current science of methane fuelled explosions and best ...

COALTECH RESEARCH ASSOCIATION NPC

CURRENT SCIENCE OF METHANE FUELLED

EXPLOSIONS AND BEST INTERNATIONAL

PRACTICE FOR PROTECTION STRUCTURES

By

R.P.van Wyk (Pr.Eng.)

May 2015 1

1 Copyright COALTECH

This document is for the use of COALTECH only, and may not be transmitted to any other party, in

whole or in part, in any form without the written permission of COALTECH.

PROJECT 2.6

Information Circular CT2014/02

CURRENT SCIENCE OF METHANE FUELED EXPLOSIONS AND BEST

INTERNATIONAL PRACTICE FOR PROTECTION STRUCTURES

By: Roedolf P.van Wyk (Pr.Eng.)

DISCLAIMERS

The findings and conclusions in this report are those of the author and do not necessarily represent

the views of Coaltech. Reference to work by any other institution is in accordance to each individual

document’s disclaimers. Mention of any company or product does not imply endorsement by

Coaltech.

DISCLAIMER

The report does not provide any recommendations on the design and

construction of seals in the South African mining community. It is beyond the

scope of this report to provide guidelines regarding the design of generic

seals. This report should only be used as a summary of information from other

publications in the field or legislation from foreign countries.

Executive Summary

Two schools of thought exist within the international ventilation community regarding protection

against explosions from within sealed of abandoned panels, namely:

Event prevention ‐ Measure, Inertisize and Control the atmosphere in the panel. (Mainly

Australia) This approached was derived after the 1994 Moura Number 2 disaster.

Event Control ‐ Construct seals capable of withstanding overpressures generated during

Methane/Coal Dust fuelled explosion. (Typically USA) US Final Rule 2008 was developed

after the 2006 Sago disaster.

The US Final Rule 2008 legislation challenged current international seal practice and expanded the

knowledge regarding the chemistry and physics of Methane/Coal Dust fuelled confined explosions.

The legislation introduced dramatic changes to US seal pressure ratings from the historic 20psi

(140kPa) value to present 50psi, 120psi and above 120psi (345kPa, 827kPa ,>827kPa) ratings.

Furthermore this new legislation specifies standards towards seal design, approval, construction and

material or atmosphere testing methods, etc.

The Australian mining sector intensively scrutinized the above US legislation. The industry

communally decided that their system of preventing an incident through risk analysis and “world

best practice” standards by which they control the sealed atmosphere better suits their industry

than the US system of avoiding failure by constructing stronger seals. The New South Wales

government has however adopted the learning from both the Sago Mine explosion report and Final

Rule 2008 in a safety bulletin published in August 2013. The bulletin addresses protection against

lightning strikes, new seal pressure ratings and procedures when the sealed atmosphere passes

through the explosive zone.

Current South African legislation regarding sealing of abandoned panels does not directly conform to

either of these international legislations. The South African seal pressure rates closely resemble the

Australian values. The Australian legislation is primarily based on two header road longwall mining

extraction methods seldom found in South Africa. The US legislation was developed in reaction to

explosions in board‐and‐pillar type mines. This mining method is typically used South African Coal

mining.

The aim of the report is to inform the reader of the current international sealing practice by

discussing recent published scientific reports regarding Methane/Coal Dust fuelled confined

explosions, the origin and implementation of U.S. and Australian sealing legislation as well as new

developments in the industry.

Table of Contents

Executive Summary i

Table of units and abreviations 8

1. Introduction 9

2. History of mine seals

United States of America 12

Australia 14

U.K. & European Practices 17

3. Theory of Methane‐Air explosions

Diffusion of Methane in Air 18

Thermodynamic analysis 19

Tunnel explosion mechanics 19

Experimental proof 24

Summary 25

4. 140kPa & 400kPa Experimental Results

Plug Seal 27

Meshblock Seal 28

Gunmesh Stopping 29

Testing Methodology 31

Structural Analysis 31

Ingwe Spec Walls 33

Conclusions 33

5. Panel Inertization 34

6. U.S.Legislation

Background 38

Legislation 39

7. US Army Corps of Engineers

Introduction 42

Protective Structure Design and Analysis Methods 43

Protective structure design 43

Analysis of Pre‐2006 NIOSH and USBM Seal Tests 44

Analysis of seal foundations 44

Analysis of seal structures 44

Guidelines for Design of Coal Mine Seals 45

Behaviour of 120‐psi Seals Subject to Methane‐Air 45

Detonation Pressure

Blast wave attenuators 46

Summary 48

8. Current Australian Sealing practice

ACARP Project C7015 50

Industry questionnaire survey 50

Relevant differences between US & Australian mines 51

Views on changed US approach 51

SIMTARS Propagation Tube Test Work 51

ACARP Report Summary 53

Comments on ACARP report 53

NSW Safety Bulletin – SB13‐04 55

9. Summary 57

References 59

Annexures

A ‐ EXAMPLES OF TYPICAL APPROVED 120psi CONTAINMENT STRUCTURE DESIGNS AND RELATED

COST COMPARISONS

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Table of Units and abbreviations

Abbreviations ACARP ‐ Australian Coal Industry Research Program CV ‐ Constant Volume DDT ‐ Deflagration to Detonation Transition DIF ‐ Dynamic Increase Factor DLF ‐ Dynamic Load Factor DME ‐ Department of minerals and Energy Affairs (South Africa) LLEM ‐ Lake Lynn Experimental Mine, Pennsylvania MSHA ‐ Mine Safety and Health Administration NIOSH ‐ National Institute of Occupational Safety and Health SIMTARS ‐ Safety In Mines Testing And Research Station USACE ‐ United States Army Core of Engineers

Units

kPa ‐ Kilo Pascals (1,000 Pa)

MPa ‐ Mega Pascals (1,000,000 Pa)

psi ‐ Pounds per Square Inche

m ‐ Meter

mm ‐ Millimeter

m/s ‐ Meter per Second

Practical Information Boxes ‐ Information provided within these information boxes are remarks by

the author aimed at simplifying the scientific text with information

specifically relating to the South African mining community.

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INTRODUCTION

Current South African mining regulations require appointed managers to control Methane risks in

mined‐out areas of coal mines and thereby reduce explosion risk from Methane build‐up by either

ventilating or sealing those areas. Continued ventilation of abandoned areas is costly and may divert

ventilating air away from other, more productive uses. Seals are barriers constructed in underground

coal mines to isolate abandoned mining panels from active workings. At present South African mining

regulations require the following:

Any abandoned underground panel must be sealed in accordance with the DME Guideline for the Compilation of a mandatory Code of Practice for the prevention Coal Dust Explosions in Underground Mines. The requirement calls for: The monitoring of the atmosphere behind the containment walls provides the manager with the input to the risk assessment of the abandoned area. Two conditions are likely to exist. 1. The atmosphere in the area has stabilized above or below the explosive range of flammable

gas. As an explosion cannot occur under these conditions, the containment walls now only require regular monitoring in order to verify safe conditions. ….Containment walls must be designed to withstand a static pressure of approximately 140kPa….

2. The atmosphere in the sealed area remains within the explosive range of flammable gas. Further action is then required, e.g.:‐

o Install approved explosion proof seal. ….”Explosion proof seals” means a seal which is designed to withstand a static pressure of typically 0,4MPa and requires an approved design endorsed by a Professional Civil Engineer.

The international coal mining industry is divided into two schools of thought on the management of sealed abandoned panels:

Event Control ‐ Construction of structurally sound seals to guard against failure (Typically practiced in the USA), vs.

Event Prevention ‐ Measure, Monitor and Inert the atmosphere to prevent an incident from occurring. (Typically practiced in Australia)

Both these schools of thought are being successfully implemented in different regions of the world to assure the safety of miners. All of this work is based on tragic lessons learned from previous explosive events. The truth of the matter is that after a tragedy the investigation committee on the incident, the industry as a whole and the lawmakers work together to prevent a reoccurrence of the event. The US department of labour in co‐operation with MSHA in 2006 appointed research teams from both NIOSH and US Army Corps of Engineers to investigate the causes of the Sago Mine tragedy as well as research international best sealing practice. The preliminary findings of both teams were published in separate reports in 2007. The NIOSH report IC9500 provided the then latest knowledge on Methane/Coal‐Dust fuelled explosions specifically towards the chemistry, thermodynamics and overpressures expected during such an event. Furthermore the document made recommendations on pressure ratings to be used for future US seal designs. This information is the cornerstone of the 2008 Final Rules for US seal design. Australian coal mining went through a similar review of coal mine safety in the mid 1990s after Moura Number 2 explosion. The Australian industry adopted a different approach to managing sealed abandoned panels compared to the US, due to differences in mine safety management and mining

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methods. Much of the recent research conducted by the Americans on the science behind Methane/Coal‐Dust fuelled explosions has challenged and expanded international understanding on the subject. As a result there has been a move for some states in Australia to consider and possibly adopt new US standards for seal pressure rating codes. However industry as a whole, including mines’ management, state inspectorates and mining unions have decided not to adopt the principles dictated of the 2008 US seal regulations. Australian approaches are formulated on a risk assessment basis under which hazards must be identified and appropriate “world’s best practice” systems adopted. The principal approach in Australia to explosive events is early prevention of hazardous situations through use of real time gas monitoring from the goaf periphery to ensure the maintenance of goaf inert atmospheric conditions. Another line of defence is having inert gas systems on hand (most commonly jet or diesel engine exhaust, nitrogen or CO2) to proactively ensure potentially explosive gas concentrations cannot form or are handled appropriately. The final approach is through use of well‐engineered seal structures constructed to segregate all worked out areas where there is any likelihood of explosive gas concentrations occurring. Seals on gassy goafs most commonly are designed to meet a 140kPa rating. Recent regulations by the New South Wales Government do recommend that for the time period during which sealed areas pass through the explosion range the mine be abandoned if only 140kPa(20psi) seals are used, but that the mine may remain open if 827kPa(120psi) are used. This pressure rating follows directly from the US final rule 2008 pressure ratings.

Coaltech aims to remain relevant and up‐to‐date with international safety and best practice standards for all activities in the coal mine industry. The recent US Final Rule 2008 legislation changes regarding the design of explosion proof bulkheads as well as research conducted in the U.S. on the science behind Methane and Coal Dust explosions came to the attention of the Coaltech board. This information prompted the board to commission a research into the current international best practice regarding containment seals. From this desktop research study based on report and other publications on the internet plus correspondence with researchers in the U.S. via email it was decided to send a fact finding team to the U.S. to investigate the relevance of the U.S. work to South African coal mining. During May 2014 a delegation consisting of a Mining Engineer, Ventilation Officer, Rock Engineer and Structural Engineer was sent to the U.S. on a ten day fact finding mission. The itinerary for the mission was mostly planned by Dr.Zipf a renowned and respected researcher and engineer involved in the development of the new US seal standards. The itinerary included official meetings with:

MSHA (Mine Safety and Health Administration) head office

MSHA Tech Support Laboratories

NIOSH (National Institute of Occupational Safety and Health) Laboratories

USACE (United States Army Core of Engineers) Laboratories

Cardno, MM&A and ECSI Consulting engineers

Minova and Strata Worldwide Construction companies

Signal Peak Energy Company, Billings mine and BHP Billiton, Farmington mine. This itinerary covered a comprehensive cross section of the US Coal mining industry involved with the 2008 Final Rule Legislation representing the legislators, researchers, design engineers, construction companies and end‐users. The Coaltech research team learned from the meetings and additional information provided by the U.S. delegates how and why the legislation changed, how the science has affected the legislation, new developments in this field as well as practical considerations regarding construction of these seals. This information gained will be discussed in further detail in this report. The report aims to provide information and is not aimed at delivering an opinion to the implementation or changes to the current South African seal standards. This report should enable readers to understand the current state of the

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art of coal mine seal design plus legislation in both the U.S. and Australia. The reader should determine to his/her own discretion the applicability of these standards to South African mining conditions. It remains the prerogative of the reader to decide the relevance of both the US and Australian schools of thought towards the sealing practice in South Africa. This document does not recommend one standard above the other, but provides a mosaic of portions from several documents published by the above institutions. The author attempted to present the information in a logical manner to a cross section of readers from related disciplines regarding the subject as well as providing a reference for readers with an in depth background in the field. This document does not provide a complete knowledge of the subject nor does it provide new information to the field, what the document aims to do is provide an overview of the subject with a logical and chronological presentation of relevant information.

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1. HISTORY OF MINE SEALS

United States of America

Zipf et.al states in IC9500(2007) that the earliest known engineering standards 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 psi) and that it was “based on the general opinion of men experienced in mine‐explosion investigations.” Evidently, the intent of the regulation was to prevent an explosion in one mine from propagating to a neighbouring mine. In Germany and Poland, authorities decided that seals should be designed to withstand 500‐kPa based on observations from moderate‐strength experimental coal mine explosions.

The U.S.Federal Coal Mine Health and Safety Act of 1969 required mined‐out areas to be ventilated or sealed with ‘explosion‐proof bulkheads’ that were constructed with ‘solid, substantial and incombustible materials.’ In the publication Explosion‐Proof Bulkheads: Present Practices, Mitchell (1971) reviewed coal mine explosions and recommended what became 20‐psi (140kPa) criterion for coal mine seals. The general premise behind Michell’s recommendation was that coal mine explosions originate in the active areas of coal mines that is ventilated and has only limited quantities of Methane or coal dust. An explosion within the sealed area was not considered, because it was commonly believed that sealed areas were inert and either contained methane‐rich or oxygen‐poor atmospheres. Mitchell noted that more than 60m from the origin of an explosion of a small amount of explosive mix in 15m of entry, the explosive pressure seldom exceeded 140‐kPa. Most sealed areas are far from the active mining area, so Mitchell concluded that a seal may be considered “explosion‐proof” if it is designed to withstand a static load of 140‐kPa.

Prior to 1992, the Code of Federal Regulations (CFR) lacked a definitive design specification for explosion‐proof mine seals. Stephan (1990) reviewed Mitchell’s work and also concluded that the explosive pressure on seals generally does not exceed 20‐psi. As a result of the Stephan report, the explosion pressure performance criterion for seals became 20‐psi in the 1992 change to Code of Federal Regulations Rule 30 CFR 75∙335(a)(2). This rule change is generally referred to as the “Alternative Seals Rule”, since it facilitated the development of alternatives to the conventional Mitchell‐Barrete seal of solid concrete blocks used up to this time. Examples of alternatives include cement foam plug seals, polyurethane foam and aggregate plug seals, Omega Block seals and wood crib block seals (Zipf et al. 2009). To determine whether an alternative seal design met the 20‐psi requirement, the Mine Safety and Health Administration (MSHA) relied on full‐scale explosion tests conducted by the National Institute for Occupational Safety and Health Administration (NIOSH) in their Lake Lynn Experimental Mine (LLEM) in Pennsylvania.

Working under the direction of MSHA and the seal manufacturers, NIOSH researchers constructed actual full‐scale alternative seals at the LLEM and subjected them to a side‐on (i.e., quasi‐static) pressure of 20‐psi that was generated by a test methane explosion. The candidate seal passed the test if it survived the explosion pressure without any visible damage such as cracking or displacement. Air leakage across the seal was then measured to determine if it also met leakage requirement.

