IC 9500 

INFORMATION CIRCULAR/2007 

Explosion Pressure Design 

Criteria for New Seals 

in U.S. Coal Mines 

Department of Health and Human Services

Centers for Disease Control and Prevention

National Institute for Occupational Safety and Health
Information Circular 9500 

Explosion Pressure Design Criteria 

for New Seals in U.S. Coal Mines 

By R. Karl Zipf, Jr., Ph.D., P.E., Michael J. Sapko, and Jürgen F. Brune, 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 

July 2007 
This document is in the public domain and may be freely copied 

or reprinted. 

DISCLAIMER 

Mention of any company or product does not constitute endorsement by the National 

Institute for Occupational Safety and Health (NIOSH). In addition, citations to Web sites 

external to NIOSH do not constitute NIOSH endorsement of the sponsoring organizations 

or their programs or products. Furthermore, NIOSH is not responsible for the content of 

these Web sites. 

ORDERING INFORMATION 

To receive documents or other information about occupational safety and health topics, 

contact NIOSH at 

NIOSH—Publications Dissemination 

4676 Columbia Parkway 

Cincinnati, OH 45226–1998 

Telephone: 1–800–35–NIOSH (1–800–356–4674) 

Fax: 513–533–8573 

e-mail: pubstaft@cdc.gov
 

or visit the NIOSH Web site at www.cdc.gov/niosh

DHHS (NIOSH) Publication No. 2007–144 

July 2007 

SAFER • HEALTHIER • PEOPLETM 
CONTENTS
 

Page 

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

1.0 Introduction ..........................................................................................................................3 

1.1 Report objective.........................................................................................................3 

1.2 Seals and ventilation systems in underground coal mining.......................................3 

1.3 Seal applications and design issues ...........................................................................9 

 1.4 Development of explosive gas and dust accumulations in sealed areas of 

 coal mines.............................................................................................................10 

1.5 Explosions in sealed areas of coal mines.................................................................13 

2.0 Comparison of seal design practices in the United States, Europe, and Australia.............16 

2.1 Origin and evaluation of the 140-kPa (20-psig) seal design criterion in the 

United States.........................................................................................................16 

2.2 Seal design practices in Europe and Australia.........................................................18 

2.2.1 Seal design practices in the United Kingdom............................................18 

2.2.2 Seal design practices in Germany..............................................................20 

2.2.3 Seal design practices in Poland .................................................................20 

2.2.4 Seal design practices in Australia..............................................................20 

3.0 Explosion chemistry and physics .......................................................................................24 

3.1 The 908-kPa (132-psia) constant volume explosion pressure.................................24 

3.2 Effect of coal dust on explosion pressure................................................................26 

3.3 Homogeneous methane-air mixtures in sealed-area atmospheres...........................28 

3.4 Explosions in tunnels...............................................................................................28 

3.5 Static, dynamic, and reflected pressure from explosions in tunnels........................31 

3.6 The 1.76-MPa (256-psia) Chapman-Jouguet (CJ) detonation wave pressure.........33 

3.7 The 4.50-MPa (653-psia) reflected detonation wave pressure................................35 

 3.8 Possible higher detonation and reflected shock wave pressures .............................36 

3.9 Measured experimental mine explosion pressures ..................................................36 

3.10 Summary of main parameters affecting gas explosion strength..............................37 

4.0 Modeling explosion pressures on seals ..............................................................................39 

4.1 Model characteristics...............................................................................................39 

4.2 Model calibration.....................................................................................................41 

4.3 Confined explosion models of large gas cloud volumes .........................................44 

4.4 Partially confined explosion models of leaking seals..............................................44 

5.0 Design pressure-time curves for seals ................................................................................50 

6.0 Minimum new seal designs to withstand the design pressure-time curves ........................59 

6.1 Dynamic structural analysis with WAC ..................................................................61 

6.2 Quasi-static analysis with a plug formula and Anderson’s arching formula...........63 

6.3 Design charts for minimum seal thickness..............................................................64 

6.4 Additional structural requirements for new seals....................................................67 

6.5 Alternative structural analyses of new seals............................................................68 

7.0 Summary of procedures for new seal design......................................................................68 

7.1 Two approaches to sealing mined-out areas............................................................68 

7.2 Design, construction, and inspection for new sealed areas .....................................69 

7.3 New research and development in seal design ........................................................71 

8.0 References ..........................................................................................................................72 
ILLUSTRATIONS 

Page 

1. Flowchart for selecting design pressure-time curve for new seals ...........................................2 

2. Typical layout of room-and-pillar mine using bleeders in ventilation system .........................5 

3. Typical layout of room-and-pillar mine using bleederless ventilation system.........................6 

4. Typical layout of longwall mining with delayed panel sealing ................................................7 

5. Typical layout of longwall mining with immediate panel sealing............................................8 

6. Coward triangle for explosive zone of methane in air............................................................11 

7. Three general types of explosive gas accumulation within sealed areas................................13 

8. Continuous atmospheric gas monitoring system in Australia.................................................22 

9. Tomlinson boiler for inertization at an Australian coal mine.................................................23 

10. Variation of absolute pressure versus methane concentration: theoretical and experi- 

 mental determinations..........................................................................................................25 

11. Variation of absolute pressure for methane-air and coal dust-air...........................................27 

12. Four stages of combustion process in a closed tunnel and the approximate pressures...........30 

13. Strong positive feedback loop between pressure increase, turbulence, and combustion rate ...31 

14. Variation of theoretical pressure increase ratio versus methane concentration for CV 

 explosion pressure, detonation wave pressure, and reflected detonation wave pressure .......38 

15. Layout of calibration models in the LLEM ............................................................................41 

16. Calibration experiments and calculations compared ..............................................................43 

17. Layout of large-volume confined explosion models ..............................................................45 

18. Calculated pressure-time histories at seal for large-volume explosions using AutoReaGas

 and FLACS..........................................................................................................................46 

19. Peak explosion pressure versus run-up length........................................................................47 

20. Layout of partially confined, partially filled explosion models..............................................48 

21. Calculated pressure-time histories at seal for “leaking seal” explosion models using 

AutoReaGas and FLACS ....................................................................................................49 

22. Peak explosion pressure versus volume size behind leaking seal: calculations and 

 experimental measurements ................................................................................................51 

23. 4.4-MPa (640-psig) design pressure-time curve and typical model calculations ...................52 

24. 800-kPa (120-psig) design pressure-time curve and typical model calculations....................53 

25. 345-kPa (50-psig) design pressure-time curve and typical model calculations......................54 

26. Illustration of design pressure-time curve application for new seal construction ..................57 

27. One-way arching failure mechanism in Wall Analysis Code (WAC)....................................62 

28. Typical resistance function for unreinforced wall with one-way arching ..............................62 

29. Design chart for minimum seal thickness with 4.4-MPa (640-psig) design pressure-time 

 curve using various construction materials .........................................................................65 

30. Design chart for minimum seal thickness with 800-kPa (120-psig) design pressure-time 

 curve using various construction materials .........................................................................66 

31. Design chart for minimum seal thickness with 345-kPa (50-psig) design pressure-time 

 curve using various construction materials .........................................................................67 
TABLES 

Page 

1. Design considerations and characteristics for each seal type ...................................................9 

2. Summary of known explosions in sealed areas of U.S. coal mines, 1986–2006....................14 

3. Comparison of worldwide seal designs: construction and related practices...........................19 

4. Characteristics of LLEM experiments for gas explosion model calibration ..........................41 

5. Comparison of calculated to measured LLEM experimental gas explosion pressures in 

 single-tunnel (D-drift) tests.................................................................................................42 

6. Comparison of calculated to measured LLEM experimental gas explosion pressures in 

 multiple-tunnel (A-, B-, and C-drift) tests ..........................................................................42 

7. Technical requirements for the recommended pressure-time curves for structural design 

 of new seals in different conditions ....................................................................................56 

8. Typical material properties for seal construction....................................................................60 

ACRONYMS AND ABBREVIATIONS USED IN THIS REPORT 

CFD computational fluid dynamics 

CFR Code of Federal Regulations 

 CH4 methane 

CJ Chapman-Jouguet 

 CO2 carbon dioxide 

CV constant volume 

DDT deflagration-to-detonation transition 

 FLACS Flame Acceleration Simulator (Gexcon AS) 

LLEM Lake Lynn Experimental Mine 

MSHA Mine Safety and Health Administration 

N2 nitrogen 

 NASA National Aeronautics and Space Administration 

 NIOSH National Institute for Occupational Safety and Health 

O2 oxygen 

PRL Pittsburgh Research Laboratory (NIOSH) 

SDOF single degree of freedom 

S.G. specific gravity 

TNT trinitrotoluene 

UCS uniaxial compressive strength 

WAC Wall Analysis Code 
UNIT OF MEASURE ABBREVIATIONS USED IN THIS REPORT 

cm centimeter 

 cm

2

 square centimeter 

 cm3 cubic centimeter 

 cm/s centimeter per second 

 ft foot

 ft/s foot per second 

g gram

 g/cm3  gram per cubic centimeter 

 g/m3 

 gram per cubic meter 

hr hour 

 in inch

 in2 

 square inch 

K kelvin

 kg kilogram

kg/m3 

 kilogram per cubic meter 

km kilometer 

 km2 

 square kilometer 

kPa kilopascal 

L liter

 m meter

m

2

 square meter 

m3 cubic meter 

 min minute

 mm millimeter 

 mm

2

 square millimeter 

 MPa megapascal 

 ms millisecond

 m/s meter per second 

 pcf pound per cubic foot 

psi pound-force per square inch 

psia pound-force per square inch absolute 

psig pound-force per square inch gauge 

sec second 

 yd3 

 cubic yard 

 °C degree Celsius 
EXPLOSION PRESSURE DESIGN CRITERIA 

FOR NEW SEALS IN U.S. COAL MINES 

By R. Karl Zipf, Jr., Ph.D., P.E.,1

 

1

Senior research mining engineer. 

 Michael J. Sapko,2

2

Principal research physical scientist (retired). 

3

and Jürgen F. Brune, Ph.D.3 

Principal research mining engineer. 

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

EXECUTIVE SUMMARY 

Seals are barriers constructed in underground coal mines throughout the United States to 

isolate abandoned mining panels or groups of panels from the active workings. Historically, 

mining regulations required seals to withstand a 140-kPa (20-psig) explosion pressure. However, 

the Mine Improvement and New Emergency Response Act (“MINER Act”) requires the Mine 

Safety and Health Administration (MSHA) to increase this design standard by the end of 2007. 

This report provides a sound scientific and engineering justification to recommend a three-tiered 

explosion pressure design criterion for new seals in coal mines in response to the MINER Act. 

Much of the information contained in this report also applies to existing seals. 

Engineers from the National Institute for Occupational Safety and Health (NIOSH) 

examined seal design criteria and practices used in the United States, Europe, and Australia and 

then classified seals into their various applications. Next, the engineers considered various kinds 

of explosive atmospheres that can accumulate within sealed areas and used thermodynamic 

calculations and simple gas explosion models to estimate worst-case explosion pressures that 

could impact seals. Three design pressure-time curves were developed for the dynamic structural 

analysis of new seals under the conditions in which those seals may be used: unmonitored seals 

where there is a possibility of methane-air detonation or high-pressure nonreactive shock waves 

and their reflections behind the seal; unmonitored seals with little likelihood of detonation or 

high-pressure nonreactive shock waves and their reflections; and monitored seals where the 

amount of potentially explosive methane-air is strictly limited and controlled. Figure 1 is a 

simple flowchart that illustrates the key decisions in choosing between the monitored or 

unmonitored seal design approaches and the three design pressure-time curves. 

For the first condition, an unmonitored seal with an explosion run-up length of more than 

50 m (165 ft), the possibility of detonation or high-pressure nonreactive shock waves and their 

reflections exists. The recommended design pressure-time curve rises to 4.4 MPa (640 psig) and 

then falls to the 800-kPa (120-psig) constant volume (CV) explosion overpressure. For 

unmonitored seals with an explosion run-up length of less than 50 m (165 ft), the possibility of 

detonation or high-pressure nonreactive shock waves and their reflections is less likely. A less 

severe design pressure-time curve that simply rises to the 800-kPa (120-psig) CV explosion 

overpressure may be employed. For monitored seals, engineers can use a 345-kPa (50-psig) 

design pressure-time curve if monitoring can ensure that (1) the maximum length of explosive 

mix behind a seal does not exceed 5 m (16 ft) and (2) the volume of explosive mix does not 

exceed 40% of the total sealed volume. Use of this 345-kPa (50-psig) design pressure-time curve 

requires monitoring and active management of the sealed area atmosphere. These design 

pressure-time curves apply to new seal design and construction. 
Figure 1.—Flowchart for selecting design pressure-time curve for new seals. 

NIOSH engineers used these design pressure-time curves along with the Wall Analysis 

Code (WAC) from the U.S. Army Corps of Engineers and a simple plug analysis to develop 

design charts for the minimum required seal thickness to withstand each of these explosion 

pressure-time curves. These design charts consider a range of practical construction materials 

used in the mining industry and specify a minimum seal thickness given a certain seal height. 

Results of these analyses show that resistance to even the 4.4-MPa (640-psig) design pressuretime curve can be achieved using common seal construction materials at reasonable thickness, 

demonstrating the feasibility and practical applications of this report. Engineers can also use 

other structural analysis programs to analyze and design seals by using the appropriate design 

pressure-time curve for the structural load and a design safety factor of 2 or more. Finally, this 

report also provides criteria for monitoring the atmosphere behind seals. 

NIOSH will continue research efforts to improve underground coal mine sealing 

strategies and to prevent explosions in sealed areas of coal mines. In collaboration with the 

U.S. National Laboratories, NIOSH will further examine the dynamics of methane and coal dust 

explosions in mines and the dynamic response of seals to these explosion loads. This upcoming 

project seeks to better understand the detonation phenomena and simple techniques to protect 

seals from transient pressures. Additional work will include field measurements of the 

atmosphere within sealed areas. Successful implementation of the seal design criteria and the 

associated recommendations in this report for new seal design and construction should 

significantly reduce the risk of seal failure due to explosions in abandoned areas of underground 

coal mines. 

2
 
 

1.0 Introduction 

1.1 Report Objective 

Seals are used in underground coal mines throughout the United States to isolate 

abandoned mining areas from the active workings. Prior to the Sago Mine disaster in 2006, 

mining regulations required seals to withstand a 140-kPa (20-psig) explosion pressure 

(30 CFR4

 75.335(a)(2)). 

4

Code of Federal Regulations. See CFR in references. 

However, Program Information Bulletin No. P06–16 issued by MSHA 

on July 19, 2006 [McKinney 2006], requires seals to withstand a 345-kPa (50-psig) explosion 

pressure. The recently enacted MINER Act requires MSHA to increase this design standard by 

the end of 2007. This report provides a sound scientific and engineering justification to 

recommend a three-tiered explosion pressure design criterion for new seals in coal mines in 

response to the MINER Act. The recommendations contained herein apply to new seal design 

and construction in U.S. coal mines. 

1.2 Seals and Ventilation Systems in Underground Coal Mining 

To control methane in mined-out areas of coal mines and thereby reduce explosion risk 

from methane buildup, current mining regulations (30 CFR 75.334) require companies to either 

ventilate or seal those areas. Continued ventilation of abandoned areas is costly and may divert 

ventilating air away from other, more productive uses. Sealing is sometimes a more economical 

and possibly a safer alternative to ventilation. The use of seals may have contributed to fewer 

explosions in active mine areas. Without sealing, large mined-out areas still require regular 

inspections and can expose miners to a variety of underground hazards. 

A ventilation system delivers fresh air to the mains, submains, gate road entries, 

production panels, and all the active areas of the mine via intake airways, while return airways 

remove contaminated air laden with dust and methane. Various ventilation control devices, 

namely stoppings, overcasts, and regulators, control and direct the airflow throughout the system. 

Fans located on the surface provide the power to move the required air quantity. In addition to 

the primary ventilation system for providing air to all the active mining faces, bleeder entries 

located around the perimeter of mining areas serve to dilute methane from all mined-out areas 

long after panels are extracted. 

When an area of an underground coal mine is mined out, operators will often choose to 

isolate the abandoned area with simple barriers called seals rather than continue to ventilate the 

area. Seals are walls constructed from solid, incombustible materials such as poured concrete, 

concrete blocks, cementitious foams, and other materials that separate abandoned panels or 

groups of panels from the active areas of the mine. MSHA data indicate that over 14,000 seals in 

over 2,200 sets exist in active coal mines throughout the United States. Historical estimates 

suggest that mining companies or their contractors built several thousand seals annually. 

In active mining, primary access to production areas occurs via a system of mains and 

submains corridors. These corridors contain the ventilation system and a conveyor system to 

remove the mined coal. Production panels are developed from these corridors. An entry is a coal 

mine tunnel driven parallel to the direction of advance or the long axis. A crosscut is a coal mine 

tunnel perpendicular to an entry. 

3
 
For room-and-pillar mining, as shown in Figures 2–3, mining companies typically 

develop 5–11 entries plus the crosscuts to mine a panel. The pillars created during advance 

mining may be extracted completely during retreat mining. A room-and-pillar system may or 

may not use “bleeders” along the outer perimeter of the panel as part of its ventilation system to 

remove methane gas from the mined-out areas. Figure 2 shows a typical layout with bleeders, 

which is the more common practice, while Figure 3 shows a typical bleederless room-and-pillar 

layout. Bleederless systems are sometimes applied when spontaneous combustion is a potential 

problem for the mine. For longwall mining, as shown in Figures 4–5, coal companies will 

typically mine a three-entry gate road system off the mains or submains to develop a longwall 

panel. The entire coal block is then extracted using retreat longwall mining. 

Once a panel or a group of panels in a mining district has been mined out, seals may be 

constructed. The sealed area may be completely open, as in the case of certain room-and-pillar 

panels, or the sealed area may be completely or partially filled with gob such as a longwall panel. 

Gob is caved rock left in the wake of full-extraction coal mining by either the longwall or roomand-pillar mining method. Depending on mining conditions, operators might seal individual 

room-and-pillar panels, individual longwall panels, or groups of panels in mining districts. 

Sealing an individual room-and-pillar panel might entail construction of multiple seals at the 

mouth and bleeder ends of the panel. Sealing several adjacent panels may occur later. Finally, 

sealing the entire room-and-pillar panel district might occur with the construction of multiple 

seals across mains, submains, and bleeder entries at a judicious location (Figure 2). When using a 

bleederless ventilation system, sealing of individual room-and-pillar panels and districts occurs 

in a similar manner, but fewer seals are required (Figure 3). 

4
 
 Figure 2.—Typical layout of room-and-pillar mine using bleeders in ventilation system. Also shown are 

typical locations for district and panel seals. 

5
 
 Figure 3.—Typical layout of room-and-pillar mine using bleederless ventilation system. Also shown are 

typical locations for district and panel seals. 

6
 
 Figure 4.—Typical layout of longwall mining with delayed panel sealing. Also shown are typical locations 

for district and panel seals. 

7
 
Sealing mined-out longwall panels is similar to room-and-pillar mining. Multiple seals 

may be constructed at the mouth and bleeder ends of the panel after a longwall panel is mined 

out and the tailgate is no longer needed. A mined-out longwall panel district may then be closed 

off by constructing seals across mains, submains, and bleeders at the proper location. This type 

of sealing is referred to as “delayed panel sealing” and is common where there is low risk of 

spontaneous combustion (Figure 4). Where spontaneous combustion is a potential problem, 

mining companies may decide to seal a longwall panel during retreat mining, called “immediate 

panel sealing” (Figure 5). In this case, seals are constructed in every crosscut nearest the gob 

down the headgate entries immediately behind the longwall face. The newly formed mined-out 

area is substantially isolated from oxygen soon after mining, thereby decreasing the risk of 

spontaneous combustion problems. Depending on the length of the longwall panel, 50–100 seals 

might be constructed as the panel is mined. 

8
 

 Figure 5.—Typical layout of longwall mining with immediate panel sealing. Also shown are typical locations for district, panel, and crosscut seals. 
1.3 Seal Applications and Design Issues 

In developing design criteria for seals, engineers must consider the seal application and 

the conditions created by those applications. Different explosion pressures and other forces that 

may act on seals in various applications should influence their design. There are four seal applications with unique characteristics: (1) panel, (2) district, (3) crosscut, and (4) fire. Figures 2–5 

illustrate the first three seal applications. Fire seals will not be considered in this report. 

For each seal application, there are three conditions to consider: (1) explosion-loading 

potential, (2) convergence loading potential, and (3) leakage potential. The explosion-loading 

potential depends mainly on the volume and geometry of the mined-out area behind the seal. 

Larger sealed volumes with longer propagation distances can lead to higher gas and coal dust 

explosion pressures. The roof and floor convergence loading potential depends mainly on the 

proximity of the seals to mined-out areas. Seals located close to fully extracted longwall or 

room-and-pillar panels are more likely to experience damage due to excessive convergence. 

Finally, the leakage potential of a seal depends on the ventilation system as well as any damage 

to the seal and surrounding rock caused by convergence loading. Seals located in areas of high 

pressure differential in the ventilation system will have greater potential for leakage of either 

fresh air into the sealed area or potentially explosive methane out from the sealed area. The level 

of each of these conditions by seal type is summarized in Table 1. 

Table 1.—Design considerations and characteristics for each seal type 

Seal 

type 

Explosion 

loading 

potential 

Convergence 

loading 

potential 

Ventilation

pressure

differential 

Leakage

potential

District ................... Very large ............ Low...................... High.................... Moderate.
 

Panel ..................... Large.................... Moderate ............. Moderate............ Moderate.
 

Crosscut ................ Small.................... High ..................... Low .................... High.

 (1) Room-and-pillar panel seals or longwall gate road seals (Figures 2–5) are the first 

seal application. These seals are constructed soon after a panel’s abandonment at the mouth and 

bleeder ends of a room-and-pillar panel or longwall panel on the tailgate side. Hundreds of 

meters of open entry are likely behind the seals and around the periphery of a room-and-pillar 

panel. In a longwall gate road, while the outer gate entries probably cave in after mining, the 

inner entries may remain open for 3 km (10,000 ft) or more in larger mines. The length of open 

entry behind these seals can lead to a potentially large volume of explosive mix, in turn creating 

a high explosion-loading potential. Panel seals have a moderate level of convergence loading. 

They also have a moderate leakage potential due to the possibility of damage from ground 

pressure and higher pressure differential from the ventilation system. Judicious placement of the 

seals, however, can minimize the risk of ground pressure and therefore of damage to the seal and 

the resulting leakage. 

