurring closer toward the headgate portion of the 

face, in face segments 3 and 2a, reducing methane emission rates during mining stoppages there 

as well [Schatzel et al. 2006]. 

Methane-related delays reduce emission rates during face 

methane monitoring when coal is not being mined. 

Graphs of the cumulative measured average methane emission rates for the ~315-m 

(~1,030-ft) panel face segments in both the H–T and T–H directions are shown in Figure 18. 

From these data, two least-squares linear regression curves were calculated to predict methane 

emissions for longer face lengths of 366, 427, and 488 m (1,200, 1,400, and 1,600 ft) in the 

Pittsburgh Coalbed. To create trend line A, all H–T and T–H passes were averaged, and then an 

overall average emission pass plot was created. Trend line A was fit to this overall average 

emission pass data (Figure 18). The trend line equations are given in metric units. It should be 

noted that the emission rate for a 315-m (1,030-ft) face is based on a projection of the cumulative 

emissions data from the face segments projected from shield 145 to shield 157, or 290 m (953 ft) 

from the headgate corner using data from all of the longwall passes. Trend line B was fit to the 

H–T passes only. Since most of the delays occurred and higher face methane emission rates 

occurred on the H–T passes, this plot may be more representative of problematic concentrations 

of face gas than trend line A, which includes the generally lower T–H pass face emission rate 

data (Figure 18). Using trend line B, the predicted face methane emissions represent increases of 

36%, 61%, and 88% for 366, 426, and 488 m (1,200, 1,400 and 1,600 ft) faces, respectively, 

compared to the base of about 315-m (1,030-ft) face using all face methane emission data. 

Since the H–T passes produced higher methane emission rates 

and more methane concentration delays than the T–H direction, 

the H–T passes were used to project emission rates to longer 

faces. In applying this technique, it is recommended that the 

operator use data from the pass direction, which produced the 

higher emission rates, to perform the least-squares projection to 

longer faces. 

36
 
Figure 18.—Methane emission prediction curves for the Pittsburgh Coalbed. 

Summary 

•	 Longwall face methane emissions monitoring data can be used as the basis for predicting 

methane emissions on longer longwall faces. 

•	 The quantity of coalbed methane removed by a GGV is significant, on the order of 70%– 

75% of gas emissions from the monitored longwall face. 

•	 The tailgate corner continues to challenge ventilation and methane control technology, 

and most of the methane delays occurred when the shearer was mining near the tailgate. 

•	 Longwall face emissions can be effectively reduced by in-seam, horizontal, methane 

drainage boreholes. Additional methane drainage techniques applied in advance of mining can also be effective, such as fracked vertical and directionally surface boreholes. 

•	 Face methane emission rates are diminished during methane-related delays and when 

coal is not being mined. 

•	 In applying this monitoring-based technique for predicting face emissions on longer 

longwall faces, it is recommended that the operator use data from the pass direction, 

which produced the higher emission rates. 

37
 
4. PREDICTING METHANE EMISSIONS FROM LONGER LONGWALL FACES 

BY ANALYSIS OF EMISSION CONTRIBUTORS 

This section describes the second of two methods of longwall face emission prediction 

techniques. NIOSH conducted a longwall methane emission and mining time study at a mine 

operating in the Pittsburgh Coalbed to assess the methane emission consequences of mining a 

longer face. Based on the results of the study, a set of site-specific mathematical formulas and 

constants were developed to characterize each of four longwall emission contributors. The 

mathematical formulas were then applied to longer longwall face mining scenarios to predict 

the methane emissions from these faces. 

Overview 

A detailed analysis of the methane sources and their individual contributions to the total 

longwall methane emissions can be obtained from methane concentration data collected at the 

beginning and end of the longwall face, along with the shearer location and other relevant 

ventilation and mining data. The methane emission contributors from the mining of a longwall 

face that were evaluated for this study were (1) gas released from the coal broken by the shearer, 

(2) gas emitted from the broken coal on the face conveyor, (3) gas emitted from the coal transported on the belt, and (4) background gas emitted from the coal face and adjoining ribs in the 

intake airway gate road entries. Once the methane contributions from the various sources have 

been defined for an actual longwall cutting sequence, the methane emissions from an ideal (i.e., 

delay-free) cut sequence can be predicted. The calculated methane emission contributions can 

then be extrapolated to longer longwall faces, taking into account the variations in coal production and transport factors, to more accurately predict future methane emission rates from longer 

longwall faces. 

Methodology 

This research used the same data set as that described in Section 3. The methane monitors 

were located at shields 20, 80, and 145 (see “Instrument locations” in Figure 14). These positions 

were chosen based on previous studies that indicated possible air interactions at the corners of 

the headgate and tailgate [Diamond and Garcia 1999]. The methane contribution interpretation 

does not take into account the possible effect of ventilation airflow interactions with the gob, 

which is consistent with the initial goal of the study [Krog et al. 2006]. 

The physical aspects, equipment, operational, and ventilation scenarios of the longwall 

panel (Figure 15) are very important in evaluating methane sources because they determine the 

mathematical formulas used in the methane contribution model. The site-specific variables for 

this study are discussed in Krog et al. [2006]. 

38
 
Assumptions 

The following assumptions were made: 

•	 Background methane emissions from the longwall face (active, not idle) are linearly 

dependent on longwall face length. 

•	 Background rib methane emission from the No. 1 belt entry are linearly dependent on the 

remaining length of the longwall panel, but are independent of longwall face width or 

activity. 

•	 There is no interaction between the face air and the air in the gob. 

•	 The stage loader located at the headgate was not incorporated in the simulation. 

•	 Methane liberation rate at the shearer is proportional to the cut coal volume. 

•	 Methane liberation rate on the face conveyor is proportional to the product of coal 

tonnage and elapsed time on the conveyor [Krog et al. 2006]. 

Construction of Formulas 

The following describes the mathematical process used in the evaluation. The longwall 

face was broken down into 61-m (200-ft) sections to correspond with the width of 30 shields. 

Since the face conveyor speed was 1.78 m/s (350 ft/min), the transport time for coal over 61 m 

(200 ft) of the face conveyor was 34 sec (61 m / 1.78 m/s), and the transport time for the 

counterflowing ventilation airflow was 24 sec (61 m / 2.54 m/s). Therefore, the time for the coal 

to be transported by the face conveyor and the time required for the counterflowing air to transit 

the same distance adds up to 58 sec, which was rounded to 1 min to match the methane concentration readings being recorded every minute. The emission times for methane from the sources 

of interest, as well as transit times for the associated ventilation airflows, can be summarized 

as follows: 

•	 Coal cut by shearer, 0–2 min 

•	 Coal on face conveyor, 1–3 min 

•	 Airflow along longwall face, 2 min 

•	 Coal on belt, 5 min 

•	 Airflow along belt, 10 min 

When the shearer is cutting coal, the liberation of methane from the coal face will be 

recorded by the methane monitor located near the tailgate at shield 145, either instantly or up to 

2 min later, depending on the shearer’s location. When the shearer is located on the tailgate side 

of shield 145, no extra methane liberated by the shearer will be recorded by the methane monitor. 

If the shearer is located within 61 m (200 ft) on the headgate side of shield 145, then methane 

liberation due to the cutting of the coal will be recorded for that minute. When the coal was cut 

nearer the headgate, it would take up to 2 min for the face air to travel to the methane sensor 

located at shield 145. 

The contribution of methane emissions from the coal on the face conveyor is more 

complicated than the above circumstances because coal cut by the shearer can be transported on 

the face conveyor for 1–3 min. To determine the transport time for methane emissions from the 

coal on the face conveyor, the counterflowing longwall ventilation airflow and the shearer 

39
 
location/direction of travel must be taken into account. For example, coal cut at shield 145 will 

release methane that will be recorded at shield 145 instantly (0 min). Coal transported on the face 

conveyor will release gas into the face ventilation airflow for a total of 3 min (0–2 min), during 

which time the counterflowing face ventilation will take up to 2 min to reach shield 145. 

Therefore, methane emitted from the coal on the face conveyor will be recorded at shield 145 

for 0–4 min in this example. 

If a methane delay is triggered by a methane monitor’s high 

methane reading, the underlying increase in methane emissions 

may have occurred several minutes earlier. To reduce methanerelated production delays, an understanding of methane 

emission sources and their magnitudes is required. 

Continuing this example, methane emissions from the coal on the belt will be recorded at 

shield 145 for 5–19 min after coal was initially cut at shield 145 (2 min for the face conveyor, 

2 min for the face airflow transit time, 1–5 min for the belt transport time, and 0–10 min for the 

belt airflow transit time). Therefore, coal mechanically cut by the shearer will affect methane 

emission levels near the tailgate instantly and for as long as 19 min after being cut [Krog et al. 

2006]. 

The contributions of methane from individual sources over time has led to a simple set of 

five linear equations for shearer, face conveyor, belt, background emissions from the coal face, 

and background emissions from the adjoining ribs in the intake gate roads, which were solved for 

each minute of the three shifts monitored. Constants in each equation were calculated by leastsquares linear regression such that the calculated results best matched the actual readings at 

shields 20 and 145. 

Results 

The calculated emission constants were consistent for the 3 days of the study, except for 

the background methane levels at shield 20. This value varied considerably over the 3 days of the 

study. The reason for the observed dramatic increase in this value is most likely the interception 

of a horizontal degas hole located upwind of the shield 20 methane sensor [Schatzel et al. 2006]. 

After considering the consistency of the emission constants for the other three components (face 

conveyor, belt, and shearer) from shields 20 and 145 individually and combined, it was decided 

to use the 3-day average for shields 20 and 145 for further evaluation of the background emission 

component [Krog et al. 2006]. 

Table 7 shows the calculated average methane emission rates for each contributor using 

the daily constants for shields 20 and 145. The “total” values in the table are the calculated 

average methane for the individual days. The “actual” values are the recorded methane for the 

individual days. 