Between 1986 and 2006, 12 known explosions occurred within sealed areas of active U.S. underground coal mines. Table 1 summarizes the known characteristics of these explosions as determined by the relevant MSHA accident investigation reports

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Mine Year Damage from explosion

Cause of explosive mix

Suspected ignition source

Estimated explosive pressure

Roadfork No.1 1986 4 Seals Destroyed Recently sealed area

Spark from rock fall

Unknown

Mary Lee no.1 1993 2 Seals Destroyed & Shaft Cap displaced

Leaking seals Lightning 14‐kPa

Oak Grove no.1

1994 3 Seals Destroyed Leaking seals Unknown Unknown

Gary 50 1995 None – Seals Survived

Leaking seals Lightning or roof fall

35‐48 kPa

Oasis 1996(May) 3 Seals Destroyed Leaking seals Lightning or roof fall

<138 kPa

Oasis 1996(June) Unknown Leaking seals Lightning or roof fall

Unknown

Oak Grove no.1

1997 3 Seals Destroyed Leaking seals Lightning >138 kPa

Big Bridge Mine

2002 1 Seals Destroyed Recently sealed area

Unknown Unknown

Sago 2006 10 Seals Destroyed

Recently sealed area

Lightning >642 kPa

Darby 2006 3 Seals Destroyed Recently sealed area

Oxygen/Acetylene torch

>152 kPa

Table 1 – Summary of known explosions in sealed areas of U.S.coal mines 1986‐2006

After the two explosions in 2006, MSHA issued on 19 July 2006, a Program Information Bulletin (PIB) No. P06‐16, titled “Use of Alternative Seal Methods and Materials Pursuant to 30CRF 75∙335(a)(2)” that required new alternative seals be designed and built to reliably withstand an overpressure of at least 50 pounds per square inch (psi)” (McKinney 2006). Although PIB No.P06‐16 raised the seal design standard from 20 to 50 psi, no guidance was given on how new seals should be designed and constructed to meet the new requirement. On 22 May, 2007, MSHA issued an “Emergency Temporary Standard (ETS) on Sealing of Abandoned Areas,” which specified new strength requirement of (1) 50‐psi (345‐kPa) overpressure for sealed areas that are monitored and maintained inert; 120‐psi (800‐kPa) overpressure if the sealed area atmosphere is not monitored and maintained inert; and greater than 120‐psi (800‐kPa) overpressure if certain conditions exist within the sealed area that may promote the development of higher explosive pressures (ETS2007). The ETS became the “Final rule on Sealing of Abandoned Areas,” which was issued on April 2008 and became fully in force by 20 October, 2008 (Final Rule 2008) The major structural engineering requirement for seals in the Final Rule did not change significantly from those in the ETS; however, many requirements were modified or clarified. As stated in the Final Rule summary, “the final rule includes requirements for seal strength, design, construction, maintenance and repair of seals, and monitoring and control of atmospheres behind seals in order to reduce the risk of seal failure and the risk of explosions in abandoned areas of underground coal mines”. It also contains provisions for training of mine personnel who conduct work on seals and sealed areas and recordkeeping requirements for archiving data important to seals and sealed areas. Thus the ETS and Final Rule differs from their historic counterparts in that they firstly consider the risk of an explosion from within the sealed of panel contrary to previous laws that was based on an explosion originating in the active side and secondly addresses issues such as design, construction, maintenance, training, etc. that was previously unspecified. MSHA goes further by implementing in

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the Final Rule controls over end‐users and installers. The mine management, designers, and contractor have to prove their compliance to the legislator. It is relevant to know that the ETS and Final Rule were introduced within a period of less than 22 months. Thus all the research and preparation required to develop the legislation was condensed into a few months. Furthermore the time and extent of public hearing remarks to the proposed legislation were abbreviated by the rushed nature of this legislation. Due to the nature of the U.S. legal system any changes to the existing legislation has to follow the same stringent route of public hearings and scrutiny as the original approval procedure. As a result there is a resistance to implementing any subsequent knowledge and industry developments in regards to seal science and practice. Thus even when new science proves a part of the legislation wrong or when a revolutionary engineering design is developed none of these are considered by the Final Rule. The inflexibility of the law system means that the Final Rule is applied in a blanket fashion to all mines irrespective to the mine’s specific gas conditions and other risk factors that are not addressed in the Final Rule. The blanket application of the law has as a direct result the construction of large seals in areas where no risk for explosions exists, thus unnecessary cost to the operator. It furthermore has resulted in a lack of motivation to develop attenuation measures aimed at reducing the pressure peaks that could result in cheaper equally effective seals. The current U.S. legislation for the design of seals in coal mines has evolved over nearly a hundred years, from its primitive empirical estimated values, to sound scientific foundations and practical implementations based on the lessons learned from historic disasters. The rigid nature of the U.S. law system will resist any major changes to the current legislation bar catastrophe or fundamental scientific fact. Australia In Australia, within Queensland according to Standards for Seals and Airlocks 1967 issued by Coal Operations Branch, Safety and Health Division, Queensland Department of Mines and Energy (QDME), four specific elements must be addressed when installing seals. These are;

design and specification,

location,

construction, and

maintenance and monitoring. Stoppings, as defined by Hartman et al (1997), are physical barriers erected between intakes, returns or abandoned mine voids to prevent air from mixing. Stoppings are classified according to construction, length of service, and purpose as temporary or permanent. Temporary stoppings are extensively used in areas where frequent adjustment to air directions are necessary. They are moderately airtight and are normally hung in active workings where changes occur rapidly in the mining and ventilation methods. They must be readily movable and are generally reusable. Permanent stoppings, also called bulkheads, are installed in places where a permanent or a long‐term control of flow is needed, such as between the main intakes and returns or belt entries. In the past these have been constructed of frame, sheet metal (prefabricated sections), masonry (stone, brick, or concrete block) or “shotcrete” sprayed on wire mesh. Because their purpose is to stop airflow for an indefinite period, they must be made airtight by tapping, plastering or caulking and resistant to cracking from blasting concussion or ground movement. Permanent stoppings are also used as fire bulkheads to seal off abandoned workings. Abandoned workings may in time hold toxic or explosive gas mixtures and so these bulkheads must both stop atmospheric mixing and be able to withstand a pressure event. A seal is a special stopping used to isolate abandoned workings and goafs or as fire bulkheads. Seals

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eliminate the need to ventilate those areas; they may also be used to isolate fire zones or areas susceptible to spontaneous combustion. According to the above survey questionnaire, prior to 1997 various materials were used to segregate different areas of the mine ventilation from one another. A summary of the materials and specific applications are shown in Table 2. It is clear from the table that no clear pressure rating was available to the industry at the time. Table 2 – Questionnaire results for materials used as seals Belt Road

Segregation Main intake / Belt Segregation from

Return

Segregate intake from Belt Air in panel Gateroads

Final Panel seal segregation from adjacent panels

Final panel seal segregation from

Mains

Brattice 2

Plasterboard 2 2 7 3 2

Sheet Metal 1 1

Mortar Blocks 1 6 4

Blocks 2 4 2

Low Density Blocks 2

Reinforced Cementitios

4 2 1 1

Composite Polymer 1

Nothing Remainder

On August 7, 1994, 11 miners and 1 contractor were killed when a methane‐air mixture ignited within a recently sealed room‐and‐pillar panel at the BHP Australia Coal Moura No. 2 coal mine in Queensland, Australia [Roxborough 1997]. The most likely ignition source was determined to be the heating caused by spontaneous combustion within the sealed area. The overpressures generated from the methane ignition resulted in the failure of several seals that were newly installed about 22hr before the ignition. As a result of this disaster, a considerable public outcry demanded that an in‐depth inquiry be conducted to determine the cause of the explosion and to recommend ways to prevent future occurrences in the Queensland coal mines. In late 1997 and early 1998, NIOSH Pittsburgh Research Laboratory (PRL) collaborated on a joint research project with Barclay Mowlem Construction Ltd. Of Queensland, Australia, to investigate the capability of various seal and stopping designs and an overcast design to meet or exceed the requirements of the Queensland Department of Mines and Energy's [1996] Approved Standard for Ventilation Control Devices. This standard was the result of deliberations and investigations by Task Group 5, which was formed by the recommendation of the Warden's Inquiry concerning the Moura No. 2 mine explosion [Roxborough 1997]. Task Group 5 was charged with the reassessment of the regulatory provisions for explosion‐resistant seals and the investigation of mine inerting techniques. Prior to the enactment of new regulations on 16 March, 2001 in Queensland, introduction of the Queensland Mines Department Approved Standard for Ventilation Control Devices provided prescriptive ratings for seals and stoppings and required live testing of seals and stoppings in an “internationally recognized mine testing explosion gallery”. Control devices in Australia are as follows: 14kPa, 35kPa, 140kPa, and 345 kPa (2psi, 5psi, 20psi, and 50 psi). The expected outcome of these standards for seals and airlocks in Queensland is that all ventilation control structures will have an overpressure rating based on an assessment of the risk and purpose of the particular control structure. These standards do not address the structural design or the material to be used in seal construction. As part of the enormous amount of research undertaken at the time after the recommendations of Task Group 5 (Oberholzer and Lyne, 2002) in establishing practical design criteria to assist mining engineers to minimize the risks of seal failure. Gateroad seal design more or less conformed to seal ratings used in the United States since 1971 where it was stated in 39 CFR 75.335 (Mine Safety and Health Administration – Title 30 Code of Federal Regulations, 1997) requires a seal to “withstand a

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static horizontal overpressure of 138 kPa (20 psi). Previous research by the former U.S Bureau of Mines (Weiss et al, 1999) indicated that it would be unlikely for overpressures exceeding 20 psi (138kPa) to occur very far from the explosion origin provided that the area on either side of the seal contained sufficient incombustible and minimal coal dust accumulations. Current legislation in Queensland depends on the purpose or intent of the seal and its location, different design criteria are recommended by QDME. These recommended design criteria are listed in Table 3. Table 3. Queensland approved standard for ventilation control devices.

Location Purpose or Intent

Type A (2 psi)

14 kPa

(Recommended)

Limited Life

Production Panel

All VCDs installed are to remain “fit for purpose” for the life of the panel and be

capable of withstanding an overpressure of 14 kPa.

Type B (5 psi)

35 kPa

(Recommended)

Main Roadways

All VCDs constructed as part of the main ventilation system are to remain “fit for purpose” for the life of that area of the mine and always capable of withstanding an overpressure of 35 kPa.

Sealed Areas For use in mines where the level of naturally occurring of flammable gas is insufficient

to reach the lower explosive limit under any circumstances.

Type C (20 psi)

140 kPa

Sealed Areas For use in all circumstances not covered by Type B and D seals.

Type D (50 psi)

345 kPa

Sealed Areas When persons are to remain underground whilst an explosive atmosphere exists in

a sealed area and the possibility of spontaneous combustion, incendive spark or

some other ignition source could exist.

Type E (10 psi)

70 kPa

Surface

Infrastructure

Surface entry stoppings for temporary emergency use and may include

‐ Surface air locks, Main fan housing

New South Wales Trade & Investment Mine Safety Bulletin SB13‐04 published on 29 August 2013 Titled: “Sealing of a goaf or mined out area in an underground coal mine and management of legacy sealed areas” states the following. “The Mine Safety and Health Administration USA report into the Sago Mine explosion now means that the possibility of a lightning strike to the surface over and surrounding such areas must be considered as a direct ignition source in addition to mine infrastructure that may be capable of conducting electrical energy into a mine.” “It is recommended that any coal operator in NSW that cannot quantitatively demonstrate that their existing sealing arrangements for goaves or mined out areas are tolerable and maintained as low as reasonably possible (ALARP), then one of the following measures be implemented to eliminate or control the risk of explosion post sealing: … b. Evacuate the mine until a sealed goaf or mined out area has passed through the explosive range and employ normal 20 psi overpressure rated mine seals.

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c. Permit the goaf or mined out area to pass through the explosive range, without evacuation of the mine, after the installation of 120 psi or 120 psi plus overpressure rated seals in all entrances to the goaf or mined out area…” UK and European Practices. The issue of explosion resistant seals has been addressed a number of times in the UK by committee. In 1942 (3), descriptions of various explosion resistant seals were given, some of which had been successfully used to contain explosions within sealed areas after sealing of fires or heatings. Generally, the seals were very long (30 feet or so), but no particular explosion rating was stated. In 1962 (4), it was assumed that seals should be designed to withstand explosion pressures in the range of 20 to 50 psi (140 to 350 kPa). Construction methods were described again, but these were based on past practice rather than any design methods or tested seals. In 1985 (5), the design of explosion resistant seals was again reviewed. Explosion resistance rating appears to have been increased to 524 kPa (76 psi) based on observed pressures developed by methane/coal dust explosions. The length of a monolithic gypsum pack was established to resist this pressure was given as : L = (H+W)/2 + 0.6 Where L = length of seal (m) H = height of seal (m) W= width of seal (m) It was also acknowledged that If it were not for the possible risk of explosion, the operation of sealing‐off would consist simply of providing a seal designed solely to prevent access of air to the fire and requiring little or no mechanical strength." There does not appear to be any requirement for explosion resistance ratings on any other ventilation structures other than seals used to control fire and spontaneous combustion. Very little information has been obtained on the standards for explosion resistant seals in European coal mining operations. West German coal mines are required to comply with a "Directive for the construction of stoppings" (6) , which requires the explosion resistant stoppings be capable of withstanding maximum static pressures of 0.5 MPa (5 bar, 75 psi). It would appear that these structures are intended to "seal off, hermetically, parts of the mine workings," to prevent the propagation of "mechanical, thermal and toxic effects" to other areas of the mine. From a search of abstracts, Cybulski et al (7), indicate that explosion pressures in sealed off areas had been recorded at more than 30 bar (450 psi). However, conceding the difficulty of building a stopping of such a strength , it is assumed that, in Poland, less strong stoppings of about 5 bar (75 psi) would be sufficient in practice. Again it is considered that explosion resistant seals are required to prevent an explosion from propagating from within a sealed area.

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3. THEORY OF METHANE‐AIR EXPLOSIONS

Diffision of Methane in Air

A common misconception exist 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 and that miners are

instructed to measure Methane concentrations in a coal mine 1ft from the roof. The density of

Methane is about 0.55 times that of air, so buoyancy effects do exist. However, in the complete 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.

It is probably this misconception that firstly gave rise to the idea that explosions from within sealed

areas are not possible. The belief was that Methane will accumulate in a layer near the roof and

become inert due to saturation levels above 17%. Thus if all the Methane released from within the

sealed panel will rise and become inert in a layer near the roof, the chance for an explosive event is

thus near zero.

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

non‐uniform concentration, and at uniform constant pressure and temperature 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 relationship is called Fick’s first law of diffusion, in which the

diffusion constant for Methane in air is 0.157 cm³/s (McCabe and Smith 1967).

Zlochower (2007a) used diffusion theory to estimate the time required for methane concentrated at

the roof of a 2.1m high coal seam to diffuse uniformly from roof to floor. The methane concentration

reaches uniformity within 10% in about 21 hours. His calculations assume a slow methane influx rate.

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.

PRACTICAL CONSIDERATION

Methane behaves differently in a ventilated and unventilated environment. Current training of mine

personnel teaches that Methane will always accumulate near the roof, but this is untrue in an

unventilated environment such as a sealed of panel.

A Methane concentrate value from within the panel is representative of the complete panel and not

only a layer near the roof of the panel.

19 | P a g e

Thermodynamic analysis

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 develop a respect for the amount of energy potential in a Methane‐air mix note that the energy

content in 1m³ of ideal Methane air mix is about equivalent to 0.75 kg of TNT.