(2) District seals (Figures 2–5) are the second application and possibly the most common 

seal application. These seals are constructed at strategic locations to remove groups of room-andpillar or longwall panels from the ventilation system. In large room-and-pillar or longwall 

mining situations, the entries behind the seals most likely remain open for distances of hundreds 

of meters, and the potential volume of explosive mix behind these seals may fill several large 

panels. The large volume of explosive mix contributes to a very large explosion-loading 

9
 
potential. Convergence loading is likely to be low given the distance of the seals from the minedout areas. Leakage potential of district seals is again moderate due to the low convergence 

loading but the high ventilation pressure differential. 

(3) Longwall gate road crosscut seals (Figure 5) may be constructed if the spontaneous 

combustion potential for the coal is high, necessitating the isolation of the mined-out areas from 

oxygen as soon as possible. These seals are constructed simultaneously with the retreating 

longwall face in the crosscut nearest the gob in headgate entries. Open area behind these seals is 

small, making the potential volume of explosive mix and the explosion-loading potential also 

small. Crosscut seals are likely, however, to have high convergence loading and therefore to 

become damaged. Despite low ventilation pressure differential, the high convergence loading 

contributes to high leakage potential.

 (4) Fire seals are used to isolate a fire from the ventilation system and may be located 

anywhere in a mine layout. Fire seals have the unique requirement that they must develop their 

design strength quickly; a cure time of less than 1 day is preferable. Fire seals are mentioned here 

for completeness, but will not be considered further in this report. 

1.4 Development of Explosive Gas and Dust Accumulations in Sealed Areas 

of Coal Mines 

Ventilation is maintained in mined-out areas during seal construction up to the point of 

final seal completion. Upon sealing, the typical coal mine atmosphere contains about 21% 

oxygen and 79% nitrogen and less than 1% methane. When ventilation to the abandoned area 

ceases, composition of that atmosphere will begin to change depending on the geologic 

characteristics of the coal. Some coals will slowly oxidize and therefore remove oxygen and 

release carbon dioxide into the atmosphere of the abandoned area. However, with few 

exceptions, all underground coalbeds liberate methane, and thus the methane concentration 

within the sealed areas will increase. Methane is explosive in air when the concentration ranges 

from 5% to 16% by volume [Cashdollar et al. 2000]. Most sealed areas will eventually enter this 

explosive range at some point in time after sealing. Methane will continue to accumulate in the 

sealed area; when the concentration exceeds 16%, that atmosphere is no longer explosive. The 

time required for the atmosphere in the sealed area to pass beyond the upper explosive limit and 

become inert ranges from about 1 day to several weeks or more depending on the mine’s 

methane liberation rate. Shallow mines, mines that have breached into old workings, and sealed 

areas with high leakage rates may never become inert. 

To illustrate the development of explosive gas accumulations in sealed areas and the selfinertization process, mine ventilation engineers use the Coward diagram shown in Figure 6 

[Coward and Jones 1952]. The range of explosive methane-oxygen mixes is shown in red; 

this explosive zone is referred to as the “Coward triangle.” Inert mixtures of methane-oxygen are 

shown in green. A sealed area atmosphere always starts at point A upon sealing, which is about 

21% oxygen and 0% methane. A desirable sealed area atmosphere from a safety perspective is 

fuel-rich and oxygen-low, which is more than 20% methane and less than 10% oxygen. Points C 

and F lie within this fuel-rich and oxygen-low inert region. 

10
 
 Figure 6.—Coward triangle for explosive zone of methane in air [Coward and Jones 1952]. Also shown are 

different paths to an inert atmosphere. 

The inertization path depends on the rate of methane emission relative to the rate of 

oxygen depletion in the gob atmosphere. For coal that emits methane but does not oxidize, the 

sealed area atmosphere follows path A–B. With some coal oxidation, which produces carbon 

dioxide and decreases the oxygen content, the atmosphere may follow path A–C toward an inert 

condition. A highly oxidizing coal, such as one prone to spontaneous combustion, may follow 

path A–D and reach an inert condition that is fuel-lean and oxygen-low. In this unique case, the 

sealed area atmosphere does not cross through the explosive zone on the path to inertization. 

The critical time occurs when the sealed area atmosphere crosses through the explosive 

zone shown in red in Figure 6. In some mines, the time that the sealed area atmosphere spends 

within the explosive zone may be as little as a day or two. However, in other mines, the time 

11
 
within the explosive zone can be several weeks to many months, or the sealed area atmosphere 

may never leave the explosive zone. 

During the time the sealed area contains a volume of explosive mix, any ignition source 

could initiate an explosion. Therefore, the normal sealing practice can create an explosive gas 

accumulation until the sealed area atmosphere either self-inerts naturally or becomes inert 

artificially via engineered procedures, such as the injection of inert gas. 

Based on the types of seals and the mining methods shown schematically in Figures 2–5, 

NIOSH researchers have identified three types of explosive gas accumulation that can form 

within a sealed area. Figures 7A and 7C show two types of explosive gas accumulation that can 

occur as a result of normal sealing practice. The first type of explosive gas accumulation is a 

large volume that is completely filled with explosive mix and is completely confined with no 

possible venting (Figure 7A). This situation arises behind district and panel seals some time after 

sealing during the inertization phase. Because the explosive mix is confined with no venting, if it 

ignites, there is no place for the expanding gases to go, and significant pressure increases within 

the sealed area will result. 

Even after a large sealed area has become inert as a result of methane concentration 

above the upper explosive limit, oxygen depletion from coal oxidation, or artificial inertization, 

sealed areas continue to present explosion hazards because air leakage around seals can create an 

explosive atmosphere around the perimeter of the sealed area. During periods of falling 

atmospheric pressure, sealed areas tend to outgas and leak potentially explosive methane gas into 

the mine ventilation system. The active-mine side of seals must therefore have sufficient airflow 

to dilute this methane influx. During periods of rising atmospheric pressure, however, oxygenladen air tends to leak into sealed areas and can create a volume of potentially explosive mix 

immediately behind the seals. In addition, the mine ventilation system itself can create a pressure 

differential across a sealed area, leading to leakage into one set of seals and leakage out of 

another set. This second type of explosive gas accumulation caused by leaking seals is depicted 

in Figure 7B. The explosive mix is partially confined and can vent into a large reservoir of inert 

atmosphere. This situation can arise behind any kind of seal, district, panel, or crosscut. If an 

ignition occurs, significant pressure increases are still possible. 

The third type of accumulation is a completely filled but partially confined and partially 

vented volume (Figure 7C). This kind of accumulation develops behind panel or crosscut seals 

adjacent to a fully extracted longwall or room-and-pillar panel either during initial selfinertization or subsequent air leakage. These seals are most often constructed close to the broken 

rock of the mined-out area (the gob). If accumulated gas ignites, the expanding gases can vent to 

some extent into the inert gob. Nevertheless, large pressure increases within the sealed area 

remain a distinct possibility. 

12
 
13
 

Figure 7.—Three general types of explosive gas accumulation within sealed areas. 

1.5 Explosions in Sealed Areas of Coal Mines 

Since 1986, 12 known explosions have occurred within the sealed areas of active 

U.S. underground coal mines. Table 2 summarizes the known characteristics of these explosions 

including the mine name, year, size of sealed area, damage, cause, suspected ignition source, 

estimated explosion pressure, and reference to relevant reports on the incident. The information 

and conclusions expressed in Table 2 and in this discussion stem from the relevant MSHA 

accident investigation reports and not from separate NIOSH studies. 

The earliest identified explosion within a sealed area occurred in 1986 at the Roadfork 

No. 1 Mine [South 1987]. The mine had recently sealed the area, and the explosion destroyed 

four seals. MSHA investigators suspect that a spark from a roof fall ignited the explosive 

mixture. 
Table 2.—Summary of known explosions in sealed areas of U.S. coal mines, 1986–2006 

Mine Year (Month) Size of 

sealed area 

Damage from 

explosion 

Cause of

explosive mix 

Suspected 

ignition 

source 

Estimated

explosion 

pressure 

Reference

Roadfork No. 1 ..... 1986 .................. Several 4 seals Recently sealed 

 room-and- destroyed. area. 

pillar panels. 

Mary Lee No. 1 ..... 1993 .................. Several 2 seals Leaking seals ........ 

square destroyed

miles. and shaft cap 

displaced. 

Oak Grove No. 1... 1994 .................. Several 3 seals Leaking seals ........ 

square destroyed. 

miles. 

Gary 50................. 1995 .................. Several None; seals Leaking seals ........ 

square survived. 

miles. 

Oak Grove No. 1... 1996 .................. Several 5 seals Leaking seals ........ 

square destroyed. 

miles. 

Oasis .................... 1996 (May) ........ Several 3 seals Leaking seals ........ 

square destroyed. 

miles. 

Oasis .................... 1996 (June) ....... Several Unknown.............. Leaking seals ........ 

square 

miles. 

Oak Grove No. 1... 1997 .................. Several 3 seals Leaking seals ........ 

square destroyed. 

miles. 

Big Ridge Mine 2002 .................. Unknown ............ 1 seal Recently sealed 

Portal No. 2. destroyed. area. 

Sago ..................... 2006 .................. 1 room-and- 10 seals Recently sealed 

pillar panel. destroyed. area. 

Darby .................... 2006 .................. 1 room-and- 3 seals Recently sealed 

pillar panel. destroyed. area. 

Jones Fork E–3 .... 2006 .................. Unknown ............ 3 seals Unknown ............... 

 destroyed. 

Spark from Unknown ........... South [1987]. 

rock fall. 

Lightning ............ 14 kPa Checca and 

(2 psig) Zuchelli 

[1995]. 

Unknown............ Unknown ........... Scott and 

Stephan 

[1997]. 

Lightning 35–48 kPa Sumpter et al. 

or roof (5–7 psig) [1996]. 

fall. 

Lightning ............ <35 kPa Scott and 

(<5 psig) Stephan 

[1997]. 

Lightning <138 kPa Ross and 

or roof (<20 psig) Schultz [1996]. 

fall. 

Lightning Unknown ........... Ross and 

or roof Schultz [1996]. 

fall. 

Lightning ............ >138 kPa Scott and 

(>20 psig) Stephan 

[1997]. 

Unknown............ Unknown ........... Kattenbraker 

[2002]. 

Lightning ............ >642 kPa Gates et al. 

(>93 psig) [2007]. 

Oxygen/ >152 kPa Light et al. 

 acetylene (>22 psig) [2007]. 

torch. 

Unknown............ Unknown ........... Under investigation. 

Blacksville No. 11 .. 1992 .................. Shaft area .......... Shaft cap 

 and collar 

 destroyed. 

Recently sealed 

area. 

Welding 

 sparks. 

6.9 MPa 

(1,000 psig) 

Rutherford et al.

[1993]. 

1

Explosion occurred under shaft cap of abandoned shaft. 

The explosion at the Mary Lee No. 1 Mine in 1993 [Checca and Zuchelli 1995] blew out 

two seals underground and displaced a shaft seal cap by 1 m (3.3 ft). Air leakage around the seals 

may have allowed an explosive mix to develop behind the seals. Production of methane gas from 

the sealed areas via surface boreholes may have increased air leakage through seals and 

contributed to the explosive mix accumulation in the sealed area. Lightning is the suspected 

ignition source. MSHA investigators estimated the explosion pressure at about 14 kPa (2 psig). 

Scott and Stephan [1997] describe explosions at the Oak Grove No. 1 Mine that occurred 

in 1994, 1996, and again in 1997. The first explosion occurred in April 1994 in a sealed area, 

which enclosed approximately 3.5 km2

 (1.35 square miles) of abandoned workings. This 

explosion destroyed 3 of the 38 seals that surrounded the mined-out area. After the explosion, the 

seals were rebuilt to the 140-kPa (20-psig) design standard. In January 1996, a second explosion 

in the sealed area destroyed five additional seals less than 600 m (2,000 ft) from the seals 

destroyed by the 1994 explosion. MSHA investigators estimated the second explosion pressure 

to be less than 35 kPa (5 psig). In July 1997, the third and most violent explosion occurred in the 

14
 
same vicinity as the previous two explosions, and three more seals were destroyed. Scott and 

Stephan [1997] estimated the propagating forces of the explosion to be greater than 140 kPa 

(20 psig). Again, air leakage around the seals may have led to an explosive mix accumulation 

behind the seals. Possible methane production from surface boreholes into the sealed area and 

high ventilation pressure differentials may have exacerbated the air leakage. Lightning seems to 

be the most likely ignition source for all three explosions. 

Sumpter et al. [1996] describe explosions that occurred sometime in 1995 at the Gary 

No. 50 Mine (now called Pinnacle Mine). Once again, air leakage around the seals caused an 

explosive mix to accumulate immediately behind the seals. Surface methane production from 

gob boreholes may have caused air leakage around seals and the development of an explosive 

mix. Several ignition sources are suspected including lightning, a roof fall, or metal-to-metal 

contact. MSHA investigators estimated the explosion pressure to be about 35–48 kPa (5–7 psig). 

Two explosions within sealed areas occurred at the Oasis Mine in 1996 [Ross and Schultz 

1996]. In May 1996, mine personnel noted an unusual spike on the fan pressure recording chart. 

Inspection of the mine revealed three destroyed seals and one damaged seal, along with elevated 

levels of carbon monoxide gas. A second explosion occurred in June 1996. Mine personnel noted 

smoke coming from an exhaust shaft and another spike on the fan pressure recording chart. 

Underground damage information was not collected. Lightning is a suspected ignition source in 

both explosions. The mine was idle at the time of both explosions. MSHA investigators 

estimated the pressure from the first explosion to be less than 138 kPa (20 psig), but pressure 

from the second explosion is unknown. 

An explosion occurred in 2002 within a recently sealed area at the Big Ridge Mine Portal 

No. 2, destroying one seal [Kattenbraker 2002]. Little else is known about the incident. 

Official MSHA accident investigations of the explosions at the Sago Mine [Gates et al. 

2007] and Darby Mine [Light et al. 2007] that occurred in 2006 have been completed. In each 

case, explosions occurred within the sealed area, which caused the catastrophic failure of seals. 

Recent MSHA inspections of the Jones Fork E–3 Mine found evidence of an explosion within a 

sealed area; however, there were no injuries associated with the event. 

The Blacksville No. 1 explosion in 1992 [Rutherford et al. 1993] differs from the other 

methane explosions described in this report in that an explosive mixture of methane-air 

developed in an unventilated, abandoned shaft under a shaft cap. Sparks from welding ignited the 

accumulated explosive mixture. The explosion pressure destroyed the reinforced concrete shaft 

cap and the shaft collar. Based on the observed structural damage, MSHA investigators 

calculated the explosion pressure necessary to cause the damage. Back-calculated pressures 

ranged from 5.4 to 20.8 MPa (782 to 3,016 psig). Rutherford et al. [1993] concluded that the 

maximum internal pressure that developed in the Blacksville shaft was around 6.9 MPa 

(1,000 psig). 

In summary, several documented explosions within sealed areas that destroyed seals 

occurred during 1986–2006 prior to the Sago Mine disaster. Significant accumulations of 

methane-air mix behind the seals led to the explosions. Investigators could not always 

conclusively determine the ignition source, although lightning was suspected in several 

instances. Explosion pressure could not be determined reliably for many of the sealed area 

explosions; however, one explosion was reported to have produced around 6.9 MPa (1,000 psig). 

At this time, no data are available on explosions within sealed areas that happened prior 

to 1986. Nagy [1981] documents 18 major explosions in underground coal mines that occurred 

during 1958–1977 and another 52 smaller explosions during 1970–1977. Review of the ignition 

15
 
source from all of these explosions indicates that all occurred in the active areas of the mine. It is 

not known if any explosions occurred within sealed areas. 

The number of explosions in the 1990s and 2000s may correlate with a trend toward 

more sealing by the U.S. underground coal mining industry. Unfortunately, quantitative data on 

the number of seals constructed annually do not exist in the record. Mitchell [1971] notes that 

prior to World War II, sealing unused and abandoned areas was a common practice. He also 

states that the few seals built during 1945–1970 were mainly in mines with high spontaneous 

combustion potential, implying a decline in the overall use of seals during this time period. 

Passage of the Federal Coal Mine Health and Safety Act of 1969, which required mines to either 

ventilate or seal all areas with “explosion-proof bulkheads,” may have contributed to an increase 

in the use of seals since 1969. Increased underground coal production may have also contributed 

to an increase in sealing. 

2.0 Comparison of Seal Design Practices in the United States, Europe, and 

Australia 

2.1 Origin and Evolution of the 140-kPa (20-psig) Seal Design Criterion in the 

 United States 

The earliest known engineering standard for seals in U.S. underground coal mines is a 

1921 regulation for sealing connections between coal mines located on U.S. government-owned 

lands. Rice et al. [1931] stated that this regulation required seals to withstand a pressure of 

345 kPa (50 psig) and that it was “based on the general opinion of men experienced in mineexplosion investigations.” Evidently, the intent of the regulation was to prevent an explosion in 

one mine from propagating to a neighboring mine. Sealing a mined-out, abandoned area may 

have been a secondary consideration. Rice et al. [1931] provided engineering designs for seals to 

meet the 345-kPa (50-psig) criterion, along with test results to substantiate the designs. 

The 140-kPa (20-psig) criterion for “explosion-proof” seals in the United States 

originates from D. W. Mitchell’s 1971 work entitled Explosion-Proof Bulkheads: Present 

Practices. Mitchell developed what became the 140-kPa (20-psig) design standard in response to 

needs of the Federal Coal Mine Health and Safety Act of 1969. This act required mined-out areas 

to be ventilated or sealed with “explosion-proof bulkheads” that were to be constructed with 

“solid, substantial, and incombustible materials.” The original act required the bulkhead “to 

prevent an explosion which may occur in the atmosphere on one side from propagating to the 

atmosphere on the other side” [Mitchell 1971]. 

It seems that prior to 1970, mining engineers believed that sealed areas required 

protection from explosions originating in the active mining area that would breach the seals and 

flood the active workings with toxic or flammable gases. Mitchell reports on work at the 

Bruceton Experimental Mine done by Rice in the 1930s. Rice found that a weak stopping with 

rock dust barriers on both faces would prevent flame propagation into the sealed area even 

though the stopping was destroyed. Mitchell did not consider the possibility of an explosion 

originating within the sealed area that could rupture the seals and destroy the active mining area 

through blast effects or with toxic gases. It was commonly believed that sealed areas were inert 

with methane concentrations far above the 16% upper explosive limit. 

Mitchell reviewed seal design standards and practices used in the United States, the 

United Kingdom, Germany, and Poland. In the United Kingdom, commissions investigating 

various coal mine explosions assumed that pressures of 140–345 kPa (20–50 psig) could develop 

16
 
and therefore a 345-kPa (50-psig) standard would provide an adequate safety margin for seals. In 

Germany and Poland, authorities decided that seals should withstand 500 kPa (73 psig) based on 

observations from moderate-strength experimental coal mine explosions. 

Mitchell also considered the hundreds of test explosions conducted in the Bruceton 

Experimental Mine from 1914 through the 1960s. Most explosions developed from 7 to 876 kPa 

(1 to 127 psig), although a few tests developed higher pressures that caused considerable 

damage, which were unrecordable with existing sensors. Mitchell noted that more than 60 m 

(200 ft) from the origin of an explosion of a small amount of explosive mix in 15 m (50 ft) of 

entry, the explosion pressures seldom exceeded 140 kPa (20 psig). Most sealed areas are far from 

the active mining areas, so Mitchell concluded that a seal may be considered “explosion-proof” 

if it is designed to withstand a static load of 140 kPa (20 psig). Again, this conclusion is derived 

from the perspective of containment of an explosion of a limited amount of explosive 

atmosphere on the active mining side. It does not consider the containment of an explosion 

within the sealed area. Mitchell reasoned that explosions from the active mining side will usually 

occur far enough away from seals such that a 140-kPa (20-psig) design standard would provide 

the desired protection. 

In addition, Mitchell examined the hazard of explosive methane gas leakage into the 

active mine atmosphere from sealed areas, which can occur during periods of falling barometric 

pressure. The additional methane drainage into the active workings could exceed the capacity of 

the ventilation system and result in an explosion hazard somewhere in the mine. However, 

Mitchell did not consider the opposite hazard created when air leaks from the active atmosphere 

into a sealed area to form an explosive mix behind the seals. 

Prior to 1992, the Code of Federal Regulations lacked a definitive engineering design 

specification for explosion-proof seals. 30 CFR 75 stated that pending the development and 

publication of more specific design criteria for explosion-proof seals or bulkheads, such seals or 

bulkheads may be constructed of solid, substantial, and incombustible material such as concrete, 

brick, cinder block, etc. Stephan [1990] sought to provide technical justification for 140 kPa (20 

psig). Based on investigations of underground coal mine explosions in active workings between 

1977 and 1990, he concluded that the explosion pressure on seals generally does not exceed 140 

kPa (20 psig). Thus, the explosion pressure performance criterion for seals became 140 kPa (20 

psig) in the 1992 rule change to 30 CFR 75.335(a)(2). NIOSH researchers also note that the CFR 

states this criterion as a “static horizontal pressure” of 140 kPa (20 psig). 

The Stephan [1990] report also recognizes that the abandoned areas can contain an 

explosive methane-air mix as the atmosphere crosses through the flammable range in the process 

of self-inertization. Stephan clearly warns that “a seal constructed to withstand an explosion 

pressure wave of 140 kPa (20 psig) may not be sufficient in these cases.” Stephan also 

recognizes that air leakage through seals can lead to an explosive mix accumulation behind seals 

and that potential ignition sources always exist, such as roof falls or spontaneous combustion. 

In summary, the original 140-kPa (20-psig) design criterion for seals is not based on 

containment of an explosion within the sealed area. The criterion apparently stems from the 

belief that the atmosphere within the sealed area was not explosive and that the real hazard from 

sealed areas arises from leakage of methane or toxic gases from sealed areas into the ventilation 

system. 

17
 
H +W

t = + 0.6 (1) 2 

 

 

 

2.2 Seal Design Practices in Europe and Australia 

Table 3 summarizes the seal design, construction, and related sealed-area practices used 

in Europe and Australia. The underground coal mining methods in each locale vary significantly, 

although all are highly mechanized. European coal mines tend to use arched, single-entry gate 

roads for longwall mining. Australian coal mines use two-entry and some three-entry gate road 

systems for longwall development. Production from room-and-pillar coal mining is very limited 

in both Europe and Australia. In contrast, the U.S. coal industry uses both room-and-pillar and 

longwall mining, and the mains, submains, and gate roads will have multiple entries. The following discussions trace the origins of seal design standards in locales outside of the United States. 