40
 
 

 

 

 

 

 

 

Table 7.—Daily methane emission contributor averages and percentages 

using daily shield 20 and 145 constants 

Shield 20 and 1

Daily results 

45 Conveyor 

(m3

/s) 

Belt 

(m3/s) 

Shearer 

(m3/s) 

Face 

(m3/s) 

Rib 

(m3/s) 

Total 

(m3/s) 

Actual 

(m3/s) 

Day 1 0.012 0.015 0.005 0.027 −0.001 0.057 0.057 

Day 2 0.005 0.014 0.004 0.028 0.002 0.052 0.052 

Day 3 0.007 0.006 0.005 0.030 0.019 0.068 0.068 

Average 0.008 0.012 0.005 0.029 0.006 0.059 0.059 

Shield 20 and 145 

Daily results 

Day 1 

Day 2 

Day 3 

Conveyor 

(cfm) 

26 

11 

15 

Belt 

(cfm) 

31 

29 

14 

Shearer 

(cfm) 

10 

8 

11 

Face 

(cfm) 

57 

60 

64 

Rib 

(cfm) 

−3 

3 

40 

Total 

(cfm) 

121 

111 

143 

Actual 

(cfm) 

121 

111 

143 

Shield 20 and 145 

Daily results 

Conveyor 

(%) 

Belt 

(%) 

Shearer 

(%) 

Face 

(%) 

Rib 

(%) 

Total 

(%) 

Actual 

(%) 

Day 1 22% 26% 8% 47% −3% 100% 100% 

Day 2 10% 26% 7% 54% 3% 100% 100% 

Day 3 10% 9% 8% 45% 28% 100% 100% 

Average 14% 20% 8% 49% 9% 100% 100% 

Table 8.—Methane contribution percentages from longwall emission contributors 

during gas delays compared to daily averages for a 315-m (1,030-ft) longwall face 

Methane contributions 

3-day averages 

Conveyor 

17% 

Belt 

19% 

Shearer 

6% 

Background 

59% 

Gas delays No. 

Cutting gas delays 23 32% 18% 9% 42% 

H–T gas delays 19 33% 18% 9% 41% 

T–H gas delays 4 26% 18% 8% 48% 

The face and rib emissions represent about 59% of the total daily emissions; however, 

this percentage is somewhat misleading because these emissions continue throughout the entire 

shift, even during mining delays. In contrast, the shearer, face conveyor, and belt emissions are 

intermittent sources (i.e., they are only a factor during active mining on the face and during coal 

transport), but they are the primary contributors to longwall face gas delays, as shown in Table 8. 

The calculated methane contributions clearly show that the H–T passes experience higher 

methane concentrations than the T–H passes [Krog et al. 2006]. The time study data for the 

3 days of the study showed 19 gas delays (shutdowns of mining equipment due to excessive 

methane concentrations) on 10 of the 11 complete H–T passes. The average location for the 

gas delays on the H–T passes was shield 119, 245 m (804 ft) or 78% of the distance from the 

headgate corner. 

41
 
 

Monitoring data show that when longwall gas delays occur, the 

methane contributions from the face conveyor and shearer 

dramatically increase in their relative contribution to the total 

face emissions. Controlling these two emission sources is 

critical to maintaining statutory limits at the working face and 

avoiding methane emission-related coal production delays. 

There were four gas delays recorded on three of the nine T–H passes. The average location for the gas delays on the T–H passes was shield 71, 146 m (480 ft) or 46% of the distance 

down the longwall face from the headgate corner [Krog et al. 2006]. 

The direction of cutting affects methane emissions on the longwall face because at the 

end of an H–T pass, the face conveyor, shearer, and belt are all contributing gas at or near their 

maximum rate. The T–H passes do not have coincidental maximums for face conveyor, shearer, 

and belt emissions, so a more consistent emission rate occurs over the entire cut sequence. This 

explains the less frequent gas delays on T–H passes. Of the 11 H–T passes over the three shifts 

that were monitored for this study, only one cut did not have a gas delay, but it did have a face 

conveyor delay. A total of 10 complete passes had gas delays. Of the nine T–H passes, three had 

gas delays. 

Figure 19 shows the calculated methane contributor components for day 1 using the 

3-day average constants of shields 20 and 145. The calculated methane contribution is the 

product of the 3-day average constants of shields 20 and 145 and the formula results for each 

minute for day 1. Additional analyses and graphs of the data are presented by Krog et al. [2006]. 

Figure 19.—Calculated individual methane contributors and total calculated methane 

emissions for study day 1. 

42
 
Analysis of face methane emission rates indicates that gas 

delays occurred during periods of high emissions. The predictive emissions provided good correlation between the measured 

and predicted methane emission rates during the delays. 

Technique for Predicting Methane Emissions 

on Longer Longwall Faces 

Coal production and transport factors have an important influence on overall methane 

emission rates on the longwall face. Coal productivity (coal volume mined per hour) will be 

increased for a longer face since the cut cycles are face length-dependent. At longer face lengths, 

the wedge/sumping times are assumed to be the same as those for the base case, but the cutting 

times will increase proportionally to the face length, minus the sumping distance. 

Assuming that the longwall face conveyor can keep up with the shearer and the shearer 

cuts at the same speed over the greater face length, it follows that the productivity of the shearer 

will increase because a greater percentage of time will be allotted to cutting than sumping. 

Therefore, the total methane liberation from the mined coal during a shift would increase, but the 

shearer’s emission rate during cutting will remain the same for a longer longwall face if the 

cutting speed remains constant. The rib emission will be linearly dependent on the remaining 

length of the panel. The background emissions for the longwall face will increase linearly with 

face length. However, if the belt line is used as a source of face air, the background emission 

from the intake rib will remain constant. 

The face conveyor, if operating at a constant speed, will transport a greater volume of 

coal per hour for a longer longwall face. The face conveyor will also transport the coal over a 

greater distance and for a longer time, thereby increasing the methane emissions from this 

component on a longer longwall face. The belt emissions are a function of the amount of coal on 

the belt, the transport time, and the amount and direction of belt air. The belt emissions can also 

reach a steady-state maximum when the entire belt is full with coal, thereby capping the upper 

limit of belt emissions regardless of panel width [Krog et al. 2006]. 

The expected peak methane emission increases for wider 

longwall panels result primarily from the coal transported on the 

face conveyor and the background emissions from the larger 

area of exposed coal on the face. Methane emissions associated 

with the cutting of coal on the face by the shearer will remain 

constant for longer longwall face lengths as long as the cutting 

speed remains constant. 

The calculated results for methane emissions on longer longwall faces are predicted for a 

location 15 m (50 ft) outby the tailgate corner before any possible interaction with the gob gas 

near the tailgate. For a 305-m (1,000-ft) wide longwall panel, methane emissions were predicted 

for two full cuts without delays. The predicted peak methane emission of 0.110 m3

/s (234 cfm) 

closely matches the maximum values recorded during the study—0.099 m3/s (210 cfm) [Krog 

43 
et al. 2006]. Figure 20 shows the predicted methane emissions for two full cuts without any 

delays for a 488-m (1,600-ft) wide longwall panel. The calculated peak methane emissions for 

the 488-m (1,600-ft) wide longwall panel are 37% higher than for the roughly 305-m (1,000-ft) 

wide longwall panel. 

Coal on the face conveyor caused the largest calculated increase in methane emission 

rates on the longer longwall faces, whereas coal cut by the shearer and on the belt, as expected, 

caused no increase (Table 9). The face conveyor’s methane emission increase is due to the 

increased length and time that the coal will be carried by the conveyor. Keeping the length of the 

remaining panel (and thus the length of the belt) constant at 1,195 m (3,920 ft) for each of the 

increased face length emission calculations precludes any extra peak methane load being emitted 

by coal on the belt. 

Two methods exist for reducing longwall face methane concentrations produced by all sources. The first is to increase the 

amount of face airflow. Second, long-term methane drainage will 

also reduce the methane contents in the coal, thereby reducing 

methane emissions from potentially all methane contributors. 

Table 9.—Calculated rates for methane emission contributors on idealized passes 

on longer longwall faces 

Face width Conveyor Belt Shearer Background Peak 

(m) (m3

/s) (m3/s) (m3/s) (m3/s) (m3/s) 

305 0.040 0.027 0.009 0.035 0.110 

366 0.048 0.027 0.009 0.040 0.124 

427 0.056 0.027 0.009 0.046 0.138 

488 0.064 0.027 0.009 0.052 0.152 

Face width Conveyor Belt Shearer Background Peak 

(ft) (cfm) (cfm) (cfm) (cfm) (cfm) 

1,000 85 58 18 73 234 

1,200 101 58 18 86 263 

1,400 118 58 18 98 292 

1,600 135 58 18 110 322 

Percent relative to 305-m (1,000-ft) longwall face 

Face width Conveyor Belt Shearer Background Peak 

305 m (1,000 ft) 100% 100% 100% 100% 100% 

366 m (1,200 ft) 120% 100% 100% 117% 112% 

427 m (1,400 ft) 140% 100% 100% 134% 125% 

488 m (1,600 ft) 160% 100% 100% 150% 137% 

44
 
 

The background emissions increase with longer face lengths because of the increase in 

exposed longwall face area. However, the methane contribution from the gate road ribs does not 

increase because the length of the gate roads remains constant in these calculations. It should 

also be noted that the longer longwall faces will theoretically have higher coal productivity 

because a greater percentage of time will be spent cutting coal instead of sumping [Krog et al. 

2006]. 

As mentioned previously, the highest predicted methane emissions for longer longwall 

faces are near the end of the H–T cuts, when the emissions from the coal on the face conveyor 

and belt are at their highest (Table 9). During the three shifts monitored for this study, 1 of the 

11 complete H–T passes did not have a gas delay, but it did have a face conveyor delay. Slowing 

down the shearer for the second half of the face traverse will reduce the peak methane emissions 

and give a more consistent emission level. The influence of shearer cutting speed on emissions 

for the 488-m (1,600-ft) panel is shown graphically in Figure 21. The figure depicts face 

methane emission rates for the shearer traveling at 14 m/min (50 ft/min) for the first 250 m 

(820 ft) of the H–T cuts, then at 12 m/min (39 ft/min), and finally at 10 m/min (33 ft/min) to 

the tailgate. 

Figure 20.—Calculated methane emissions for two full cuts without delays for a 488-m 

(1,600-ft) longwall face. 

45
 
Figure 21.—Calculated influence of variable shearer transit speed pass for a 488-m (1,600-ft) longwall face. 

The reduction in methane emission rates resulting from slowing the shearer cutting speed 

is dramatic, with only 86% (0.131/0.152/ m3/s (278/322 cfm)) of the peak methane emissions 

being encountered. The 5 extra minutes required to cut the H–T pass due to the reduced shearer 

transit time (Figure 21) is still less than the observed average gas delay of 7 min for H–T passes. 

By slowing the shearer cutting speed on H–T passes, the calculated peak methane emission 

values based on the cutting direction are now within 7% of each other, compared to the 20% 

difference for the full-speed H–T pass (Figures 20–21). 

The conversion of the belt entry to return airflow and the conversion of the No. 3 return 

entry to intake could increase total airflow at the longwall face, as well as eliminate the belt coal 

methane emission component from the total methane emission load reaching the tailgate corner 

of the face. The primary drawback to this conversion is that the No. 3 entry ribs tend to have a 

higher background methane emission rate than that of the No. 1 entry ribs due to the virgin coal 

along the No. 3 entry’s rib. 