Zipf et.al in IC9500 states that thermodynamic theory provides us with information regarding the

pressures and temperatures that can be achieved with Methane‐air mix explosions. Possible

temperature within the explosion of 2670 K and a pressure increase of 800‐kPa is predicted in theory.

Combustion of nonstoichiometric Methane‐air mixes produces lower temperature and pressure

increases. Experimental data prove that these values are accurate especially in places where the ratio

of surface area to enclosed volume is small such as found in mine entries. Natural gas in coal mines

usually consist 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. 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.

Thermodynamic theory provides us with similar information relating to Coal‐Dust‐air mix explosions

as with Methane‐air mix explosions. Coal dust explosion data presented by Hertzberg and Cashdollar

[1987], Wiemann [1986] and Cashdollar[1996] show that rapid combustion of coal dust in air will

develop a Constant Volume (CV) explosion pressure similar to that for Methane‐air.

Tunnel explosion mechanics

The above results for CV explosions are based on 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 centre of the vessel, (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

FACT 1

Combustion of stoichiometric (≈10%) Methane‐Air mix in a

closed volume raises the absolute pressure from 101 to 908 kPa

FACT 2

Combustion of fuel‐rich coal dust‐Air mix in a closed volume

raises the absolute pressure from 101 to 790‐890 kPa, which is

only slightly less than combustion of Methane‐Air mix.

20 | P a g e

development of much higher explosion pressures. Numerous authors have described the combustion

process as it initiates from 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 2007]

Consider a mine entry closed at both ends and filled with Methane‐Air mix as shown in Figure 1.

Ignition occurs at the far right end and the flame propagates to the left. Upon ignition the initial

laminar flame speed is only 3 m/s; however a slow deflagration accelerates and the turbulent flame

speed may increase to 300m/s. The pressure in the burner gas behind the flame front increases to

908‐kPa CV explosion pressure. The combustion front acts as a piston, compressing the unburned gas

in front of it. The leading edge of the acoustic wave propagates to the left at the local sound speed of

about 341 m/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 ahead of the flame front

is termed “pressure piling”

Figure 1 – Slow Deflagration diagram and Equivalent pressure profile along tunnel

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 roof and rib falls, ground support, machinery and wall roughness are possible forms of

tunnel roughness that will enhance turbulent flow. The increased turbulent flow in the unburned gas

ahead of the flame 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 speed, the combustion process becomes

fast deflagration as shown in Figure 2. Combustion of precompressed unburned gases leads to

pressures greater than the 908‐kPa CV explosion pressure. These transient pressure waves will

equilibrate and the overall pressure inside the closed tunnel will eventually settle balance to 908‐kPa.

Wrinkled flame frontTurbulent flow

Pressure wave front

Unburned Gas

Slow deflagration wave1.0

2.0

3.0

4.0

5.0

MPa

21 | P a g e

Figure 2 – Fast Deflagration diagram and Equivalent pressure profile along tunnel

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). When

detonation occurs, the pressure wave front and the flame front become one (Figure 3). In a

detonation, the transient pressure rises in a few microseconds to about 1.76‐MPa Methane‐Air, but

then quickly equilibrates to 908‐kPa CV explosion pressure.

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 auto‐explosions occur within these pockets. These small shocks

compress and preheat the unburned gas so intensely that they auto‐ignite the mixture. The small

compression waves coalesce into a larger amplitude shock. The detonation thus becomes self‐driven

by the auto‐explosions occurring at the shock front and propagates away from the DDT point at the

detonation pressure for as long as combustible material is available.

MPa

5.0

2.0

1.0

3.0

4.0

Fast deflagration wave

Pressure wave front Wrinkled flame frontTurbulent flow

Burned Gas Unburned Gas


The rougher the tunnel surface or the more equipment standing in the roadway, the more likely it

becomes that an explosion will develop. This statement on face value goes against popular believe

that a smooth channel is required to develop higher explosive pressure values.

To understand the above phenomenon the reader must consider the hydraulic principle that a

laminar surface condition has increased surface drag compared to turbulent surface conditions.

(Dimples on a golf ball make the ball fly further than a smooth surface ball.)

The second principle to consider is that the increased roughness increases turbulence in the air

flow resulting in better particle dispersement and contact between potential and actual explosive

particles. (Stratified air mixture in modern internal combustion engines.)

22 | P a g e

Figure 3 – Detonation diagram and Equivalent pressure profile along tunnel

A fundamental parameter for gaseous detonation 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 30cm for methane in air, while Shepherd [2006] gives a range of 25‐35cm. Kuznetsov et

al.[2002] report a cell size of 20cm for stoichiometric mix of Methane‐Air and that this cell size

increases to about 30‐45cm as the mix composition deviates from stoichiometric methane in air.

Because the smaller dimension of typical coal mine tunnels exceeds 1m 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. For coal mine headings with and equivalent diameter of about

2m, the run‐up length to DDT could range from 100 to 200m. 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 effects the onset of DDT and can only decrease the run‐up length. Pending

further research, NIOSH scientists selected 50m as the minimum run‐up length for detonation of

Methane‐Air in a heading.

MPa

4.0

3.0

5.0

2.0

1.0

Detonation wave

FACT 3

If detonation occurs in an ideal Methane‐Air mix at 1 standard

atmosphere, the detonation pressure developed is 1.76‐MPa.

Wrinkled flame frontDetonation and

combustion front

Burned Gas Unburned

Gas

23 | P a g e

Figure 4 – Reflected wave diagram and Equivalent pressure profile along tunnel

If a detonation wave impacts a solid wall such as a mine seal, a reflective shock wave forms and

propagates in the opposite direction back through the combustion products. Several classical works

on fluid dynamics of combustion present analysis of this reflective detonation wave pressure. Landau

and Lifshitz [1959, 1987] derived a relationship between the incident and reflective shock pressure as

ϒ ϒ ϒ

ϒ (1)

Where ϒ = the specific heat ratio of the combustion products

Assuming that ϒ = 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. When the 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 x 1.76 or 4.50‐MPa.

Reflected detonation wave

Burned Gas

MPa

5.0

4.0

3.0

2.0

1.0

Reflected Detonation wave

FACT 4

A Methane‐Air detonation wave reflects from a solid surface at

a pressure of 4.50‐MPa.

24 | P a g e

Experimental proof

The theoretical calculations above gives a CV explosion pressure of 908‐kPa detonation pressure of

1.76‐MPa and reflective pressure of 4.50‐MPa. Test explosions conducted at experimental mines in

the United States and Europe confirms 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 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 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 at the Maja Mine in Poland. The amount of explosive mix ranged from 10 to 1,000m³ and the

length of the gas zone ranged from 4.3m to the full 57m length of the experimental tunnel. Two tests

in which the explosive mix completely filled the tunnel produced peak pressures greater than 3.2‐

MPa. Pressure piling clearly occurred during these particular tests. Flame speed was measured

1,200m/s corresponding to 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 0.2‐MPa to 1.5‐MPa. These experimental results showed a clear relationship between the length

of the explosive mix zone and the maximum explosive pressure. A gas zone length more than 50m

long can develop peak explosion pressures of more than 2.0‐MPa which in turn lead to detonation.

In test #1397 conducted at Experimental Mine Barbara in Poland, Cybulski [1975] back‐calculated

explosion pressures in excess of 4.1‐MPa. The experimental explosion was initiated in coal dust about

200m from the closed end of a tunnel. Three measurements of pressure wave speed ranged from

1,600 m/s to 2,000 m/s which clearly suggest detonation. Unfortunately, sensors could not measure

the pressure directly, however the explosion punched a 1.4m² hole into a 32mm thick steel door.

In his Ph.D. dissertation, Genthe [1968] examined peak explosive pressure, flame speed and the length

of an explosive mix zone in order to determine their relationship. Experimental explosions with


The maximum combustion overpressure that can be developed during a Coal‐Dust‐Fuelled

explosion is 790kPa. The maximum combustion overpressure that can be developed during a

Methane‐Fuelled explosion is 1760kPa.

This fact is contrary to the popular South African mining believe that a Coal‐Dust‐Fuelled (CDF)

explosion develops higher pressures than what a Methane‐Fuelled (MF) explosion can develop.

The chief reason for the lower pressure in a CDF explosion is the amount of energy required from

the explosion process to move and disperse coal particles into the atmosphere prevents the

development of detonation conditions. Due to the well dispersed particle concentrate in MF

explosions as explained by diffusion the MF explosion can develop detonation.

25 | P a g e

subsonic flame speeds less than about 330 m/s led to explosion pressures less than 1.0‐MPa.

Explosions which developed supersonic flame speeds of up to 1,200 m/s produced peak pressures of

up to 1.8‐MPa. The length of the explosive mix zone also correlated to high peak explosion pressures.

Similar to the previously described results from Cybulski et al. [1967], an explosion with a gas zone

length of 50m produced peak explosion pressure of 1.8‐MPa which could be indicative of detonation.

In the article Zipf, R.K., Gamezo, V.N., Sapko, M.J., Marchewka, W.P., Mohamed, K.M., Oran, E.S.,

Kessler, D.A., Weiss, E.S., Addis, J.D., Karnack, F.A., Sellers, D.D., “Methane‐Air Detonation

Experiments at NIOSH Lake Lynn Laboratory,” Journal of Loss Prevention in the Process Industries,

doi:10.1016/j.jlp.2011.05.003, available online, 8 pp., May 2011 proves that experimental explosion

pressures are close to theoretical pressures. Methane‐air mixtures varied between 4% and 19%

Methane by volume. Average pressures recorded behind the first shock pressure peak varied between

1.2 MPa & 1.7 MPa. The detonation propagated with an average velocity between 1,512m/s to

1,863m/s. The measured detonation velocities and pressures are close to their corresponding

theoretical Chapman‐Jouguet (CJ) detonation velocity (DCJ) and detonation pressure (PCJ). Outside of

these detonation limits, failed detonations produced decaying detached shocks and flames

propagating with with velocities of 1/2 DCJ. Using a direct method to initiate detonation in test

mixtures, detonation is sustainable in methane in air mixtures over a range from 5.3% to about 15.5%.

This range almost encompasses the entire normal combustion limits of 5% to 16% methane in air as

reported by Cashdollar et al.(2000)

In the article by Oran, E.S., Gamezo, V.N., Kessler, D.A., ‘Deflagrations, Detonations, and the

Deflagration‐to‐Detonation Transition in Methane‐Air Mixtures’, April 2011 the Naval research

laboratory for computational physics and fluid dynamics investigate if given a large enough volume of

flammable mixture of NG and air, such as may exist in a coal mine, can a weak spark ignition develop

into a detonation? During the tests it was shown that it is possible to compute DDT in Methane‐Air

mixtures from first principles by solving the RNSE, given enough numerical resolution.

Summary

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 pressures.

Calculations in previous sections of this report describe this worst‐case scenario ‐ the combustion of

a confined, stoichiometric Methane‐Air mix. Pressure was shown to increase from atmospheric

pressure to 908‐kPa. 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 wave 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. A DDT may occur resulting in a detonation wave,

which has a pressure of 1.76‐MPa at 1 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.

26 | P a g e

An inhomogeneous, poorly mixed or layered explosive gas cloud will generate lower explosion

pressure. However, according to previous discussions, diffusion leads to homogeneous, well‐mixed

and non‐layered explosive mixtures within the still atmosphere of a sealed area. 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 for an explosion.

Experimental work in mines has proven the above theoretical values as being realistic and relevant.

27 | P a g e

4. 140kPa & 400kPa WALL EXPERIMENTAL RESULTS

Over the past 40 years thousands of explosion tests have been conducted on different types of seals

and plugs. Although the pressures used to design these seals are lower than the current U.S. three

tiered pressure rates, the results from these tests remain relevant and applicable. The author has

included the results from a small number of tests on Gunnite type seals in the Lake Lyne Experimental

Mine (LLEM) from the report “Evaluation of Reinforced Cementitious Seals” [Weiss, Cashdollar,

Mutton, Kohli, Slivensky – 1999]. The results are shown due to their relevance to seals currently used

in South African.

Three types of seals (Plug Seal, Meshblock Seal and Gunmesh Stopping) are briefly discussed regarding

their construction methodology, material characteristics, anchoring mechanism.

Plug Seal

The specific seal construction consisted of two of 75mm thick Gunmesh and shotcrete stoppings

providing the outer walls of the 1200mm thick plug seal. The interior was filled with an injected lower

density core of Aquablend with a designed compressive strength of 3.45‐MPa. A description of the

construction technique of a Gunmesh stopping is presented in the “Gunmesh Stoppings” section later

in the report. Aquablend is the trade name for a low‐density, pumpable, cementitious product. The

Gunmesh walls provide a permanent shutter for the wet‐mix core filling material. Steel spacers located

at 1300mm from the floor and spaced across the crosscut at 600mm centres provide lateral support

to the two stopping walls, which were subjected to hydraulic head by the Aquablend wet mix. The

wet‐mix slurry core is placed using an air‐driven Langley Placer in a continuous process.

Three 32mm diameter injection ports were cast into the inside face Gunmesh shutter stopping. These

ports were located 400mm from the mine roof. One port was located 900mm from the left rib, the

second port was located near the middle of the crosscut and the third port was located 900mm from

the right rib. Plastic extension pipes (air bleeders) were located within the stopping walls 300mm from

the mine roof. These pipes were angled towards the mine roof to the highest cavities to ensure

complete filling to the roof. The Aquablend was injected simultaneously through all three injection

ports. As the Aquablend reached the roof and came out of the bleeder pipes, these pipes were

progressively closed. The last injection port was pressurized until refusal of the placer at 1.38‐MPa

slurry pressure. This ensured that the slurry level was in direct contact with the mine roof.

28 | P a g e

Figure 5 – Sectional layout diagram of plug seal

Meshblock seals

Meshblock seals range in thickness from 175mm to 325mm. Roof and floor bolts were installed at

600mm centres and rib bolts were installed at 1m centres, which formed a vertical plane at the centre

line of each seal. These 24mm diam steel bolts were 1.2m long and fully encapsulated with polyester

resin capsules within the 600mm deep 30mm dim holes. The floor of the LLEM facility is consists of a

concrete surfacebed. The concrete floor was chiselled to a depth of approximately 20mm, providing

a key and a level footing for each seal. It must be noted that the test environment in the LLEM is one

of solid, non‐yielding strata.

Figure 6 – Photo of Meshblock installation and diagrammatic section of Meshblock wall

75 751200

Air bleeder pipes

32mm dia injection ports

Gunmesh stopping Wall

Aquablend plug seal

Reinforcing starter

Meshblock

29 | P a g e

The meshblock formwork consisted of a U‐shaped frame formed as a formed grid of 4mm diam

steel‐wire framework (square grid pattern on 152mm centres). A 3mm aperture steel mesh screen

encloses the sides and is an integral part of this formwork, enabling the shotcrete nozzleman to

examine the flowing shotcrete material. The Meshblocks were laid horizontally in rows in which the

ends were butted to each other and secured by plastic or wire ties. Normally two rows of Meshblock

were erected at a time and cast with shotcrete. The cycle was repeated until seal completion. There

was a 45mm overlap on each successive layer of Meshblock. The sides of the Meshblock form were

secured by five steel clips that were attached to the wire grid to keep the seal width consistent. Care

must be taken to ensure that the interval between casting successive layers does not exceed ½hr in

order to prevent the forming of a cold joint. All Meshblock seals were constructed in a continuous

manner until completion. Each of these three seals was sprayed with the Quikrete MB500 shotcrete,

which is a mixture of cement and minus 5mm aggregate.