2.2.1 Seal Design Practices in the United Kingdom 

Early research in the United Kingdom sought means for suppressing spontaneous 

combustion fires in mined-out areas. After sealing an area to suppress a gob fire, an explosion of 

flammable gases distilled from the coal can occur. Fire control seals must resist the anticipated 

forces developed by the explosion. Beginning in 1942, and reissued in 1962 and again in 1985, 

a committee of the U.K. Institution of Mining Engineers issued a report on Sealing Off Fires 

Underground to provide ventilation system design guidance for possible fire control with seals 

[Mason and Tideswell 1933; Willett et al. 1962; Hornsby et al. 1985]. Succeeding committees 

state that “it is desirable in designing explosion-proof stoppings (i.e., seals) to assume that 

pressures of 140 to 345 kPa (20 to 50 psig) may be developed.” These reports recommended seal 

designs, mostly using gypsum, to resist the assumed explosion pressures. In addition, these 

reports recommend “pressure balancing” to control the oxygen influx to sealed areas, along with 

monitoring practices for these areas. With reference to explosion testing at the former U.K. 

Buxton facility, the Sealing Off Fires Underground report [Hornsby et al. 1985] recommended 

an explosion design pressure of 524 kPa (76 psig) and a formula for calculating the required 

thickness of an explosion-proof seal, given as 

where t = the required seal thickness, m, 

H = roadway height, m, 

and W = roadway width, m. 

In any case, minimum seal thickness is 3 m (10 ft). This formula assumes the use of “Hardstop” 

for the seal, which is a gypsum product with a compressive strength of about 4 MPa (600 psi) 

after 2 hr and 12–14 MPa (1,700–2,000 psi) after 24 hr. Recent explosion tests on full-scale seals 

validated this design formula and showed that the formula contained an implicit safety factor of 

at least 2 [Brookes and Nicol 1997; Brookes and Leeming 1999; U.K. Health and Safety 

Executive 1998]. 

18
 
Table 3.—Comparison of worldwide seal designs: construction and related practices

Country Mining method Design standard Year Problems Formula Typical width × height Typical thickness Material Inert? Monitor? 

UnitedKingdom 

Single-entrylongwall 

0.5 MPa(73 psig) × 2 

Pre-1960 No seals

destroyed t 

HW 

= 

+ 

+ 

2 

0 6. 

6 × 3 m

(20 × 10 ft) 

4–5 m(13–16 ft) 

Gypsum Set up to

inertize 

Tube bundle 

Germany Single-entry

longwall 

0.5 MPa(73 psig) × 2 

Pre-1960 No seals

destroyed t 

a

bz 

= 

0 7.

σ 

6 × 5 m

(20 × 16 ft) 

3–6 m(10–20 ft) 

2/3 fly ash 1/3 cement 

No Initially, 

as needed 

Poland Single-entry

longwall 

0.5 MPa(73 psig) × 2 

Pre-1960 No seals

destroyed 

Full-scale test 6 × 5 m

(20 × 16 ft) 

3–6 m(10–20 ft) 

Varies GAG jet 

engine 

As needed 

Australia Two-entry

longwall 

345 kPa(50 psig) × 1or 140 kPa(20 psig) × 1 

1999 Moura

No. 2(1994) 

Structural analysis 6 × 3 m

(20 × 10 ft) 

Rarely used: 0.3–1.5 m(1–5 ft) 

Varies Many mines Tube bundle 

UnitedStates 

Longwall and room-and-pillar 

140 kPa(20 psig) × 1 

1971 Seals

destroyed 

Full-scale test 6 × 2 m

(20 × 7 ft) 

0.5–1 m(1.5–3.5 ft) 

Varies 1 mine 1 mine 

19 
 0.7 a

t = (2)

σ bz 

 

 

 

2.2.2 Seal Design Practices in Germany 

Michelis and Kleine [1989] describe regulatory standards in Germany for designing and 

constructing explosion-proof seals in underground coal mines. The official “Directives for the 

Construction of Stoppings” require that seals withstand a static pressure of 500 kPa (72 psig). 

Seal designs are based on a static load of 1 MPa (144 psig), which is a safety factor of 2. This 

standard has apparently been in place since the 1940s and possibly earlier. Similar to the U.K. 

seal design standards, the German standard also includes a formula to calculate the required seal 

thickness, given as 

where t = seal thickness, m, 

a = largest roadway dimension (width or height), 

and σbz = flexural strength of the seal material, MPa. 

Genthe [1968] developed this formula based on an arching analysis. Seal construction material is 

a mixture of 2/3 flyash and 1/3 cement with the possible addition of an accelerator. The flexural 

strength of this material ranges from about 1 to 2 MPa (150 to 300 psi), and its compressive 

strength is about 5 MPa (750 psi). 

Full-scale testing of seals at the Tremonia Experimental Mine verified the design formula 

in typical conditions. A safety factor of 2 may be implicit to the formula. 

2.2.3 Seal Design Practices in Poland 

Cybulski et al. [1967] discussed a series of test explosions conducted in the 1 Maja Mine, 

which generated pressure greater than 3 MPa (450 psig) and caused great damage to a test seal. 

These researchers believed it difficult or impractical to construct a seal robust enough to 

withstand these observed pressures. They reasoned that in practice only small volumes of 

explosive methane-air could accumulate in the face area of an active longwall operation and 

therefore the maximum explosion pressure at a seal does not exceed 500 kPa (72 psig). This 

design standard seems to correlate with those of Germany and the United Kingdom. 

Examination of the Polish technical literature did not identify a design formula for seal 

thickness. Full-scale testing at Experimental Mine Barbara is used to validate various seal 

designs. Lebecki et al. [1999] describe several such validation tests. These tests will apply a 

pressure of about 1 MPa (145 psig) to a candidate seal in order to ensure that the design has a 

safety factor of about 2. 

2.2.4 Seal Design Practices in Australia 

After the Moura No. 2 disaster, which killed 11 miners in 1994 [Roxborough 1997], 

Australian regulatory authorities and the Australian coal mining industry implemented major 

safety changes with regard to seals and sealed areas of coal mines. The Moura No. 2 explosion 

resulted from the ignition of a methane-air mixture within a room-and-pillar panel that was 

sealed about 22 hr prior to the explosion. Queensland regulations now recognize two types of 

20
 
seals: type C and type D [Oberholzer and Lyne 2002]. The seal regulations in New South Wales 

have requirements similar to those in Queensland [Gallagher 2005]. 

A type D seal must withstand a 345-kPa (50-psig) explosion overpressure and is required 

“when persons are to remain underground while an explosive atmosphere exists in a sealed area 

and the possibility of spontaneous combustion, incendive spark or some other ignition source 

could exist” [Lyne 1996]. Alternatively, if monitoring of the sealed area atmosphere 

demonstrates that an explosive atmosphere does not exist, then a type C seal designed to 

withstand a 140-kPa (20-psig) overpressure is permitted. In adopting these pressure design 

criteria for type C and type D seals, Australian authorities recognized that explosion pressures up 

to 1.4 MPa (200 psig) had been observed in experimental mine explosions. However, these 

experts believed that it is not practical to build structures to withstand this pressure throughout a 

multiheading mine [Lyne 1996]. 

Using a type C seal, designed for a 140-kPa (20-psig) overpressure, requires stringent 

monitoring of the sealed area atmosphere. NIOSH researchers note that the Queensland standard 

for a type C seal does not allow for any amount of explosive mix behind a seal. When using the 

type C seal, detection of any explosive mix within a sealed area requires the immediate 

withdrawal of all mining personnel until the problem is corrected, usually by injecting inert gas 

behind the seal. 

The Australian standards allow the mine operators broad latitude to adopt whichever 

technology or materials they wish to employ. However, the seal design must meet four key 

elements: 

1.	 Full-scale testing at an internationally recognized mine testing explosion gallery must 

validate the design and specifications for a seal. 

2.	 The seal design must consider site-specific factors such as design life, geotechnical 

conditions, repair possibility, and water head. 

3.	 Management must ensure that the actual seal installation meets all design 

specifications. 

4.	 Management must inspect and maintain all seals according to design specifications. 

Initially, the new Australian seal standards relied on full-scale testing to validate seal 

designs. Tests conducted in the late 1990s on a few seal designs provided key validation data for 

structural analysis computer programs, and now these analysis programs have become the means 

to evaluate new seal designs as opposed to additional full-scale testing. 

As mentioned earlier, the use of type C seals designed to withstand a 140-kPa (20-psig) 

explosion overpressure requires routine gas sampling and analysis to ensure that the sealed area 

atmosphere contains no explosive mix. Demonstrating this lack of explosive mix requires a 

monitoring system along with a management plan to collect the requisite data, analyze and 

interpret it in a timely manner, and take the necessary actions, such as withdrawal of people or 

inertization, if required. Queensland regulatory authorities have issued standards for the 

monitoring of sealed areas that provide guidance for the location of monitoring points, along 

with the sampling frequency [Lyne 1998]. 

With regard to the traditional Coward triangle graph representing the methane-air 

explosive zone, the Queensland monitoring standard defines an explosive risk buffer zone whose 

boundaries are less than 2.5% or greater than 22% methane and greater than 8% oxygen. This 

standard requires “a regular sampling regime such that a maximum change in the methane 

21
 
 

 

Figure 8.—Continuous atmospheric gas monitoring system in Australia. Top left: Monitoring shed over 

mine showing borehole and sample tubes. Top right: Close-up of sample tube bundle. Bottom right: Sample 

tube pumps. Bottom left: Inside monitoring shed showing manifold and gas analyzer. 

concentration of 0.5% methane absolute can be detected between samples” [Lyne 1998]. In 

many situations, a sampling frequency every few hours is common practice. 

To meet the required sampling frequency, most Australian longwall mines have deployed 

tube-bundle systems for continuous gas monitoring similar to that shown in Figure 8. The top left 

of Figure 8 shows a typical monitoring shed located on the surface above a longwall mine. The 

monitoring tubes enter the mine via a borehole to the left of the shed. Typical tube-bundle 

systems will monitor from 20 to 40 points or more, with about half located in the active mining 

areas and the other half in the sealed areas. The top right of Figure 8 shows a close-up of a 

seven-tube bundle. The pumps shown in the bottom right of Figure 8 draw air samples 

continuously from each monitoring point. The bottom left of Figure 8 shows where the sample 

tubes enter the monitoring shed for analysis. Inside the monitoring shed is a solenoid-valvemanifold system activated by a programmable logic controller. Samples are automatically 

directed to an on-line gas analyzer and analyzed for carbon monoxide, carbon dioxide, methane, 

and oxygen. It is assumed that nitrogen and argon comprise the balance. A typical tube-bundle 

system provides a gas analysis at each monitoring point every 1–3 hr. Real-time data are 

displayed at the mine’s control center, where trained operators can respond as necessary. 

22
 
In addition to monitoring to ensure that the sealed area does not contain any explosive 

mix, many Australian coal mines artificially inert sealed areas. Artificial inertization is mainly 

employed at mines with high risk of spontaneous combustion. Two major systems are in use at 

this time: nitrogen gas injection and the Tomlinson boiler. Nitrogen injection systems may use 

molecular membranes to separate nitrogen from the atmosphere. While these systems are 

adequate for routine nitrogen injection at a low flow rate, they may lack sufficient capacity for 

injection during an emergency, such as a fully developed spontaneous combustion event. The 

Tomlinson boiler, shown in Figure 9, burns diesel fuel and air in a combustion chamber, and the 

resulting exhaust gases are cooled and compressed for injection into a sealed area. The inert gas 

is mainly nitrogen and carbon dioxide with trace amounts of carbon monoxide and 1%–2% 

oxygen. 

23
 

Figure 9.—Tomlinson boiler for inertization at an Australian coal mine. 

Since the Moura No. 2 disaster, which resulted from an explosion within a recently sealed 

area, the Australian regulatory authorities and mining industry have developed sealed area 

management systems to ensure that potentially explosive methane-air mixes do not accumulate 

undetected within sealed areas. A key component of this management system is monitoring with 

real-time data acquisition systems coupled to simple data analysis, display, and warning systems. 

In addition to monitoring, some mines may employ artificial inertization of their sealed areas to 

control potentially explosive mixes. 
 

 

 

 

 

 

3.0 Explosion Chemistry and Physics 

3.1 The 908-kPa (132-psia) Constant Volume Explosion Pressure 

The chemical reaction for an ideal stoichiometric mix of about 10% by volume methane 

in air is given by 

CH4 + 2O2 → CO2 + 2H2O + Energy (A) 

To give mining engineers a sense for the amount of energy in a methane-air mix, the energy 

content in 1 m3

 of ideal methane-air mix is about the same as 0.75 kg of TNT. 

Starting at standard atmospheric temperature and pressure, combustion of a methane-air 

mixture in a closed constant volume under ideal adiabatic conditions will raise the final 

temperature and pressure inside that closed constant volume. Thermodynamics enables 

calculation of that temperature and pressure increase. 

The ideal gas law is 

pv = nRT (3) 

where p = total pressure, 

v = specific volume, 

n = number of moles of gas, 

R = universal gas constant, 

and T = the absolute temperature. 

For a closed constant volume system under ideal adiabatic conditions, the initial and final 

temperatures and pressures are related as 

p f / pi = Tf / Ti (4) 

Thermodynamic equilibrium programs such as Cheetah [Fried et al. 2000] or NASA-Lewis 

[McBride and Gordon 1996] predict that the final temperature for methane-air combustion is 

about 2,670 K. For an initial temperature of 298 K, the temperature increase ratio is thus 

2,670/298, or 8.96, and therefore the ratio of final to initial pressure is also about 8.96 

[Zlochower 2007b]. 

Assuming that the initial total pressure is 101 kPa (14.7 psia), the final total pressure 

(absolute pressure) is 908 kPa (132 psia). This pressure is called the constant volume explosion 

pressure, or CV pressure, and sometimes these numbers are rounded to 900 kPa (135 psia). The 

pressure increase (gauge pressure) is therefore 807 kPa (117 psig). This pressure is called the 

constant volume explosion overpressure, or CV overpressure; again, sometimes these numbers 

are rounded to 800 kPa (120 psig). 

NOTE: In this section of the report, pressure is always absolute pressure as opposed to 

gauge pressure. 

24
 
FACT 1 

Combustion of stoichiometric (≈10%) methane-air mix in a 

closed volume raises the absolute pressure from 101 to 

908 kPa (14.7 to 132 psia). 

Combustion of nonstoichiometric methane-air mixes produces lower temperature and 

pressure increases. Figure 10 (derived from Cashdollar et al. [2000]) shows the variation of 

absolute pressure throughout the flammable range of methane concentration in air. The 

maximum absolute pressure occurs at about 10% methane in air, slightly above stoichiometric 

proportions of 9.5%, but that pressure is substantial over a considerable range surrounding the 

stoichiometric mix. Natural gas in a coal mine usually consists of 90% or more methane, but 

it may also contain other alkanes such as ethane, propane, butane, and pentane. These higher 

hydrocarbons may increase the energy release and the pressure somewhat. Natural gas may also 

contain minor amounts of carbon dioxide, which will lead to a slightly lower pressure increase. 

However, these effects are considered negligible at this time. The experimental data shown in 

Figure 10 are slightly less than theoretical calculations due to incomplete combustion and heat 

losses in the experiments. As it is not possible to predict the composition of an explosive 

methane-air mix within a sealed area, prudent engineering requires that we plan for the highest 

potential explosion pressure, i.e., the pressure developed by the ideal stoichiometric mix. 

25
 

 

7 

6 

5 

4 

3 

Flammability of Methane (CH4) in 120 liter test chamber 

10 1.0 

Experimental data from 120-L chamber 

Theoretical pressure from NASA-Lewis code 

140 

9 0.9 

8 0.8 

120 

0 2 4 6 8 10 12 14 16 18 

Methane, % 

Explosion pressure - bar

0.7 

0.6

Explosion pressure - MPa 

100 

Expolsion pressure - psi 

80 

0.5 

60 0.4 

0.3 40 

2 

1 

0 

0.2 

20 

0.1 

0.0 0 

Figure 10.—Variation of absolute pressure versus methane concentration: theoretical and experimental 

determinations [Cashdollar et al. 2000]. 
Razus et al. [2006] compiled results from CV explosion tests of methane-air done by 

researchers around the world. They showed that the absolute pressure ranged from 700 to 870 

kPa (101.5 to 126 psia), which is somewhat less than the theoretical CV explosion pressure, 

but in good agreement with the results by Cashdollar et al. [2000]. Several factors contribute to 

lower experimental pressures compared to theoretical reaction pressures calculated by the 

thermodynamic equilibrium programs NASA-Lewis and Cheetah. Incomplete combustion 

accounts for part of the discrepancy; however, heat loss to the reaction vessel (i.e., nonadiabatic 

conditions) accounts for the majority. 

In the laboratory tests summarized by Razus et al. [2006], the explosion vessels ranged in 

volume from 4.2 to 20 L, although one test was from a 204-m3

 vessel. The ability to achieve 

adiabatic conditions depends largely on the ratio of surface area for heat loss compared to the 

combustion volume. At the laboratory scale, this ratio is large, so achieving adiabatic conditions 

is difficult. Experiments conducted by NIOSH researchers with 20-, 120-, and 1,000-L explosion 

vessels have shown that larger vessels more closely approximate adiabatic conditions. With coal 

mine entries, the ratio of surface area to enclosed volume is small, so the adiabatic assumption is 

appropriate and the CV explosion pressure at that scale will approach the theoretical value of 

908 kPa (132 psia). 

NIOSH researchers also examined the effect of moisture on the CV explosion pressure. 

Thermodynamic equilibrium calculations by Zlochower [2007b] show that changing the initial 

conditions from dry air at 15 °C to moisture-saturated air decreases the CV explosion pressure 

from 934 to 925 kPa (135 to 134 psia). Thus, adding moisture to the system has negligible effect 

on the pressure that develops in confined explosions. 

The CV explosion pressure of 908 kPa (132 psia) will dissipate as the hot explosion gases 

cool or the gases escape through fractures in the surrounding rock. However, this pressure 

dissipation occurs over a long time scale and, as discussed later in the report, the full high 

pressures must be considered in the structural design of seals. 

3.2 Effect of Coal Dust on Explosion Pressure 

Coal dust explosion data presented by Hertzberg and Cashdollar [1987], Wiemann 

[1986], and Cashdollar [1996] show that the rapid combustion of coal dust in air will develop a 

CV explosion pressure similar to that for methane-air. In a coal dust explosion, volatilization of 

the fuel dust occurs rapidly within the flame front, leading to the evolution of various gaseous 

hydrocarbons, which react similarly to methane gas. Thus, the CV explosion pressure for coal 

dust-air is similar to methane-air, but slightly less. 

Figure 11 [Cashdollar 1996] shows that methane-air reaches its maximum absolute 

pressure of almost 908 kPa (132 psia) at a concentration of about 65 g/m3

, which is about 10% 

methane by volume. The theoretical maximum indicated on this figure is consistent with the 

complete calculations shown in Figure 10. The experimental data are slightly less than 

theoretical calculations due to incomplete combustion and heat losses in the experiments. The 

mix becomes fuel-rich and nonflammable above a concentration of about 150 g/m3

 or 16% by 

volume [Cashdollar et al. 2000]. 

Figure 11 also shows the theoretical maximum absolute explosion pressure for coal dust, 

which ranges from about 790 to 890 kPa (115 to 129 psia) based on calculations using the 

NASA-Lewis code [Zlochower 2007b]. The best-fit line describing the experimental data is also 

less than theoretical expectations due to heat losses in the experiments. Coal dust, however, does 

26
 
not have a similar rich limit, and instead it reaches a maximum pressure and levels off at 

concentrations of about 200–300 g/m3

. The energy release from a coal dust explosion is only 

limited by the available oxygen in the reaction vessel or the sealed area of a coal mine, if enough 

dust is available. 

 

 

8 

7 

6 

5 

4 

3 

2 

9 

10 

Comparison of Gas and Dust Flammability 

20-L Chamber data 

1.0 

0.9 

0.8 

Methane gas experimental

data (average) 

Pittsburgh coal dust

experimental data (average) 

Theoretical maximum explosion 

pressure for CH4 - Air 

Theoretical maximum explosion pressure 

range for coal dust - Air 

0 50 100 150 200 250 300 350 400 

Concentration, g/m3 

140 

120 

Explosion pressure - bar 

Explosion pressure - MPa 0.7 

0.6 

100 

Explosion pressure - psi 

80 

0.5 

0.4 

0.3 

60 

40 

0.2 

20 

1 0.1

0 0.0 0

Figure 11.—Variation of absolute pressure for methane-air and coal dust-air [Cashdollar 1996]. 

FACT 2 

Combustion of fuel-rich coal dust and air mix in a closed 

volume raises the absolute pressure from 101 kPa (14.7 psia) 

to about 790–890 kPa (115–129 psia), which is only slightly 

less than combustion of methane-air mix. 

Similar to methane, coal dust explosibility also depends on the oxygen concentration. 

Cashdollar [1996] shows that coal dust in air is no longer explosive below an oxygen 

concentration of 10%. 

27
 
3.3. Homogeneous Methane-Air Mixtures in Sealed-Area Atmospheres 

A common misconception exists that methane layering will develop within the still air of 

a sealed area, similar to the way it can occur in ventilated working areas of the mine. This 

misconception arises because it is common knowledge that methane tends to collect along the 

roof, especially in pockets, and that miners are instructed to measure methane concentrations in a 

coal mine about 1 ft from the roof and ribs. The density of methane is about 0.55 times that of 

air, so buoyancy effects do exist. However, in the completely still air within a sealed area, 

diffusion processes will dominate over buoyancy effects, which will lead to development of a 

homogeneous methane-air mix within a few days or less. 

Factors such as methane liberation rate and location of methane source (ribs, floor, or 

roof) also affect the rate of mixing. If methane enters the sealed area slowly from the roof, then 

the mixing rate with air is controlled only by diffusion. If the influx is rapid, then a convective 

mass transfer component along with diffusion will enhance component mixing. If methane enters 

from the floor, then mixing is even faster. 