46
 
Mine operators have several options to reduce or dilute methane 

emissions expected from wider longwall panels: 

1. Increasing ventilation airflow quantities to the longwall 

face 

2. Reducing the shearer transit speed, especially on H–T 

passes 

3. Using or increasing methane drainage techniques to 

reduce emissions from all the considered sources by 

reducing the methane content of the coal

4.	 Implementing ventilation design changes, e.g., not 

coursing the belt entry’s ventilation airflow to the face 

Summary 

•	 High methane emission rates may have occurred minutes before their detection and 

response by sensors. 

•	 Methane contributions from the face conveyor and shearer dramatically increase in their 

relative contribution to the total face emissions during gas delays. 

•	 Peak methane emission increases for wider longwall panels will result primarily from the 

coal transported on the face conveyor and the background emissions from face coal. 

•	 Mine operators have several options to reduce or dilute methane emissions expected from 

wider longwall panels: 

1.	 Increasing ventilation airflow quantities to the longwall face 

2.	 Reducing the shearer transit speed, especially on H–T passes 

3.	 Using or increasing methane drainage techniques 

4.	 Implementing ventilation design changes 

47
 
5. DEVELOPMENT OF NUMERICAL MODELS TO INVESTIGATE 

PERMEABILITY CHANGES, DISTRIBUTIONS, AND GAS EMISSIONS 

AROUND A LONGWALL PANEL 

Overview 

Once a longwall face begins retreating, the formation of the longwall gob begins. Underground longwall mining of coal causes disturbance of the overlying rock mass. The caved rock 

(gob) behind the retreating longwall face can contain high void ratios, providing high permeability flow paths to the methane. The disturbance can increase the rock mass permeability 

through a reduction on the stress, as well as formation of new fractures in the rock. Methane gas 

contained in the disturbed rock mass can migrate toward the low-pressure mine workings and 

present an explosion hazard. This section describes the application of a finite-difference program 

to predict permeability changes within the rock mass using empirical relationships between 

fracture permeability and stress. 

The extent of rock failure is determined using a geomechanical model that considers 

both rock matrix and bedding plane failure. The caved rock (gob) is modeled as a compressible, 

granulated material. The calculated two-dimensional horizontal and vertical permeabilities 

around the longwall face are averaged and used as one of the inputs to the reservoir model. The 

modeling approach provides a basis for estimating methane inflow and optimizing control 

measures. The reservoir simulator models methane desorption from the coal matrix, methane 

release from the rock layers, and flow toward the mine excavations. The reservoir model was 

calibrated against records of methane flow at a study mine in southwestern Pennsylvania. Good 

correlation between actual gas production and model outputs has been achieved [Esterhuizen and 

Karacan 2005]. 

This section also describes a newly developed methodology to determine both horizontal 

and vertical variations in the permeability of the gob. Estimating the permeability distribution 

within the gob poses challenges due to its complexity. Variations of the permeability in the 

vertical direction are based on a model of caving and block rotation, which considers the effect 

of block dimensions and fall height on the void ratio. Gob compaction by the overburden and 

associated permeability changes are determined from a three-dimensional (3-D) geomechanical 

model, which simulates the gob as a strain-hardening granular material. The resulting 3-D 

permeability distribution in the gob is then transferred to a reservoir model. This section demonstrates the application of the method and shows that reasonable results are obtained when compared to empirical experience and measurements [Esterhuizen and Karacan 2007]. 

Model Development 

A two-staged approach has been followed to develop models of methane emissions and 

flow around longwall mines. The first stage has been to make use of the Itasca Consulting 

Group’s [2000] FLAC2D (Fast Lagrangian Analysis of Continua) finite difference code to 

simulate the geomechanical response of the rock mass to longwall mining. The program was 

used to calculate the stress changes, extent of rock fracturing, and bedding plane shear. The 

output of the FLAC models was used to calculate likely permeability changes based on empirical 

relationships. The permeability distribution was then used to develop inputs for the 3-D 

Generalized Equation-of-State Model (GEM) compositional reservoir simulator by Computer 

48
 
 

 

Modelling Group [2003]. The reservoir simulator was used to develop a model of several 

longwall panels in a study mine against which the model was calibrated by history matching of 

GGV methane production. The model has been used to evaluate the relative merits of methane 

drainage options. 

Permeabilities measured by Hasenfus et al. [1988] in strata above the Pittsburgh Coalbed 

demonstrated that the permeability can vary by several orders of magnitude in different sections 

of a vertical borehole. Hasenfus et al. measured hydraulic conductivities of 7 × 10−6

 cm/s 

(3 × 10−6 in/s) in sandstone near the ground surface. Brutcher et al. [1990] tested the conductivity 

of a sandstone aquifer and found the values to vary between 10−4

 and 10−6 cm/s (10−4 and 

10−6

 in/s) while shale conductivity was one order of magnitude lower. Booth and Spande [1992] 

reported hydraulic conductivities of 9 × 10–5 cm/s (4 × 10−5 in/s) for sandstone in southern 

Illinois. Matetic et al. [1995] reported permeabilities of 7 × 10−6

 cm/s (3 × 10−6 in/s) in shale 

materials near surface and 7 × 10−5 cm/s (3 × 10−5 in/s) for sandstone in southeastern Ohio. 

The field-measured hydraulic conductivities all fall within published ranges for 

sandstones and shales, and the upper limits fall within the range that one would expect for 

jointed rock. Coal measure rocks in the Eastern United States are typically poorly jointed but 

contain bedding planes that act as discontinuities, which allow horizontal flow. Vertical flow is 

constrained, especially by thin clay layers. Changes in stress can produce large variations in the 

permeability of laboratory- and field-scale rock. The permeability of the field-scale rock mass is 

affected by the closure or opening of fractures under changing stresses. The equivalent permeability can be related to the fracture aperture and fracture spacing [Bai and Elsworth 1993; Hoek 

and Bray 1981; Louis 1969]. In addition to stress changes, fracturing occurs in the rock mass in 

the vicinity of a longwall panel. Three zones can be distinguished in the roof rocks: the caved 

zone, fractured zone, and bending zone [Peng and Chiang 1984; Singh and Kendorski 1981]. 

FLAC2D models were used to simulate rock behavior along a longitudinal section 

through the centerline and a section across the width of a typical longwall panel at the study 

mine. The longitudinal sections were used to assess stress changes and rock failure around the 

advancing longwall face. The cross-section was used to assess rock behavior around the edges of 

the longwall, which included the behavior over the chain pillars left between longwall panels. 

The model dimensions were typically 400 m (1,300 ft) wide by 350 m (1,150 ft) deep to simulate 

a longitudinal section through a longwall panel [Esterhuizen and Karacan 2005]. 

Rock Permeability Calculations 

Permeability changes were calculated for both stress changes and rock fracturing. 

Permeabilities were calculated independently for the horizontal and vertical directions. Equations 

3 and 4 were used to determine the stress-affected horizontal and vertical permeabilities, 

respectively: 

−0.25(σ yy −σ yy 0 ) Kh h = K 0 × e (3)

−0.25(σ −σ ) K xx xx 0 = K × e v v0 (4) 

where K is permeability, σxx and σyy are the horizontal and vertical stresses, and the “0” subscript 

indicates initial conditions [Esterhuizen and Karacan 2005]. 

49
 
 

 

The permeability of fractured rock was determined from published values of jointed rock 

and fractured rocks [Hoek and Bray 1981]. Model elements that fail in compression are assumed 

to experience an increase in permeability of 100 md above their current permeability in both the 

horizontal and vertical directions, regardless of rock type. Similarly bedding shear failure is 

assumed to increase the permeability by 50 md in the bedding (horizontal) direction. The 

fracture- and shear-related permeability changes are added to the current permeabilities at the 

time of failure. The new permeability is then subject to further variation as a result of stress 

changes using the logic described above. 

The gob is formed by rock fragments that fall from the roof strata into the void created by 

the removal of the coalbed. The bulking of the gob is affected by the fall height of the rock fragments as well as the size and shape of the fragments. When the fall height is greater than the 

lateral dimension of the rock fragments, they are more likely to rotate and come to rest in a 

jumble, which produces relatively large void spaces. As caving proceeds upward, the caved rock 

occupies an ever-increasing proportion of the free space, thus reducing the fall height of the 

subsequent fragments. As the fall height reduces, the potential for fragments to rotate diminishes 

and the amount of bulking is reduced. 

The variation of the bulking of the gob in the vertical direction was estimated using a 

procedure suggested by Munson and Nenzley [1980]. The procedure assumes that maximum 

bulking of the caved rock will occur when the fall height exceeds about twice the block width. 

The maximum bulking factor was assumed to be 75% after tests on simulated gob materials 

[Pappas and Mark 1993]. The bulking factor (S) is expressed as 

V + V S = r v

(5) Vr

where Vr is the rock volume and Vv is the void volume. It was further assumed that a smooth 

transition will occur from the maximum bulking factor to zero when the fall height is zero 

[Esterhuizen and Karacan 2007]. 

The resulting variation in the bulking factor with vertical distance above the floor of a 

1.8-m (6-ft) high coalbed was calculated using this relationship for various block widths. Near 

the floor of the mined coalbed, the bulking factor remains at about 75%, but there is a rapid drop 

in the bulking factor at about 2.7–4.6 m (9–15 ft) above the floor, depending on the block width. 

This lower zone is known as the fully caved zone; the upper zone is the partially caved zone. 

The effect of gob compaction by the weight of the overlying strata was introduced by 

assuming the bulking factor will reduce in direct proportion to the amount of compaction. For 

example, 10% compaction will reduce the bulking factor by 10% of the current value at all 

points, regardless of the distance above the coalbed floor. This is a simplifying assumption, 

which does not significantly affect the permeability results. 

The Carman-Kozeny equation for flow-through porous media was used to estimate the 

permeability of the gob (K) as follows: 

K0 ⎛

⎜

⎝

⎜

n

3 ⎞

⎟

⎠

⎟ K =
 

0.241
 (1− n)

2 (6)
 

50
 
 

where K0 is the base permeability of the broken rock at the maximum porosity, and n is the 

porosity. The value of K0 was taken as 1 × 106

 md, which places it in the “open jointed rock” 

range according to Hoek and Bray [1981]. This permeability value also falls within the same 

range of permeabilities calculated from experimental friction factors for flow-through crushed 

stone compiled by Stephenson [1979]. The K0 value of 1 × 106

 md results in realistic GGV 

production when used in reservoir modeling of longwalls [Karacan et al. 2007b]. 