As the Meshblock structure was build upwards, the floor steel bolts were extended vertically towards

the roof. Steel bolts overlap 600mm from the vertical extended reinforcing. Normally the roof bolts

were installed and the lower bolt holes were aligned by string‐line and drilled so that the vertical steel

reinforcing formed straight lines.

The measured compressive strength of the shotcrete ranged from 38 to 41‐MPa after 7 days and 46

to 60MPa after 28 days. Note that the shotcrete is not applied horizontally through a spraying process

into the void created by the Meshblock shutters, but is rather placed into the void vertically using

shotcrete equipment. The main difference of this method is that the material is not added as vertical

layers onto a backing, but is applied in horizontal layers inside a form. The advantages of this method

is that voids do not form in the concrete behind reinforcing and the concrete application layers run

perpendicular to the direction of bending moment stresses thus reducing the impact of dry joints.

Furthermore the amount of rebound material is reduced

Gunmesh stoppings

Gunmesh is a product supplied by Tecrete Industries. The Tecrete MB500 is a mixture of cement and

minus 5mm aggregate shotcrete supplied in 25kg bags and applied with the REED lova 215 gunite

machine.The roof, rib and floor was cleaned of loose debris back to solid material. The concrete floor

of the LLEM facility was keyed approximately 20mm to form a level base. The bolt pattern in the

Gunmesh stoppings required 24mm diam by 1200mm long bolts in the roof, rib and floor spaced at

1m centres. The bolts were fully encapsulated 600mm into the solid rock forming a vertical plane.

30 | P a g e

Figure 7 – Photo of Gunmesh installation and diagrammatic section of Gunmesh wall

The Gunmesh formwork consisted of a 4mm diam galvanized wire framework (150mm apertures)

sheet in 1.2 x 3.0m sections. A galvanised steel mesh with 3mm apertures was welded integral with

this heavier wire framework. This composite sheet was attached to an additional square grid pattern

of welded 4mm diam galvanised wire bars held apart from the composite sheet by cross braces of the

same material, thus forming a lattice of ≈50mm thickness open at one side. This sheet was tied to the

roof and floor bolts. The Gunmesh sheet edges were overlapped 100mm and secured together with

plastic cable ties. Once the formwork was in place and attached to the peripheral bolts, it was in‐filled

from the open side with shotcrete. The vertical roof and floor bolts were linked by attaching steel bolts

of the same diameter. The bolt sections were overlapped 0.5m with the extended section of the

grouted roof and floor bolts. Care must be taken that there is total coverage of the steel bolts with no

shadow of dry or overspray shotcrete material and that the Gunmesh cage is attached to and envelops

the steel bolts. The Gunmesh stopping was sprayed shotcrete with no delay until the specified nominal

thickness was achieved. The compressive strength of the shotcrete used was 37MPa at 7 days and

50MPa at 28 days.

31 | P a g e

Testing Methodology

All the explosion tests on the various seals were conducted in the LLEM which is located outside of

Pittsburgh Pennsylvania. The underground entries consist of approximately of 7,620m long workings

developed in the mid‐1960’s for the commercial extraction of limestone and 2,286m of entries

developed by the former USBM in 1980‐82 for research [Mattes et al. 1983]. Each of the seals are

constructed in the crosscuts along a 500m long tunnel. Nearly 19m³ of natural gas is injected into the

closed end of the tunnel that is sealed off by a plastic diaphragm within a 210m ignition zone. An

electric fan with an explosion‐proof motor housing is used to mix the natural gas with the air in the

ignition zone. Three electrically activated matches across the face of the entry are used to ignite the

flammable Gas‐Air mixture. Barrels filled with water are placed in the ignition zone to create

turbulence.

To achieve an explosion pressure pulse significantly in excess of 138kPa, coal dust is used along the

tunnel length. The coal dust is loaded onto shelves suspended from the mine roof at 3m increments

along the tunnel length. The mass of coal dust is increased to produce higher explosive pressures.

Pressure data is gathered by strain gauge pressure transducers fixed to the side of the test tunnel and

optical sensors to detect flame arrival. The pressure transducers are from Dynisco, Viatran or Ginesco

and are rated at 0‐100 psia, with 0‐5V output, infinite resolution and response time <1ms. The sensor

data gathered during the explosion tests were relayed from the data gathering stations to an

underground instrumentation room. A high speed, 64 channel, PC‐based computer data acquisition

system was used to collect and analyze the data. This system collected the sensor data at a rate of

1,500 samples per second over a 5 second period. The data was then processed using LabView, Excel

and PSI‐Plot software. The report pressure data was averaged over 10ms to achieve a 15‐point

smoothing.

Structural analysis

Table 3 summarizes the explosion test results on the five test walls. Each of the wall types,

construction dimensions and peak explosion pressures are indicated. This information alone does not

explain the full extent of how each seal performed. An example hereof is that the Meshblock seal in

crosscut 2 survived a peak pressure of 425‐kPa while the equal thickness and type of wall in crosscut

3 was damaged by a peak pressure of 300‐kPa.

The author of this report is a Professional Structural engineer. Based on the author’s expertize in

structural engineering he deducted that the test seals can be analysed as reinforced flat slabs with

rotational freedom and translation fixture to all slab edges. The methodology of this analysis is to

enter the structures dimensions and maximum blast pressure, as a static pressure, into finite element

slab analysis software. This software produces the maximum bending moment that the slab will

experience in both span directions. From the bending moment the pressure in the concrete is

calculated using elastic bending equations. These equations produce a more conservative pressure

than what can be expected from full plastic deformation of the concrete section. With the small data

set available, a more conservative approach is required and thus the elastic deformation equations

will suffice.

32 | P a g e

Table 4 – Explosion test summary

No. Seal Type Description Seal

Dimensions Explosion Test Number

# 347 #348 #349

1 Plug Seal w = 5430mm h = 1950mm t = 1200mm As = 0mm²

PEP = 150kPa BM = 80kN.m TCP = 0.3MPa

PEP = 330kPa BM = 173kN.m TCP = 0.7MPa

340kPa BM = 178kN.m TCP = 0.7MPa

2 Meshblock Seal w = 5760mm h = 2260mm t = 325mm As = 818mm²

PEP = 150kPa BM = 80kN.m TCP = 18MPa

PEP = 315kPa BM= 165kN.m TCP = 38MPa




PEP = 300kPa BM = 227kN.m TCP = 52MPa Cracked

595kPa BM = 466kN.m TCP = 106MPa Destroyed



PEP = 370kPa BM = 200kN.m TCP = 154MPa Destroyed

Destroyed

5 Gunmesh Stopping w = 5790mm h = 2220mm t = 75mm As = 818mm²

PEP = 105kPa Cracking & Spawning due to pressure

PEP = 345kPa Destroyed

Destroyed

PEP – Peak Explosion Pressure

BM – Bending moment

TCP – Transformed Cross section Pressure

The analytical method produces theoretical compressive stresses within the slab section that can be

compared to the cube strength of each wall at the specific age of the concrete. Observations made

during the tests seem to substantiate this logic. The first observation is from the first explosion test

on the Gunmesh wall in cross cut 5 namely that: ”A chipped out section of the center of this stopping

on the C‐drift wall indicated localized compression failure of the shotcrete, with the entire structure

very close to failure.” Secondly a picture of the back side of the Meshblock seal in cross cut 4 taken

after the second explosion shows the position of cracks on the face of the seal. The shape and position

of these cracks closely matches the shape and position of cracked two way span slab analysis

predictions used by structural engineers to design suspended flat reinforced concrete slabs, such as

used in the author’s analysis. Furthermore all of the structures were constructed of cementitious

material and reinforced in a manner similar to how reinforced flat concrete slabs are designed and

constructed.

The values of Transformed Cross section Pressure (TCSP) in Table 4 indicate that the seals survived the

explosion pressures that produced TCSP values below or slightly above the concrete cube strength of

the walls. This proves the direct correlation between concrete strength with associated blast pressures

and the compressive stresses developed within a cementitious two way span slab with static pressure

loading. Even from these limited number of experimental data points the method seems to have value.

33 | P a g e

Ingwe spec walls

One of the more popular seal designs currently used in South African coal mines is know as “Ingwe

Spec” walls. This dry‐crete reinforced concrete thin structural member with doweled shear connection

to the surround shares many structural similarities with Meshblock seals. Assuming that the above

physical explosion tests on Meshblock seals are valid, then it stands to reason that the above results

are applicable to similar cementitious walls. Thus the observed results from explosions against

Meshblock seals can be used to back calculate the explosive pressure that an Ingwe spec wall will be

able to withstand in a similar experimental explosion.

With this assumption as basis the author calculated the explosive pressure that a Type‐I Ingwe spec

wall will withstand. Type‐I walls was used because it represents the typical mine header opening size

in South African mines. The analysis assumes that the compressive concrete strength and all

dimensions are according to the construction specifications and no construction defects are present.

According to the author’s back calculation of the “Ingwe spec” construction specification compared

to the known experimental results of Meshblock seals, the maximum explosive pressure that a “Type‐

I Ingwe Spec” wall can withstand is ±110kPa. This value is supported by similar results from static finite

element analysis of the same Ingwe spec wall. Both the back calculated and finite element analysis

calculated value of 110kPa is only a quarter of the suggested 400kPa design load.

It is worrying that both the finite element analysis and back calculations are based on perfectly

constructed edge conditions. The experimental seals were constructed within the LLEM with concrete

surfacebed floor and surrounding limestone roof plus sidewalls having high compressive strength that

does not deteriorate over time. In a real coal mine the roof, floor and side wall strata usually consist

of materials with weaker compressive strength compared to the experimental mine conditions.

Furthermore the compressive strength of the strata in a coal mine reduces over time as the coal

oxidizes or the minerals in the strata are exposed to water. The difference in edge constraint values

between actual conditions and theoretical calculation values has as a result that the actual wall

strength will be less both in bending and shear capacity compared to the theoretical calculations.

Conclusions

There exists a vast amount of experimental data relating to seal performance in experimental

Methane‐Air explosion tests. Even a basic structural analysis, such as attempted above, produces

valuable equations that can be used to predict the performance of a seal. These results were used by

the U.S. Army Core of Engineers (USACE) to calibrate their single degree of freedom structural analysis

software Wall Analysis Code (WAC), which underlines the value of this historic information.

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5. PANEL INERTIZATION

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 from

5% to 16% by volume. 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 limit and become inert ranges from about 1 day to several weeks

or more depending on the mine’s Methane liberation rate.

To illustrate the development of explosive gas accumulations in sealed areas and the self‐inertization

process, mine ventilation engineers use the Coward diagram shown in Figure 5. The range of explosive

Methane‐Oxygen 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. Point C and F lie within this fuel‐rich and oxygen‐low inert region.

The inertization path depends on the rate of Methane emission relative to the rate of Oxygen

depletion in the 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

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.

During the time the sealed area contains a volume of explosive mix, an adequate sized 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.

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Figure 8 – Coward triangle for explosive zone on Methane in air [Coward and Jones 1952]. Also

shown are different paths to an inert atmosphere.

Figure 6A shows a circumstance of gas accumulation that can occur as a result of normal sealing

practice. In this instance a large volume of explosive gas accumulates that completely fills the volume

and is completely confined with no venting possibility. 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 become inert as a result of Methane concentration above the upper

explosive limit, Oxygen depletion form 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, Oxygen‐laden air tends to leak into sealed areas and

can create a volume of potentially explosive mix immediately behind 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 instance of explosive gas accumulation

cause by leaking seals is depicted in Figure 6B.

3025201510 5

25

20

15

10

5

0

OXYGEN (%)

METHANE (%)

Normal Air

A

B

D

FE

C

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Figure 9 – Two general types of explosive gas accumulation within sealed area

As mentioned earlier in Section 2 of this report, the use of type C seals (Australia) designed to

withstand a 140‐kPa 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 management plan to collect the requisite data, analyze and

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

inertization, if required.

With regard to the traditional Coward triangle graph representing the Methane‐Air explosive zone,

the Queensland monitoring standards 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 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. Monitoring tubes enter the mine via a borehole. Typical tube‐

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

area and the other half in the sealed areas. Pumps draw air samples continuously from each

monitoring point to the monitoring shed for analysis. Inside the shed is a solenoid‐valve‐manifold

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. Real‐time data are displayed at the mine’s control

centre, where trained operators can respond as necessary.

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

A – Volume filled with explosive atmosphere

(Completely confined with no venting)

B – Leaking seals with explosive atmosphere

(Partially completely confined with venting into inert atmosphere)

KEY

Seals

Explosive atmosphere

Inert atmosphere

37 | P a g e

injection at a low flow rate, they lack sufficient capacity for injection during an emergency, such as a

fully developed spontaneous combustion event. The Tomlinson boiler 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.


A sealed of Methane rich panel presents an explosion risk due to the possibility that the

atmosphere may enter the explosive range.

This risk can be effectively managed by artificially steering the sealed panel atmosphere outside of

the explosive range. This is accomplished by continual effective atmosphere monitoring and

inertization by means of Nitrogen or CO₂ flooding.

A sealed of panel is a potential bomb. Atmosphere management effectively disarms the bomb. This

type of management reduces the initial size of the containment walls, but demands a long term

continual monitoring and inertization investment.

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6. U.S. Legislation

U.S. Federal legislation known as ‘Final Rule 2008’ (30 CFR Part 75 : Sealing of Abandoned Areas:

Final Rule) is the current document used for the design of all new seals in U.S. coal mines. The final

rule includes requirements for seal strength, design, construction, maintenance and repair of seals

and monitoring and control of atmospheres behind seals in order to reduce risk of seal failure and

the risk of explosions in abandoned areas of underground coal mines. It also addresses the level of

overpressure for new seals.

Background

In order to understand how the Final Rule 2008 is implemented it is imperative for the reader to

understand the fundamental relationship between the U.S. mining industry and to the legislator

MSHA. MSHA (Mine Health and Safety Administration) is a department under the direction of the

U.S. Federal Department of Labour tasked with the administration of health and safety at all mines

in the U.S.A. MSHA has a department under its direction responsible for evaluation of technical and

engineering related mine health and safety issues known as ‘MSHA Tech Support’.

The mining community throughout the U.S. is divided into twelve geographical districts. At the head

of each mining district is a District Manager responsible for all district mining activities according to

federal and district legislation, including ventilation control. All mines have to submit their planned

support and ventilation plans to their designated District Manager for approval before work may

commence. The ventilation plans must, along with all other ventilation layouts, include details and

layouts of sealed off abandoned panels with specific reference to the type of seals to be used. The

type of seals the mine intends to use must feature on the list of seals approved by ‘Tech Support’.

This list of approved seals is posted on the MSHA single source web site for approved seals. The

District Manager will refer the ventilation plan to ‘Tech Support’ for approval off all technical

matters regarding seals. If both the layout and the type of seals to be used are to the satisfaction of

the District Manager and ‘Tech Support’ approval for the ventilation plan will be given in writing.

Only when the written approval has been obtained from the District Manager may construction

commence under the guidelines set out in the Final Rule 2008.

Before a new seal design may be listed on the approved seals list by MSHA it must undergo

meticulous scrutiny from ‘Tech Support’. The new seal design must conform in all aspects to the

Final Rule 2008 regarding strength, design, construction, maintenance and repair of seals. The seal

design with all its required specifications is then published on the MSHA single source website for all

users, installers and designers to peruse.