The diffusion rate can be calculated from the kinetic theory of gases. If a gas component 

is present in nonuniform concentration, and at uniform constant temperature and pressure in the 

absence of external fields, that component diffuses to render its concentration uniform. The rate 

that the diffusion process occurs under most conditions is proportional to the concentration 

gradient times the diffusion constant. This elementary diffusion relationship is called Fick’s first 

law of diffusion, in which the diffusion constant for methane in air is 0.157 cm2

/s [McCabe and 

Smith 1967]. The Loschmidt diffusion apparatus is used to determine the diffusion coefficient 

for binary mixtures of gases. Methane is placed in the top 50-cm-long tube and air is placed in 

the bottom 50-cm-long tube, which are separated initially by a valve. The valve is opened and 

the gases are allowed to mix through diffusion [Shoemaker and Garland 1967]. In the apparatus, 

methane will mix with air and form an average 50–50 mixture in both the top and bottom tubes 

in 55 min. 

Zlochower [2007a] used diffusion theory to estimate the time required for methane 

concentrated at the roof of a 2.1-m (7-ft) high coal seam to diffuse uniformly from roof to floor. 

The methane concentration reaches uniformity to within 10% in about 21 hr. His calculations 

assume a slow methane influx rate. As mentioned earlier, if the methane influx is rapid or from 

the floor, then convective mass transfer will only enhance component mixing and decrease the 

mixing time. 

In summary, contrary to common misconceptions about methane layering, the completely 

still air within a sealed area will develop a fairly uniform mixture of methane and air within a 

matter of days after sealing. Diffusion processes dominate buoyancy effects, and the mixing 

process is only enhanced by any convective mass transfer. 

3.4 Explosions in Tunnels 

The prior analysis for the basic 908-kPa (132-psia) CV explosion pressure contains three 

key assumptions: (1) the reaction vessel is small and spherical so that dynamic effects due to 

pressure waves are negligible, (2) the ignition occurs at the center of the vessel, and (3) the flame 

speed remains small and well below the speed of sound (subsonic). However, methane-air 

ignitions in mines propagate along mine entries (tunnels), and the physics is much more complex 

than a simple reaction vessel. These complexities can lead to the development of much higher 

28
 
explosion pressures. Numerous authors have described the combustion process as it initiates 

from an ignition point in a fuel-air mixture and develops into an explosion [Zucrow and Hoffman 

1976; Baker et al. 1983; Zeldovich et al. 1985; Landau and Lifshitz 1987; Lewis and von Elbe 

1987; Gexcon 2007b]. 

Consider a mine entry closed at both ends and filled with methane-air mix, as shown in 

Figure 12. Ignition occurs at the far right end, and the flame propagates to the left. Four stages in 

the combustion process are detailed in the figure: (1) slow deflagration, (2) fast deflagration, 

(3) detonation, and (4) reflection of a detonation wave from head-on impact with the closed end. 

Above each stage of combustion is a pressure profile along the tunnel. Upon ignition, the initial 

laminar flame speed is only 3 m/s (10 ft/s); however, a slow deflagration accelerates, and the 

turbulent flame speed might increase to about 300 m/s (1,000 ft/s). The pressure in the burned 

gas behind the flame front increases to the 908-kPa (132-psia) CV explosion pressure. The 

combustion front acts as a piston, compressing the unburned gas in front of it. The leading edge 

of this acoustic wave propagates to the left at the local sound speed of about 341 m/s (1,120 ft/s). 

In between this wave front and the flame front, the unburned gas acquires velocity to the left and 

the static pressure inside this region will increase. This pressure increase ahead of the flame front 

is termed “pressure piling.” 

As the velocity of the unburned gas ahead of the flame front increases, the turbulence in 

the flow will increase. The degree of turbulence depends on both the flow velocity and the 

roughness of the tunnel. Obstructions from roof and rib falls, ground support, machinery, and 

wall roughness are possible forms of tunnel roughness that will enhance turbulent flow. The 

flame front will evolve from a simple planar front at low flame speeds to a progressively more 

complex wrinkled flame front as the turbulence increases. The increased turbulent flow in the 

unburned gas ahead of the flame front will increase the combustion rate, and the flame front will 

begin to catch up to the pressure wave front. At higher but still subsonic flame front speeds, the 

combustion process becomes a fast deflagration. Combustion of precompressed unburned gases 

leads to pressures greater than the 908-kPa (132-psia) CV explosion pressure. For example, if 

pressure piling has increased the pressure to 300 kPa ahead of the flame front, then the pressure 

immediately behind the flame front will be 300 kPa × 9, or 2.7 MPa (392 psia). However, these 

transient pressure waves will equilibrate, and the overall pressure inside the closed tunnel will 

eventually settle down to 908 kPa (132 psia). 

Flow dynamics play a complex role in accelerating the combustion process as a result of 

increasing turbulence. Figure 13 illustrates a strong positive feedback loop that exists between 

flame propagation speed, turbulence, and combustion rate. Combustion of methane-air mix leads 

to expansion, increased pressure, and increased velocity of combustion products and the 

unburned methane-air mix. The increased flow velocity leads to increased flame propagation 

speed, increased turbulence in the methane-air mix, and finally increased combustion rate. Thus, 

as shown in Figure 13, the feedback loop closes with even faster expansion rate, along with 

development of higher pressure and velocity. 

29
 
Figure 12.—Four stages of combustion process in a closed tunnel and the approximate pressures. 

30
 
Figure 13.—Strong positive feedback loop between pressure increase, turbulence, and combustion rate. 

3.5 Static, Dynamic, and Reflected Pressure From Explosions in Tunnels 

The pressure and energy in the gas flow ahead of the flame front shown in Figure 12 

consist of two parts: a quasi-static component and a dynamic, or kinetic, component. The quasistatic pressure component arises from the gas temperature and acts equally in all directions. The 

magnitude of the quasi-static pressure component was discussed earlier, where it was shown to 

rise to a pressure of 908 kPa (132 psia). For engineering design, one must generally consider the 

total stress acting on a structure, which is the sum of the quasi-static and dynamic components. 

As shown in Figure 12, as the hot gases behind the flame front expand, the flame front 

and the gas ahead of the flame front are pushed forward or to the left in this example. This 

pressure wave ahead of the flame front develops rapidly from a blast wave with a finite rise time 

into a shock wave with a rapid rise time. Glasstone and Dolan [1977], Kinney [1962], Landau 

and Lifshitz [1987], and Zucrow and Hoffman [1976] present equations to describe shock waves 

and the factors controlling their strength. These relationships are derived from the RankineHugoniot conditions that are based on conservation of mass, momentum, and energy at the shock 

wave front. 

31
 
 

 

 

 

The magnitude of the wind or dynamic (velocity) pressure is given by 

 1 2pV = ρ V	 (5) 2 

where pV = dynamic (velocity) pressure, 

ρ = gas density, 

and V = gas velocity. 

The dynamic pressure at the shock front is related to the quasi-static overpressure pS by 

 ⎛

⎜

⎝

5 ⎞

⎟

⎠

⎜

⎛

⎜

⎝

p

2

S 

7 po + pS 

⎟

⎞

⎟

⎠

p = (6) V 2
 

where po = initial pressure. 

In a deflagration, the quasi-static overpressure ranges from 0 to almost 807 kPa (117 psig), and 

the initial pressure is 101 kPa (14.7 psia); therefore, the dynamic pressure ranges from 0 to about 

1,000 kPa (145 psia). Even at a modest quasi-static overpressure of 400 kPa (58 psig), the 

dynamic component of pressure is about 360 kPa (52 psia). Thus, the quasi-static and the 

dynamic pressure are both significant components of the total pressure for design purposes. 

When a shock wave strikes a structure such as a seal head on, reflected overpressure on 

the seal is given by [Glasstone and Dolan 1977; Kinney 1962; Landau and Lifshitz 1987; 

Zucrow and Hoffman 1985]: 

p R =
 2 pS 

⎛

⎜

⎝

⎜

7 po + 4 pS 

7 po + pS 

⎞

⎟

⎠

⎟ (7) 

where pS = quasi-static overpressure, 

 po = initial pressure, 

and pR = reflected overpressure. 

Equation 7 applies to a nonreactive shock or blast wave in which no chemical reactions are 

occurring. For low values of the quasi-static overpressure pS, the reflected overpressure pR is 

greater by a factor of at least 2. However, stronger shocks with greater quasi-static overpressure 

can lead to reflected overpressure of eight times the incident quasi-static overpressure. For 

example, if the quasi-static pressure is 807 kPa (117 psig), then the reflected overpressure is 

about 4.1 MPa (595 psig). 

As mentioned before, the quasi-static pressure and the dynamic (velocity) pressure form 

the total pressure. Proper structural analysis of seals must consider the total gas pressure and not 

just the static component. In certain situations, the quasi-static component might act alone on a 

seal. However, in most cases, seals must withstand a total pressure consisting of both a quasistatic and dynamic (velocity) component. 

The terms “static” and “dynamic” as used in the above discussions are misnomers since 

static would imply no time dependence or motion, whereas dynamic typically implies time 

32
 
dependence. The static and dynamic (velocity) pressures suggested in Figure 12 are both 

changing in time and space. In the analysis for the explosion pressure on seals, the static pressure 

(pS) refers to the time-dependent static gas pressure that acts equally in all directions, whereas the 

dynamic (velocity) pressure pV refers to the time-dependent velocity pressure that acts in the 

same direction as the gas expansion velocity. 

3.6 The 1.76-MPa (256-psia) Chapman-Jouguet (CJ) Detonation Wave Pressure 

33
 

If the flow ahead of the flame front is sufficiently turbulent, the flame speed may increase 

from subsonic to supersonic in a process known as deflagration-to-detonation transition (DDT). 

The flame speed for a deflagration is by definition subsonic or less than about 341 m/s (1,120 

ft/s). With pressure piling effects, a deflagration generally creates transient explosion pressures 

less than about 2.0 MPa (290 psia). For a methane-air detonation, the detonation wave (a shock 

wave) propagates at about 1,800 m/s (5,900 ft/s) or about Mach 5.3 [GexCon 2007b; Shepherd 

2006]. When detonation occurs, the pressure wave front and the flame front become one (Figure 

12). In a detonation, the transient pressure rises in a few microseconds to about 1.76 MPa (256 

psia) for methane-air, but then quickly equilibrates to the 908-kPa (132-psia) CV explosion 

pressure as before. 

During a DDT event, the flame front travels at supersonic velocity, and the pressure wave 

no longer disturbs the unburned gas ahead of the flame front. Pockets of reactive gas within the 

fast-moving reaction zone are formed, and small autoexplosions occur within these pockets. 

These small shocks precompress and preheat the unburned gas so intensely that they autoignite 

the mixture. The small compression waves then coalesce into a larger amplitude shock. 

A detonation relies on shock heating and pressurization of the unburned gas to initiate the 

reaction immediately behind the shock wave. The detonation thus becomes self-driven by the 

autoexplosions occurring at the shock front and propagates away from the DDT point at the 

detonation pressure for as long as combustible material is available. 

A fundamental parameter for gaseous detonations is cell width, which is a measure of the 

physical dimensions of the cells comprising the detonation wave front. In general, for a 

detonation wave to develop and sustain itself in a pipe or mine tunnel, the diameter must exceed 

the detonation cell size [Peraldi et al. 1986; Dorofeev et al. 2000; Gamezo 2007]. Bartknecht 

[1993] reports a detonation cell size of 30 cm (1 ft) for methane in air, while Shepherd [2006] 

gives a range of 25–35 cm (0.8–1.2 ft). Kuznetsov et al. [2002] report a cell size of about 20 cm 

(0.66 ft) for a stoichiometric (≈10%) mix of methane-air and that this cell size increases to about 

30–45 cm (1.0–1.5 ft) as the mix composition deviates from stoichiometric to 8%–12% methane 

in air. Because the least dimension of typical coal mine tunnels exceeds 1 m (3.3 ft) and is more 

than the detonation cell size for methane-air, detonation of methane-air is therefore possible in 

most coal mines and has been documented experimentally [Cybulski 1975]. 

Another parameter associated with detonation is the run-up length, which is the distance 

from the ignition point to where DDT first occurs. Zeldovich et al. [1985] and Lewis and 

von Elbe [1987] describe the phenomena and discuss early work to understand the factors 

controlling DDT. More recently, Oran and Gamezo [2007] describe a 10-year theoretical and 

numerical effort that has led to the successful simulation of DDT from first principles. However, 

most practical knowledge of DDT still comes from observation and empiricism. 
Zeldovich et al. [1985] and Zeldovich and Kompaneets [1960] describe work by 

Shchelkin, who studied the onset of DDT in tubes during the 1940s. Shchelkin reported the run-up 

length as 40–50 times the tube diameter for smooth-walled tubes. Steen and Schampel [1983] 

review several empirical criteria based on tube diameter and Reynolds number to predict the 

onset of DDT and concluded that it was not possible to make a firm prediction. Lee [1984], 

Bartknecht [1993], van Wingerden et al. [1999], and Kolbe and Baker [2005] state that for 

smooth pipes, the run-up length may range from 50 to 100 times the pipe diameter. The most 

important factor governing run-up length is turbulence that accelerates combustion. As reported 

by Zeldovich et al. [1985] and Zeldovich and Kompaneets [1960], Shchelkin found that for 

rough pipes, the onset to DDT decreased to 10–20 times the tube diameter. Smirnov et al. [2003] 

also describe how turbulence tends to promote the onset of DDT. In more recent experimental 

work on DDT, Dorofeev et al. [2000] reported the minimum length to DDT as seven times the 

detonation cell size. This criterion leads run-up lengths of about 3 m, which is a much smaller 

length to DDT than any other criterion. 

For mine tunnels with an equivalent diameter of about 2 m (6 ft), the run-up length to 

DDT could range from 100 to 200 m (300 to 600 ft). Roughness of the tunnel walls or blockages 

in the tunnel from mining machinery or roof support structures contribute to increased flow 

turbulence, which in turn affects the onset to DDT and can only decrease the run-up length. 

Pending further research, NIOSH scientists selected 50 m (150 ft) as the minimum run-up length 

for detonation of methane-air in a tunnel. NIOSH scientists will conduct additional research to 

better understand run-up length and the factors that control it. 

If detonation of methane-air occurs, the pressure developed by the detonation wave can 

be predicted from first principles. Based on conservation of mass, momentum, and energy for a 

detonation wave, Fickett and Davis [1979], Baker et al. [1983] and Weir and Morrison [1954] 

present the following equation for the detonation wave pressure of any gaseous mixture: 

⎛

⎜

⎝

2

D ⎞

⎟

⎠

1+ γ
 ⎜ ⎟

P1 (1+ γ 2 ) 

P2 

1 

c1

=
 (8) 
 

where P1, P2 = pressures ahead and behind the detonation wave, 

 γ1, γ2 = specific heat ratios of reactants and products, respectively, 

 c1 = sound speed, 

and D = detonation wave speed. 

Based on thermodynamic equilibrium calculations with the NASA-Lewis code for methane-air 

[Zlochower 2007b], the detonation wave speed is about 1,815 m/s (5,950 ft/s), and the sound 

speed is about 352.6 m/s (1,157 ft/s). The specific heat ratio for the reactants is about 1.39 and 

for the products about 1.17. The computed pressure ratio is therefore 17.4. Assuming that the 

pressure (P1) of the reactants ahead of the detonation wave is 101 kPa (14.7 psia), the detonation 

wave pressure (P2) is about 1.76 MPa (256 psia). This pressure is also known as the ChapmanJouguet (CJ) detonation pressure. 

Additional thermodynamic calculations with the Cheetah [Fried et al. 2000] and NASALewis [McBride and Gordon 1996] codes also predict a value of 1.76 MPa (256 psia) for the 

CJ detonation pressure. Vasilev [2006] and Shepherd [2006] also report similar values for the 

34
 
CJ detonation pressure and the detonation wave speed for methane-air. Equation 8 stems from 

classic detonation theory developed independently by Zeldovich, von Neumann, and Döring 

known as the ZND detonation model. Zucrow and Hoffman [1976], Fickett and Davis [1979], 

and Zeldovich et al. [1985] provide complete discussion of detonation theory and the ZND 

model. Note that Equation 8 applies to the detonation of gaseous reactants as opposed to solid 

reactants. A slightly different form of Equation 8 exists for the case of solid reactants, and the 

predicted CJ detonation wave pressure is also similar. 

FACT 3 

If detonation occurs in an ideal methane-air mix at 1 standard 

atmosphere, the detonation pressure developed is 1.76 MPa, 

or 256 psia (CJ detonation pressure). 

Again, as indicated in Figure 12, when detonation occurs, the pressure rises over 

microseconds to 1.76 MPa (256 psia) but then decays to the 908-kPa (132-psia) CV explosion 

pressure. When detonation occurs, unreacted gases ahead of the flame front remain at the 

original static pressure and at rest until the detonation wave arrives and the reaction occurs. This 

CJ detonation pressure is a kind of static pressure in that it acts equally in all directions. 

Since the gas velocity ahead of the detonation wave is 0, the dynamic pressure is also 0 

until the detonation wave arrives. 

3.7 The 4.50-MPa (653-psia) Reflected Detonation Wave Pressure 

If a detonation wave impacts a solid wall such as a mine seal, a reflected shock wave 

forms and propagates in the opposite direction back through the combustion products. Several 

classical works on the fluid dynamics of combustion present analyses of this reflected detonation 

wave pressure. Landau and Lifshitz [1959, 1987] derived a relation between the incident and 

reflected shock pressure as 

 PR 5γ +1+ 17γ 2 + 2γ +1 

= (9) PI 4γ 

where γ = the specific heat ratio of the combustion products. 

Assuming that γ = 1.28, the ratio of reflected to incident detonation wave pressure is 2.54. The 

prior derivation found that the pressure of a methane-air detonation wave is 1.76 MPa (256 psia). 

When this wave reflects from a solid surface such as a seal, the reflected shock wave pressure 

and the transient peak pressure on the seal is 2.54 × 1.76 or 4.50 MPa (653 psia). Equation 9 

applies to a reactive, reflected shock wave, i.e., a detonation wave, as opposed to Equation 7, 

which applies to a nonreactive, reflected shock wave. However, both relationships lead to 

reflected shock wave pressures up to about 40 times greater than ambient atmospheric pressure. 

35
 
FACT 4 

A methane-air detonation wave reflects from a solid surface 

at a pressure of 4.50 MPa (653 psia). 

3.8 Possible Higher Detonation and Reflected Shock Wave Pressures 

At least two situations can arise that could produce even higher detonation and reflected 

shock wave pressures. At the moment of DDT, some pressure piling may remain just ahead of 

the newly formed detonation wave. As the detonation wave propagates through this 

precompressed methane-air mix, higher detonation pressures may develop locally, well in excess 

of the steady-state CJ detonation pressure. Fortunately, this pressure is highly localized and 

short-lived if DDT occurs early during combustion. Under these conditions, the supersonic 

detonation wave will quickly pass through a precompressed gas zone and the pressure returns to 

the steady-state CJ detonation pressure of 1.76 MPa (256 psia) [Shepherd et al. 1991; Dorofeev 

et al. 1996; Kolbe and Baker 2005; Shepherd 2006]. 

3.9 Measured Experimental Mine Explosion Pressures 

The theoretical calculations above give a CV explosion pressure of 908 kPa (132 psia), 

detonation pressure of 1.76 MPa (256 psia), and reflected detonation wave pressure of 4.50 MPa 

(653 psia), with possibilities for even higher pressures still. Test explosions conducted at 

experimental mines in the United States and abroad confirm the reality of these pressures. 

Nagy [1981] summarized decades of methane and coal dust explosion research conducted 

in the Bruceton Experimental Mine. In all cases, these tests were open-ended, i.e., the explosive 

mixture is partially confined and able to vent, unlike the totally confined environment within a 

sealed area. A few of the larger tests developed peak pressures of 1.04 MPa (150 psig) and 

indicate that some pressure piling occurred as the explosion propagated. Early work at the 

Tremonia Mine in Germany [Schultze-Rhonhof 1952] developed pressures of 1 MPa (145 psig) 

in similar open-ended experiments, supporting the U.S. findings. 

Cybulski et al. [1967] described nine experimental methane-air explosion experiments in 

a 57-m-long tunnel (187 ft) at the 1 Maja Mine in Poland. The amount of explosive mix ranged 

from 70 to 1,000 m3

 (2,500 to 35,300 ft3), and the length of the gas zone ranged from 4.3 m 

(14 ft) to the full 57-m (187-ft) length of the experimental tunnel. Two tests in which the 

explosive mix completely filled the tunnel produced peak pressures greater than 3.2 MPa (450 

psig). Pressure piling clearly occurred during these particular tests. Flame speed was measured at 

1,200 m/s (3,936 ft/s) corresponding to about Mach 3.5, which suggests the possibility that 

detonation occurred. Other tests in which the tunnel was not completely filled with explosive 

mix developed peak pressures in the range of 0.2–1.5 MPa (30–225 psig). These experimental 

results showed a clear relationship between the length of the explosive mix zone and the 

maximum explosion pressures. A gas zone length more than 50-m (165-ft) long can develop 

peak explosion pressures of more than 2.0 MPa (290 psig), which in turn may lead to detonation. 

In test No. 1397 conducted at Experimental Mine Barbara in Poland, Cybulski [1975] 

back-calculated explosion pressures in excess of 4.1 MPa (595 psig). The experimental explosion 

was initiated in coal dust about 200 m (656 ft) from the closed end of a tunnel. Three 

36
 
measurements of pressure wave speed ranged from 1,600 to 2,000 m/s (5,250 to 6,560 ft/s), 

which clearly suggest detonation. Unfortunately, sensors could not measure the pressure directly; 

however, the explosion punched a 1.4-m2

 hole into a 32-mm-thick steel door. The shear force 

necessary to punch this hole indicates an explosion pressure of at least 4.1 MPa (595 psig). 

In his Ph.D. dissertation, Genthe [1968] examined peak explosion pressure, flame speed, 

and the length of an explosive mix zone in order to determine their relationships. Experimental 

explosions with subsonic flame speeds less than about 330 m/s (1,100 ft/s) led to explosion 

pressures less than 1.0 MPa (145 psig). Explosions that developed supersonic flame speeds of up 

to 1,200 m/s (3,940 ft/s) produced peak pressures of up to 1.8 MPa (270 psig). The length of the 

explosive mix zone also correlated to higher peak explosion pressures. Similar to the previously 

described results from Cybulski et al. [1967], an explosion with a gas zone length of 50 m 

(165 ft) produced peak explosion pressure of 1.8 MPa (261 psig), which could be indicative 

of detonation. 