The results indicate that the permeability is 1 × 106

 md in the lower part of the gob, 

where bulking is at the maximum. This high-permeability zone extends to about 1.5 times the 

mining height. The permeability rapidly decreases at points that are more than about 3 m (10 ft) 

above floor. The permeability soon drops to 100 md at a height of about 6.1–9.1 m (20–30 ft), 

which places it in the permeability range of jointed rock. This height corresponds with the 

empirical observation that cave rock extends about four to six times the height of the mined 

coalbed [Mucho et al. 2000]. 

Distribution of Gob Permeability in a Mined Panel 

The compaction of the gob behind the longwall face and the associated permeability 

changes were determined through the use of the FLAC3D numerical modeling program [Itasca 

Consulting Group 2005]. The program allowed realistic modeling of stress redistribution about a 

longwall panel and was able to model rock fracture and gob compaction. The gob was modeled 

using the double-yield material type in FLAC3D, which represented materials in which there 

may be significant irreversible compaction in addition to shear yielding. The material parameters 

were selected to simulate a gob material that has an initial bulking factor of 0.75, displaying an 

exponential increase in load as it is compacted by the weight of the overlying strata. Further 

details of the modeling approach and typical input parameters are presented by Esterhuizen and 

Karacan [2005]. 

Permeability varies considerably within the gob. Near the edges 

of the gob, model results show that the permeability is relatively 

high, with a maximum of 1 × 106 md estimated near the corners 

of the mined panel. The permeability in the central part of the 

gob seems to be relatively constant at about 1 × 105

 md. These 

values are most meaningful for mines operating in the Pittsburgh 

Coalbed. 

The FLAC3D model can be set up to simulate the progressive extraction of coal by a 

retreating longwall panel. The void behind the face and the roof rocks up to four to six times the 

mining height are filled with gob material. The overburden rocks in the model will subside and 

compact the gob until a state of equilibrium is reached. The compaction of the gob was obtained 

by querying the FLAC3D model at selected points. Knowing the compaction distribution, it was 

possible to calculate the remaining void space at each point. It was then a matter of calculating 

the porosity and associated permeability values using the Carman-Kozeny equation. It was 

assumed that the horizontal permeability was twice the vertical permeability owing to the shape 

of rock fragments. 

51
 
The data show that the typical gob gas venthole locations near 

the tailgate margin of the panel, in this case a distance of about 

90 m (300 ft), are well-positioned to intersect the higher 

permeability zones above the gob. 

Summary 

•	 Permeability values determined for the gob decrease from the margin of the gob to the 

center of the gob by about a factor of 10 to about 1 × 105 md. 

•	 The design of GGVs should include the interception of the higher permeability zone near 

the tailgate gate road. 

52
 
6. METHANE EMISSION CONTROL DURING MINING OF LONGWALL PANELS
 

USING GOB GAS VENTHOLES 

Overview 

The caving of material filling in the void space formed by the removal of the extracted 

coalbed and the induced fracturing into the mine roof and floor produce an extensive, voluminous coalbed methane reservoir. The coalbed methane accumulating in the growing gob soon 

becomes the largest single gas source in the underground mining environment when using the 

longwall mining method [Curl 1978; Schatzel et al. 1992]. Although ventilation pressure 

differentials on the active panel are maintained to keep the gob gas from accumulating on the 

longwall face, gob gas typically contributes to the coalbed gas concentrations at the tailgate side 

of the face. The longwall gob gas can also contribute to gas accumulations in the gate roads and 

bleeder system. 

High methane emissions originating from the active face areas and from the fractured 

formations overlying and underlying the mined coalbed can adversely affect both safety and 

productivity in underground coal mines. Since ventilation alone may not be sufficient to control 

the methane levels in the longwall mining environment, GGVs have become a standard supplementary methane control option in many mines. These ventholes intercept and capture gas 

released from the subsided strata before it can enter the ventilation system, thus reducing emissions into the mine entries. 

Background 

The recovery of methane from longwall gob areas requires consideration of both rock 

mechanics and fluid dynamics principles in the context that gas flow through strata is mainly 

controlled by the permeability of the rock units involved. This in turn is a function of the stress 

disturbances caused by mining activity. To improve gas capture at a reasonable cost, it is 

important to understand the behavior of the entire GGV system, including the venthole 

placement and completion strategies, and the reservoir properties of the gob. In order to 

investigate the effect of GGV completion parameters, a typical multipanel longwall mine was 

modeled using Computer Modelling Group’s [2003] GEM. 

The mine selected for this study operates in the Pittsburgh Coalbed in Greene County, 

Pennsylvania. Overburden depths in the area range between 150 and 270 m (500 and 900 ft). 

Longwall panels in the old mining districts of this mine were 253 m (830 ft) wide and were 

increased to about 305 m (1,000 ft). The alteration of permeability fields in and above the panels 

as a result of the mining-induced disturbances has been estimated from mechanical modeling of 

the overlying rock mass using FLAC [Itasca Consulting Group 2000]. Model calibration was 

performed through history-matching the gas production from GGVs in the study area. Figure 22 

shows the 3-D grid model that was constructed for this study. The figure shows only the 

Waynesburg Coalbed (top layer) and Pittsburgh Coalbed (bottom layer). The other layers 

between the Waynesburg and Pittsburgh Coalbeds have been removed for a better visualization 

of the wellbores used in the model [Karacan et al. 2006]. 

53
 
 Figure 22.—Three-dimensional representation of grid model of the study area that shows the Waynesburg 

(top layer) and Pittsburgh (bottom layer) Coalbeds. Also shown are GGVs, their placement, completion 

representation, and elements of a pseudo-ventilation system. 

Most GGVs at the cooperating site are drilled within a short distance (10–30 m (30–100 

ft)) of the coalbed being mined and cased with steel pipe. Commonly, the bottom section of the 

casing (generally about 61 m (200 ft)) is slotted and placed adjacent to the expected gas production zone in the overburden strata. The diameter of the casing may change, but is usually around 

18–20 cm (7–8 in). 

Effect of Completion Parameters of Gob Gas Ventholes on Their Performance 

The standard casing diameter for the GGVs in the study area was 18 cm (7 in). The 

effects of different slotted casing diameters (25 and 10 cm (10 and 4 in)) on methane production 

were evaluated. The length of the slotted casing and its setting depth above the top of the Pittsburgh Coalbed were held constant at their original design values, 61 and 12 m (200 and 40 ft), 

respectively. 

The modeling results (Table 10) predict that the cumulative methane production using the 

25-cm (10-in) casing will increase 4.9%, compared to the 18-cm (7-in) standard-diameter casing. 

The amount of methane produced with the 10-cm (4-in) casing was about 6.7% less than that 

produced with the 18-cm (7-in) diameter casing. However, the amount of mine air produced with 

the 10-in casing was 12% more, which resulted in a lower predicted methane concentration in the 

produced gas stream. The average calculated methane concentration in the produced gas stream 

at the end of simulated mining period (910 days) was 60% with 25-cm (10-in) casing, compared 

to 62% and 64% with 18- and 10-cm (7- and 4-in) diameter casing, respectively [Karacan et al. 

2006]. 

54
 
The predicted increase in cumulative methane production with the larger-diameter 

wellbore was due to the increase in the open-to-flow area of the wellbore and the calculated 

productivity index (a mathematical means of expressing the ability of a reservoir to deliver fluids 

to the wellbore, stated as the volume delivered per drawdown pressure) in the wellbore models. 

Also, with larger-diameter wellbores, the pressure losses were less compared to smaller-diameter 

wellbores. The predicted reduction in methane concentration with the 25-cm (10-in) diameter 

casing is most likely the result of more mine air being captured due to an expanded pressure sink 

and associated depletion radius created by the production of gas from the larger-diameter casing. 

However, since the total predicted gas production (methane and air) was higher for the 25-cm 

(10-in) diameter casing, it still resulted in higher cumulative methane production, even though 

the methane concentration was less (Table 10). 

Table 10.—Predicted effect of casing diameter on cumulative methane and total gas production 

for 910 days of simulated mining 

Casing 

diameter, 

cm (in) 

Cumulative 

methane produced, 

×106 m3 (MMscf) 

Change in 

methane 

produced 

compared to 

7-in casing, % 

Cumulative gas 

produced 

(methane + air), 

×106 m3 

(MMscf) 

Change in air 

compared to 

7-in casing, % 

Calculated 

methane 

concentration, 

% 

10 (4) 11.1 (392) −6.7 17.3 (609) −15 64 

18 (7) 11.9 (420) — 19.2 (677) — 62 

25.4 (10) 12.5 (440) +4.9 20.6 (728) +12 60 

Increasing gob gas venthole diameter results in more methane 

production. However, methane concentration may be diluted due 

to the inclusion of more mine air. Differing coalbed reservoir 

parameters may modify the optimum gob gas venthole 

configurations. 

To evaluate the influence of the length of the completion interval on GGV performance, 

the length of the slotted casing section was changed in the model to 30 and 76 m (100 and 

250 ft), compared to the original 61-m (200-ft) length. The casing diameter was kept at 18 cm 

(7 in), and the setting depth of 12 m (40 ft) above the top of the Pittsburgh Coalbed was maintained. The modeling results predict that the cumulative methane production will increase with 

increases in slotted casing length (Figure 23). 

Methane production with 76 m (250 ft) of slotted casing was 13.0 × 106 m3 (460 MMscf), 

compared to 12.1 × 106

 m3 (392 MMscf) with the standard 61 m (200 ft) of slotted casing. This 

difference corresponds to a 9.5% increase in methane capture from the four panels modeled 

(Figure 22). However, when the slotted casing length was shortened to 30 m (100 ft), the predicted methane production decreased to 8.91 × 106

 m3 (315 MMscf), which was about 25% less 

than that captured with 61 m (200 ft) of slotted casing. 

55
 
Figure 23.—Simulated cumulative methane production from ventholes with varying lengths of slotted casing. 

Increasing slotted casing length results in increased methane 

production. Knowledge about key reservoir parameters in the 

overlying strata (e.g., coalbed methane content, formation 

permeability) can be very useful input for configuring gob gas 

venthole slotted casing lengths. 

The effect of casing setting depth (distance from the top of the mining layer) on gas 

production was investigated by modeling alternative completion depths of 19.3, 7.6, and 4.6 m 

(65, 25, and 15 ft), compared to the original 12-m (40-ft) depth. The 7.6-m (25-ft) completion 

depth generally corresponded closely to the caved zone height, which was modeled as 7.3 m 

(24 ft) above the Pittsburgh Coalbed. Consequently, the 4.6-m (15-ft) completion depth resulted 

in the venthole being drilled into the caved zone. The 19.3-m (65-ft) completion depth corresponds to a depth slightly below the Sewickley Coalbed. For these scenarios, the casing diameter 

and slotted casing lengths were kept at their original design values, 18 cm and 61 m (7 in and 

200 ft), respectively. 