All work must be carried out according to the specifications set out by the specific seal design. All

test result of material samples have to be reported to the District Manager who conveys the

information to ‘Tech Support’ for approval. All documents have to be stored or shared according to

the Final Rule 2008 with the District Manager. Inspections of all work before, during and after

completion will be performed by qualified representatives of the District Manager and reports will

be submitted in this regard.

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Legislation

Below follows the text of the law known as ‘30 CFR § 75.335 Seal strengths, design applications, and

installation’. This text should be read in conjunction with all addendums, tables, graphs, figures and

official explanations that form part of the legislation. The excerpt below is quoted to indicate the

amount of detail and encompassing scope of the legislation.

“(a) Seal strengths. Seals constructed on or after October 20, 2008 shall be designed, constructed,

and maintained to withstand‐‐

(1)(i) At least 50‐psi overpressure when the atmosphere in the sealed area is monitored and

maintained inert and designed using a pressure‐time curve with an instantaneous overpressure of at

least 50 psi. A minimum overpressure of at least 50 psi shall be maintained for at least four seconds

then released instantaneously.

(ii) Seals constructed to separate the active longwall panel from the longwall panel previously mined

shall be designed using a pressure‐ time curve with a rate of pressure rise of at least 50 psi in 0.1

second. A minimum overpressure of at least 50 psi shall be maintained; or

(2)(i) Overpressures of at least 120 psi if the atmosphere in the sealed area is not monitored, is not

maintained inert, the conditions in paragraphs (a)(3)(i) through (iii) of this section are not present,

and the seal is designed using a pressure‐time curve with an instantaneous overpressure of at least

120 psi. A minimum overpressure of 120 psi shall be maintained for at least four seconds then

released instantaneously.

(ii) Seals constructed to separate the active longwall panel from the longwall panel previously mined

shall be designed using a pressure‐ time curve with a rate of pressure rise of 120 psi in 0.25 second. A

minimum overpressure of 120 psi shall be maintained; or

(3) Overpressures greater than 120 psi if the atmosphere in the sealed area is not monitored and is

not maintained inert, and

(i) The atmosphere in the sealed area is likely to contain homogeneous mixtures of methane between

4.5% and 17.0% and oxygen exceeding 17.0% throughout the entire area;

(ii) Pressure piling could result in overpressures greater than 120 psi in the area to be sealed; or

(iii) Other conditions are encountered, such as the likelihood of a detonation in the area to be sealed.

(iv) Where the conditions in paragraphs (a)(3)(i), (ii), or (iii) of this section are encountered, the mine

operator shall revise the ventilation plan to address the potential hazards. The plan shall include seal

strengths sufficient to address such conditions.

(b) Seal design applications. Seal design applications from seal manufacturers or mine operators shall

be in accordance with paragraphs (b)(1) or (b)(2) of this section and submitted for approval to

MSHA's Office of Technical Support, Pittsburgh Safety and Health Technology Center, P.O. Box 18233,

Cochrans Mill Road, Pittsburgh, PA 15236.

(1) An engineering design application shall‐‐

(i) Address gas sampling pipes, water drainage systems, methods to reduce air leakage, pressure‐

time curve, fire resistance characteristics, flame spread index, entry size, engineering design and

analysis, elasticity of design, material properties, construction specifications, quality control, design

references, and other information related to seal construction;

(ii) Be certified by a professional engineer that the design of the seal is in accordance with current,

prudent engineering practices and is applicable to conditions in an underground coal mine; and

(iii) Include a summary of the installation procedures related to seal construction; or

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(2) Each application based on full‐scale explosion tests or equivalent means of physical testing shall

address the following requirements to ensure that a seal can reliably meet the seal strength

requirements:

(i) Certification by a professional engineer that the testing was done in accordance with current,

prudent engineering practices for construction in a coal mine;

(ii) Technical information related to the methods and materials;

(iii) Supporting documentation;

(iv) An engineering analysis to address differences between the seal support during test conditions

and the range of conditions in a coal mine; and

(v) A summary of the installation procedures related to seal construction.

(3) MSHA will notify the applicant if additional information or testing is required. The applicant shall

provide this information, arrange any additional or repeat tests, and provide prior notification to

MSHA of the location, date, and time of such test(s).

(4) MSHA will notify the applicant, in writing, whether the design is approved or denied. If the design

is denied, MSHA will specify, in writing, the deficiencies of the application, or necessary revisions.

(5) Once the seal design is approved, the approval holder shall promptly notify MSHA, in writing, of

all deficiencies of which they become aware.

(c) Seal installation approval. The installation of the approved seal design shall be subject to approval

in the ventilation plan. The mine operator shall‐‐

(1) Retain the seal design approval and installation information for as long as the seal is needed to

serve the purpose for which it was built.

(2) Designate a professional engineer to conduct or have oversight of seal installation and certify that

the provisions in the approved seal design specified in this section have been addressed and are

applicable to conditions at the mine. A copy of the certification shall be submitted to the District

Manager with the information provided in paragraph (c)(3) of this section and a copy of the

certification shall be retained for as long as the seal is needed to serve the purpose for which it was

built.

(3) Provide the following information for approval in the ventilation plan‐‐

(i) The MSHA Technical Support Approval Number;

(ii) A summary of the installation procedures;

(iii) The mine map of the area to be sealed and proposed seal locations that include the deepest

points of penetration prior to sealing. The mine map shall be certified by a professional engineer or a

professional land surveyor.

(iv) Specific mine site information, including‐‐

(A) Type of seal;

(B) Safety precautions taken prior to seal achieving design strength;

(C) Methods to address site‐specific conditions that may affect the strength and applicability of the

seal including set‐back distances;

(D) Site preparation;

(E) Sequence of seal installations;

(F) Projected date of completion of each set of seals;

(G) Supplemental roof support inby and outby each seal;

(H) Water flow estimation and dimensions of the water drainage system through the seals;

(I) Methods to ventilate the outby face of seals once completed;

(J) Methods and materials used to maintain each type of seal;

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(K) Methods to address shafts and boreholes in the sealed area;

(L) Assessment of potential for overpressures greater than 120 psi in sealed area;

(M) Additional sampling locations; and

(N) Additional information required by the District Manager.”

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7. U.S. ARMY CORPS OF ENGINEERS

Introduction

Researchers at the US Army Corps of Engineers (USACE) Engineer Research and Development Centre (ERDC) are recognized experts in the engineering design of structures that can resist explosion effects and protect military personnel. Because of their recognized expertise, the Office of Mine Safety and Health Research (OMSHR) researchers developed an interagency research agreement with USACE in 2006 for guidance in coal mine seal design. The USACE research objective was to apply protective structure design technology used by the defence establishment to the design of seals in the US coal industry and to transfer that technology to the US mining industry.

During July 2014 USACE released a document titled “Structural Analysis and Design of Seals for Coal Mine Safety” TR‐14‐31 (G.W.McMahon, J.A.Rullan‐Rodriguez, M.S.Holmer, R.E.Walker, J.L.O’Daniel, J.R.Britt, R.K.Zipf) that discusses methods to structurally analyse coal mine seals and proposes methods to be used by engineers to design coal mine seals. This publication presents the result of comprehensive data acquisition regarding the calibration and use of structural analysis software for concrete structures during methane air explosion events. Furthermore the report provides numerical solutions regarding shear resistance of walls and plugs using either moiling or steel anchors. These numerical solutions are tabulated and summarised in several graphs that can be used by a civil engineer in the design of either a concrete wall or plug in accordance to the design specifications given in the ‘Final Rule 2008’.

The TR‐14‐31 report continues further to evaluate the structural performance of concrete walls designed according to the ‘Final Rule 2008’ using the Wall Analysis Code (WAC) single degree of freedom structural analysis software during a worst case methane air explosion event as suggested by the NIOSH document IC9500.

Lastly TR‐14‐31 explores alternative methods to protect walls and plugs against the shock wave pressures generated during a methane air explosion event. This section discusses blast wave attenuators both in terms of experimental and numerical performance analysis. These structures are not allowed according to ‘Final Rule 2008’, but are discussed due to the large protection properties these structures display to reduce damage to seals during an explosive event.

This report will not elaborate beyond the information provided in the USACE report, but will highlight important issues raised and discussed in TR‐14‐31.

The document TR‐14‐31 is available on the Coaltech reports website for inspection and downloading.

Important concepts to design engineers


The following issues must be considered when designing a seal:

The seal must work in conjunction with the surround strata. The surrounding strata material

is not an engineered material. The strata’s strength is position and time variable.

A seal structure loaded beyond its design capacity should fail in a ductile manner and not

catastrophically brittle. Explosive concrete compression failure, buckling of elements,

overall shear failure or modular disintegration modes of failure must be avoided.

Use well understood and known engineering materials to reduce the amount of uncertainty

inherent in this design.

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According to the authors of TR‐14‐31 numerous techniques and important engineering facts are presented in the document; however, there are three simple overarching concepts that designers of coal mine seals must consider based on the experience of the protective structure design community. First, seal design involves both the design of the seal structure itself and the foundation needed to support or restrain the seal structure. A seal design will fail to perform as intended if it breaks free from the surrounding foundation rock or coal and moves as a unit due to a weak, inadequate foundation design. Second, if a seal is loaded beyond its design capacity, it should fail gradually in a ductile failure mode and not catastrophically through a brittle failure mode. Catastrophic failure through buckling or shear failure must be avoided. Third, seal designs should use materials with known, well‐understood, and controllable material properties. For this reason, reinforced concrete is the material of choice for most protective structures in military applications. The rock properties of most seal foundations are unknown and difficult to quantify. For this reason, design engineers should use rock bolt anchors for the seal foundation rather than relying on unknown frictional properties of the surrounding rock mass.

Protective Structure Design and Analysis Methods

This chapter in TR‐14‐31 introduces protective structure design concepts that were pioneered by military engineers for the design of structures to resist explosion effects. The principles of protective structure design can be applied to the design of coal mine seals. The chapter discusses three analysis methods that were applied to analyse coal mine seals.

The simplest analysis method is the equivalent static method, in which the given dynamic design problem is transformed into an equivalent static design problem through the use of a dynamic load factor (DLF). This factor converts the dynamic load into an equivalent static load for subsequent analysis and design. The strength of the materials used in the structure are scaled by a dynamic increase factor (DIF) that accounts for the increase in strength that most materials exhibit when subjected to a dynamic load.

The most widely applied method for dynamic structural analysis is numerical solution of single‐degree‐of‐freedom (SDOF) systems. These methods transform the structure into an equivalent mass with an equivalent stiffness and numerically integrate the equation of motion to calculate the displacement response of the structure. The Wall Analysis Code (WAC) (Slawson 1995) is a well‐known and widely‐accepted example of an SDOF analyzer.

Finite element methods were also applied to coal mine seal analysis. When used to conduct a fully‐dynamic analysis of a structure subjected to a dynamic load, these methods compute the internal dynamic stresses in the structure directly for subsequent design consideration.

Protective structure design

Prior to the mid‐20th century, the design of facilities to resist explosions was empirical; that is, it was based on studies of past catastrophic events. Beginning in the 1960s, military engineers developed quantitative procedures for PSD that are described in several design manuals. These military documents present design procedures for structures that may be subjected to explosions and similar dynamic loads. The design of coal mine seals is a similar problem, i.e., a seal is subject to dynamic loading from a methane‐air explosion. The designer must calculate the dynamic response of the seal structure and its foundation, and finally, the designer must specify construction details for each design.

The study reported that form TR‐14‐31 applies the protective structure design methods described in the Unified Facilities Criteria (UFC) 3‐340‐02 (Department of Defense 2008), which is organized into six chapters. (All chapter references below refer to UFC 3‐340‐02)

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• Chapter 1 discusses the components of an explosion protection system and the tolerance of humans to explosion forces. • Chapter 2 describes blast loads from various types of explosions, primary and secondary fragments from explosions, and various kinds of shock loads. • Chapter 3 presents the principles of dynamic analysis of structures, including resistance functions, dynamically equivalent systems, and dynamic responses. • Chapter 4 describes reinforced concrete design including the dynamic strength of materials, the design of slabs, beams, and columns, and design for flexure, diagonal shear, and direct shear. • Chapter 5 discusses structural steel design including the properties of steel, the design of beams and columns, and design to resist failure modes such as tensile failure, shear failure, and buckling. • Chapter 6 examines special materials for explosive facility design such as masonry, precast concrete and earth‐covered structures.

Analysis of Pre‐2006 NIOSH and USBM Seal Tests

The study reported that forms TR‐14‐31 used various analysis tools to examine structural tests of coal mine seals conducted at the NIOSH Lake Lynn Laboratory (LLL). Adequate data already exist to verify the analysis and design of seals constructed from conventional steel and concrete materials. However, little data exist for the analysis and design of seals constructed with cement foam, polyurethane foam and aggregate, or mine waste rock (gob). The single‐degree‐of‐freedom model Wall Analysis Code (WAC) was applied to test data of the cement foam plug seal and polyurethane foam and aggregate plug seal types. Based on full‐scale test data of various seals, resistance functions were developed that describe the response of a seal to a load, either static or dynamic. The resistance functions are based on a shear failure mode around the perimeter of the seal and apply to plug‐type seal designs. The resistance functions were verified against the existing test data and then used to design similar structures to resist the new MSHA design loads stated in the Final Rule 2008. These resistance functions were then built into the Wall Analysis Code for Mine Seals (WAC‐MS)

Analysis of Seal Foundations

Both the seal structure and its foundation must resist the design pressure‐time loads indicated by the curves in the seal regulations. The design explosion load applied to the face of a seal must be transferred through the seal structure into the surrounding coal, roof, and floor rock. This chapter in TR‐14‐31 considers two methods to anchor a seal, i.e., hitches and rock bolt anchors. A hitch is a shallow excavation in the rock surrounding the seal and is designed to hold the structure in place when it is subjected to a design load. The bearing strength of the rock provides the resisting force to anchor the seal. However, the mechanical properties of the surrounding rock mass are highly variable, difficult to measure, and may be inadequate. Therefore, the recommended method to anchor a seal is through rock bolts embedded in the seal structure and extending into the surrounding rock. These steel anchors can provide sufficient shear or tensile resistance to hold the seal in place under a design load. This section presents methods to analyse the anchorage capacity of hitches and rock bolt anchors for seal foundations.

Analysis of Seal Structures

This chapter of TR‐14‐31 discusses how seal structures can fail and then presents detailed analyses for three kinds of seal structures, i.e., reinforced concrete walls, concrete plugs, and plugs made from mine waste rock or gob. The various failure modes for a reinforced concrete wall are described including flexure of the wall, diagonal shear in the wall, and shear along the foundation interface. A reinforced concrete wall analysis is presented that follows methods presented in UFC 3‐340‐02 (DOD 2008). An explosion test on a reinforced concrete wall is also described. The wall is analysed using the WAC, and the calculated response of the wall is compared to the measured response from the full‐scale dynamic test. The comparison demonstrates that analyses using SDOF tools such as WAC are

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very accurate for most practical engineering purposes. A thick concrete plug anchored to the surrounding rock with rock bolts is analysed with the FEM. The analyses showed the stress magnitudes within the concrete plug and the combined shear and tensile loads on the rock bolt anchors. Finally, a gob plug seal is analysed to show the expected stresses and anchorage requirements for this unique structure.