37
 

3.10 Summary of Main Parameters Affecting Gas Explosion Strength 

Several factors can influence the explosion pressure that develops within a sealed 

abandoned area of a coal mine. Some can be controlled through engineering or monitoring; 

others cannot. Because many of these factors cannot be controlled, conservative engineering 

practice dictates that mining engineers plan for the worst-case explosion pressures. 

Calculations in previous sections of this report describe this worst-case scenario— 

the combustion of a confined, stoichiometric methane-air mix of about 10% methane by volume. 

Pressure was shown to increase from atmospheric pressure to 908 kPa (132 psia). The combustion rate of methane-air in a tunnel may be enhanced by turbulence that is induced by roughness 

or obstructions in the tunnel. As turbulence increases, the combustion rate also increases, which 

leads to more turbulence in a strong feedback loop. Pressure waves develop ahead of the flame 

front, and these waves may evolve into nonreactive shock waves, which can reflect from solid 

surfaces such as seals with large pressure. As shown by Equation 7, the reflected overpressure 

from a nonreactive shock wave ranges from at least 2 to more than 10 times the initial pressure. 

A DDT may occur resulting in a detonation wave, which according to Equation 8 has a pressure 

of 1.76 MPa (256 psia) at 1 standard atmosphere initial conditions. When detonation waves 

reflect from solid objects such as mine seals, the reflected pressure from a reactive shock wave 

can induce transient pressures of 4.50 MPa (653 psia), according to Equation 9. Under certain 

conditions, even higher pressures are possible. 

An inhomogeneous, poorly mixed, or layered explosive gas cloud will generate lower 

explosion pressure. However, according to discussions in section 3.3, diffusion leads to 

homogeneous, well-mixed, and nonlayered explosive mixtures within the still atmosphere of a 

sealed area. The location of the ignition point also has an effect that can either increase or 

decrease the explosion pressure, but there is no apparent way to engineer the ignition source. 

Five additional major factors affect the pressures developed during a gas explosion: (1) the 

concentration of methane in air, (2) the overall volume of explosive mix, (3) the degree of filling 

of the volume with explosive mix, (4) the degree of confinement of the explosive mix, and 

(5) the degree of venting possible from an explosion. 

Departure from the ideal mix used in the above calculations results in lower explosion 

pressures. However, both a 6% methane-air mix near the lower flammability limit and a 14% 

mix near the upper flammability limit develop a 500-kPa (73-psia) explosion pressure (Figure 
10). Thus, a methane-air mix develops variable but substantial explosion pressure over most of 

its flammable range. Without detonation, reflected nonreactive shock waves can produce 

pressures at least twice the initial pressure. Detonation and reflected detonation wave pressures 

are also substantial over most of the flammable range, as shown in Figure 14. 

The overall volume of explosive methane-air mix also affects the level of explosion 

pressures. Larger sealed areas have longer run-up lengths and increased possibility for DDT and 

the resulting higher transient pressures. Information available at this time indicates that any 

sealed volumes with a run-up length greater than about 50 m (165 ft) behind the seal are at risk 

of developing the higher pressures that result from a detonation [Lee 1984; Bartknecht 1993]. 

38
 

 

 

 

 

35 

30 

25 

20 

Thermodynamic Equilibrium Calculations 

50 5.0 

For Methane - Air mixtures 

NASA-LEWIS Code 

700 

45 

40 4.0 

4.5 

600 

CV-constant volume pressure 

CJ detonation pressure 

Reflected detonation pressure 

5 6 7 8 9 10 11 12 13 14 15 

Volume % CH4-air 

500 3.5 

Pressure rise - bar

Pressure rise - MPa 

Pressure rise - psi 

3.0 

400 

2.5 

300 

2.0 

15 

10 

5 

0 

1.5 200 

1.0 

100 

0.5 

0.0 0

Figure 14.—Variation of theoretical pressure increase ratio versus methane concentration for CV explosion 

pressure, detonation wave pressure, and reflected detonation wave pressure. 
The degree of filling of the sealed volume refers to what proportion of the sealed volume 

is the explosive mix or, in other words, how much of the sealed volume is filled with explosive 

mix. A newly sealed panel that is crossing through the explosive range may be 100% filled with 

explosive mix. An explosion within a confined sealed area that is 100% filled with explosive mix 

will develop 100% of the CV explosion pressure of 908 kPa (132 psia), while a volume that is 

only 40% filled will only see a 363-kPa (53-psia) explosion pressure. After a sealed area has 

become inert, a leaking seal may develop an explosive mix behind it, but the degree of filling 

may be only a few percent of the total sealed volume. 

The degree of confinement influences the level of explosion pressure. An unconfined 

explosion occurs in open air, and the reaction gases can expand freely in all directions. An openair ignition of an explosive gas mix is an example of an unconfined explosion. A completely 

confined explosion occurs when the gases cannot expand freely in any direction. An explosion 

within a sealed area that is 100% filled with explosive mix is an example of a completely 

confined explosion. Most sealed areas should be considered completely confined, and the 

explosion pressure will equilibrate to the CV explosion pressure of 908 kPa (132 psia) 

independent of the size of the sealed volume. An explosion behind district and panel seals may 

meet the completely confined condition after sealing while the sealed area atmosphere crosses 

through the explosive range during initial inertization. 

Some explosions are partially confined in that the gases can expand freely along an entry. 

For example, the ignition of a limited-volume explosive mix behind a seal where the gases can 

expand into inert atmosphere deep in the sealed area is a partially confined explosion. 

The degree of venting influences the level of explosion pressure. Venting and confinement represent opposite ends of the spectrum in that a completely confined explosion is unvented 

and an unconfined explosion is completely vented. However, the degree of venting for partially 

confined explosions can vary. Using the prior example, an explosion of a limited volume of mix 

behind a leaking seal where the gases can expand freely into inert atmosphere is a partially 

confined and completely vented explosion. The explosion pressure developed will vary depending on the degree of venting from the explosion area. An example of partial venting is a crosscut 

seal where the gob may restrict gas expansion. 

4.0 Modeling Explosion Pressures on Seals 

4.1 Model Characteristics 

The prior discussions on explosion pressures placed general bounds on peak explosion 

pressures possible. However, NIOSH researchers sought additional information on the pressuretime history that could develop in a methane-air explosion. Experimental mine explosions can 

generally only study comparatively small volumes of explosive mix. Most experiments 

worldwide fill less than 20 m (65 ft) of tunnel with methane-air mix, although a few tests have 

filled as much as 58 m (190 ft) of tunnel with explosive mix. Accordingly, NIOSH researchers 

used two reputable gas explosion computer models to extrapolate small-volume gas explosion 

data to larger gas explosions typical of what could happen in a coal mine. 

The two gas explosion models are AutoReaGas, available from Century Dynamics, Inc. 

[2007], in the United Kingdom and FLACS (Flame Acceleration Simulator), available from 

GexCon [2007a] of the Christian Michelson Research Institute in Norway. AutoReaGas and 

FLACS are specialized computational fluid dynamics (CFD) models for numerically solving the 

39
 
partial differential equations governing a gas explosion. These models are used extensively in the 

oil, gas, and chemical industries to assess risks, consequences, and mitigation measures for 

various gas explosion scenarios. In particular, they have seen application to offshore oil and gas 

production facilities since the Piper-Alpha oil platform disaster that occurred in the North Sea, 

U.K., in 1988. A few European research groups have made attempts to use these models to study 

gas explosions in mines, but to date such work is very limited. The NIOSH work described 

herein probably represents the most extensive use of these models in a mining industry 

application. For a complete discussion of most gas explosion model capabilities and limitations, 

see the reviews by Lea and Ledin [2002] and Popat et al. [1996]. 

Gas explosion numerical models, such as AutoReaGas and FLACS, consist essentially of 

three elements: (1) the Reynolds-averaged Navier-Stokes equations, (2) a turbulence model, and 

(3) an empirical turbulent flamelet model. The Reynolds-averaged Navier-Stokes equations 

describe the fluid flow and are expressions for conservation of mass, momentum, and energy for 

a differential volume in terms of pressure, temperature, gas density, and velocity components. 

Coupled to the conservation equations is an equation of state, which is usually approximated 

with the ideal gas law such as pv = nRT. In gas explosion models, the Navier-Stokes equations 

are modified to consider the changing concentration of both reactants and products. 

The second major element in gas explosion models is a turbulence model to describe the 

dissipation rate of turbulence kinetic energy. Most CFD models, including AutoReaGas and 

FLACS, use an empirical k–ε turbulence model. Simply stated, the k–ε turbulence model relates 

the dissipation rate (ε) of turbulence kinetic energy (k) to the production of turbulence kinetic 

energy from Reynolds stresses and the removal of turbulence kinetic energy due to dissipative 

effects. ε depends on the velocity fluctuations in the flow, which in turn depends on a length 

scale, 1/K, where K is a wave number. ε(K) follows a power-law spectrum where little energy 

dissipation occurs in large eddies with small K and most energy dissipation occurs in small 

eddies with large K. At a critical length scale, lK, the organized motion cascades to small eddies 

whereupon kinetic energy is converted into heat. The k–ε turbulence model contains several 

empirically determined constants that are well known for many practical applications. 

The third element in these models is a combustion model to describe the concentration 

change rates of reactant and product species and the associated energy release rate. Most CFD 

models use empirical reaction rate models. AutoReaGas uses an empirical correlation between 

reaction rate and flame speed. FLACS uses a “β flame model” that correlates turbulent burning 

velocity with turbulence parameters. In both models, an increase in turbulence kinetic energy 

results in an increase in the reaction rate. 

In most applications of the AutoReaGas and FLACS models in the oil, gas, and chemical 

industries, the computed and measured explosion pressures do not exceed about 500 kPa (72 

psia). These models do not properly consider the physics of detonation or DDT. The zone size of 

these initial gas explosion models is on the order of 0.5 m (1.5 ft) and thus too coarse to capture 

shock waves and other rapidly rising pressure waves. The coarse grid size tends to “smear” such 

phenomena. Thus, at the extremely high pressures that could occur in a mining explosion, the 

models are not correct; however, they will correctly indicate the pressure buildup to these high 

pressures. Despite these shortcomings at high pressures, such models still provide useful insights 

into many practical applications of interest at lower pressures. The full extent of this initial gas 

explosion modeling application in mining is documented by Draper and Fairlie [2006] using 

AutoReaGas and by Hansen [2006] using FLACS. 

40
 
 

 

 

 

A-22 

A-132 

A-183 

A-233 

A-283 

A-355 

A-453 

A-550 

A-649 

A-807 C-757 

0 30 

Scale, m 

12m 

6m 

24m 

30m 

23m 

15m 

E drift 

A drift

B drift

C drift 

B-108 

B-70 

B-30 

B-10 

B-158 

B-211 

B-257 

B-329 

B-427 

B-526 

B-626 

B-782 

C-13 

C-84 

C-134 

C-184 

C-234 

C-304 

C-403 

C-501 

C-598 

Bulkhead 

Ignition Point 

Stage 1 Models ­

B drift calibration tests 

LLEM experiment numbers: 

484 - 12.2m methane zone, end ignition 

485 - 18.3m methane zone, end ignition 

486 - 18.3m methane zone, center ignition 

0 30 

Scale, m 

IGNITION POINT 

Stage 1 Models ­

D drift calibration tests 

LLEM experiment numbers: 

D - 1 

D - 2 

D - 3 

D - 4 

D - 5 

D - 6 

D - 7 

D - 8 

D - 9 

D - 10 

468 - 3.66m methane zone, end ignition 

469 - 8.23m methane zone, end ignition 

470 - 12.2m methane zone, end ignition 

Figure 15.—Layout of calibration models in the LLEM. Left: B-drift calibration tests; right: D-drift calibration tests. 

4.2 Model Calibration 

Initial gas explosion model calculations attempted to duplicate measured pressure versus 

time histories from six tests done in the Lake Lynn Experimental Mine (LLEM). Figure 15 

(right) shows the test and model geometry for three single-tunnel experiments in D drift of the 

LLEM, and Figure 15 (left) shows the same for three multiple-tunnel experiments in A, B, and 

C drifts. As shown in Table 4, each test involved a larger amount of explosive methane-air mix. 

The length of the gas clouds ranged from about 3.7 to 18.3 m (12 to 60 ft). 

Table 4.—Characteristics of LLEM experiments 

for gas explosion model calibration 

Test No. 

Length of 

methane zone (m) 

(about 10% methane) 

Approximate 

methane 

volume (m3

) 

Ignition point 

468 ................ 3.66 4.25 0.15 m from D-drift end. 

469 ................ 8.23 9.91 0.15 m from D-drift end. 

470 ................ 12.2 15.21 0.15 m from D-drift end. 

484 ................ 12.2 16.14 0.15 m from B-drift end. 

485 ................ 18.3 23.64 0.15 m from B-drift end. 

486 ................ 18.3 23.64 9.20 m from B-drift end. 

41
 
Tables 5 and 6 compare calculated pressure to measured pressure at different gauge 

points for the three single-tunnel experiments in D drift and the three multiple-tunnel 

experiments in A, B, and C drifts, respectively. The pressures calculated by AutoReaGas agree 

with experiment to within ±47%, and the FLACS calculations agree with experiment to within 

±24%. For gas explosion models of this kind, Lea and Ledin [2002] consider this level of 

agreement between calculated and measured pressures acceptable for engineering purposes. 

The calculations with FLACS were done “blind” beginning only with the problem geometry 

and no foreknowledge of the measured pressures. 

Table 5.—Comparison of calculated to measured LLEM experimental gas explosion pressures 

in single-tunnel (D-drift) tests 

Measured AutoReaGas FLACS 

Test No. Gauge No. peak pressure 

(kPa) 

Calculated 

pressure 

(kPa) 

Percent 

difference 

Calculated 

pressure 

(kPa) 

Percent 

difference 

468 ............ D–1 21.53 27.42 27.36 18.70 –13.14

 D–10 18.59 23.24 25.01 17.60 –5.33 

469 ............ D–1 58.87 53.98 –8.31 57.50 –2.33

 D–10 48.13 46.76 –2.85 51.90 7.83 

470 ............ D–1 75.03 74.29 –0.99 76.30 1.69

 D–10 74.80 62.09 –16.99 71.00 –5.08 

Table 6.—Comparison of calculated to measured LLEM experimental gas explosion pressures 

in multiple-tunnel (A-, B-, and C-drift) tests 

Measured AutoReaGas FLACS 

Test No. Gauge No. peak pressure Calculated Percent Calculated Percent 

(kPa) pressure 

(kPa) difference pressure (kPa) difference 

484 ............ B–10 83.89 44.59 –46.85 71.20 –15.13

 B–526 29.43 19.75 –32.89 28.66 –2.62 

485 ............ B–10 101.30 109.64 8.23 97.64 –3.61

 B–526 34.14 29.17 –14.56 42.34 24.02 

486 ............ B–10 48.12 65.64 36.41 57.06 18.58

 B–526 24.88 22.14 –11.01 23.10 –7.15 

Figure 16 shows typical measured versus computed pressure-time histories for both the 

AutoReaGas and the FLACS models. For these small-volume gas explosions, experiment and 

model compare well. The magnitudes of the peak pressures compare well, along with the shape 

or width of the pressure waves. However, these models do not compute arrival time of the 

pressure waves accurately. The first arrival of the calculated pressure wave is slower than that 

measured. This difference arises from the nature of the actual ignition. The models assume a 

single-point ignition, whereas in the actual tests, an electric match that emitted a shower of 

sparks started the explosion simultaneously in many different locations. In summary, despite the 

offset in timing, the gas explosion models reproduced the measured experimental data well. 

42
 
 

1.2 0.12 17 

15 1.0 0.10 

LLEM - 3.05m 

LLEM - 32.9m 

LLEM - 160.3m 

0.00 0.20 0.40 0.60 0.80 1.00 1.20 1.40 1.60 1.80 2.00 

Time 

13 

0.8 0.08 11 Pressure - MPa Pressure - MPa Pressure - MPa 

Pressure - bar Pressure - bar Pressure - bar

9 0.06 

Pressure - psi Pressure - psi Pressure - psi 

0.6 

0.4 

0.2 

7 

0.04 

5 

3 0.02 

1 

0.0 0.00 

-1 

-0.2 -0.02 -3

1.2 0.12 17 

AutoReaGas - 3.05m 

AutoReaGas - 32.9m 

AutoReaGas - 160.3m 

0.00 0.20 0.40 0.60 0.80 1.00 1.20 1.40 1.60 1.80 2.00 

Time 

15 1.0 0.10 

13 

0.8 0.08 11 

9 0.6 0.06 

7 

0.4 0.04 

5 

0.02 3 0.2 

1 

0.0 0.00 

-1 

-0.2 -0.02 -3

1.2 0.12 

FLACS - 3.05m 

FLACS - 32.9m 

FLACS - 160.3m 

0.00 0.20 0.40 0.60 0.80 1.00 1.20 1.40 1.60 1.80 2.00 

Time 

17 

15 1.0 0.10 

13 

0.8 0.08 11 

9 0.6 0.06 

7 

0.4 0.04 

5 

3 0.2 0.02 

1 

0.0 0.00 

-1 

-0.2 -0.02 -3

Figure 16.—Calibration experiments and calculations compared. Top: Lake Lynn Experimental Mine 

calibration data; center: calculations from AutoReaGas model; bottom, calculations from FLACS model. 

43
 
4.3 Confined Explosion Models of Large Gas Cloud Volumes 

Having calibrated the models successfully, the next group of models examined larger 

and larger volumes of completely confined explosive mix similar to the first type of gas 

accumulation shown in Figure 7A. The model geometry, shown in Figure 17, is based on the 

same LLEM model employed earlier. Each model has infinitely strong seals placed in the A, B, 

and C drifts 41, 71, 161, 228, or 300 m (135, 233, 528, 748, or 984 ft) from the end of B drift. 

A stoichiometric (10%) methane-air mix fills the entire model volume, and ignition occurs at the 

end of B drift. 

Figure 18 shows the computed pressure-time history at seal B for the larger and larger 

volumes of explosive mix using the AutoReaGas model (Figure 18, top) and the FLACS model 

(Figure 18, bottom). With the 41-m cloud, the pressure rises to about the 908-kPa (132-psia) 

CV explosion pressure over 0.5 sec and then remains at that level as expected. The computed 

pressure-time history shows some reflections, but their magnitude is small. With the 71-m 

(233-ft) cloud, the pressure rises to about 1.0 MPa (145 psia) and then settles down to the 

908-kPa (132-psia) CV explosion pressure. With the larger clouds (161, 228, and 300 m), 

the pressure rises very quickly in less than 0.1 sec to 2–3 MPa (290–435 psia), but then 

equilibrates to the 908-kPa (132-psia) CV explosion pressure as expected. 

As mentioned earlier, these high pressures of more than 1.0 MPa (145 psia) by the 

AutoReaGas and FLACS models are not accurate since detonation may have occurred, and these 

models do not capture DDT or detonation. However, the models are correct in indicating that 

very high pressures have developed. 

Figure 19 summarizes the peak explosion pressures computed for seals A, B, and C by 

the AutoReaGas and FLACS models for larger explosive mix volumes and longer explosion 

lengths. Also shown in this figure are the 908-kPa (132-psia) CV explosion pressure, the 1.76­

MPa (256-psia) CJ detonation pressure, and the 4.50-MPa (653-psia) reflected detonation wave 

pressure. Beyond a length of 100 m (330 ft), the computed pressures are more than 2.0 MPa 

(290 psia), and detonation is very likely. These calculations suggest that gas clouds with run-up 

lengths less than 50 m (165 ft) may not develop pressures much beyond 1.0 MPa (145 psia) and 

may be less likely to detonate unless there are turbulence-intensifying obstructions, such as mine 

timbers, standing supports or other forms of roughness. 

4.4 Partially Confined Explosion Models of Leaking Seals 

This group of models considers an explosive mix that forms directly behind a seal due to 

air leakage, similar to the second type of gas accumulation shown in Figure 7B. This explosive 

mix is only partially confined and able to vent freely into inert atmosphere deeper into the sealed 

area. The model geometry shown in Figure 20 is again based on the LLEM. The model has 

infinitely strong seals in the A, B, and C drifts at 228 m (748 ft) from the beginning of B drift. 

A 10% methane-air mix fills the volume for 15, 30, or 60 m (49, 98, or 197 ft) behind the seals. 

The ignition point is right behind the B-drift seal, which is the worst possible case. 

Figure 21 shows computed pressure-time history at seal B for the various explosive mix 

volumes considered using the AutoReaGas model (Figure 21, top) and the FLACS model 

(Figure 21, bottom). Computed pressures at the B seal range from 100 to 500 kPa (15 to 73 psig) 

and are within the normal operating boundaries of these models. 

44
 
A-22 

A-132 

A-183 

A-233 

A-283 

A-355 

A-453 

A-550 

A-649 

A-807 C-757 

0 30 

Scale, m 

12m 

6m 

24m 

30m 

23m 

15m 

E drift 

A drift

B drift

C drift 

B-108 

B-70 

B-30 

B-10 

B-158 

B-211 

B-257 

B-329 

B-427 

B-782 

C-13 

C-84 

C-134 

C-184 

C-234 

C-304 

C-403 

C-501 

C-598 

Bulkhead 

Ignition Point 

Stage 2 Models ­

Larger explosion volume models 

B-626 

Seal location for 41m explosion volume 

B-526 

Seal location for 71m explosion volume 

Seal location for 300m explosion volume 

Seal location for 161m explosion volume 

Seal location for 228m explosion volume 

Figure 17.—Layout of large-volume confined explosion models. 

45
 
Pressure vs Time History at Seal B - Various Cloud Sizes (AutoReaGas) 

35 3.5 500 

30 3.0 41m cloud 

71m cloud 400 

161m cloud 

25 2.5 300m cloud 

Pressure - bar 20 

15

Pressure - MPa 2.0 300 

1.5 

200 

10 1.0 

100 

5 0.5 

0 0.0 0 

0.50	 0.60 0.70 0.80 0.90 1.00 1.10 1.20 1.30 1.40 1.50 

Time - seconds 

Pressure vs Time at Seal B - Various Cloud Sizes (FLACS) 

35 3.5 500 

30	 3.0 41m cloud
 

71m cloud
 400 

161m cloud 

25 2.5 228m cloud 

Pressure - bar 20 

15

Pressure - MPa 2.0 

1.5 

300m cloud 

300 

200 

Pressure - psi Pressure - psi 

10 1.0 

100 

5 0.5 

0 0.0 0 

1.00 1.20 1.40 1.60 1.80 2.00 2.20 2.40 2.60 2.80 3.00 

Time - seconds 

 Figure 18.—Calculated pressure-time histories at seal for large-volume explosions using AutoReaGas (top)

and FLACS (bottom). 