Raising the slotted casing setting depth to 19.8 m (65 ft), compared to 12 m (40 ft), above 

the Pittsburgh Coalbed resulted in a 4% predicted cumulative methane production increase. The 

predicted cumulative methane production declined by about 5% and 29% when the casing was 

set to within 7.6 and 4.6 m (25 and 15 ft) of the top of the mining layer, respectively. The total 

gas production increased by 4.9% with the 19.8-m (65-ft) slotted casing setting depth scenario 

(mostly because of the increase in methane concentration) and decreased by 5% for the 7.6-m 

(25-ft) setting depth (mostly because of the decrease in methane concentration). The total gas 

production increased by about 10% with the 4.6-m (15-ft) slotted casing setting depth. In the 

56
 
 

4.6-m (15-ft) setting depth scenario, the lower slots of the casing were in the caved zone influenced by the mine ventilation system, where flow resistance was small. Therefore, the ventholes 

pulled 74% more mine air, compared to the operator’s standard 12-m (40-ft) setting depth. Since 

most of the produced gas was mine air at the 4.6-m (15-ft) setting depth, the average methane 

concentration in the cumulative produced gas at the end of mining was about 40%, compared to 

60%–70% average methane concentration calculated for other depths (Figure 24). 

A real-world example of the gas quality consequences of completing GGVs into the 

caved zone is illustrated with measured gas concentration data from two ventholes continuously 

monitored during a field study (Figure 25). For this site, the height of the caved zone was estimated to be 12 m (40 ft). This is higher than the site shown in Figure 22 due to the presence of 

the sandstone paleochannel. The first venthole on the study panel was completed to a depth of 

14.3 m (47 ft) above the top of the Pittsburgh Coalbed, generally within the standard depth range 

for the mine site. However, the second venthole drilled on the study panel was inadvertently 

drilled deeper to a depth of 11 m (35 ft) above the top of the Pittsburgh Coalbed, which is in the 

caved zone. As shown in Figure 25, the methane concentration in the produced gas from the 

venthole completed into the caved zone averaged about 30% less than that of the standard completion depth above the caved zone because of the increased production of mine ventilation air 

[Karacan et al. 2007b]. 

Figure 24.—Simulated methane concentrations in the gas production stream from ventholes completed to 

varying depths above the Pittsburgh Coalbed. 

57
 
 Figure 25.—Actual methane concentrations measured in the gas production stream from two GGVs 

completed to different depths above the Pittsburgh Coalbed on the same longwall panel. 

Locating the bottom of gob gas ventholes into the caved zone is 

counterproductive for methane control because mine ventilation 

air is then removed from the borehole, limiting the potential 

removal of methane. 

The increase in mine ventilating air from the GGVs is a problem for several reasons. The 

first concern is that if the venthole is producing at its maximum capacity, then the borehole is 

producing a portion of the ventilation air and in favor of the maximum quantity of methane 

available in the subsided strata, which can then migrate to the underground workplace where it is 

a potential explosion hazard. In addition, when coalbeds and associated caved zone strata are 

prone to spontaneous combustion, the flow of additional mine air into this zone may pose an 

increased risk of a mine fire of spontaneous combustion origin [Smith et al. 1994]. There are also 

economic issues associated with producing higher levels of mine ventilation air from the GGVs. 

Obviously, there are costs associated with providing ventilation air to the underground workings, 

as there are with the drilling and operation of the ventholes. Economically, it is counterproductive to incur cost to first introduce the ventilation air to the mine, and then incur additional 

cost to remove it from the mine via the GGVs. Finally, for those mining operations capturing gob 

gas for commercial sale, it is very important to maintain as high a methane concentration as 

possible in the produced gas stream, or additional expenses will be incurred to remove the 

nonhydrocarbon gases prior to selling it to the pipeline. 

58
 
An important factor in optimizing coalbed methane production 

from gob gas ventholes is surface exhauster operation. In the 

case of methane-powered exhausters, keeping the methane 

concentration high extends the production life of the gob gas 

venthole, since most operators will shut in boreholes when 

concentrations drop below 25%. Many exhauster motors 

powered by methane cease to run reliably when concentrations 

drop to levels approaching 25%, which significantly limits their 

effectiveness in controlling gob gas underground emissions. 

Summary 

•	 Locating the bottom of GGVs into the caved zone limits the removal of methane by using 

borehole capacity to remove some mine ventilation air. 

•	 Increasing slotted casing length results in increased methane production from a GGV. 

Other site-specific factors need to be considered in increasing the casing length, such as 

the geology where the borehole will be anchored. 

•	 Many exhauster motors powered by methane cease to run reliably when concentrations 

drop to levels approaching 25%, which limits their effectiveness in controlling gob gas 

contributions to underground emissions. Powering the surface exhauster with an alternate 

fuel source will allow coalbed methane production to continue without combustion 

performance concerns. To maintain safe operating conditions, a 25% methane concentration cutoff is commonly used for all exhauster operation. 

59
 
 

7. THE APPLICATION OF GOB GAS VENTHOLES TO CONTROL METHANE 

IN WIDER LONGWALL PANELS AND GOBS 

Overview 

The competitive nature of the domestic and international coal industry is constantly 

pushing mining companies to increase coal productivity and the size of their longwall panels. 

Increasing panel widths can result in increased methane emissions due to the larger volumes of 

gas released from the surrounding and overlying strata. The design and locations of GGVs are 

important to control the extra amount of gas resulting from increased amount of fractured strata. 

The grid model shown in Figure 26 was developed to estimate the increase in expected 

levels of methane emissions and to investigate alternative GGV completion and placement 

scenarios on a larger panel [Karacan et al. 2005]. These data were produced in a new mining 

district in the Pittsburgh Coalbed. Mining in the district had begun with 381-m (1,250-ft) wide 

panels and was progressing to 442-m (1,450-ft) wide panels. 

Figure 26.—Study mine for wider longwall faces: grid model of a new mining district in the 

Pittsburgh Coalbed. 

Effects of Increased Longwall Panel Width on Gob Methane Emissions 

Due to the uncertainty of the methane emission consequences associated with mining of 

the larger panel size, the field area was modeled to estimate the expected increase in gas flow 

from the gob and to investigate methane control options. The primary question to be answered 

from a mine safety and methane control perspective was whether the current number and 

configuration of GGVs would adequately control the projected increase in gob gas on the larger 

longwall panels [Karacan et al. 2005]. 

The four GGVs on the first panel in this new mining district were drilled to varying 

depths of 14, 11, 9, and 12 m (47, 35, 30, and 40 ft) above the top of the Pittsburgh Coalbed. 

60
 
 

Two of the ventholes were completed into or at the top of the caved zone, as opposed to the 

preferred distance of at least 12 m (40 ft) above the top of the coal. The simulation results for the 

381-m (1,250-ft) wide panel with the four GGVs were compared with those obtained with a 

panel width of 442 m (1,450 ft). The same venthole placement (91 m (300 ft) from the tailgate 

side of the panel) and completion and production histories were maintained (Figure 27). 

The performance of the four individual GGVs on the simulated wider panel was very 

similar to that of the original panels. This was expected since the wellbore flow model was not 

dependent on panel width. Due to subsidence variations, the most likely reservoir parameter that 

could be influenced by increased panel width was the reservoir permeability associated with the 

productivity index of the ventholes. However, for supercritical panels (as are both of these 

simulated base cases), subsidence should be complete and stress conditions should be similar 

irrespective of the panel width, as confirmed by the FLAC computations [Karacan et al. 2005]. 

Modeling the increase in panel width resulted in 1.3 × 106

 m3 (47 MMscf) of additional 

methane liberation from the coal mined on the longwall face and 3.88 × 106

 m3 (137 MMscf) 

from the overlying disturbed strata, for a total of 5.21 × 106

 m3 (184 MMscf) of additional 

methane over the 268 days of mining simulated for this study. The 1.3 × 106

 m3 (47 MMscf) 

of additional methane liberation on the face will flow to the ventilation system unless it is 

predrained from the mined coalbed. Alternatively, the ventilation airflow may have to be 

increased to dilute the additional gas. Depending on the availability of additional gob gas 

drainage capacity, some of the additional 3.88 × 106 m3 (137 MMscf) of methane originating in 

the overlying strata may flow to the ventilation system. This would represent the potential of 

up to about 0.168 m3/s (355 cfm) of additional methane entering the underground workplace. 

Increasing the longwall panel width increases the quantity of 

methane present due to the increased fractured reservoir 

volume. However, this size increase does not positively affect 

the performance of GGVs. Thus, to avoid additional methane 

emissions to the ventilation system associated with increased 

panel size, an optimized methane control system may be 

required. 

Comparison of Ideally Completed and Operated Gob Gas Boreholes 

With Those Operating in the Field 

Simulations were conducted to replace the four simulated actual ventholes with four 

optimal5

 ventholes on the 442-m (1,450-ft) wide panel. As shown in Figure 28, optimal ventholes produced gas that was 85%–90% methane during the entire mining period, whereas the 

methane concentration of the gas produced by the actual wells in the study area declined significantly with time, with the methane concentration in the 65%–70% range for most of the mining 

period. The difference in methane concentrations between the two scenarios is due to the fact 

that some of the actual ventholes were drilled into or very close to the caved zone; thus, their 

produced methane was diluted with mine ventilation air. 

5

Regularly maintained and continuously operated. 

61
 
 

 

Tailgate 

Headgate 

1250-ft 

1450-ft 

Scenario A 

1450-ft 

Scenario B 

Scenario C 

1450-ft 

Scenario D 

Scenario E 

1450-ft 

Scenario F 

Base Case Base Case 

Optimal wells 

Actual wells Mining direction 

380 m 

442 m 

442 m 

442 m 

442 m 

Figure 27.—Schematic representation of GGV configurations for the 

simulation of methane control options on wider longwall panels. 

Figure 28.—Comparison of the performance of actual GGVs at the study site with the optimal GGVs for the 

442-m (1,450-ft) wide panel. 

62
 
Optimal wells were also predicted to produce about 50% more methane than the actual 

wells. This, again, demonstrates the importance of implementing proper venthole completion 

practices and paying close attention to the maintenance and operation of the GGVs, i.e., keeping 

them running to optimize their methane control capability. 