Guidelines for Design of Coal Mine Seals

This chapter of TR‐14‐31 describes detailed design procedures for reinforced concrete seals and unreinforced concrete plug seals. The procedures are illustrated with numerical examples, and design charts are presented for a wide range of conditions.

Reinforced concrete seal design follows the following general steps.

• A. Design Inputs 1. Specify design load 2. Specify allowable failure 3. Specify safety factor 4. Determine dynamic load factor (DLF) 5. Specify coal mine geometry 6. Determine support conditions 7. Specify material properties 8. Determine dynamic increase factor (DIF)

• B. Foundation Design (seal anchorage) 1. Determine approximate yield line location 2. Determine approximate shear forces 3. Determine number of rock bolt anchors 4. Determine minimum seal thickness based on foundation design

• C. Seal Structure Design 1. Estimate reinforcing steel requirements 2. Determine diagonal shear reinforcement 3. Determine actual vertical moment capacity 4. Determine actual horizontal moment capacity 5. Determine static properties of the design 6. Determine actual yield line location and ultimate resistance 7. Determine direct shear capacity of concrete 8. Determine dynamic response of seal

Behaviour of 120‐psi Seals Subject to Methane‐Air Detonation Pressure

A reinforced concrete mine seal must be designed to have a static ultimate resistance of 288 psi in order to respond elastically to the 120‐psi instantaneous rise time pressure‐time curve in the Final Rule (2008). The designs created in previous chapters of TR‐14‐31 satisfy these strength requirements. However, this design pressure‐time curve is a simplified curve and not an actual methane‐air detonation pressure‐time curve. The following sections in TR‐14‐31 analyse how the reinforced concrete designs presented in previous chapters will respond to an actual methane‐air detonation pressure‐time curve.

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Blast wave attenuators

TR‐14‐31 states that a blast wave attenuator is “a stationary device used to reduce or lessen the blast wave effects by reducing the interior increase in pressure” (ASCE 1997). They are used to protect buildings in petrochemical facilities or underground rooms in military complexes from the effects of a nearby explosion. As an example of blast wave attenuator construction, sand and rocks may be placed in the ventilation shafts of an underground bunker in order to decrease blast wave effects while allowing ventilation airflow. Alternatively, a system of air intake baffles may be used that have little effect on airflow but that dramatically reduce the transmission of a short‐duration blast pressure into the structure (Baker and Harrell 1992).

Figure 10. Gob plug seal concept by Lusk, Unrug, and Perry (2009) composed of (1) blasted roof rock, (2) a

light‐weight construction seal, and (3) filler material between the seal and blasted rock.

Figure 11. “Shot‐rock” concept by Sapko et al. (2010) composed of (1) blasted roof rock and (2) an optional

borrowed gob foundation

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Figure 12. “Fully‐stowed” concept by Sapko et al. (2010) composed of stowed gob material held in place with

wire mesh screen attached to the roof, ribs, and floor rock with rock‐bolt anchors

Blast attenuators are used to limit the effects of short blast durations (ASCE 1997) which reaches 617 psi for few milliseconds, but rapidly decreases to less than 100 psi after less than 10 msec. As shown by Lusk, Unrug, and Perry (2009) and Sapko et al. (2010), blast wave attenuators can significantly reduce the peak explosion pressure on a coal mine seal and may reduce the seal’s structural requirements with a potential savings in construction costs.

At least three concepts have been developed for blast wave attenuators in underground coal mines and are shown diagrammatically in figures 7 to 9. Each of these concepts were studied experimentally, and a limited amount of test data were collected on gob exposed to blast waves to document its performance as a complete coal mine seal or as a seal component.

Sapko et al. (2010) obtained measurements of the blast‐wave attenuation capabilities of blasted rock. Three full‐scale tests of attenuators were conducted at the NIOSH Lake Lynn Laboratory (LLL) in an entry measuring about 20 ft wide by 7 ft high and using a methane‐air explosion to generate the applied blast wave. The base of each test attenuator was blasted, run‐of‐mine limestone rock less than 18‐in. diameter topped with crushed limestone less than 6‐in. diameter.

Table 5 summarizes the dimensions of the three test attenuators and the pressure measurements upstream and downstream of the attenuators. In Tests 1 and 3 with no gap below the roof, the attenuator reduced the down‐stream static pressure by a factor of 67, or a reduction of 98.5%. This reduction is consistent in magnitude with experiments by Baker and Harrell (1992), who examined the effectiveness of baffles in suppressing air shocks entering intake airways of hardened structures. In Test 2, there was a 2.5‐ft gap at the roof. Since the height of the entry was about 7 ft, the attenuator only reduced the cross‐sectional area of the entry by about 64%. However, this reduction in cross‐sectional area reduced the down‐stream static pressure by a factor of about 2.4. These experiments demonstrate that a simple rock pile placed from floor to roof and at least 10 ft long at its base can reduce the magnitude of short duration blast waves by a factor of 67.

Table 5. Summary of blast‐wave attenuator tests at NIOSH–LLL

Test Number

Attenuator Thickness Upstream Static (side-on) Pressure

(psi)

Downstream Static (side-on) Pressure

(psi) 1 42 ft at base tapering to 15 ft at roof

No gap at roof 47 0.7

2 48 ft at base tapering to 32 ft at top 2.5-ft gap at roof

50 21

3 42 ft at base tapering to 5 ft at roof No gap at roof

57 0.85

Effectiveness of blast‐wave attenuators – numerical simulations The experiments described above used either a small, high‐explosive charge or a methane‐air

explosion to generate a blast wave. However, these test explosions do not capture the long‐

duration, constant‐volume explosion pressure that can be expected from an actual methane‐air

explosion within a sealed area in an underground coal mine. Therefore, in the Sapko et al. (2010)

study, J. R. Britt used a numerical gas explosion program called SHAMRC to calculate the pressures

upstream and downstream of an attenuator resulting from a large methane‐air explosion in a mine.

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Figure 13. SHAMRC‐calculated reflected pressures on the upstream side (at 510‐ft range) and downstream

side (at 640 ft) of the model attenuator (Sapko et al. 2010).

Figure 10 shows the calculated pressures upstream and downstream of the model attenuator

directly aligned with the blast source in C‐drift. The upstream static (side‐on) pressure on the

attenuator is about 240 psi, which produced a reflected pressure on the attenuator of almost 700

psi. The reflected pressure on the seal located downstream of the C‐drift attenuator was initially

about 20 psi. The reduction in pressure magnitude from upstream to downstream of the attenuator

was a factor of about 35, which is consistent with the experimental results. However, as shown in

Figure 10, the pressure on the seal downstream of the attenuator increases steadily to about 60 psi

after 2 sec, and eventually reached the constant‐volume explosion pressure of about 120 psi for a

stoichiometric methane‐air mixture.

Summary

The USACE document TR‐14‐31 presents an important body of information regarding improved

engineering design for protective structures against Methane and Coal‐Dust fuelled explosive

events. The document provides:

Simplified design requirements to engineers using well know engineering materials, such as

reinforced concrete.

Critical analysis of current structural design methods based on historical experimental data.

New cost effective alternative solutions in the form of blast wave attenuators.

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This document diverts from many current design practices in the U.S., but considering the vast

knowledge base on blast protection measures and structures available to the USACE it would be

foolish to ignore the recommendations and design considerations of this document.

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8. CURRENT AUSTRALIAN SEALING PRACTICE

This portion of the work is a pure desktop study of research articles and other information published

on the internet. The authors of the work were not interrogated regarding scientific proof or on the

interpretation of the work.

ACARP project C17015

Most of this section refers to the document published in 2011 titled “Australian Sealing Practice and use of Risk Assesment Criteria (ACARP) project C17015” by Hsin Wei Wy (Gillies Wu Mining Technology), A.D.S.Gillies (Missouri University of Science and Technology), J.W.Oberholzer (SIMTARS), R.Davis (SIMTARS). This document states that the purpose of the research and this resultant document was:”…to examine views of Australian operating mines on the industry’s approach to the use of seals; also the new US approaches to sealing and their possible application to Australian conditions…”

The document states that while many approaches to underground coal mining in Australia and the US are similar, Australian approaches to the management of hazardous mine atmospheres differ significantly. Australian risk management approach to handling hazardous situations implies adoption of international industry best practices. The Australian industry has gone through a debate on how the new US information on seal behaviour and new regulations should be incorporated, if at all, into Australian practice. However the Australian coal mining industry as a whole have decided not to adopt the principal dictates of the 2008 US seal regulations.

The second part of the ACARP research project being led by SIMTARS is undertaking further physical testing study of the risk of explosions in sealed areas. The propagation tube study of the consequences of explosions is being conducted to both determine the nature of the explosion overpressure that a structure can be subjected to and also the nature of the pressure pulse that will impact on the structure.

Industry questionnaire survey

The industry questionnaire survey that the document refers to was conducted in 2008 on a large number of Australian mines to establish how mine managers were handling seal design and implementation. In brief the questionnaire sought the following information:

Ventilation network details, such as main fans, underground monitoring systems, etc.

Specific questions regarding whether sealed areas pass through the explosive range.

Possible dimensions for explosion propagation, propensity for propagation and probability of explosion.

Current approach to installing Ventilation Control Devices (VCD) and seals.

Ground stress relationships and seal integrity and time dependent stress on seal during life of seal.

Additional issues: Sources of explosion Should seals be designed as impervious membrane or as explosion barrier? How should seals be designed and tested? Should design be by structural analysis or physical destruction testing? Pressure balancing Contractors vs Company labour installing VCD

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Relevant differences between US and Australian mines

Some of the relevant physical and operational difference highlighted in the report between US and Australian mines are:

Most (±80%) US longwall gate roads have three headings compared to two headings in Australian mines. The middle heading in US mines can be expected with a gas initiation to lead to an explosion disturbance characterized by a long run‐up distance.

US mines have a significantly lower take up of electronic monitoring. Only one US mine makes use of “Tube bundle” gas monitoring. Little proactive use of inertisation in US mines Limited use of ventilation network programs in US mines Limited usage of trained Ventilation Officers in US mines

The report raised the following cultural and implementation differences between US and Australian mines:

Currently many changes are occurring in the handling of mine atmospheres and potential flammability conditions as a result of the Sago Mine disaster.

There are a large number of small mines in the US MSHA adopts a system of “prescriptive regulations” and in general there is lower acceptance

of risk assessment approaches. There is a perceived lack of trust between managers and inspectors. US underground mining industry is more diverse and larger than the Australian industry, as a

result the industry is less cohesive and there is less available data, less sharing of information and less frequent industry forums.

Views on changed US approach

According to the results summarized in the document the Australian industry agrees that it should not adopt the new MSHA regulations, but should rather stick with what appears to work best for Australian conditions. The Australian industry should use appropriate risk levels for seal design. The proposed 840kPa seal design in the US is considered excessive and the US move to this rating is an overreaction. It is the believe in Australia that regardless of application of seal pressure rating requirements it is impossible to contain some explosions in the highly variable mine environment. The introduction of prescriptive US seal pressure ratings does not appear to have been formulated on any risk assessment basis. There are impressions that the US is coming from a lower standard compared to current Australian practice.

Australian approaches to health and safety management are formulated on a risk assessment basis under which hazards must be identified and appropriate “world’s best practice” systems adopted. The principal approach in Australia to goaf management is early prevention of hazardous situations through use of real time gas monitoring from the goaf periphery to ensure the maintenance of goaf inert atmospheric conditions. The comment has been made that the US approach of principally and almost exclusively considering “seal rating” is one of “guarding against failure rather than adopting an approach of prevention”.

SIMTARS Propagation Tube Test Work

The Queensland group SIMTARS is examining consequences of explosions through testing being conducted in a propagation tube as part of the effort to determine the risk of explosions in sealed off areas. This is being investigated not only to determine the nature of the explosion overpressures that a structure can be subjected to but also the nature of the pressure pulses impacting on the structure.

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An analysis of possible scenarios in a mine was made and indicated that there were nine different situations where a methane explosion could occur in a mine. The most probable of these scenarios was a high length to diameter ratio roadway that would be full or partially filled with an explosive mixture. If an explosion occurred in the workings of the mine the roadway would not be enclosed and in the case of an explosion occurring behind a seal the roadway would be enclosed. Tests in the propagation tube (as shown in Figure 11) were designed so that varying parts of the tube were filled with an explosive mixture and the tube was left open or closed off with structure that withstood or failed under the pressure.

Fi

Figure 14 – Layout of the explosion propagation tube at SIMTARS

Even though it could be argued that mine scale larger galleries with greater volume are more suitable for this type of study the propagation tube is nevertheless deemed appropriate for the following reasons.

It has been proven that the maximum constant volume pressure is determined by the temperature of the burning gases in the container and not by its volume. The nature of the volume or space in the container might however influence the temperature that can be reached.

The level of instrumentation on the tube allows significant information with regard to the pressures to be gathered.

The tube allows a high rate of testing and multiple daily firings to be conducted. The tube has a design strength of 2MPa and can be closed with a strong structure to allow a

contained constant volume explosion with a stoichiometric mixture. The tube at 30m long and 0.5m diameter has high length to diameter ratio. This allows simulation

of compression of unburned gases before the explosion front. Due to availability, natural gas it is being used as the fuel in preference to pure methane.

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ACARP Report Summary

From survey results analysis and recent Australian debate on the topic it is concluded that seal design should start from premise that it is impossible to build a perfect seal.

Seal designs must be determined using priorities from risk assessment of particular situations. Risk levels should meet ALARA (as low as reasonably achievable) with health and safety conditions expectations of less than 1 death per million miner days of work Less than 10 deaths per 10 million miner days of work Never more than 10 miner deaths

There is no known evidence of a mine atmosphere explosion detonations; every mine explosion has remained within the limits of a deflagrations.

Seal should be rated to “seal” and not on structural applied pressure loading (to keep goaf gases out of ventilation air and oxygen out of goafs).

Monitoring of goaf atmosphere and requirements for inert gases is critical. Mines with low gas levels should not face onerous conditions. Mines with potentially

explosible gases need to monitor, respond and control. It is believed that “one rule is not appropriate for all situations”.

Seals must be competent engineered structures that normally meet 140kPa pressure rating. More understanding of mine strata geomechanics is needed; structural analysis should take

account of the properties and behaviour of the strata surrounding the seal and maintain a low leak interface with coal seam and surrounding strata.

More understanding of goaf gases ignition potential is needed. More information is needed on the variability of gas concentration data across the extent of a goaf; it cannot be assumed that gas composition is the same along the length and breadth of individual goafs.

The comment has been made that the US approach of principally and almost exclusively considering “seal rating” is one of “guarding against failure rather than adopting an approach of prevention”. There appears to be a consensus among mine operators, inspectorates and union leaders that Australia should not blindly go down the path of copying US current and post Sago sealing practices.

Comments on ACARP report

Since the publication of the ACARP report in 2011 three other reports by R.K.Zipf et.al. (NIOSH) in 2012 titled “Methane‐Air detonation Experiments at NIOSH lake Lynn Laboratory” ,G.W.McMahon et.al. (US Army Corps of Engineers) in 2014 titled “Structural Analysis and Design of Seals for Coal Mine Safety” and G.W.McMahon et.al. (US Army Corps of Engineers) in 2014 titled “CFD Study and Structural analysis of the Sago Mine Accident” disprove many statements and assumptions made in the ACARP report. These statements include:

“There is no known evidence of a mine atmosphere explosion detonations; every mine explosion has remained within the limits of a deflagrations.” According to the CFD study report by the USACE “Belt hanger calculations show the load required to initiate bending of the belt hangers was about 150psi.” The maximum pressure change that a deflagration explosion can develop is 120psi. This new evidence confirms that the atmospheric pressure generated during the Sago Mine incident was beyond the maximum theoretic deflagration limit, thus it follows that to generate the calculated pressure either localized or general detonation had to occur.