46
 
Peak Explosion Pressure versus Length of Explosion 

AutoReaGas FLACS 

908 kPa CV explosion pressure 1.76 MPa CJ detonation wave pressure 

4.5 MPa reflected detonation wave pressure 

Length of Explosion - ft
 

0 200 400 600 800 1000
 

0 

1 

2 

3 

4 

5 

6 

Peak Explosion Pressure - MPa 

0 

100 

200 

300 

400 

500 

600 

700 

800 

Peak Explosion Pressure - psi 

Detonation Possible 

Detonation 

Unlikely 

60 

Peak Explosion Pressure - bar 

50 

40 

30 

20 

10

 0 

0 50 100 150 200 250 300 350
 

Length of Explosion - m
 

Figure 19.—Peak explosion pressure versus run-up length. 

47
 
A-22 

A-132 

A-183 

A-233 

A-283 

A-355 

A-453 

A-550 

A-649 

A-807 C-757 

0 30 

Scale, m 

12m 

6m 

24m 

30m 

23m 

15m 

E drift 

A drift

B drift

C drift 

B-108 

B-70 

B-30 

B-10 

B-158 

B-211 

B-257 

B-329 

B-427 

B-782 

C-13 

C-84 

C-134 

C-184 

C-234 

C-304 

C-403 

C-501 

C-598 

30m explosion volume 

Bulkhead 

Ignition Point 

Stage 3 Models ­

Leaky seal models 

Seal 

B-626 

60m explosion volume 

B-526 

15m explosion volume 

Figure 20.—Layout of partially confined, partially filled explosion models. 

48
 
 

 

80 

Pressure vs Time History at Seal B - Various Cloud Sizes (AutoReaGas) 

7 0.7 100 

0.4 

15m cloud 

30m cloud 

60m cloud 

0.50 0.60 0.70 0.80 0.90 1.00 1.10 1.20 

Time - seconds 

6 0.6 

5 0.5 

Pressure - bar 4 

3 

Pressure - MPa 0.3 

2 0.2 

20 

1 0.1 

0 0 0 

Pressure vs Time at Seal B for Various Cloud Sizes (FLACS) 

0 

0.1 

0.2 

0.3 

0.4 

0.5 

0.6 

0.7 

Pressure - MPa 

0 

20 

40 

60 

80 

100 

Pressure - psi 

15m cloud 

30m cloud 

60m cloud 

5 

4 

3 

0 

2 

1 

6 

7 

Pressure - bar 

1.00 1.10 1.20 1.30 1.40 1.50 1.60 1.70 

Time - seconds 

Pressure - psi 

Figure 21.—Calculated pressure-time histories at seal for “leaking seal” explosion models using 

AutoReaGas (top) and FLACS (bottom). 

40 

60 

49
 
Figure 22 shows the computed peak explosion pressures for the 15-, 30-, and 60-m 

(50-, 100-, and 200-ft) gas clouds from the models for the A, B, and C seals. Also shown are the 

measured peak explosion pressures versus gas cloud length for the six calibration experiments 

presented in Table 4. As shown in Figure 22, a simple linear relationship exists between 

explosive mix length and the peak pressure developed at the seal, up to about 30 m (100 ft). 

As the explosive mix length becomes larger and longer, the peak explosion pressure on the seal 

increases. The model calculations extrapolate well from the known LLEM experiments. This 

simple relationship provides practical guidance for both monitoring and the allowable amount of 

explosive mix that can exist behind a seal of given strength. 

5.0 Design Pressure-Time Curves for Seals 

50
 

Previous derivations based on the chemistry and physics of explosions placed bounds on 

the peak pressures that can develop on a seal. The gas explosion models confirmed the 908-kPa 

(132-psia) CV explosion pressure that will develop from any confined gas explosion. The largevolume gas explosion models hinted at the much larger explosion pressures that can develop as a 

result of pressure piling, reflected pressure waves, or detonation. The limited-volume gas 

explosion models of partially confined explosions demonstrate that if proper engineering can 

limit the volume of explosive mix behind a seal, it is possible to limit the explosion pressure that 

could develop. 

Considering the three types of seals discussed in this report and the three types of 

explosive gas accumulations shown in Figure 7, NIOSH engineers developed three design 

pressure-time curves for different seal types under different mining conditions. (In this section 

of the report and hereafter, pressure is always gauge pressure unless otherwise noted.) In the 

4.4-MPa (640-psig) design pressure-time curve shown in Figure 23, the pressure first rises to 

4.4 MPa (640 psig) over 0.001 sec, falls to 800 kPa (120 psig) after 0.1 sec, and then remains at 

that level. The initial pressure rise over 1 ms is consistent with that of detonation waves or 

nonreactive shock waves. Several computed pressure-time histories from the large gas explosion 

models indicate that the initial pressure peaks equilibrate to the 800-kPa (120-psig) 

CV explosion overpressure after 0.1 sec. The 4.4-MPa (640-psig) design pressure-time curve 

encompasses these gas explosion model simulations, which is a conservative engineering 

approach. 

The 800-kPa (120-psig) design pressure-time curve, shown in Figure 24, rises to 800 kPa 

(120 psig) over 0.25 sec and then remains at that level. This pressure rise rate is more 

conservative than the computed rise time for the pressure-time histories from the small-volume, 

confined gas explosion models. This rise time is also consistent with laboratory-scale 

experimental 10% methane-air explosions ignited at the center of a 3.7-m (12-ft) diameter sphere 

reported by Sapko et al. [1976]. 

Finally, the 345-kPa (50-psig) design pressure-time curve, shown in Figure 25, rises to 

345 kPa (50 psig) over 0.10 sec and remains there. Again, this pressure rise rate is more 

conservative than gas explosion model calculations of similar situations. 
 

 

Peak Explosion Pressure versus Length of Explosive Cloud Behind Seal 

Length of explosive cloud - ft
 

0 50 100 150 200
 

0.0 

0.1 

0.2 

0.3 

0.4 

0.5 

0.6 

0.7 

Peak Explosion Pressure - MPa 

0 

20 

40 

60 

80 

100 

Peak Explosion Pressure - psi 

AutoReaGas 

FLACS 

LLEM Experiments 5 

4 

3 

0 

2 

1 

6 

7 

Peak Explosion Pressure - bar 

0 10 20 30 40 50 60 70 
 

Length of explosive cloud - m
 

Figure 22.—Peak explosion pressure versus volume size behind leaking seal: calculations and experimental measurements. 

51
 
NIOSH Design Pressure-Time Curve #1 - Unlimited, Confined Volume
 

Pressures are Overpressure
 

0.0 

0.5 

1.0 

1.5 

2.0 

2.5 

3.0 

3.5 

4.0 

4.5 

Explosion Overpressure - MPa 

0 

50 

100 

150 

200 

250 

300 

350 

400 

450 

500 

550 

600 

650 

Explosion Overpressure - psi 

161m cloud - AutoReaGas 

161m cloud - FLACS 

Design pressure-time curve #1

 0

 5 

10 

15 

20 

25 

30 

35 

40 

45 

Explosion Overpressure - bar 

0.00	 0.10 0.20 0.30 0.40 0.50 0.60 0.70 0.80 0.90 1.00 

Time - seconds 

Figure 23.—4.4-MPa (640-psig) design pressure-time curve and typical model calculations. 

52
 
 7

 6

 5

 4

Explosion Overpressure - ba

 3

r 

Explosion Overpressure - psi 

NIOSH Design Pressure-Time Curve #2 - Limited, Confined Volume
 

Pressures are Overpressure
 

Explosion Overpressure - MPa 

10 1.0

140 

9 0.9

120 

8 0.8

41m cloud - AutoReaGas 

41m cloud - FLACS 

Design pressure-time curve #2

0.00 0.10 0.20 0.30 0.40 0.50 0.60 0.70 0.80 0.90 1.00 

Time - seconds 

0.7 100 

0.6 

0.5 

0.4 

0.3 

2 0.2 

20 

1 0.1

0 0.0 0

Figure 24.—800-kPa (120-psig) design pressure-time curve and typical model calculations. 

40 

60 

80 

53
 
 

40 

3.0 

NIOSH Design Pressure-Time Curve #3 - Limited Volume, Unconfined
 

Pressures are Overpressure
 

4.0 0.40 Explosion Overpressure - MPa 

3.5 0.35 50 

30m cloud - AutoReaGas 

30m cloud - FLACS 

Design pressure-time curve #3 

0.00 0.10 0.20 0.30 0.40 0.50 0.60 0.70 0.80 0.90 1.00 

Time - seconds 

0.30 

Explosion Overpressure - bar 

Explosion Overpressure - psi 

2.5 0.25 

30 2.0 0.20

1.5 0.15 

20 

1.0 0.10 

10 

0.5 0.05

0 0.00 0 

Figure 25.—345-kPa (50-psig) design pressure-time curve and typical model calculations. 

In developing these design pressure-time curves, NIOSH engineers considered the 

following key facts and limitations: 

1.	 For sealed areas of sufficient volume to have an explosion run-up length greater than 

50 m (165 ft) in any direction, detonation of methane-air becomes a possibility. The 

design pressure-time curve must include the 4.50-MPa (653-psia) reflected detonation 

wave pressure followed by the 908-kPa (132-psia) CV explosion pressure. In 

addition, sealed areas with run-up length greater than 50 m (165 ft) may also develop 

high-pressure nonreactive shock waves and their reflections. Most sealed areas of a 

coal mine are confined volumes with no venting possibility. Effectively, the seal will 

see an overpressure of 4.4 MPa (638 psig). 

2.	 For sealed areas with all possible explosion run-up lengths less than 50 m (165 ft), 

detonation and development of nonreactive shock wave reflections are less likely. 

Run-up length is the distance from the ignition point of an explosive mix to where 

high-pressure shock waves of sufficient intensity can develop and induce DDT. Prior 

to DDT, an explosion propagates with subsonic velocity as a deflagration, and after 

DDT, it propagates with supersonic velocity as a detonation. For additional 

discussion of run-up length and DDT, see section 3.6. 

54
 
3.	 For a confined volume of explosive mix with no venting possible, the design 

pressure-time curve should encompass the 908-kPa (132-psia) CV explosion pressure. 

Effectively, the seal must resist 800 kPa (120 psig). Again, most sealed areas of a coal 

mine are confined volumes with no venting possibility. 

4.	 For a partially confined volume of explosive mix with venting, the maximum 

pressure in the design pressure-time curve may be 345 kPa (50 psig) if the length of 

the explosive mix volume behind the seal is limited to 5 m (16 ft) or less. A properly 

managed sealed area atmosphere requires a well-engineered monitoring and 

inertization system to ensure that the length of explosive mix behind a seal does not 

exceed the design limit. 

5.	 For a confined volume of explosive mix with no venting possible, the explosive mix 

can fill up to 40% of the volume and the design pressure-time curve may be 345 kPa 

(50 psig) if the length of the explosive mix volume behind the seal is limited to 5 m 

(16 ft) or less. Note that 40% of the CV explosion overpressure is about 345 kPa 

(50 psig). 

For additional discussion of explosive volume fill, confinement, and venting, see 

section 3.10. 

The most important factor in designing seals and sealing an area centers on an up-front 

management decision of whether to monitor and actively manage the sealed area atmosphere or 

to seal the area and not monitor or manage the sealed area atmosphere in any way. The design 

pressure-time curves presented herein reflect this important management decision. Table 7 

presents the technical criteria governing the use of the design pressure-time curves for the 

structural design of seals in two different scenarios. Scenario 1 pertains to unmonitored seals 

with no monitoring and no inertization. Scenario 2 applies to monitored seals with a managed 

atmosphere behind the seals and inertization as necessary. The associated Figure 26 illustrates 

these scenarios and the technical criteria within schematic mine layouts. 

55
 
Scenario 1: Unmonitored seals Scenario 2: Monitored seals 

Seal type • No monitoring • Managed atmosphere behind seals 

• No inertization • Inertization as necessary 

Panel and district seals • Sealed volume > 50 m (165 ft) long • Monitoring criteria at 5 m (16 ft)

• Run-up length > 50 m (165 ft) >20% CH4 and <10% O2 

• DDT possible • Explosive volume fill < 40% 

• Confined, not vented • DDT less likely 

• Explosive volume fill . 100% • Partially confined and vented 

• Use 4.4-MPa (640-psig) design curve • Use 345-kPa (50-psig) design curve 

• See Figure 23 • See Figure 25 

Panel and district seals • Sealed volume < 50 m (165 ft) long • Monitoring criteria at 5 m (16 ft)

• Run-up length < 50 m (165 ft) >20% CH4 and <10% O2 

• DDT less likely • Explosive volume fill < 40% 

• Partially confined and vented • DDT less likely 

• Explosive volume fill . 100% • Partially confined and vented 

• Use 800-kPa (120-psig) design curve • Use 345-kPa (50-psig) design curve 

• See Figure 24 • See Figure 25 

Crosscut seals • Sealed volume < 50 m (165 ft) long • Monitoring criteria at 5 m (16 ft)

• Run-up length < 50 m (165 ft) >20% CH4 and <10% O2 

• DDT less likely • Explosive volume fill < 40% 

• Partially confined and vented • DDT less likely 

• Explosive volume fill . 100% • Partially confined and vented 

• Use 800-kPa (120-psig) design curve • Use 345-kPa (50-psig) design curve 

• See Figure 24 • See Figure 25

 NOTE.—Not meeting the requirements for limiting the run-up length, the explosive mix volume, and the 

venting of a possible explosion or the monitoring criteria necessitates use of the 4.4-MPa (640-psig) design 

pressure-time curve for seal design. Run-up length is the distance from ignition point to DDT. See section 5.0. 

For explanation of explosive volume fill, confinement, and venting, see section 3.10. 

Table 7.—Technical requirements for the recommended pressure-time curves for structural design 

of new seals in different conditions 

56
 
 

 

Figure 26.—Illustration of design pressure-time curve application for new seal construction. Scenario 1 

depicts unmonitored seals with no monitoring and no inertization. Scenario 2 depicts monitored seals with a 

managed atmosphere behind the seals and inertization as required. Note that not meeting the requirements 

for limiting the run-up length, the explosive mix volume, and the venting of a possible explosion or the 

monitoring criteria necessitates use of the 4.4-MPa (640-psig) design pressure-time curve for seal design. 

Run-up length is the distance from ignition point to DDT. See section 5.0. For explanation of explosive 

volume fill, confinement, and venting, see section 3.10. 

57
 
Table 7 and Figure 26 consider panel and district seal types along with crosscut seal types 

for scenario 1, the unmonitored-sealed-area-atmosphere approach, or scenario 2, the monitored and 

managed-sealed-area-atmosphere approach. The application criteria presented below and in Table 7 

are mutually exclusive and lead to the logical categorization shown. However, if doubt exists, the 

seal design engineer should always use the 4.4-MPa (640-psig) design pressure-time curve. 

Scenario 1A: Unmonitored Panel and District Seals With Detonation or Reflected 

Shock Wave Possibilities 

For unmonitored panel and district seals where the length of the sealed volume exceeds 

50 m (165 ft) in any direction, engineers should use the 4.4-MPa (640-psig) design pressure-time 

curve shown in Figure 23. Because the potential explosion run-up length is more than 50 m 

(165 ft), detonation is a real possibility and thus the possibility of a reflected detonation wave, 

i.e., a reflected reactive shock wave. In addition, sealed areas with run-up length greater than 

50 m (165 ft) may also develop high-pressure nonreactive shock waves and their reflections. The 

sealed area for this case is completely confined, not vented in any way, and 100% filled with 

explosive mix (Figure 26A). The situation depicted here may occur in many sealed areas, 

especially right after sealing during the initial inertization phase. 

Scenario 1B: Unmonitored Panel and District Seals Without Detonation or Reflected 

Shock Wave Possibilities 

For unmonitored panel and district seals where the length of the sealed volume does not 

exceed 50 m (165 ft) in any direction, engineers can use the 800-kPa (120-psig) design pressuretime curve shown in Figure 24. Because the potential explosion length is less than 50 m (165 ft), 

detonation or high-pressure nonreactive shock waves and their reflections are less likely to 

develop. A potential explosion will still reach the 800-kPa (120-psig) CV explosion overpressure. The sealed area for this case is completely filled with explosive mix and is confined, 

but it can vent somewhat into the broken rock of a mined-out area, i.e., the gob (Figure 26B). 

This situation is also common and may arise when sealing a full-extraction panel, either longwall 

or room-and-pillar. 

Scenario 1C: Unmonitored Crosscut Seals Without Detonation or Reflected Shock Wave 

Possibilities 

For unmonitored crosscut seals, the length of the sealed volume will not likely exceed 

50 m (165 ft) in current mining practice. As before, detonation along with nonreactive shock 

waves and their reflections are less likely, and engineers can use the 800-kPa (120-psig) design 

pressure-time curve shown in Figure 24. The sealed volume is completely filled with explosive 

mix, is confined, and can vent somewhat into the gob (Figure 26C). This situation arises 

commonly at longwall mines extracting spontaneous combustion-prone coal. 

Scenario 2D: Monitored Panel and District Seals 

For monitored panel and district seals where the length of the sealed volume exceeds 

50 m (165 ft) in any direction, if monitoring can ensure that the maximum length of explosive 

mix behind a seal does not exceed 5 m (16 ft) and the volume of explosive mix does not exceed 

40% of the total sealed volume, engineers can use the 345-kPa (50-psig) design pressure-time 

58
 
curve shown in Figure 25. The limited volume explosive mix is partially confined and able to 

vent into the inert atmosphere beyond (Figure 26D). This situation will arise in the atmosphere 

behind a panel or district seal that first becomes inert and then, due to subsequent air leakage, 

develops a localized explosive mix. 

Scenario 2E: Monitored Panel and District Seals 

For monitored panel and district seals where the length of the sealed volume is less than 

50 m (165 ft) in any direction, if monitoring can ensure that the maximum length of explosive 

mix behind a seal does not exceed 5 m (16 ft) and the volume of explosive mix does not exceed 

40% of the total sealed volume, engineers can again use the 345-kPa (50-psig) design pressuretime curve shown in Figure 25. This situation will develop behind seals to a full-extraction panel 

that later leak (Figure 26E). 

Scenario 2F: Monitored Crosscut Seals 

For monitored crosscut seals where the length of the sealed volume is less than 50 m 

(165 ft) in any direction, if monitoring can ensure that the maximum length of explosive mix 

behind a seal does not exceed 5 m (16 ft) and the volume of explosive mix does not exceed 40% 

of the total sealed volume, engineers can use the 345-kPa (50-psig) design pressure-time curve 

shown in Figure 25. This situation will develop behind crosscut seals in spontaneous 

combustion-prone longwall mines (Figure 26F). 

In summary, NIOSH engineers developed three explosion pressure design pressure-time 

curves to describe the structural loading on mine seals resulting from a methane-air explosion in 

the sealed area of a coal mine under several different conditions. If these conditions are not met, 

the engineer responsible for a seal design should use the conservative 4.4-MPa (640-psig) design 

pressure-time curve. 

6.0 Minimum New Seal Designs to Withstand the Design Pressure-Time Curves 

The explosion pressure design criteria for new seals developed in the preceding sections 

serve as a basis for the structural design. In this section, NIOSH engineers present examples for 

possible approaches to new seal designs using simplified structural engineering methods. The 

main purpose of these structural analyses and the minimum seal thickness charts presented 

herein is not to provide final design recommendations for seals to resist the three proposed 

explosion pressures, but to provide approximate seal design thickness so that engineers can judge 

the merits of alternative sealing strategies. With the new proposed explosion pressure design 

criteria, the mining industry faces choices of sealing with thicker seals and not monitoring, or 

sealing with thinner seals and monitoring, or not sealing and continuing to ventilate. To weigh 

these alternatives and decide on a strategy, these approximate design charts provide engineers 

with approximate values for seal thickness using the range of typical seal construction materials. 

Again, these design charts are not intended as a substitute for proper structural design of seals 

with site investigation, structural analysis, and quality control during construction. Furthermore, 

they should not be accepted by MSHA as a substitute for the required structural engineering. 

Due to the complex nature of the structural interface between the mine roof and floor 

rock strata, the coal ribs, and the seal, a general design for a mine seal is not possible. The 

59
 
fundamental design assumptions change from application to application so that each seal design 

will have to be engineered for a specific application and location in a given mine. 

The following considerations should serve as conceptual ideas for new seal designs and 

demonstrate that it is possible to engineer a mine seal to withstand these possible explosion 

pressures. The two structural engineering approaches used, one-way arching and plug-type failure, 

only demonstrate two possible failure modes that are both dependent on the structural reactions of 

the surrounding strata. There are other structural engineering approaches to designing such seals, 

but a detailed discussion of these methods goes beyond the scope of this study. 

The design pressure-time curves developed in the prior section depart significantly from 

the 140-kPa (20-psig) explosion pressure design criterion found in recent U.S. mining 

regulations and the 345-kPa (50-psig) standard currently in force. NIOSH engineers conducted 

structural analyses with these design pressure-time curves to develop practical design charts 

using three separate design approaches: 

1.	 Dynamic structural analysis using the Wall Analysis Code (WAC) developed by the 

U.S. Army Corps of Engineers for designing protective structures subject to blast 

loads. 

2.	 Static plug analysis using quasi-static approximations to the dynamic design pressuretime curves. 

3.	 Static arching analysis using the same quasi-static load approximations. 

These three significantly different analysis methods generated similar seal thickness 

requirements and confidence in the design charts presented herein for evaluating alternative 

sealing strategies. 

In conducting these structural analyses, NIOSH engineers considered nine typical 

construction materials covering the range of materials readily available to the mining industry. 

Table 8 summarizes these material properties, which range from high-strength, lowdeformability to low-strength, high-deformability materials. Each material has potential 

application depending on the particular circumstances of the seal. The high- and mediumstrength materials are characterized by assumed compressive strengths ranging from 3.5 to 

24 MPa (500 to 3,500 psi). The low-strength materials are characterized by assumed shear 

strengths ranging from 0.13 to 0.50 MPa (18 to 72 psi). 