Optimally drilled and continuously operated GGVs produced 

more methane at higher concentrations compared to their 

counterparts that were diluted by ventilation air and operated 

discontinuously. Although some operators prefer to keep GGV 

completion depths near the mined coalbed, these operators risk 

short-lived and intermittent GGV production. 

Since the performance (gas production potential) of individual GGVs was not influenced 

significantly by an increased panel width, the additional 0.168 m3/s (355 cfm) of methane 

released from the overlying strata as a result of mining the larger panel will potentially enter the 

underground workplace if additional methane drainage capacity is not provided. Thus, the constructed model was used to evaluate multiple scenarios to optimize the number and locations of 

the GGVs on the wider panel and to minimize the volume of additional methane entering the 

ventilation system. 

The simulated alternative GGV placement and completion scenarios investigated for this 

study, as shown in Figure 27, were: (A) moving the four actual ventholes to locations 152 m 

(500 ft) from tailgate side, i.e., 61 m (200 ft) closer to the centerline of the panel than on the first 

panel; (B) adding four optimal boreholes continuously operating with 19-kPa (2.7-psi) suction 

pressure and completed to 12 m (40 ft) above the Pittsburgh Coalbed located between each 

actual venthole; (C) adding four optimal ventholes located 91 m (300 ft) from the gate roads on 

the headgate side of the panel and positioned diagonally to the actual ventholes; and (D) adding 

four optimal ventholes located 91 m (300 ft) from the gate roads on the headgate side and 

positioned directly opposite the actual ventholes on the tailgate. Also, two scenarios with all 

optimal ventholes were tested (E and F) for their production performance and for reducing 

methane emissions into the ventilation system. 

The data presented in Table 11 compare the cumulative methane production volumes 

from each of the GGV configuration scenarios shown in Figure 27. The lowest predicted cumulative methane production (scenario A) is obtained when the four ventholes with the actual 

completions are located 152 m (500 ft) from the tailgate entry of the 442-m (1,450-ft) panel. 

This configuration produced 0.37 × 106

 m3 (13 MMscf) less methane compared to the base case 

production from the 381-m (1,250-ft) panel, where the ventholes were the traditional 91 m 

(300 ft) from the gate road, which illustrates the importance of near-margin venthole placement 

[Diamond et al. 1994]. 

The highest predicted cumulative methane production is achieved when additional 

optimal ventholes are used (Table 11). To demonstrate, scenario D produces 3.91 × 106 m3 (138 

MMscf) more methane compared to the four actual ventholes on the 3,810-m (1,250-ft) panel 

base case. Scenario C was the next highest incremental producer of methane at 3.37 × 106

 m3 

(119 MMscf). Although scenarios C and D are similar, scenario D is probably the higher 

producer because the tailgate and headgate ventholes are closer to each other (since they are 

directly opposite each other on the panel) and are intercepted by mining at the same time. This 

scenario results in a quicker overlap of the venthole drainage radii, which enhances gas 

63
 
 

 

 

 

 

desorption from the overlying coalbeds associated with the subsided strata. Also, when headgate 

and tailgate ventholes are intercepted at the same time, the headgate ventholes start producing 

earlier and stay on production longer compared to the diagonal location in scenario C, which 

results in more methane production. Within the scenarios evaluated, the third highest incremental 

methane producer is scenario B, with 3.19 × 106 m3 (113 MMscf) more methane than the four 

actual ventholes on the 381-m (1,250-ft) panel. 

The application of GGVs located on the headgate site of a longwall panel may seem 

counterintuitive to many seasoned methane control experts who have traditionally used the 

tailgate side of the panel. Other than the numerical modeling result, there are other supporting 

lines of evidence for this approach. Tracer gas testing performed by the NIOSH Pittsburgh 

Research Laboratory (PRL) in longwall gobs showed very long-ranged effects of a GGV over 

the length of a longwall panel. By splitting GGV locations between the headgate and tailgate 

near-margin locations, fewer boreholes are competing for gas in the fractured overburden, 

enhancing production potential. Tailgate locations are always the preferred site for cross-measure 

boreholes where the underground ventilation system assists in transporting methane-air mixtures 

through the mining-induced fracture system. For headgate-side GGVs to be effective, they must 

be completed above the caved zone, and induced mine fractures must not extend to ventilation 

pathways. The first report of the near-margin GGV design produced by the USBM used holes on 

the headgate side of a longwall panel due to land accessibility restrictions and achieved favorable 

results [Diamond et al. 1994]. 

One of the main considerations in GGV design and operation is to locate and drill the 

ventholes optimally and operate them continuously. This will minimize the number of ventholes 

that need to be drilled to produce the same amount of methane that a larger number of less than 

optimally drilled and operated ventholes will produce. Thus, in addition to the GGV placement 

scenarios (A through D), scenarios E and F were simulated to determine the minimum number of 

optimal ventholes that would produce the same maximum amount of methane as in the highestproducer configuration of actual and optimal ventholes. In scenario E, six optimal ventholes 

were placed along the tailgate side of the panel. In scenario F, two of the optimal ventholes were 

placed on the headgate side directly opposite the first two ventholes on the tailgate side. 

Table 11.—Cumulative predicted methane production difference 

from gob gas ventholes on a 442-m (1,450-ft) wide panel 

Venthole 

design 

Cumulative methane production 

difference compared to a 381-m 

(1,250-ft) wide face, 

×103 m3 (MMscf) 

442-m (1,450-ft) −25 

Base case −0.9 

Scenario A −370 (−13) 

Scenario B 3,200 (113) 

Scenario C 3,370 (119) 

Scenario D 3,910 (138) 

Scenario E 2,440 (86) 

Scenario F 3,790 (134) 

64
 
The predicted methane production performance of venthole configurations in scenarios E 

and F using six optimal GGVs were compared to the performance of scenario D, the highest predicted methane producer (3.91 × 106

 m3 (138 MMscf)) in the previous simulations. Table 11 

shows that scenario E, where six optimal ventholes were located along the tailgate side of the 

panel, produced about 2.44 × 106 m3 (86 MMscf) more methane than the 381-m (1,250-ft) base 

case, whereas scenario F, with two of the optimal wells on the headgate side, produced 3.79 × 

106

 m3 (134 MMscf) more methane. Thus, the six optimal ventholes of scenario F produced 

almost as much methane as the eight (four actual and four optimal) ventholes of scenario D. 

The predicted performance differences between scenarios E and F are due to the location of the 

ventholes and the length of time they stayed on production. 

Analyses show that six optimally completed, placed, and 

operated ventholes could produce the same amount of methane 

on the wider panel as eight nonoptimal wells. Increasing the 

production of gob gas decreases the methane emissions into the 

mining environment by reducing the amount of gas unremoved 

by methane drainage. 

Summary 

•	 Increasing the longwall panel width increases the quantity of methane present in the gob, 

but does not positively affect the flow and methane concentration produced from GGVs. 

•	 To avoid additional methane emissions to the ventilation system associated with 

increased panel size, an optimized methane control system may be required. 

•	 Optimally drilled GGVs produced more methane at higher concentrations compared to 

ventholes that were diluted by ventilation air. 

•	 When completing GGVs to depths near the mined coalbed, operators risk short-lived and 

intermittent venthole production. 

•	 Six optimally completed, placed, and operated ventholes can produce the same amount of 

methane on the wider panel as eight nonoptimal wells. 

65
 
8. INDUCED FRACTURING AND COALBED GAS MIGRATION 

IN LONGWALL PANEL OVERBURDEN: 

THE NIOSH BOREHOLE MONITORING EXPERIMENT 

Overview 

The predictive capability of numerical models can be improved by a better understanding 

of underground processes using site-specific field measurement data. To better understand the 

longwall mining effects on the methane reservoir overlying a longwall panel and to provide data 

for ongoing PRL modeling efforts, a borehole monitoring experiment (BME) was designed and 

implemented on an active longwall panel. The mine operates in the Pittsburgh Coalbed in Greene 

County, Pennsylvania. The goal of the BME is to provide recommendations to operators related 

to improving methane control measures through field observations of this study and through 

field-monitoring-assisted reservoir modeling technique enhancements. This section discusses 

the BME and the most important measurements, results, and recommendations associated with 

this effort. 

Description of the Borehole Monitoring Experiment on an Active Longwall Panel 

The boreholes were drilled in a new mining district where a prior longwall panel had 

been previously described by numerical models. The BME provides a direct comparison of field 

site observations to previous numerical experiments and also allows construction of refined 

models of the mined area for more accurate and detailed analysis. The drill site was selected 

based on an active panel, and all borehole activity was completed (except for monitoring) prior 

to undermining by the longwall face. The BME site location was chosen to be away from either 

end of the panel to avoid mechanical influences caused by the draping of overburden in this 

region. The borehole drilling depths and monitoring intervals were initially chosen to address a 

different stratigraphic zone in each hole. The experimental design specified that drilling was to 

be completed 2 months in advance of undermining the borehole location. 

The BME was planned for a panel that was 442 m (1,450 ft) wide. The three test 

boreholes were arranged in a line that was equidistant from the tailgate gate roads. The distance 

to the tailgate panel margin was 101 m (330 ft), the same distance that the mine used for its 

GGVs so that they would be in the same mechanical behavior zone and in a similar stress field. 

One of the operator’s GGVs was to be drilled 1,930 m (6,325 ft) from the completion end of the 

panel and was to be included in the BME monitoring activities. The distances between each of 

the monitoring boreholes was 15 m (50 ft), and the third borehole was 76 m (250 ft) from the 

nearby GGV. 

The three test boreholes were drilled and completed at different depths to monitor initial 

reservoir and mechanical properties in different strata horizons and subsequent property changes 

during the mining of the longwall panel. The top of the Pittsburgh Coalbed was at 252 m (827 ft) 

from the surface. According to this plan, the first, or shallowest, borehole (BH–1) was drilled to 

a total depth of 220 m (721 ft). The second, or midrange borehole (BH–2), was drilled to a depth 

of 230 m (755 ft). The deepest borehole (BH–3), which was also closest to the GGV, was drilled 

to 245 m (803 ft). 

Based on the local geology and the depths of the boreholes shown in Figure 29, BH–1 

was intended to monitor the Sewickley Coalbed, BH–2 to monitor mostly shale sequences, and 

66
 
BH–3 (the deepest) to monitor the shale and sandstone horizons that would be retained in the gob 

during and after undermining. Boreholes were drilled with a 15-cm (6-in) diameter bit. The 

drilling of the boreholes was started and completed when the longwall face was 760 m (2,500 ft) 

away from the BH–1 location. After drilling was completed, the deepest borehole (BH–3) was 

logged open-hole with density, gamma ray, and sonic tools to identify the formations, refine 

the drilling depths, and calculate porosity, density, and some of the mechanical properties of 

the rock. 