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In the NIOSH detonation experiments it was found that detonation can be achieved with methane‐air mixtures varying between 4% and 15% methane by volume that encompasses the entire normal combustion limits. This was achieved using a larger diameter test pipe (105cm) with baffle restrictions along the length with Br of 0.3 or 0.6 and a length‐to‐diameter ratio of 70. This larger equipment compared to the SIMTARS propagation tube proved that detonation is more likely to occur than what was previously believed. The larger diameter allows detonation cells to develop near the outer mixture limits that cannot be achieved in the SIMTARS equipment. Furthermore the disturbances along the length of the pipe better represents underground condition than the smooth pipe used in the SIMTARS tests. The 524kPa peak pressure measure in the SIMTARS experiments across a small mixture range of 8.5% to 9.5% methane air mixture is significantly lower than the peak shock pressure range between 1,2MPa and 1,7MPa measured by the NIOSH experiments.

“…it is impossible to contain some explosions in the highly variable mine environment …” The USACE have developed several design tools such as the WAC single degree of freedom structural analysis software, that was calibrated according to pressure and deflection measurements of actual explosions, that can calculate to a high degree of certainty the expected performance of a seal during an explosion. This software in conjunction with ventilation control structures such as blast attenuation walls and sound engineering practice can be combined to design seals that will survive a worst case explosion event.

The ACARP document highlights several differences between US and Australian mining practice that should be considered in relation to South African mining conditions.

13 of the 14 respondents to the questionnaire use long wall mining methods and the remaining mine uses the board‐and‐pillar mining method, but do not install any seals within their operations. This distribution of extraction methods is not representative of South African mining conditions where most mines use board‐and‐pillar mining plus make extensive use of seals in closing off of abandoned panels.

The ACARP document specifically refers to the “…middle heading in US mines can be expected with a gas initiation to lead to an explosion disturbance characterized by a long run‐up distance.” Thus it identifies that this header presents a significantly different risk due to the higher potential to develop a detonation along the longer run‐up length. This middle heading is analogue to typical board‐and‐pillar mining environment where long run‐up distances are inherent to the process. Furthermore is has to be remembered that the Sago Mine incident occurred in a board‐and‐pillar type mine similar to South African mining practice.

The ACARP document states that: o US mines have a significantly lower take up of electronic monitoring. o Only one US mine makes use of “Tube bundle” gas monitoring. o Little proactive use of inertisation in US mines o Limited use of ventilation network programs in US mines o Limited usage of trained Ventilation Officers in US mines

All the above statements are true for typical South African mines as well. To implement an Australian approaches to health and safety management formulated on a risk assessment basis under which hazards must be identified and appropriate “world’s best practice” systems adopted the South African mining industry will have to invest in the above systems and technologies.

The ACARP project C17015 report should not be read in isolation neither should its findings be assumed to suite the South African mining industry. As shown above newly published work in the

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understanding of Methane fuelled explosions render many assumptions and statement made in this document incorrect. More importantly a reader of the ACARP report must bear in mind the differences between Australian and South African mining practice to understand why the sealing practice used in Australia is supported in the report. It follows from the ACARP report that if a South African mine would like to adopt the Australian sealing practice governed by risk management of hazardous mine atmospheres several improvement will have to be made towards technology and management processes to achieve international industry best practices. NSW Safety Bulletin SB13‐04

New South Wales Trade & Investment Mine Safety Bulletin SB13‐04 published on 29 August 2013 Titled: “Sealing of a goaf or mined out area in an underground coal mine and management of legacy sealed areas” states the following. The Mine Safety and Health Administration USA report into the Sago Mine explosion now means that the possibility of a lightning strike to the surface over and surrounding such areas must be considered as a direct ignition source in addition to mine infrastructure that may be capable of conducting electrical energy into a mine. It further goes on to make the following recommendations. It is recommended that any coal operator in NSW that cannot quantitatively demonstrate that their existing sealing arrangements for goaves or mined out areas are tolerable and maintained as low as reasonably possible (ALARP), then one of the following measures be implemented to eliminate or control the risk of explosion post sealing:

a. Maintain the goaf or mined out area inert before and after sealing, and employ normal 20 psi overpressure rated mine seals.

b. Evacuate the mine until a sealed goaf or mined out area has passed through the explosive range and employ normal 20 psi overpressure rated mine seals.

c. Permit the goaf or mined out area to pass through the explosive range, without evacuation of the mine, after the installation of 120 psi or 120 psi plus overpressure rated seals in all entrances to the goaf or mined out area.

d. Ensure all ratings of seals are for a tested design that meets the overpressure rating awarded. Design testing must be for full scale seals. Testing and rating of seals must be undertaken by an organisation approved by the Chief Inspector.

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9. SUMMARY

Design of protective structures and measures against Methane and Coal‐Dust fuelled explosion

events have evolved significantly from first official 50 psi standard in 1921 “based on the general

opinion of men experienced in mine‐explosion investigations” to the current standards based on time

dependant overpressure experimental and thermodynamically calculated values or risk based

atmosphere management. This report aimed to provide information regarding the international

history of mine seal design to explain the current legacy of mine seals.

The report presents the current accepted science of Methane and Coal‐Dust fuelled explosions. The

information presented explains the relevance of the theoretical and experimental observations in

relationship to the coal mine environment. The concept of diffusion of Methane into the still

atmosphere is discussed to highlight the current misconception that Methane will become inert

along a thin layer near the roof of a sealed of panel. The complete ignition to detonation process is

explained according to current accepted science in order to predict time vs pressure wave/pulse

values for different underground environments.

Previously it was thought that detonation or even ignition of sealed off panels was so unlikely that it

was not worth considering. This assumption was the basis of the 20psi (140kPa) seal standards in the

U.S. (1971‐Mitchell) and later implemented in other mining countries including South Africa and

Australia. The latest science shows that detonation of methane fuelled explosion in sealed of panels

is indeed possible as was tragically proven in both SAGO and Darby mine explosions. As a result the

U.S. Federal Department of Labour introduced through MSHA new legislation known as the Final

Rule 2008 for the design and management of seals in coal mines. This legislation is quoted in the

report to show the extent of the design, implementation and maintenance requirements demanded

by the new US law.

After the Moura Number 2 mine explosion in 1994 the Austalian coal mining industry experienced a

similar industry wide re‐evaluation of panel sealing methodology. Task Team 5 under the guidance

of Oberholzer developed a different approach to sealing of abandoned panels compared to the US.

The approach in the Australian mines is governed by risk assessment and management of the sealed

atmosphere. As a result the pressure ratings required from their seals is significantly lower than the

equivalent US seals, but the latest recommendations in the Australian industry do use the US seal

pressure ratings during specific conditions. It must be recognised and considered in light of the

South African mining industry that the Australian sealing standards are primarily based on Longwall

mining methods with two road headers. This type of mining method is not common in South African

mines.

Not all mines exhibit equal risk in terms of developing an explosive atmosphere within a sealed of

panel. As a result the report briefly discusses the option of continual monitoring and management of

the sealed atmosphere by inertization as is typical in Australian mines.

Against the comprehensive knowledge and experience of the U.S.Army Corps of Engineers in the

design of structures to protect against explosions, their document TR‐14‐31 provides a valuable

source of information in the design of coal mine seals. TR‐14‐31 was developed by a group of

engineers and scientists with the view of providing a useable and comprehensive design guide to

structural and mining engineers in the development of seals for coal mines. Furthermore the

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document provides a critical review of current design standards in the US as well as proposes

alternative methods to protect against underground coal mine explosion events.

This report provides a summary of the history, current science and engineering behind

coal mine seals. This report does not provide conclusions, nor recommendations or

guidelines regarding the design of seals or management of sealed off panels.

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REFERENCES

1. Federal Register, [2008], “Rules and Regulations Sealing of Abandoned Areas‐Final Rule”, Title

30 CFR Part 75.335 CFR, Code of Federal Regulations, Washington DC: U.S. Government Printing

Office, Office of the Federal Register 73(76), Friday, April 18, 2008.

2 Mine Safety and Health Administration, [2010A], “Guidelines for the Seal Design Application”,

http://www.msha.gov/Seals/GuidelinesSealDesignApplications.pdf access February 2010,

11pp.

3. Mitchell D.W., [1971], “Explosion‐Proof Bulkheads – Present Practices”, U.S.Department of the

Interior, Bureau of Mines, Pittsburgh, PA: RI 7581, pp 1‐16.

4. Zipf, R.K., E.S. Weiss, S.P. Harteis, and M.J. Sapko, [2009], “Compendium of Structural Testing

Data for 20‐psi Coal Mine Seals”, IC 9515, U.S. Department of Health and Human Services,

National Institute for Occupational Safety and Health, 143 pp.

5. Hyde, D., Walker, R.E., O’Daniel, J.L., and McMahon, G.W., [2010], “Wall Analysis Code for Mine

Seals”, Proceedings of the 80th Shock & Vibration Symposium, San Diego, CA.

6. Slawson TR, [1995]. Wall Response to Airblast Loads: The Wall Analysis Code (WAC). Structures

Laboratory, U.S. Army Engineer Waterways Experiment Station, Vicksburg, MS.ATTN CEWES‐SS.

7. E.S. Weiss, K.L. Cashdollar, I.V.S. Mutton, D.R.Kohli, and W.A. Slivensky, [1999], “Evaluation of

Reinforced Cementitious Seals”, U.S. Department of Health and Human Services, National

Institute for Occupational Safety and Health, RI 9647

8. A.P. Cook, J.N. van der Merwe, [2000], “Design, construction and testing of underground seals”,

Itasca Africa (Pty)Ltd, Safety in Mines Research Advisory Committee.

9. C.Strydom, [1999], “Method Statement for Explosion Resistant Underground Containment

Stoppings Using the Space Frame / Gunite Wall”, Ingwe Coal Corporation, South Africa

10. R.K. Zipf Jnr. (Ph.D., P.E.), Micheal J.Sapko, and J.F. Brune (Ph.D.), [2007], “Explosion Pressure

Design Criteria for New Seals in U.S. Coal Mines”, U.S. Department of Health and Human

Services, National Institute for Occupational Safety and Health, IC9500

11. Gordon W. McMahon, Jose A. Rullán‐Rodríguez, Matthew S. Holmer, Robert E. Walker, James

L. O’Daniel, James R. Britt, and Richard K. Zipf, [2014], “Structural Analysis and Design of Seals

for Coal Mine Safety”, U.S.Army Corps of Engineers, Engineering Research and Development

Centre

12. C.Classen, Goaf Inertisation and Sealing Utilising Methane from In‐Seam Gas Drainage System,

11th Underground Coal Operators’ Conference, University of Wollongong & the Australian

Institute of Mining and Metallurgy, 2011, 369‐374

13. D.R.Humphreys, J.McCracken, “Task Group 5 – Final Report” 622.8684 TAS 1998

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14. Ian Anderson, Mine Safety Report No. SB13‐04, “Sealing of a goaf or mined out area in an

underground coal mine and management of legacy sealed areas”, August 2013

15. Mutton, IV and Remennikov, A, “Designing explosion rated ventilation seals for coal mines using

high‐fidelity physis‐based computer modelling”, in Aziz, 10th Underground Coal

Operators’Conference, University of Wollongong & the Australian Institute of Mining and

Metallurgy, 2010, 312‐325

16. Oran, E.S.; Gamezo, V.N.; Kessler, D.A.; “Deflagrations, Detonations, and the Deflagration‐to‐

Detonation Transition in Methane‐Air Mixtures”, Laboratory for Computational Physics and

Fluid Dynamics, Naval Research Laboratory, NRL/MR/6400‐11‐9332, April 2011

17. Zipf, R.K.Jr.; Gamezo, V.N.; Sapko, M.J.; Marchewka, W.P.; Mohamed, K.M.; Oran, E.S.; Kessler,

D.A.; Weiss, E.S.; Addis, J.D.; Karnack, F.A.; Sellers, D.D.; Office of Mine Safety and Health

Research (OMSHR), National Institute for Occupational Safety and Health (NIOSH), Laboratory

for Computational Physics and Fluid Dynamics (LCPFD), Naval Research Laboratory (NRL),

“Methane‐Air Detonation Experiments at NIOSH Lake Lynn Laboratory”, 2011

18. Cashdollar, K.L.; Weiss, E.S.; Harteis, S.P.; Sapko, M.J.; Urosek, J.E.; “Results of In‐Mine Research

in Support of the Investigation of the Sago Mine Explosion”, Report for Investigations 9678,

Department of Health and Human Services, September 2009

19. L. Kim Davis, Robert E. Walker, G.W.McMahon, J.Robert Britt, James L. O’Daniel, “CFD Study

and Structural Analysis of the Sago Mine Accident” – Final Draft, Geotechnical and Structures

Laboratory – US Army Corps of Engineers, May 2007

20. Department of Mineral Resources Republic of South Africa – “Guidelines for the Compilation of a Mandatory Code of Practice for the Prevention of Coal Dust explosions in Underground Coal Mines”, Feb 2002

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ANNEXURE A

EXAMPLES OF TYPICAL APPROVED 120psi

CONTAINMENT STRUCTURE DESIGNS

AND RELATED COST COMPARISONS

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COALTECH RESEARCH ASSOCIATION NPC

EXAMPLES OF TYPICAL APPROVED 120psi

CONTAINMENT STRUCTURE DESIGNS

AND RELATED COST COMPARISONS

By

R.P.van Wyk (Pr.Eng.)

March 2015 1

1 Copyright COALTECH

This document is for the use of COALTECH only, and may not be transmitted to any other party, in

whole or in part, in any form without the written permission of COALTECH.

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IMPORTANT NOTICE

THE DESIGNS AND COST COMPARISONS IN THIS DOCUMENT ARE EXAMPLES ONLY.

COALTECH DOES NOT PRESCRIBE OR RECOMMEND ANY OF THESE DESIGNS, THEY ARE MERELY EXAMPLES AIMED AT STIMULATING IDEAS AND DESIGNS FOR THE VARIOUS MINES AND MINING COMPANIES.

Mr ROEDOLF VAN WYK OF MANTELLA TRADING 310 WAS APPOINTED AS A DESIGN CONSULTANT BY COALTECH TO DRAW UP THIS DOCUMENT. COALTECH DOES NOT RECOMMEND OR SUPPORT ANY SPECIFIC SUPPLIER, CONSULTANT OR GROUP OF CONSULTANTS.

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Introduction

The current standard in U.S. and Australian coal mines calls for the construction of 120psi (830kPa)

seals in high explosion potential areas. This report investigates the estimated cost of constructing

these 120psi standard seals in South African mines.

Currently in South African coal mines only 140kPa and 400kPa seals of varying design are being

constructed. This report will make no reference to these current designs and will only examine the

cost of 120psi (830kPa) seals. At this point in time the only approved designs that comply to the new

120psi specifications are those being constructed in the U.S. under the control of the U.S. Mines

Health and Safety Association (MSHA). Several new designs have been proposed by the U.S. Army

Corps of Engineers (USACE) in co‐operation with MSHA. The cost of constructing these MSHA and

USACE seal designs will be investigated in this report.