Table 8.—Typical material properties for seal construction 

60 

Compressive Shear strength Density strength 

HIGH-STRENGTH, HIGH-DENSITY, LOW-DEFORMABILITY MATERIALS 

1 ................ 24 MPa (3,500 psi) — 2,400 kg/m3

 (150 pcf)
 

2 ................ 17 MPa (2,500 psi) — 1,900 kg/m3

 (120 pcf)
 

3 ................ 10 MPa (1,500 psi) — 2,400 kg/m3

 (150 pcf)
 

MEDIUM-STRENGTH, MEDIUM-DENSITY, MEDIUM-DEFORMABILITY MATERIALS 

4 ................ 8 MPa (1,200 psi) — 1,760 kg/m3

 (110 pcf)
 

5 ................ 5 MPa (750 psi) — 1,600 kg/m3

 (100 pcf)
 

6 ................ 3.5 MPa (500 psi) — 1,600 kg/m3

 (100 pcf)
 

LOW-STRENGTH, LOW-DENSITY, HIGH-DEFORMABILITY MATERIALS 

7 ................ — 0.50 MPa (72 psi) —
 

8 ................ — 0.25 MPa (36 psi) —
 

9 ................ — 0.13 MPa (18 psi) —
 
For structural analysis, the recommended design pressure-time curves may have a quasistatic approximation that can apply in practical situations. The 800-kPa (120-psig) pressure-time 

curve (Figure 24) and the 345-kPa (50-psig) pressure-time curve (Figure 25) remain at these 

pressures for a long duration. For the 800-kPa (120-psig) pressure-time curve, static pressure of 

800 kPa (120 psig) is equivalent; similarly for the 345-kPa (50-psig) curve, static pressure of 345 

kPa (50 psig) is equivalent. Furthermore, the rise time for these pressure-time curves is 0.25 and 

0.1 sec, respectively, which is much more than the transit time for a stress wave across a seal. 

NIOSH engineers estimate that this transit time ranges from 0.0001 to 0.010 sec, which is much 

less than the rise times of these two design pressure-time curves. 

NIOSH engineers did not attempt to approximate the 4.4-MPa (640-psig) design 

pressure-time curve shown in Figure 23 with an equivalent static load. Additional studies are 

required to develop a reliable quasi-static approximation to this pressure-time curve. 

NIOSH engineers also note that repeated pressure waves will likely impact a seal 

structure, as shown by gas explosion model computations in Figure 18. These multiple waves 

arise from pressure wave reflections due to the complex mine geometry. A possibility exists that 

these repeated pressure waves could resonate with a natural frequency of the structure; however, 

NIOSH engineers view this scenario at this time as unlikely. While the period of these repeated 

pressures waves could be similar to the natural period of a seal structure, the number of waves is 

limited and their magnitude is decreasing. 

6.1 Dynamic Structural Analysis With WAC 

WAC is a single-degree-of-freedom (SDOF) structural dynamics model that solves the 

equation of motion to determine the displacement-time history at midheight of a wall. Failure 

occurs if this displacement exceeds a given limit. Following Slawson [1995], the equation of 

motion for a SDOF system is 

M · y′′(t) + Cd · y′(t) + R(y(t)) = F(t) (10) 

where M = equivalent or “lumped” mass of the system, 

 Cd = damping coefficient taken as 5% of the critical value, i.e., very lightly damped, 

 y(t) = displacement of the mass as a function of time, t, 

 y′(t) = velocity of the mass or first derivative of displacement, 

 y′′(t) = acceleration of the mass or second derivative of displacement, 

 R = structural resistance as a function of displacement, 

and F = the structural load as a function of time, i.e., one of the design pressure-time 

 curves developed earlier. 

For a resistance function, NIOSH engineers used the “unreinforced wall with one-way 

arching” option within WAC. In this option, the supports are rigid at the roof and floor, while the 

walls are unrestrained. The fundamental assumption underlying the arching analysis is that the 

seal has rigid contact with the roof and floor and that movement along these surfaces does not 

happen in a shear or plug failure mode. The design engineer will need to verify that this 

assumption holds true before proceeding with this WAC analysis. In the arching failure 

mechanism, the wall is assumed to crack horizontally at midheight and at the roof and floor upon 

application of the blast load. As shown in Figure 27, the two blocks remain rigid, rotate through 

an angle θ, and develop arching forces to resist the blast loading. The wall will begin to crush at 

61
 
H/2 

t 

H/2 

Blast pressure 

θ

Figure 27.—One-way arching failure mechanism in Wall Analysis Code (WAC). 

Resistance force 

DisplacemDisplacement 

Figure 28.—Typical resistance function for unreinforced wall with one-way arching (after Slawson [1995]). 

the points indicated, and the magnitude of the resisting forces will depend on the compressive 

strength of the wall material. Figure 28 (after Slawson [1995]) shows a typical resistance 

function for an unreinforced wall with one-way arching. 

62 
 

 

 

 

 

The arching model for wall behavior applies best when the wall thickness to wall height 

ratio ranges from about 1/15 to 1/4 [Coltharp 2006]. For lower thickness-to-height ratios, 

a flexural failure mechanism dominates, whereas for higher ratios, a shear failure mechanism 

along the wall edges becomes more dominant. Most of the analyses presented herein meet this 

criterion for the arching failure mechanism. 

As a failure criterion, NIOSH engineers selected an allowable rotation angle θ of 1°. 

The displacement at failure in the SDOF model calculations is 

 H 

yFail = tanθ (11) 2 

where H = wall height, 

and θ = allowable rotation angle. 

For a 3-m (10-ft) high wall, the displacement at failure is about 2.5 cm (1 in). This displacement 

is consistent with prior testing at the NIOSH Pittsburgh Research Laboratory. 

Guidelines for the use of WAC suggest a 1° rotation angle to provide a “medium level of 

protection.” At this level of protection, a wall subject to blast loading has cracked and displaced 

substantially, but it has survived. The wall may require repair and may not survive additional 

blast loadings. NIOSH engineers therefore selected an allowable rotation angle θ of 1° since that 

level of protection best meets the intended purpose of a seal. Finally, to achieve an additional 

safety factor of 2 with WAC, NIOSH engineers scaled the computed minimum seal thicknesses 

by a factor of square root of 2 . This scaling effectively doubles the applied load on the structure. 

6.2 Quasi-static Analysis With a Plug Formula and Anderson’s Arching Formula 

As mentioned earlier, NIOSH engineers used two additional quasi-static approaches to 

compute minimum seal thickness. The first approach analyzes the seal as a simple plug loaded 

by a pressure load on the face and restrained by shear forces around the perimeter. Safety factor 

for plug failure is 

 SS (2W + 2 H )tS SFPF = (12) PS W H 

where SS = either the shear strength of the seal material, shear strength of the surrounding 

 rock, or shear strength of the interface, whichever is less, 

PS = static pressure load, 

W = seal width, 

H = seal height, 

and tS = seal thickness. 

Solving for seal thickness, we obtain 

63
 

 P W H SF 

tS = 

S PF (13) SS (2W + 2 H ) 
2

⎛

⎜

⎝

t S ⎞

⎟

⎠ PS = 0.72 n f k (14) H
 

PS t (15) S = H 

0.72 n f k 

For a simple plug failure analysis to apply best, the thickness-to-height ratio of the seal 

should exceed 1. Table 8 shows the shear strength for the three low-strength, high-deformability 

seal materials considered in this analysis. 

Based on Anderson’s [1984] simple three-hinged arch theory, Sapko et al. [2005] 

developed the following formula relating the ultimate pressure-bearing capacity of a seal, PS, 

to the compressive strength of the seal material and the seal dimensions: 

where fk = the compressive strength of the seal material, as given in Table 8, 

and n = an empirical factor ranging from 0.75 to 1.25. 

Solving for seal thickness, we obtain 

For Anderson’s arching analysis to apply, the thickness-to-height ratio of the seal should 

fall within the range of 1/15 to 1/4, similar to the preferred range with WAC. 

6.3 Design Charts for Minimum Seal Thickness 

Based on a seal width of 6.1 m (20 ft) and the materials shown in Table 8, NIOSH 

engineers calculated a minimum seal thickness versus height for the three design pressure-time 

curves using WAC, plug analysis, and Anderson’s arching analysis. As mentioned earlier, the 

minimum seal thicknesses computed by WAC are scaled by a factor of 2 , which effectively 

applies a safety factor of 2 to the design load. A safety factor of 2 is applied explicitly in the plug 

analysis. Computed minimum seal thicknesses from both analyses are combined to form the 

design charts shown in Figures 29, 30, and 31 for the 4.4-MPa (640-psig), 800-kPa (120-psig), 

and 345-kPa (50-psig) design pressure-time curves, respectively. These very different analyses 

merged well to form these design charts. In transitioning between methods, NIOSH engineers 

had to decide between the two analysis methods, recognizing that a WAC analysis applies best 

when the seal thickness-to-height ratio is less than 1/4, whereas plug analysis applies best when 

that ratio exceeds 1. Accordingly, NIOSH engineers selected the WAC analysis when the ratio 

was less than 1/2 and plug analysis when the ratio exceeded 1/2. However, this selection was 

made at a safety factor of 1 and not 2. 

Figure 29 shows seal solutions for the 4.4-MPa (640-psig) design pressure-time curve 

(Figure 23), Figure 30 shows the same for the 800-kPa (120-psig) design pressure-time curve 

(Figure 24), and Figure 31 shows possibilities for the 345-kPa (50-psig) design pressure-time 

curve (Figure 25). Withstanding the 4.4-MPa (640-psig) design pressure-time curve presents the 

greatest challenge; however, as shown in Figure 29, in a 2-m (80-in) high coal seam, a 1-m 

(40-in) thick seal with material strength of 24 MPa (3,500 psi) or a 1.2-m (48-in) thick seal with 

material strength of 17 MPa (2,500 psi) will resist this worst-case design pressure-time curve. 

Such a seal might require about 15 m3

 (20 yd3) of material to construct. As mentioned in prior 

64
 
 

UCS=24 MPa (3500 psi) 2.40 S.G. (150 pcf) - WAC UCS=17 MPa (2500 psi) 1.92 S.G. (120 pcf) - WAC 

UCS=10 MPa (1500 psi) 2.40 S.G. (150 pcf) - WAC UCS=8 MPa (1200 psi) 1.76 S.G. (110 pcf) - WAC 

20 40 60 80 100 120 

Seal height - inches 

140 160 180 

0.00 

1.00 

2.00 

3.00 

4.00 

5.00 

Minimum seal thickness - m 

0 

20 

40 

60 

80 

100 

120 

140 

160 

180 

Minimum seal thickness - inches 

4.4 MPa (640 psi) 

0.50 1.00 1.50 2.00 2.50 3.00 3.50 4.00 4.50 5.00 

Seal height - m 

Figure 29.—Design chart for minimum seal thickness with 4.4-MPa (640-psig) design pressure-time curve 

using various construction materials. (UCS = uniaxial compressive strength; S.G. = specific gravity.) 

discussions, this design pressure-time curve applies to unmonitored district or panel seals. Seal 

designs to withstand the 4.4-MPa (640-psig) design pressure-time curve using lower-strength and 

lighter-weight construction materials are not presented pending analyses with methods other than 

WAC. Use of such lower-strength and lighter-weight materials will require very thick pluglike 

seals with thickness greater than the seal height, as suggested by Figure 29. 

As shown in Figure 30, numerous options exist to withstand the 800-kPa (120-psig) 

design pressure-time curve. For a 2-m (80-in) high coal seam, various materials about 1.0- to 

5.0-m (40- to 200 in) thick could meet the challenge. As shown in Figure 31, many currently 

used seal construction materials offer possibilities to withstand the 345-kPa (50-psig) design 

pressure-time curve. 

65
 
 

 

UCS=24 MPa (3500 psi) 2.40 S.G. (150 pcf) - WAC UCS=17 MPa (2500 psi) 1.92 S.G. (120 pcf) - WAC UCS=10 MPa (1500 psi) 2.40 S.G. (150 pcf) - WAC 

UCS=8 MPa (1200 psi) 1.76 S.G. (110 pcf) - WAC UCS=5 MPa (750 psi) 1.60 S.G. (100 pcf) - WAC UCS=3.5 MPa (500 psi) 1.60 S.G. (100 pcf) - WAC 

Shear Strength=0.50 MPa (72 psi) - Plug Analysis Shear Strength=0.25 MPa (36 psi) - Plug Analysis 

Seal height - inches 

20 40 60 80 100 120 140 160 180 

0.00 

0.50 

1.00 

1.50 

2.00 

2.50 

3.00 

3.50 

4.00 

4.50 

5.00 

Minimum seal thickness - m 

0 

10 

20 

30 

40 

50 

60 

70 

80 

90 

100 

110 

120 

130 

140 

150 

160 

170 

180 

190 

Minimum seal thickness - inches 

800 kPa (120 psi) 

0.50 1.00 1.50 2.00 2.50 3.00 3.50 4.00 4.50 5.00 

Seal height - m 

Figure 30.—Design chart for minimum seal thickness with 800-kPa (120-psig) design pressure-time curve 

using various construction materials. (UCS = uniaxial compressive strength; S.G. = specific gravity.) 

66
 
 

UCS=24 MPa (3500 psi) 2.40 S.G. (150 pcf) - WAC UCS=17 MPa (2500 psi) 1.92 S.G. (120 pcf) - WAC UCS=10 MPa (1500 psi) 2.40 S.G. (150 pcf) - WAC 

UCS=8 MPa (1200 psi) 1.76 S.G. (110 pcf) - WAC UCS=5 MPa (750 psi) 1.60 S.G. (100 pcf) - WAC UCS=3.5 MPa (500 psi) 1.60 S.G. (100 pcf) - WAC 

Shear Strength=0.50 MPa (72 psi) - Plug Analysis Shear Strength=0.25 MPa (36 psi) - Plug Analysis Shear Strength=0.13 MPa (18 psi) - Plug Analysis 

Seal height - inches 

20 40 60 80 100 120 140 160 180 

0.00 

0.50 

1.00 

1.50 

2.00 

2.50 

3.00 

3.50 

4.00 

4.50 

5.00 

Minimum seal thickness - m 

0 

20 

40 

60 

80 

100 

120 

140 

160 

180 

Minimum seal thickness - inches 

345 kPa (50 psi) 

0.50 1.00 1.50 2.00 2.50 3.00 3.50 4.00 4.50 5.00 

Seal height - m 

Figure 31.—Design chart for minimum seal thickness with 345-kPa (50-psig) design pressure-time curve 

using various construction materials. (UCS = uniaxial compressive strength; S.G. = specific gravity.) 

6.4 Additional Structural Requirements for New Seals 

The design charts for minimum seal thickness contain a safety factor of 2. With concrete 

and similar seal construction materials, NIOSH engineers also recommend the use of steel 

reinforcement bar to (1) better anchor the seal structure to the surrounding rock and (2) increase 

the flexural strength of the seal. Reinforcing steel within the seal also helps ensure that the 

structure fails in a gradual, ductile mode rather than a catastrophic, brittle mode. 

Seals may also be hitched into solid roof, rib, and floor rock to ensure that the seal 

adheres adequately to the surrounding strata. Numerous methods exist to accomplish adequate 

hitching, such as excavating into the roof, rib, or floor to solid material, steel reinforcement 

anchors, and other methods. 

An additional recommended change in current practice pertains to the use of water traps 

in seals to drain possible water accumulation. NIOSH engineers recommend the discontinuance 

of water traps in seals, since water traps conflict with the main purpose of a seal, namely, 

explosion protection. The available head in a water trap is insufficient to resist the recommended 

design pressure-time curves. If water accumulation is anticipated in the low point of a sealed 

67
 
area, then engineers should design and install a pumping system to remove the water without 

compromising the intended explosion protection purpose of the seal. 

6.5 Alternative Structural Analyses of New Seals 

The structural analyses of seals presented herein used the dynamic WAC and a simple 

static plug analysis. Using these simple methods, NIOSH engineers developed design charts for 

recommended minimum seal thickness using typical construction materials and for 

recommended minimum number of anchorage reinforcement bar. Again, the main purpose of 

these design charts is not to provide final design recommendations for seal thickness, but to 

provide approximate seal thickness so that engineers can judge the merits of alternative sealing 

strategies. Analysis with more sophisticated methods may lead to better, more economic seal 

designs. 

The structural analysis method should consider all likely failure modes, including 

flexural, compressive, or shear failure through the seal material along with shear failure through 

the rock or at the rock-seal interface. The structural loads requiring consideration include the 

explosion pressure loading, convergence loading, and water pressure behind the seal. The 

analysis should include the effect of both structural reinforcement within the seal and structural 

linkages to the surrounding rock. The analysis should consider all relevant mechanical properties 

of the seal material, any reinforcement, and the surrounding rock strata. The analysis should also 

use minimum material property values that the seal and surrounding rock strata will meet and 

exceed during actual construction. Finally, considering the uncertainties associated with the seal 

foundation, seal construction materials, and construction practices, NIOSH engineers 

recommend applying a safety factor of 2 in the structural analysis. However, if adequate quality 

control practices exist, a lower safety factor may be considered. 

7.0 Summary of Procedures for New Seal Design 

7.1 Two Approaches to Sealing Mined-out Areas 

An explosive methane-air mix that can accumulate within the sealed areas of a coal mine 

poses a serious safety hazard to all underground mining personnel. If the sealed area atmosphere 

should explode, the CV explosion pressure of 908 kPa (132 psia) is the minimum pressure for 

which mining engineers must plan. Pressure piling can drive the pressure beyond this level. For 

large-volume explosive gas accumulations having a length of more than 50 m (165 ft) in any 

direction, a methane-air mix can detonate, in which case the detonation wave will reach 1.76 

MPa (256 psia). When a detonation wave reflects from a seal, the reflected detonation wave 

pressure is 4.50 MPa (653 psia). In addition, sealed areas with run-up length greater than 50 m 

(165 ft) may also develop high-pressure nonreactive shock waves and their reflections. 

Considering the explosion pressures that can develop, NIOSH engineers developed three 

design pressure-time curves for the dynamic structural analysis of seals. These design pressuretime curves apply to two different design approaches for sealed areas. 

Figure 1 is a simple flowchart that illustrates the key decisions in choosing between the 

two design approaches and the design pressure-time curves. The first decision is either to 

monitor and manage the sealed area atmosphere or to leave the sealed area completely 

unmonitored after seals achieve their design strength. Monitoring and managing the sealed area 

68
 
atmosphere to ensure that the possible explosive mix behind a seal does not extend more than 

5 m (16 ft) behind a seal and that this volume of explosive mix does not exceed 40% of the 

sealed volume behind a seal allows the use of 345-kPa (50-psig) seals. The monitored sealed area 

atmosphere approach and the 345-kPa (50-psig) seal design standard is consistent with current 

practices in Europe and Australia. However, an unmonitored sealed area with no control over the 

possible explosive mix that could develop behind a seal will require 4.4-MPa (640-psig) seals if 

detonation and nonreactive shock waves and their reflections are possibilities because the run-up 

length exceeds 50 m (165 ft). If detonation or nonreactive shock waves and their reflections are 

less likely to develop because the run-up length is less the 50 m (165 ft), then a 800-kPa (120­

psig) seal is allowed. 

Scenario 1 as shown in Table 7 and Figure 26 applies to unmonitored seals with no 

monitoring and no inertization after sealing is completed and the seals achieve their design 

strength. As specified in Table 7, if the run-up length within the sealed area exceeds 50 m 

(165 ft) in any direction, then engineers should apply the 4.4-MPa (640-psig) design pressuretime curve. If the run-up length does not exceed 50 m (165 ft), then the 800-kPa (120-psig) 

design pressure-time curve may apply. 

Scenario 2, the monitored, managed-seal-area-atmosphere approach, applies when 

continuous monitoring ensures that an explosive mix no larger than 5 m (16 ft) long does not 

develop behind a seal and that the volume of explosive mix does not exceed 40% of the sealed 

volume. Limiting the potential volume of explosive mix through monitoring and possible 

inertization will limit the pressure rise of a potential explosion and allow the use of the 345-kPa 

(50-psig) design pressure-time curve. 

In the unmonitored approach shown in scenario 1, atmospheric monitoring behind the 

seals and artificial inertization of the sealed area atmosphere are not required after sealing is 

done and the seals reach design strength. However, during seal construction and initial selfinertization, monitoring of the sealed area must ensure that an explosive mix does not develop 

until the seal achieves its design strength. If an explosive mix develops prematurely, appropriate 

action must be taken immediately until the sealed area atmosphere becomes inert and the seal 

reaches its design strength. 

69
 

7.2 Design, Construction, and Inspection for New Sealed Areas 

NIOSH engineers recommend a four-phase approach to ensure the desired level of 

seal performance: (1) information gathering, (2) seal engineering, (3) seal construction, and 

(4) postsealing inspection. 

(1) During the information-gathering phase, a licensed professional engineer should: 

•	 Choose appropriate seal locations and indicate these locations on a mine map. 

•	 Assess the convergence loading potential of each site. 

•	 Estimate the ventilation pressure differential across the seals and across the sealed 

area. 

•	 Estimate the air leakage potential at each seal site. 

•	 Estimate the water pressure that could develop behind the seals. 

•	 Assess atmospheric monitoring requirements during and after sealing, and specify 

the location and frequency of samples to be analyzed. 

•	 Certify all information with a signature and professional engineer’s stamp. 
(2) In the seal engineering phase, a licensed professional engineer should: 

•	 Assess the explosion potential from the sealed area behind each seal. This 

assessment should consider the volume of the sealed area, the maximum run-up 

length for a possible explosion, the degree of filling with explosive mix, the 

degree of confinement in the sealed area, and the degree of venting possible from 

a worst-case explosion. 

•	 Choose which design approach to follow when sealing. The choice is either the 

unmonitored approach or the monitored, managed-seal-area-atmosphere 

approach. 

•	 Choose an explosion design pressure-time curve using the criteria specified in 

Table 7. 

•	 Design the seal and specify all dimensions, construction materials, reinforcement, 

foundation requirements, and any grouting of the surrounding rock. The structural 

analysis should consider flexural, compressive, and shear failure of the seal 

material and possible shear failure through the surrounding rock or the rock-seal 

interface. The seal design must resist the explosion design pressure-time curve, 

resist any water pressure, and limit air leakage. 

•	 Design the ventilation system surrounding the sealed area to minimize air leakage 

into the sealed area. 

•	 Design a monitoring system and develop a monitoring plan commensurate with 

the selected design approach. For the unmonitored approach, some monitoring is 

required during seal construction to ensure that an explosive mix does not 

accumulate within the sealed area before the seal reaches its design strength. The 

monitored, managed-seal-area-atmosphere approach requires continuous 

monitoring of the sealed area throughout the remaining life of the mine to ensure 

that no more than 5 m (16 ft) of explosive atmosphere could exist behind the seal. 