The boreholes were cased with 13-cm (5-in) steel casing. They were cemented using 

conventional grout and cement baskets, except for the bottom 6.1–9.1 m (20–30 ft). These 

sections were cased with slotted casing and were the primary monitoring zones for each hole. 

The length of slotted casing was 9.1 m (30 ft) in BH–1 in order to monitor both splits of the 

Sewickley Coalbed. The last 2.4 m (8 ft) of the slotted casing of BH–3 was cut and left openhole to keep the casing as high as possible above the gob (Figure 29). 

The experimental boreholes were configured to be completed in a manner similar to the 

mine operator’s GGVs. Both borehole designs include flame arrestors, shut-in valves, and a long, 

vertical PVC pipe stack. However, unlike the experimental boreholes, the GGVs were cased with 

61-m (200-ft) slotted casing at the bottom of all boreholes. The operator used a dump grouting 

procedure on the GGVs instead of the circulating grout method used on the test boreholes. The 

test boreholes were also different from the operator’s GGVs in that the BME boreholes were 

kept shut in throughout the mining duration and a powered exhauster was not attached. 

After the cementing operation, all of the boreholes were blown dry of any water 

remaining after completing the cementing operation by tripping the drill tubing down into the 

boreholes and using compressed air from the drill rig. Similar to the GGVs, test borehole 

wellheads were equipped with 15-cm (6-in) diameter flange for installation of the flame arrestor 

and the wellhead valve. Borehole monitoring instrumentation and associated hardware were also 

attached to the borehole stack. 

67
 
Figure 29.—Stratigraphy and downhole configurations for NIOSH BME. (TD = total depth.) 

Methodology and Instrumentation 

An important part of this study was to perform measurements of in situ permeability 

before and after the undermining of the NIOSH drill sites. To make these measurements, an 

experimental method was designed. These tests also provided valuable data on water production 

from the slotted casing sections. The initial, pre-undermining findings present data on the initial 

water saturations of the formations of interest. Initial permeability determinations were first 

conducted by using a rising-head test before installing the wellheads and shutting in the wells for 

long-term monitoring. 

For this test, boreholes were equipped with submersible, downhole transducers that were 

positioned within the downhole monitoring zones. The downhole transducers were to be installed 

underwater in the boreholes to record transient changes of the water head in the boreholes. 

However, there was no water rise recorded within a few days of borehole monitoring. Thus, a set 

of slug tests was designed and performed in each borehole using water to determine formation 

permeabilities. The water head drop was monitored for about a week on each borehole using the 

submersible, downhole transducers until the rate of water head change reached a steady state. 

After the conclusion of the initial slug tests, the downhole pressure transducers were repositioned 

just above the slotted sections. 

The boreholes were then filled with additional quantities of water so that the change in 

water head could be observed. This water monitoring interval was included to provide input on 

strata disturbances, their initiation with respect to the position of the longwall face, and the 

68
 
transient fracture permeabilities induced by mining. The submersible transducers recorded 

changes of water head until the water drained completely from the boreholes. A final set of slug 

tests was to be run after the completion of the panel for determining final permeabilities. 

In addition to submersible transducers, the wellheads on BME boreholes were equipped 

with surface pressure transducers for continuous data-recording pressure changes at the tops of 

the boreholes. Tiltmeters were also installed on the BME wellhead stacks for recording 

subsidence and strata response profiles. Conventional positional surveys were scheduled to track 

the movement of the surface. The longwall face position was recorded daily. Surface pressure 

and methane concentration readings were taken regularly to monitor the changes in the wellbores 

as a result of fracturing of strata. Pressure, flow rate, and concentration readings were also 

monitored at the nearby GGV to evaluate the underground interactions with the NIOSH 

boreholes. Figure 30 shows the wellhead configuration of one of the experimental boreholes. 

Figure 30.—Wellhead arrangement for monitoring boreholes. 

A distance of 305 m (1,000 ft) before and after the NIOSH boreholes were undermined 

was selected as the primary experiment test zone. This distance was based on the overburden 

depth and typical subsidence profiles from the Northern Appalachian Basin. All instrumentation 

was installed, the boreholes were shut in, and monitoring had begun while the longwall face was 

366 m (1,200 ft) away from the first borehole. 

69
 
 

Analysis and Interpretation of the Field Data 

Figure 31 shows the change in water head pressure in the first borehole (Sewickley 

Coalbed) and second borehole (shale and limestone zone between the Pittsburgh and Sewickley 

Coalbeds). Unfortunately, soon after starting to record data from the deepest borehole (Pittsburgh 

Coalbed), the communication with the downhole pressure transducer was lost. Thus, Figure 31 

does not include data from the third borehole. Figure 31 shows the effect of mining disturbances 

on water level changes in the boreholes. The data show that the initial water head drop in the first 

and second boreholes occurred before the boreholes were undermined. This suggests that the 

initial water head decrease is associated with fracturing and/or shearing of the strata ahead of the 

face by about 3 days in BH–1 and about 2 days in BH–2. 

Figure 31.—Water head change as a function of time in boreholes monitoring Sewickley (BH–1) and 

shale-limestone (BH–2) intervals. 

Figure 31 has other implications with regard to impacts of longwall mining on reservoir 

properties. For instance, the water level drop confirms that there are new permeability pathways 

created either by the shearing or by fracturing of strata. Also, the changing rate of water head 

drop confirms that permeability is not constant and varies (increases and decreases) as strata 

respond to changing stress conditions during mining. 

70
 
Water head drop data suggest that mining-induced disturbances 

forming the GGV fracture network can occur 24–46 m (80–150 ft) 

ahead of the mining face. This distance may vary when GGV 

slotted pipe configurations are used, which are typically 5–10 

times longer than those used in the BME. 

The data obtained during initial slug tests were evaluated by using the confined aquifer 

slug test model presented by Dawson and Istok [1991]. The calculated permeabilities for the 

monitoring zones in BH–1, BH–2, and BH–3 were 2.8, 0.1, and 0.2 md for the Sewickley 

Coalbed, for shale and limestone, and for shales, respectively. These evaluations also showed 

that these values are very close to the initial permeability values being used by NIOSH 

researchers in the numerical reservoir and fracture mechanics modeling these formations. 

In monitoring water levels in the boreholes, the loss of water can 

be rapid up to about 32 m (100 ft) above the mined coalbed, 

which can limit the duration of monitoring. The rate of water loss 

was not related to the depth of the borehole. Shearing and 

deformation in the overburden is typically severe, and much of 

the annulus is generally modified. 

To quantify permeability changes in the overlying strata during mining, a slug test model 

for confined, anisotropic aquifers of infinite or semi-infinite in radial extent was used [Dawson 

and Istok 1991]. The model was used for calculating instantaneous permeabilities that are represented by two consecutive data points recorded in 1-min intervals and also for calculating the 

average permeabilities during intervals that can be recognized by abrupt changes in the rate of 

water head drop. Figure 32 shows the calculated permeabilities for the changes in the Sewickley 

Coalbed in BH–1. Figure 32A is a plot of calculated instantaneous permeabilities, while Figure 

32B shows a running average of the same data. 

The data show that as soon as the strata are affected by mining disturbances, the 

permeability increases and water head starts to drop suddenly. The value of this initial 

permeability increase is 700 md. After this initial increase, permeability decreases to about 

100-md levels. This permeability change may indicate that initial reservoir disturbance is due to 

strata or bedding plane shear and fracturing, where permeability initially increases during initial 

movements of the strata. The permeabilities drop to a lower and uniform value (~100 md) after 

that. This permeability persists until about a half-day before the borehole location is undermined 

by the longwall face, which causes permeability to increase mostly due to shearing, horizontal 

fracturing, and vertical fracturing. After the borehole is undermined, permeability increases to 

larger values (>1,000 md) due to larger-scale fractures. 

Figure 32B shows water head change and permeability evolution averaged in some 

distinctive segments during mining for BH–1. This figure also shows that the highest permeabilities measured in the borehole occur just after disturbances first affect the coalbed, during 

initial ground movement, and during undermining of the borehole location. The average fracture 

permeability in the Sewickley horizon during this time interval was calculated as 540 md. 

71
 
 Figure 32.—A: BH–1 instantaneous change in the permeabilities of monitored horizons during mining; 

B: BH–1 average permeabilities at certain intervals of water head drop during mining. (Hw = height of water; 

Ho = initial height of water.) 

BH–2, designed to monitor a 6.1-m (20-ft) section in the shale and limestone zone 

between the Sewickley and Pittsburgh Coalbeds, was analyzed using the same model. Figures 

33A and 33B present the instantaneous and average permeabilities, respectively, within different 

segments identified based on the changes in the rate of water head drop for BH–2. Although this 

borehole is deeper and closer to mining than BH–1, this interval does not show the sudden water 

head loss as observed in the first borehole when mining influence reached the borehole location. 

Instead, the change in the water level is more gradual. This may be due to the combination of 

different structural and mechanical properties of this horizon compared to the BH–1 test horizon, 

which included the Sewickley Coalbed. 

72
 
The data in Figure 33A show that permeabilities in BH–2 gradually increase to the 100– 

200 md range, possibly because of bedding plane movements. A sudden drop in water head 

follows this period with an associated permeability increase to about 600 md. This increase 

coincides with the approach of the longwall face (15 m (50 ft) away) and may suggest 

communication between these two wells. After this permeability increase, the values begin to 

decrease to the 200–300 md range, followed by a gradual increase. The permeability increases 

coincide with the longwall face approach until the borehole location is undermined. After 

undermining, large-scale horizontal and vertical fractures are created, and the permeability 

increases to about >1,000–1,500 md. 

Permeabilities measured prior to undermining were in the ~1-md 

range. Permeability increases following undermining were 

dramatic, with increases of about 100–500 times and instantaneous increases of up to about 1,000 times. 

Figure 33B shows the same observations for BH–2 by presenting the average permeabilities in various segments during water head drop. This figure also shows that permeability 

starts increasing gradually as the mining face advances toward the borehole location until it is 

undermined, after which the permeabilities increase suddenly. This behavior suggests that before 

undermining, the fractures in the formation develop gradually. The average fracture permeability 

in this zone during mining was calculated as 560 md. 

Formation permeabilities gradually increase as mining-induced 

disturbances progress to the borehole locations. However, the 

biggest change occurs with the interception of borehole 

locations. 

Surface Measurements of Borehole Gas 

The change in methane concentrations and static pressures in the boreholes were monitored at the surface after the boreholes were shut in. These concentration and pressure measurements are shown in Figures 34A and 34B, respectively. As of July 2006, all three boreholes were 

intercepted during the week just before the miner’s vacation, when the longwall did not operate. 