The MSHA and USACE designs can be grouped into three structural types based on the intrinsic

method on which each structure functions. These structural types are:

Flexural walls (Flexural bending of structure)

Mass Plugs (Internal arching of thick section)

Attenuation barriers (Energy dissipation through compaction of material)

Each of these seal construction types vary according to:

Equipment requirement

Personnel required

Services required

Work methodology

These requirement headings are discussed and elaborated on for each type of seal. This discussion

will assist the uninitiated the reader in understanding the logistical and practical requirement to

construct each type of seal. The ultimate aim is to guide the reader to determine the most suitable

type of structure for his requirements and not be preferential towards any one design.

In order to make a directly cost comparison between each type of the seals, a set off physical

constants is assumed as typical to most coal mine environments in South Africa. If the reader wishes

to apply these seal costs to a specific mine then it has to be noted that these assumed physical

conditions may vary from the actual mine conditions. These costs may be further influence by

availability of materials, physical attributes relating to roadway opening, strata conditions and

construction skill sets available. The quoted costs are applicable to only one set of physical

conditions and are only an indication of cost, not an actual tender.

It thus remains the responsibility of the mine to employ a Structural Engineer or other suitable

designer to determine the suitability of each design to actual mining conditions. A fair market tender

will determine actual cost for each type of seal.

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Design assumptions

The following assumptions will be made regarding the seal design to enable a fair comparison

between each design type:

Roadway opening : 7m wide x 4m high

Design Load : 830kPa (120psi) with dynamic load factor of 2 plus a safety factor

of 1.5 according to MSHA Final Rule specifications. Each type of material will have a specific

safety factor assigned to it according to industry standards.

Strata conditions : Competent Hard Roof, Competent Hard Floor, Coal sidewalls (CM

Section). “Hard” classified as having a constrained compressive strength of greater than

50MPa.

Strata preparation : The strata in the roof, floor and sidewalls are to be protected from

oxidation by injection of polycarbonate resin grout under pressure into 2m deep perimeter

surrounding the wall position.

Area preparation : Access road between the shaft entrance and the work site is well

constructed and maintained. All roof and sidewall support along these access roads is

adequately supported. Work area where wall must be constructed to be swept clean with

LHD.

Travel distance : Travelling time from the shaft bottom to the wall is less than 30

minutes.

Services : Water and 1000V electricity supply is available within 100m of wall.

Surface conditions : Access is available on surface for installation of a borehole and

delivery of materials.

Ready mixed concrete : A ready mixed batching plant of good reputation is available within

a 30km radius of the wall.

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1. Flexural Wall

Layout drawing

Structural Definition

Flexural walls are constructed of high compressive strength concrete in conjunction with high

strength steel reinforcement similar to a reinforced concrete retaining wall. The wall typically

has a thickness less than the smallest roadway dimension. The wall is anchored by either being

sunk into the floor and sidewalls of the surrounding strata or by several steel bars drilled into the

strata that protrude into the wall. Due to shrinkage of concrete after placement the wall must be

pressure grouted with a cementicious material post concrete placement around the perimeter

to ensure that direct contact takes place between the wall and strata.

Due to flexing / bending of the structure the wall redistributes the applied forces towards the

perimeter of the structure in compressive and tensile stress perpendicular to the direction of the

applied load. Around the perimeter of the structure the applied force is transferred to the

surrounding strata through direct shear by reinforcing or pressure onto moiled strata. These

structures are as a result sensitive to construction tolerances towards dimensions and material

properties.

Advantages

o This type of structure is well known to engineers, therefor its structural behaviour

can be well modelled, thus the assurance from these numerical models is high.

USACE have developed extensive models to simulate the performance of these walls

according to actual methane explosive events.

o Due to the high strength of the materials and the specific use thereof this type of

structure is thinner than other structures.

o The structure is positively fixed to the surrounding strata

Disadvantages

o Structure is dependent on direct contact between the concrete structure and the

roof or floor. Deterioration of either roof or floor structural properties will affect the

performance of the wall.

Reinforcing to either side of the

concrete wall

Excavated Shear Key in

Footwall and Side walls

Steel starter bars in Roof and

Foot wall

TYPICAL SECTION THRU WALL

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o Positive fixture of the wall to the surrounding strata requires either moiling of the

strata or installation of a large number of shear steel bars. Both these activities raise

the cost, complexity, time and risk of the construction.

o This type of structure requires accurate high quality work, specialized skills and

material control.

Design specifications

Wall Thickness : 1.5m

Material strength required : 30MPa concrete at 56 days

Reinforcing required : 0.15% in each face in each direction plus distribution steel

at 500mm c/c parallel to surface

Moiling depth required : 750mm deep to side walls and floor

Starter bar details : Y25 (1m long) at 300mm c/c in roof and floor

Construction requirements

Equipment required :

Non‐Flamproof LDV people transporter (1 per 8 persons)

Flameproof Tractor c/w trailer to transport Reinforcing, Shuttering, Steel

pipes for concrete pump, moving concrete pump, catchment bin and

compressor.

Flameproof piston type concrete pump complete with pipework and

catchment bin.

Flameproof electrical compressor

Pneumatic Jackhammers for moiling key (2 per compressor)

Scaffolding to support people during: Moiling, Shutter construction,

Reinforcing installation, Concrete placement, Shutter stripping, Cement

grouting

Gopher to drill holes in roof for Starter Bars

Pneumatic Jackhammer to drill holes in floor for Starter Bars

Shuttering system to suite underground environment (Waterproof)

Telephone for communication to surface at placement borehole

Pneumatic mixer and pump to place cement grout into void between wall

top and the roadway roof.

Personnel required :

Site manager with applicable knowledge into construction of concrete

structures underground (1 per mine)

Miner with Blasting certificate or Competent A certificate according to mine

requirement (1 per shift)

Labourers qualified to operate Jackhammers, Gophers, construct scaffolding

and work at height. (At least 6 per shift)

Labourers experienced in fixing reinforcing according to construction

drawings. (At least 6 per shift)

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Labourers experienced in constructing steel or timber shuttering to fit

underground coal environment. (At least 6 per shift)

Labourers experienced in managing and cleaning steel pipelines for concrete

placement. (At least 4 per 100m of pipeline)

Technicians qualified to operate and maintain diesel equipment. (Tractor &

LDV’s) (At least one per machine per shift)

Technicians qualified to operate and maintain electrical equipment.

(Compressor & Concrete Pump) (At least one per machine per shift)

Services required :

Borehole with 8” steel sleeved casing within 100m of wall position.

Mine to indicate position where wall must be constructed

Mine Rock Engineer to confirm condition of surrounding strata

Work methodology :

Install roofbolts into sidewall outside key position to form breaker line.

Excavate moiling key in sidewalls by excavation using pneumatic

jackhammers according to construction drawing specifications. Work on

scaffold for key above 1m high.

Clean floor of excavated sidewall material.

Excavate moiling key in floor by excavation using pneumatic jackhammers

according to construction drawing specifications.

Clean out excavated material.

Drill hole in roof and floor with either a jackhammer of gopher according to

the construction drawing specifications for starter bars.

Install and fix reinforcement steel according to the construction drawings

onto the starter bars. (Use scaffolding)

Construct structurally sound and waterproof shuttering to both sides of the

wall. (Use scaffolding)

Prepare concrete pump and pipeline for placement of ready mixed concrete.

Place ready mixed concrete down borehole and pump into shutter to

completely fill the shutter void to the roof. Control ready mixed concrete

composition regarding pumpability, segregation during placement, strength,

setting time, hydration tempo.

Strike formwork when allowed by engineer and apply curing procedures to

wall.

Fill cavity between top of wall and roadway roof with high strength non‐

shrink grout after the concrete wall has settled to its final size.

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2. Mass Plug

Layout drawing

Structural definition

Mass plug seals are in effect a large volume of low compressive strength, time independent

material placed within an opening to create a barrier. The thickness of the plug is typically

greater than the least of the roadway opening dimensions. Depending on which material the

plug is constructed off the contact perimeter may need to be pressure grouted with a

cementitious material to ensure positive contact between the strata and the plug material.

Mass plug seals are robust simple structures that have a low dependence on construction

tolerances and material strength. The wall redistributing the applied forces towards the

perimeter of the structure in compressive stress through an internal formed arch with radius

dependent on the thickness to smallest opening dimension and the material strength of the seal.

Around the perimeter of the structure the applied force is transferred to the surrounding strata

through direct shear between the strata and concrete due to local interlocking or friction.

Advantages

o Due to the simple nature of this type of structure it is easy to construct, requiring only

shutters to both ends with material infill between. This ease of construction reduces

labour and equipment requirements thus total project cost.

o The structure requires a low tolerance regarding material strength and dimensions.

o Several different materials may be used that have adequate compressive strength,

limited material contraction and lack of material degradation over time to be used as

filling material.

Disadvantages

o The shear interaction between in‐situ strata and concrete is not completely understood

and thus not accurately quantified. This unknown value of shear greatly influences the

thickness of the structure and thus final cost.

o The large volume of material required to construct this wall must be transported from

surface to the wall. This carting of material either by tractor loads or down a borehole

influences the cost and time of construction.

H

Unreinforced low strength

concrete plug


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Wall Thickness : 1.2H (Typically 1.0 H – 2.0 H)

Material strength required : 20 MPa Concrete at 28 day (Typical 5MPa – 20MPa)

Reinforcing required : No Reinforcing

Moiling depth required : No Moiling

Starter bar details : No Starter bars



Non‐Flamproof LDV people transporter (1 per 8 persons)

Flameproof Tractor c/w trailer to transport one load of Shuttering, Steel

pipes for concrete pump at start of project.

Scaffolding to support people during: Shutter construction, Cement grouting

Shuttering system to suite underground environment (Waterproof)

Telephone for communication to surface at placement borehole


Site manager with applicable knowledge into construction of concrete

structures underground (1 per mine)



Labourers experienced in constructing steel or timber shuttering to fit

underground coal environment. (At least 6 per shift)

Labourers experienced in managing and cleaning steel pipelines for concrete

placement. (At least 4 per 100m of pipeline)

Services required :

Borehole with 8” steel sleeved casing within 100m of wall position.



Work methodology :

Clean floor at wall position to a broom finish.

Construct structurally sound and waterproof shuttering to both sides of the

wall. (Use scaffolding)

Prepare concrete pump and pipeline for placement of ready mixed concrete.

Place ready mixed concrete down borehole and pump into shutter to

completely fill the shutter void to the roof. Control ready mixed concrete

composition regarding pumpability, segregation during placement, strength,

setting time, hydration tempo.

Fill cavity between top of wall and roadway roof with high strength non‐

shrink grout after the concrete wall has settled to its final size.

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3. Goaf plug seal

Layout drawing

Structural definition

Goaf plug seals are not approved by MSHA final rule standards. The reason for this is that the

structure loses its attenuation ability after being exposed to an explosion, thus rendering it

redundant or less effective after an explosive event.

A goaf plug seal consists of several structures of varying materials and functions. These

structures are:

o Inbye load distribution wall – This wall functions as a containment for the rock fill material. Its

main function is to prevent the dump rock from settling and creating a gap between the roof

and dump rock. An increases gap to the roof will result in a decrease in the attenuation ability of

the structure. The secondary function of the wall is to reduce the volume of material from a

vertical end compated to the volume stacked at the material’s natural angle of repose. This

structure must have adequate strength to resist placement of material against it with a LHD.

Alternatively this structure may be omitted if another method to prevent sagging of the rock is

available or the effect of rock fill settling towards reduced attenuation functions is included in

the design of the outbye ventilation wall.

o Dump rock fill – The USACE is familiar with the use of sand bags to protect structures by

reducing the impact of an explosive event. The sand bags work by reducing overpressure or

attenuating overpressure due to the internal movement and compaction of sand grains that

result in a loss of energy and a slowed flow of air through the sand barrier. The material that is

used in this barrier must as a result have a large void ratio, thus equal graded material works

best.

Inspired by the success of simple sandbag walls NIOSH researchers in co‐operation with USACE

developed goaf attenuation barriers to be used to protect blast wall structures in underground

coal mines. These developers found that placing an adequate volume of equal graded rock

material reduces the overpressure outbye of the structure similar to sand bag walls. This

attenuation of pressure results in both a slower rise and lower maximum overpressure outbye of

the dump rock. The attenuation rate is dependant of the thickness of the barrier, grading of the

material and the gap between the top of the fill and the roadway roof. The maximum reduction

that the attenuation structure can accomplish is from worst case detonation pressure (4,4MPa

reflective wave) down to a slow rise 800kPa pressure. This reduction in pressure influences the

H

Dump rock fill (Goaf)


Outbye ventilation wall

Inbye load distribution wall

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design of the outbye ventilation wall by negating the Dynamic Load constant and additional

Safety factors due the unknown load application resulting in a thinner cheaper wall.

o Outbye Ventilation wall – This structure can be constructed as either a plug seal or flexural wall

with a reduced design pressure compared to a similar wall with no attenuation structure inbye

of it.

Advantages

o The attenuation function of the rock dump greatly reduces the impact force on the

ventilation wall. This reduction in force will result in a lighter ventilation structure with

subsequent reduced cost.

o The ready availability of equal graded dump rock in the mine may reduce the cost of

constructing the dump rock attenuator.

Disadvantages

o The structure is constructed from at least two separate structures and possibly a third

structure. This number of structures increases the final cost of construction.

o If correctly graded rock is not available on‐site then the material needs to be imported

at a considerable cost.

o Even though the design pressure on the ventilation wall is reduces, the subsequent

reduction in cost may not justify the additional cost of constructing the attenuation

barrier if cheap rock fill material is not available.

o An attenuation barrier reduces the overpressure due to the internal movement and

compression of the rock. Once this reduction in void spaces has occurred the structure

changes its structural behaviour closer to a mass plug seal than an attenuation

structure. There is currently no quantitive information available on how an attenuation

barrier will behave during subsequent explosive events. Thus the structure poses a

significant risk in future until this behaviour is understood and quantified.


Wall specifications:

Inbye gunnite wall – 400mm thick 25MPa spray applied concrete on Ref 617

welded mesh reinforcing and 1m long Y25 shear bars at 1m c/c all round

Goaf Fill – 2 H

Outbye ventilation wall – 800mm thick 25MPa concrete wall with 0.15%

reinforcing in both directions to both faces of wall. Wall to be fixed to

surround by moiling 750mm deep into floor and side wall.



Similar to flexural wall.

Flameproof diesel LHD. In order to stack the ballast rock to the roof it is

imperative that the bucket be equipped with a hydraulic push plate. This is

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not an industry standard and will require a dedicated machine for the

purpose.


Site manager with applicable knowledge into construction of underground

structures (1 per mine)



Similar to flexural wall

Services required :

Borehole with 20” steel sleeved casing within 100m of wall position for

ballast stone if not available on site.



Work methodology :

Clean floor at wall position to a broom finish.

Construct a gunnite type wall similar to Ingwe spec 400kPa wall as load

distribution wall.

Dump equal graded hard weather resistant dump rock with 80% of material

not passing a 63mm sieve down borehole.

Move dump rock from borehole with LHD and stack against inbye wall to fill

roadway opening from floor to roof. LHD will have to use a hydraulic push

plate in the bucket to stack the ballast up to the roof.

Construct a wall similar to the flexural wall, but with thickness of only

800mm as ventilation wall.

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ANNEXURE 1

COST CALCULATIONS:

Flexural Wall

Mass Plug

Goaf Plug Seal

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