The monitoring system design must specify the location of monitoring points and 

the frequency of monitoring. The required sampling frequency must consider the 

estimated air leakage through a seal to ensure that an explosive mix does not 

develop in between samples. 

•	 Certify all design documents with a signature and professional engineer’s stamp. 

(3) During seal construction, a licensed professional engineer should: 

•	 Perform quality control to ensure that actual construction follows the specified 

design. This quality assurance program should document that all seal dimensions, 

construction material properties, and the seal foundation meet the required design 

standards. 

•	 Certify the actual seal construction as done according to specification in the 

approved plan with a signature and professional engineer’s stamp. 

(4) Finally, regular postsealing inspection by mining personnel should: 

•	 Follow the continuous monitoring plan for the sealed area atmosphere if the 345­

kPa (50-psig) design pressure-time curve and the managed-sealed-areaatmosphere approach were chosen. 

•	 Monitor the structural integrity of seals and conduct structural repairs as 

necessary. 

•	 Check for any unplanned air leakage and eliminate leaks as necessary. 

•	 Check for any unplanned water accumulation behind the seal and conduct repairs 

as necessary. 

70
 
7.3 New Research and Development in Seal Design 

Over the next 3 years, NIOSH will complete a research program aimed at preventing 

explosions within sealed areas of mines and developing sealing technologies to better protect 

mining personnel. The research program may have four broad areas: 

1.	 Fundamental understanding of gas and dust explosions in abandoned and sealed areas 

of coal mines. 

2.	 Design procedures for sealing abandoned areas, including estimation of potential 

explosion forces, structural design of seals, and risk assessment procedures to define 

the gas and dust explosion threat. 

3.	 Management systems to control explosive mixtures in abandoned and sealed areas, 

including atmospheric monitoring and inertization systems for gob areas. 

4.	 Education of miners, mining engineers, and mine managers about the extreme 

hazards posed by methane in abandoned and sealed areas of coal mines and methods 

to manage the hazard. 

NIOSH researchers will collaborate with the U.S. National Laboratories to further 

examine the dynamics of methane and coal dust explosions in mines. Using CFD programs, 

researchers will seek understanding of DDT, along with the detonation phenomena and physical 

factors that control it. Large-scale explosion tests in the LLEM will provide calibration data for 

the numerical models and confirm or deny model predictions. NIOSH researchers will continue 

to use commercially available gas explosion models for additional practical insights into 

explosion processes. 

NIOSH researchers will also examine further the dynamic response of seals to gas and 

coal dust explosion loading, again in collaboration with the U.S. National Laboratories. This 

work seeks techniques to protect seals from transient pressures. Additional research will produce 

design guidelines for all aspects of seal design, including site selection, geotechnical 

considerations, construction practices, maintenance, inspection procedures, as well as the 

structural response. Again, in collaboration with the U.S. National Laboratories, NIOSH will 

develop procedures to assess the risk associated with sealing abandoned areas of coal mines. 

Additional work will include field measurements of the atmosphere within sealed areas. 

NIOSH will become a mining industry resource and leading proponent of the use of atmospheric 

monitoring and inertization systems for sealed areas of coal mines. NIOSH researchers may 

collaborate with industry partners to develop improved sealed area atmospheric monitoring 

systems and promote the adoption of such technology by the mining industry. Finally, NIOSH 

researchers will educate miners, mining engineers, and mine managers about the extreme hazards 

that can arise from any abandoned and sealed area of a coal mine. 

In summary, the design procedures in this report treat mine seals as safety-critical 

structures, whose failure could create a life-threatening situation. Accordingly, mine seals and 

their related monitoring, inertization, and ventilation systems require the highest level of 

engineering and quality assurance. Successful implementation of the seal design criteria and 

recommendations in this report should reduce the risk of seal failure due to explosions in 

abandoned areas of underground coal mines. 

71
 
8.0 References 

Anderson C [1984]. Arching action in transverse laterally loaded masonry walls. Struct 

Eng 62B(1):22. 

Baker WE, Cox PA, Westine PS, Kulesz JJ, Strehlow RA [1983]. Explosion hazards and 

evaluation. New York: Elsevier. 

Bartknecht W [1993]. Explosionsschutz: Grundlagen und Anwendung (in German). 

Berlin, Germany: Springer-Verlag. 

Brookes, DE, Leeming JR [1999]. The performance of explosion proof stoppings. In: 

Proceedings of the 28th International Conference of Safety in Mines Research Institutes (Sinaia, 

Romania, June 7–11, 1999). Vol. I. Petroşani, Romania: National Institute for Mining Safety and 

Explosion Proof Protection, pp. 59–69. 

Brookes DE, Nicol AM [1997]. Design criteria for explosion-proof stoppings: test on 

large-scale stoppings. Buxton, U.K.: Health & Safety Laboratory, Dust Explosion Section, 

ref DE/97/09. 

Cashdollar KL [1996]. Coal dust explosibility. J Loss Prev Process Ind 9(1):65–76. 

Cashdollar KL, Zlochower IA, Green GM, Thomas RA, Hertzberg M [2000]. 

Flammability of methane, propane, and hydrogen gases. J Loss Prev Process Ind 13:327–340. 

Century Dynamics, Inc. [2007]. AutoReaGas: the cost-effective validated software for 

gas explosion modeling. [http://www.century-dynamics.com/dc_products/autoreagas/overview/ 

autoreagas.pdf]. Date accessed: June 2007. 

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

Office of the Federal Register. 

Checca EL, Zuchelli DR [1995]. Lightning strikes and mine explosions. In: Wala AM, 

ed. Proceedings of the Seventh U.S. Mine Ventilation Symposium. Lexington, KY: University 

of Kentucky, pp. 245–250. 

Coltharp D [2006]. Personal communication. 

Coward HF, Jones GW [1952]. Limits of flammability of gases and vapors. Pittsburgh, 

PA: U.S. Department of the Interior, Bureau of Mines, Bulletin 503. 

Cybulski WB [1975]. Coal dust explosions and their suppression. Translated from Polish. 

Warsaw, Poland: National Center for Scientific, Technical, and Economic Information. NTIS 

No. TT 73–54001. 

Cybulski WB, Gruszka JH, Krzystolik PA [1967]. Research on firedamp explosions in 

sealed-off roadways (abstract in English and German; complete paper in German only). In: 

Proceedings of the 12th International Conference of Mine-Safety Research Establishments 

(Dortmund, Germany, September 11–15, 1967). Vol. I, paper No. 22. 

Dorofeev SB, Kochurko AS, Sidorov VP, Bezmelnitsin AV, Breitung WM [1996]. 

Experimental and numerical studies of the pressure field generated by DDT events. Shock 

Waves 5:375–379. 

Dorofeev SB, Sidorov VP, Kuznetsov MS, Matsukov ID, Alekseev VI [2000]. Effect of 

scale on the onset of detonations. Shock Waves 10(2):137–149. 

Draper EJ, Fairlie GE [2006]. AutoReaGas modeling of gas explosions in mines. 

Unpublished report prepared by Century Dynamics Ltd. for NIOSH. 

Fickett W, Davis WC [1979]. Detonation. Berkeley, CA: University of California Press. 

Fried L, Glaesemann K, Souers PC, Howard WM, Vitello P [2000]. A thermochemicalkinetics code. Cheetah 3.0. Livermore, CA: Lawrence Livermore National Laboratory. 

72
 
Gallagher R [2005]. Research needs in regard to design, performance criteria, 

construction, maintenance assessment and repair of coal mine seals. In: Proceedings of the 31st 

International Conference of Safety in Mines Research Institutes (Brisbane, Queensland, 

Australia, October 2–5, 2005). Brisbane, Queensland, Australia: Safety in Mines Testing and 

Research Station (Simtars), pp. 236–242. 

Gamezo VN [2007]. Personal communication. 

Gates RA, Phillips RL, Urosek JE, Stephan CR, Stoltz RT, Swentosky DJ, et al. [2007]. 

Report of investigation, fatal underground coal mine explosion, January 2, 2006. Sago mine, 

Wolf Run Mining Company, Tallmansville, Upshur County, West Virginia, ID No. 46–08791. 

Arlington, VA: U.S. Department of Labor, Mine Safety and Health Administration. 

Genthe M [1968]. Untersuchungen und Versuche zur Frage der Explosionssicherheit von 

Vordämmen bei der Grubenbrandbekämpfung (Research on explosion-proof bulkheads for mine 

fire control) (in German) [Dissertation]. Essen, Germany: Verlag Glückauf GmbH. 

Gexcon [2007a]. FLACS (Flame acceleration simulator). [http://www.gexcon.com/ 

index.php?src=flacs/flacs.html]. Date accessed: June 2007. 

Gexcon [2007b]. Gas explosion handbook. [http://www.gexcon.com/index.php?src=gas/ 

handbook.html]. Date accessed: June 2007. 

Glasstone S, Dolan PJ, eds. [1977]. The effects of nuclear weapons. 3rd ed. U.S. 

Department of Defense and the Energy Research and Development Administration. 

Hansen OR [2006]. FLACS modeling for NIOSH: simulations of existing and planned 

tests at Lake Lynn experimental mines. Unpublished report prepared by Christian Michelsen 

Research AS for NIOSH. 

Hertzberg M, Cashdollar KL [1987]. Introduction to dust explosions. In: Cashdollar KL, 

Hertzberg M, eds. Industrial dust explosions. West Conshohocken, PA: American Society for 

Testing and Materials (ASTM), Special Technical Publication (STP) 958, pp. 5–32. 

Hornsby CD, Hallam P, Allan JA, Walker G, Boyle JC, Criddle SJ, et al. [1985]. Sealing 

off fires underground. Memorandum from the Committee of the Institution of Mining Engineers, 

U.K. 

Kattenbraker SR [2002]. Explosion at Big Ridge mine portal #2, February 1, 2002. 

Memorandum of July 2, 2002, for the record from Steven R. Kattenbraker, Mine Safety and 

Health Administration. 

Kinney GF [1962]. Explosive shocks in air. New York: Macmillan. 

Kolbe M, Baker QA [2005]. Gaseous explosions in pipes. In: Proceedings of the ASME 

Pressure Vessels and Piping Division Conference (Denver, CO, July 17–21, 2005). American 

Society of Mechanical Engineers. 

Kuznetsov M, Ciccarelli G, Dorofeev, S Alekseev V, Yankin Y, Kim TH [2002]. DDT in 

methane-air mixtures. Shock Waves 12:215–220. 

Landau LD, Lifshitz EM [1959]. Fluid mechanics. Course of theoretical physics, 

volume 6. Oxford, U.K.: Butterwork-Heinemann. 

Landau LD, Lifshitz EM [1987]. Fluid mechanics. Course of theoretical physics, 

volume 6. 2nd ed. Oxford, U.K.: Butterwork-Heinemann. 

Lea CJ, Ledin HS [2002]. A review of the state of the art in gas explosion modeling. 

Buxton, U.K.: Health & Safety Laboratory, 

Lebecki K, Kajdasz Z, Napieracz T, Cybulski K [1999]. Tests on bulkheads resistance to 

methane explosion. In: Proceedings of the 28th International Conference of Safety in Mines 

Research Institutes (Sinaia, Romania, June 7–11, 1999). Vol. III. Petroşani, Romania: National 

Institute for Mining Safety and Explosion Proof Protection, pp. 5–19. 

73
 
Lee JHS [1984]. Dynamic parameters of gaseous detonations. Annu Rev Fluid Mech 

16:311–336. 

Lewis B, von Elbe G [1987]. Combustion, flames and explosions of gases. 3rd ed. 

Orlando, FL: Academic Press, Inc. 

Light TE, Herndon RC, Guley AR Jr., Cook GL, Odum MA, Bates RM Jr., et al. [2007]. 

Report of investigation, fatal underground coal mine explosion, May 20, 2006. Darby mine No. 

1, Kentucky Darby LLC, Holmes Mill, Harlan County, Kentucky, ID No. 15–18185. Arlington, 

VA: U.S. Department of Labor, Mine Safety and Health Administration. 

Lyne B [1996]. Approved standard for ventilation control devices, including seals and 

surface airlocks. Brisbane, Queensland, Australia: Queensland Department of Mining and 

Energy, Safety and Health Division, Coal Operations Branch, QMD 96 7396. 

Lyne B [1998]. Approved standard for monitoring of sealed areas. Brisbane, Queensland, 

Australia: Queensland Department of Mining and Energy, safety and Health Division, Coal 

Operations Branch, QMD 98 7433. 

Mason TN, Tideswell FV [1933]. Gob fires. Part 1: Explosions in sealed-off areas in nongassy seams. Safety in Mines Research Board Paper No. 75. London: His Majesty’s Stationery 

Office. 

McBride BJ, Gordon S [1996]. Computer program for calculation of complex chemical 

equilibrium compositions and applications. II. Users manual and program description. Cleveland, 

OH: National Aeronautics and Space Administration, Lewis Research Center, NASA Reference 

Publication 1311. 

McCabe WL, Smith JC [1967]. Unit operations of chemical engineering. 3rd ed. McGraw 

Hill. 

McKinney R [2006]. Program information bulletin No. P06–16 issued July 19, 2006, 

from R. McKinney, Administrator for Coal Mine Safety and Health. 

Michelis J, Kleine W [1989]. Development of components designed to resist explosion 

pressures of approximately 1 MPa for use in ventilation structures in underground mines. In: 

Proceedings of the 23rd International Conference of Safety in Mines Research Institutes 

(Washington, DC, September 11–15, 1989). Pittsburgh, PA: U.S. Department of the Interior, 

Bureau of Mines, pp. 859–857. 

Mitchell DW [1971]. Explosion-proof bulkheads: present practices. Pittsburgh, PA: U.S. 

Department of the Interior, Bureau of Mines, RI 7581. NTIS No. PB 205 507. 

Nagy J [1981]. The explosion hazard in mining. Pittsburgh, PA, U.S. Department of 

Labor, Mine Safety and Health Administration, IR 1119. 

Oberholzer JW, Lyne BJ [2002]. A review of the requirements for the testing of the 

strength of ventilation structures to be used in Queensland mines. In: Proceedings of the 

Queensland Mining Industry Health and Safety Conference, pp. 105–112. 

Oran ES, Gamezo VN [2007]. Origins of the deflagration-to-detonation transition in gasphase combustion. Combust Flame 148(1–2):4–47. 

Peraldi, O, Knystautas R, Lee JH [1986]. Criteria for transition to detonation in tubes. In: 

Proceedings of the 21st International Symposium on Combustion (University of Munich, 

Germany). Pittsburgh, PA: Combustion Institute, pp. 1629–1637. 

Popat NR, Catlin CA, Arntzen BJ, Lindstedt RP, Hjertager BH, Solberg T, et al. [1996]. 

Investigations to improve and assess the accuracy of computational fluid dynamic-based 

explosion models. J Hazard Mater 45(1):1–25. 

74
 
Razus D, Movileanu C, Brinzea V, Oancea D [2006]. Explosion pressures of 

hydrocarbon-air mixtures in closed vessels. J Hazard Mater 135(1–3):58–65. 

Rice GS, Greenwald HP, Howarth HC, Avins S [1931]. Concrete stoppings in coal mines 

for resisting explosions: detailed tests of typical stoppings and strength of coal as a buttress. 

Pittsburgh, PA: U.S. Department of Commerce, Bureau of Mines, Bulletin 345. 

Ross E Jr., Schultz MJ [1996]. Report of investigation (underground coal mine). 

Noninjury coal mine explosion. Mine No. 1 (I.D. No. 46–07273), Oasis Contracting, Inc., 

Quinland, Boone County, West Virginia, May 15, 1996, and June 22, 1996. U.S. Department of 

Labor, Mine Safety and Health Administration. 

Roxborough FF [1997]. Anatomy of a disaster: the explosion at Moura No. 2 coal mine, 

Australia. Min Technol 79(906):37–43. 

Rutherford JW, Painter RJ, Urosek JE, Stephan CR, Dupree WA Jr. [1993]. Report of 

investigation, underground coal mine explosion, Blacksville No. 1 mine, I.D. No. 46–01867. 

Consolidation Coal Company, Blacksville, Monongalia County, West Virginia, March 19, 1992. 

Arlington, VA: U.S. Department of Labor, Mine Safety and Health Administration. 

Sapko MJ, Furno AL, Kuchta JM [1976]. Flame and pressure development of large-scale 

CH4-air-N2 explosions: buoyancy effects and venting requirements. Pittsburgh, PA: U.S. 

Department of the Interior, Bureau of Mines, RI 8176. NTIS No. PB 261 331. 

Sapko MJ, Weiss ES, Harteis SP [2005]. Methods for evaluating explosion-resistant 

ventilation structures. In: Gillies ADS, ed. Proceedings of the Eighth International Mine 

Ventilation Congress (Brisbane, Queensland, Australia, July 6–8, 2005). Carlton, Victoria, 

Australia: Australasian Institute of Mining and Metallurgy, pp. 211–219. 

Schultze-Rhonhof H [1952]. Major experimental firedamp explosions at an abandoned 

mine. In: Proceedings of the Seventh International Conference of Directors of Safety in Mines 

Research (Buxton, U.K., July 7–12, 1952). Vol. 3, paper No. 25. 

Scott DS, Stephan CR [1997]. Accident investigation report (underground coal mine). 

Noninjury methane explosion. Oak Grove mine (I.D. No. 01–00851), U.S. Steel Mining 

Company L.L.C., Adger, Jefferson County, Alabama, July 9, 1997. Birmingham, AL: U.S. 

Department of Labor, Mine Safety and Health Administration. 

Shepherd JE [2006]. Structural response of piping to internal gas detonation. In: 

Proceedings of the ASME Pressure Vessels and Piping Division Conference (Vancouver, British 

Columbia, Canada, July 23–27, 2006). American Society of Mechanical Engineers. 

Shepherd JE, Teodorczyk A, Knystautas R, Lee JHS [1991]. Shock waves produced by 

reflected detonations. Progr Astronaut Aeronaut 134:244–264. 

Shoemaker DP, Garland CW [1967]. Experiments in physical chemistry. 2nd ed. New 

York: McGraw Hill. 

Slawson TR [1995]. Wall response to airblast loads: the wall analysis code (WAC). 

Prepared for the U.S. Army Energy Research and Development Center, Vicksburg, MS. 

Smirnov NN, Nikitin VF, Dushkin VR, Kulchitskiy AV [2003]. Premixed gaseous flame 

acceleration due to instability induced by geometrical characteristics of combustion chambers. 

In: Proceedings of the 54th International Astronautical Congress of the International 

Astronautical Federation. Reston, VA: American Institute of Aeronautics and Astronautics, Inc. 

South JS [1987]. Report of investigation (underground coal mine). Non-injury methane 

explosion. Roadfork No. 1 mine (I.D. 15–10753), Sidney Coal Company, Incorporated, Sidney, 

Pike County, Kentucky; discovered: October 7, 1986. Pikeville, KY: U.S. Department of Labor, 

Mine Safety and Health Administration. 

75
 
Steen H, Schampel K [1983]. Experimental investigations on the run-up distance of 

gaseous detonations in large pipes. In: Proceedings of the Fourth International Symposium on 

Loss Prevention and Safety Promotion in the Process Industries. London: Pergamon Press, 

pp. E23–E33. 

Stephan CR [1990]. Construction of seals in underground coal mines. Report No. 

06–213–90. Pittsburgh, PA: U.S. Department of Labor, Mine Safety and Health Administration, 

Pittsburgh Safety and Health Technology Center, Industrial Safety Division. 

Sumpter JE, Hudak MJ, Kalich MG, Cook LE, Dupree WA Jr., Schultz MJ [1996]. 

Report of investigation (underground coal mine). Noninjury coal mine explosion. Gary No. 50 

mine (ID No. 46–01816), U.S. Steel Mining Company, Inc., Pineville, Wyoming County, 

West Virginia, between June 9–16, 1995. Mount Hope, WV: U.S. Department of Labor, Mine 

Safety and Health Administration. 

U.K. Health and Safety Executive [1998]. Explosion-proof stoppings. Int Min Miner 

1(10):256. 

van Wingerden K, Bjerketvedt D, Bakke JR [1999]. Detonations in pipes and in the open. 

In: Proceedings of the Petro-Chemical Congress (June 23–24, 1999). 

Vasilev A [2006]. Estimation of critical conditions for the deflagration-to-detonation 

transition. Combust Explos Shock Waves 42(2):205–209. 

Weir A, Morrison RB [1954]. Equilibrium temperatures and compositions behind a 

detonation wave. Ind Eng Chem 46(5):1056–1060. 

Wiemann W [1987]. Influence of temperature and pressure on the explosion 

characteristics of dust/air and dust/air/inert gas mixtures. In: Cashdollar KL, Hertzberg M, eds. 

Industrial dust explosions. West Conshohocken, PA: American Society for Testing and Materials 

(ASTM), Special Technical Publication (STP) 958. 

Willett HL, Blunt J, Coulshed AJG, Tideswell FV [1962]. Sealing off fires underground. 

Memorandum from the Committee of the Institution of Mining Engineers, U.K. (July 4, 1962), 

pp. 709–760. 

Zeldovich YB, Kompaneets AS [1960]. Theory of detonation. New York: Academic 

Press. 

Zeldovich YB, Barenblatt GI, Librovich VB, Makhviladze GM [1985]. The mathematical 

theory of combustion and explosions. New York: Plenum Publishing. 

Zlochower IA [2007a]. Personal communication on diffusion calculations. 

Zlochower IA [2007b]. Personal communication on NASA-Lewis thermodynamic 

equilibrium calculations. 

Zucrow MJ, Hoffman JD [1976]. Gas dynamics. Vol. 1. New York: John Wiley and 

Sons, Inc. 

Zucrow MJ, Hoffman JD [1985]. Gas dynamics. Vol. 2: Multidimensional flow. Malabar, 

FL: Krieger Publishing Co. 

76
 
Delivering on the Nation’s promise: 

Safety and health at work for all people 

through research and prevention 

To receive documents or more information about occupational safety and health topics, 

contact NIOSH at 

1 –800 –35 – NIOSH (1 –800 –356 – 4674)

Fax: 513 –533 – 8573

E-mail: pubstaft@cdc.gov 

or visit the NIOSH Web site at www.cdc.gov/niosh 

DHHS (NIOSH) Publication No. 2007–144 

SAFER ● HEALTHIER ● PEOPLE tm