The GGV was intercepted after longwall mining resumed. The coal production stoppage associated with the miner’s vacation and the date when the nearest GGV began gob gas production are 

also indicated in Figure 33. 

73
 
 Figure 33.—A: BH–2 instantaneous change in the permeabilities of monitored horizons during mining; 

B: BH–2 average permeabilities. (Hw = height of water; Ho = initial height of water.) 

The monitoring of downhole underwater pressures began before mining disturbances 

reached the borehole locations. This period is characterized by almost constant methane levels 

and static borehole pressures (BH–1) until the mining-related disturbances reach the borehole 

locations. The data show that until shearing occurred, which resulted in an initial water head 

drop, methane concentration was around 75%, which is the initial reading shown. Gas concentration then increased to about 85% within a few days and pressure increased to 20 cm (8 in) of 

water gauge in the borehole. These increases suggest that the coalbed was producing methane 

that was migrating into the borehole. However, when mining disturbances reached the BH–1 

borehole location, concentration decreased to 40% and pressure dropped to 10 cm (4 in) of water 

74
 
gauge until interception (shaded portion in Figure 34), which resulted in a further decrease. 

However, after this borehole location was intercepted by the longwall face and all the water was 

drained out, methane concentration started to increase to 90%–100% because of fracturing of the 

coalbed and the absence of water in the boreholes. 

The interception of BH–2, which monitored shale and limestone layers, showed a 

different behavior. Initial methane readings in this borehole before undermining were around 

10% and the shut-in pressures were low, indicating that there was not significant methane flow 

into the borehole, probably due to low permeabilities and the presence of water in the borehole. 

However, after it was undermined and the water drained out, methane concentration increased to 

90%–100% in the borehole with a sudden pressure fluctuation. This is possibly due to development of horizontal and vertical fractures and the associated permeability increase in the monitoring zone. The loss of water from the boreholes into mining-induced fractures following 

undermining resulted in the end of meaningful transducer pressure measurements and therefore 

the termination of the downhole permeability data. 

It is also interesting to note the similar behavior of methane concentration change in 

BH–1 and BH–2 after undermining. This behavior suggests that these two boreholes monitored 

two different horizons that were about 11 m (35 ft) away from each other, and they started to 

communicate through horizontal and vertical fractures. This observation shows that the layers 

within 24 m (80 ft) of the top of the coalbed being mined are either fractured or there is enough 

shearing or opening of natural fractures that the formations can interact with each other. 

However, the fact that the mine ventilation pressures (−8 to −10 cm (−3 to −4 in) water gauge) 

have not been continuously recorded at these elevations suggests that there was either no direct 

communication with the mine or that the positive gas pressure into the boreholes from the 

monitored zones was high enough to compensate the negative pressure influence of mine 

ventilation. 

The behavior of pressure and concentration data produced by BH–3, an interval about 

6.1 m (20 ft) above the Pittsburgh Coalbed, is completely different from that of the other two 

boreholes. In BH–3, the decrease of methane concentration and shut-in pressure after mining 

showed that this borehole started to communicate with the mine atmosphere. After undermining, 

the shut-in pressure decreased to mine ventilation pressures and stayed as such for the rest of the 

monitoring. Also, methane concentrations decreased to very low values from 35%–45% levels 

before undermining. After undermining, concentration began slowly increasing to about 5% 

methane. Then the GGV exhauster, 76 m (250 ft) away, began to operate with a high flow rate 

following undermining. This resulted in an initial decrease in borehole methane concentration 

followed by a concentration rise in the gob, as recorded in BH–3. 

The start of the nearby GGV operation is also noticeable in the other two BME boreholes 

(Figure 34). Although this venthole did not operate continuously and successfully because of 

mechanical problems after an initially high production rate, the data confirm that the venthole 

was communicating with the monitoring boreholes through fractures and bedding plane separations, as previously observed by Mucho et al. [2000]. The methane concentration increase in the 

gob (or in BH–3) may be due to an increase in methane height in the gob as a result of the 

negative pressure generated by the venthole blower. This rise may also be due to drawing gas 

from other horizons into a more permeable gob following undermining. This may also explain 

the decrease in BH–1 and BH–2 shut-in pressures and the decrease in measured methane 

concentrations associated with a concentration increase in BH–3, before the recovery again to 

70% methane. 

75
 
 Figure 34.—Methane concentration (A) and static shut-in pressures (B) measured in the boreholes during 

progress of mining. 

Downhole pressure and methane concentration data suggest 

that experimental boreholes 1 and 2 (32 and 22 m (100 and 72 ft)), 

respectively, above the mined coalbed) interact and behave 

similarly to overburden conditions. Borehole 3 (7.3 m (24 ft) 

above the mined coalbed) responds to mine ventilation pressure, 

and later to gob pressure conditions. 

76
 
Summary 

•	 The GGV fracture network can form 24–46 m (80–150 ft) ahead of the mining face. This 

distance may vary when conventional GGV slotted pipe configurations are used. Much 

shorter lengths were used during monitoring. 

•	 The loss of water from GGVs can be rapid—from about 32 to 22 m (100 to 22 ft) above 

the mined coalbed. The rate of water loss was not related to borehole depth. 

•	 Overburden permeabilities within the same overburden test zones were in the ~1-md 

range prior to undermining and increased to about 100–500 times following undermining, 

with higher instantaneous peaks. 

•	 The biggest increase in permeability occurred when the longwall face reached the monitoring locations. 

•	 The borehole test interval within the fractured rock responded to overburden gas pressure, and the test interval in the caved zone responded to mine ventilation conditions. 

77
 
PRACTICAL GUIDELINES FOR CONTROLLING LONGWALL COALBED METHANE 

The following practical guidelines are recommended for controlling longwall coalbed 

methane. All predictions are based on determinations made for the Pittsburgh Coalbed in 

southwestern Pennsylvania. The senior authors may be contacted at the NIOSH Pittsburgh 

Research Laboratory for specific cases and application of the methods described. 

•	 It is recommended to use shielding degasification boreholes to decrease emission 

rates by at least 25% for development entries. Drill these boreholes as close as 

practically possible (~27 m (90 ft)) to the entries and operate them for at least 

6 months to achieve a 25%–50% decrease in emission rates (Section 1). 

•	 Equations for predicting methane emission rates into development entries are presented in Table 3 (Section 1). These relationships assume a supercritical longwall 

panel developed in the Pittsburgh Coalbed. 

•	 For panel widths greater than 305 m (1,000 ft), a trilateral borehole configuration is 

recommended for effective draining of methane (Figure 9B). Short across-panel holes 

at close spacing can also be effective at degassing a panel, but not at shielding 

development entries (Figure 9D). The required number of holes will depend on seam 

anisotropic reservoir conditions, but at least 12 holes for a 3,400-m (11,000-ft) long 

panel are recommended to achieve the same degasification as the trilateral configuration (Section 2). 

•	 If less than 12 months is available for premining gas drainage, degasification should 

be continued until the borehole is approached by mining. This approach maximizes 

the quantity of removed methane and reduces methane emission rates (Section 2). 

•	 To avoid shearer coal production delays, it is recommended that continuous GGV 

production be assured while GGVs are within about 500 ft of the working face. In 

many mines, the quantity of coalbed methane removed by a GGV is significant, 

potentially 75% of the volume of gas emissions on the longwall face. A similar 

finding was observed by prior NIOSH research [Mucho et al. 2000] based on tracer 

gas injection. If an operating GGV stops producing gas, the gas that was being 

removed will enter the ventilation system (Section 3). 

•	 Assuming a well-caved gob, increasing the longwall face length by X% will increase 

the rate of methane emissions by, at most, two-thirds of X%. For example, based on 

Pittsburgh Seam experimental data, an increase in longwall face length of 25% will 

increase methane emissions rates by (2/3) × (25%), or 17% (Section 4). 

•	 The length of the slotted casing section of a GGV will strongly influence its level of 

gas production. To effectively design the slotted casing section of a GGV, it is 

recommended to: 

○	 Review the local geology to identify the locations of gas-bearing strata; and 

○	 Set the top of the slotted section at the highest gas-producing stratigraphic horizon 

(Section 6). 

78
 
•	 Completing a GGV into the caved zone is counterproductive (Figures 24–25) and 

increases the likelihood of intermittent production from increased-width, supercritical 

panels. Therefore, the drilling or completion depth of GGVs should be at least 14 m 

(45 ft) above the mined coalbed for longwall panels, particularly in the Northern 

Appalachian Basin. This assumes that the caving zone is six to nine times the seam 

height and production of coalbed gas from the GGV is dominated by the overlying 

and underlying coalbeds (Sections 6 and 7). 

•	 It is recommended that operators emphasize continuous GGV production since it will 

potentially produce 40%–50% more coalbed gas than GGVs operating intermittently 

(Section 7). 

•	 Increasing the longwall panel width increases the quantity of methane present 

because of the increased fractured reservoir volume. However, this increase does not 

enhance the performance of GGVs. As panel width increases, the effectiveness of 

GGVs completed near the tailgate margin will not extend as close to the headgate 

side. Drilling GGVs on the headgate side or near the panel centerline can produce 

coalbed gas from this portion of the panel. It is estimated that the central compressed 

portion of the gob has about one-tenth of the permeability found near the gate road 

tensional stress zone (Sections 5 and 7). 

•	 In favorable cases, GGVs drilled on the headgate side can be effective. Completion 

depths must isolate the borehole from communication with the ventilation network 

(Figure 28). These findings are based on supercritical panel designs (Section 7). 

•	 Mining-induced fracturing was observed to occur 24–46 m (80–150 ft) ahead of 

the mine face. Boreholes and exhausters should be installed before this occurs 

(Section 8). 

•	 Data were reviewed for GGV configurations completed from 7 to 32 m (24 to 106 ft) 

above the mined coalbed for supercritical panels in the Northern Appalachian Basin. 

It is recommended that GGVs be completed toward the top of this interval and be 

designed to include the Sewickley Coalbed. Permeabilities overlying the Pittsburgh 

Coalbed before undermining were very low. Increases in permeability were dramatic 

following undermining, with increases of about 100–500 times the premining values 

and instantaneous increases of up to about 1,000 times these values. These measurements did not differ significantly despite differences in boreholes configurations. 

Fracture permeability pathways remain high to the mined coalbed toward the top of 

the described interval, yet the likelihood of drawing ventilation air into the borehole is 

minimized. The post-undermining borehole pressure data also support this finding 

(Section 8). 

79
 
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