Search Documents

By Yoginder Paul Chugh, Jason E. Tinsley

"Over the last five years, the authors have compiled a database of Short Encapsulation Pull Test (SEPT) evaluations from Interior Basin (Basin) coal mines. Typically, SEPT data are analyzed to obtain peak load carrying capacity per unit length of resin encapsulation designated as the ""Grip Factor,"" or GF. In a previous paper, ""An Analysis of Short Encapsulation Bolt Pull Test (SEPT) Data From Interior Basin Coal Mines,"" authors analyzed tangent value of the load-displacement curve as Anchorage Stiffness at 50% (AS-50) of the peak load, or GF value, to assess reinforcement potential of an installed bolt system. This paper attempts to correlate the secant AS VALUES (of SEPT on 6 ft (1.8 m) long installed bolts at a lower load of about four tons from several different mines in the Basin with their MSHA-approved primary support requirements. Since the current regulatory environment requires plan changes dynamically to ensure that roof falls are minimized, the authors hypothesize that the practiced roof control plan must meet the minimum requirement of adequately supporting immediate roof strata. The analyses are based on 181 SEPT studies performed by Minova professionals. Analyses indicate a macrolevel correlation for both roof support density and roof support cost that may have the potential to be used by mine operators to compare their performance with other mines in the area. Limitations of the analyses include ignoring mining depth, mining height and required secondary supports as part of the requirements for a roof control plan. The authors recommend additional analyses to develop this concept further to develop guidelines that can be confidently used by coal companies.INTRODUCTIONPartially- and fully-grouted bolts are extensively used in coal mines for roof support as part of MSHA-approved roof control plans. Since grouted bolts have a larger area of contact with the rock, they can develop much greater anchorage capacity than conventional expansion shell (point-anchored) bolts. Load transfer characteristics and interaction mechanics between steel rebar, grout and rock mass around the bolt hole for grouted bolts are generally assessed in the laboratory or field though short-encapsulation pull tests (SEPT), which have been extensively researched by Aziz and Jalalifar (2005), Aziz (2004a), Serbousek and Signer (1987), Signer (1990), and Mark et al. (2002), among others. In a SEPT, about 12 in. (30 cm) of bolt length is typically encapsulated in resin and a pull test is performed to collect load-displacement data prior to failure and, in some cases, post-failure. A typical load-displacement data set for a bolt is shown in Figure 1. SEPT characteristics are designated by peak load carrying capacity, or grip factor, which is the peak load carrying capacity per unit length of resin encapsulation. SEPT is routinely used to assess anchorage performance characteristics of different lithologies overlying a coal seam to identify type of bolt, bolt diameter, length and spacing across and along an excavation. Anchorage stiffness, or AS, at any load may also be calculated as the tangent or secant value of the slope of the load displacement curve at a particular load level. Anchorage Stiffness characteristics of bolts from SEPT data have typically not been considered in any depth until recently (Chugh et al., 2014; Chugh et al., 2015)."

Jan 1, 2016

By Kourosh Shahriar, Kazem Oraee, Ezzeddin Bakhtavar

"In this paper, a relationship between dependant parameters and the crown pillar thickness is first introduced. This relationship defines geotechnical problems caused by thin pillars and economic considerations created by pillars that are thicker than the optimum size. For this purpose, a dimensional analysis as an effective physico-mathematical tool was used. This technique restructures the original dimensional variables of a problem into a set of dimensionless products using the constraints imposed upon them by their dimensions. A model is hence introduced that calculates the optimum pillar thickness. The relationship introduced here and the method applied can be used by mining engineers in all situations where a combined open-pit and block caving method is deemed to be the most appropriate mining method.INTRODUCTIONMany deposits can be mined entirely with the open-pit method; others must be worked underground from the very beginning. In addition, there are the near-surface deposits with considerable vertical extent. Although they are initially exploited by the open pit method, there is often a ""transition depth"" where decision has to be made about changing to underground methods.Some of the biggest open-pit mines worldwide will reach their final pit limits in the next 10 to 15 years. Furthermore, there are many mines planning to change from open-pit to underground mining due to increasing extraction depths and environmental requirements. In this way, it is likely that block and/or panel caving will enable the operations to continue achieving a high production rate at low costs as an underground method (Bakhtavar et al., 2009).In these cases, it is often necessary to consider a crown pillar beneath the transition depth (open-pit floor) before starting an underground caving stope method (Figure 1).There are generally two kinds of crown pillar: a ""surface crown pillar"" and ""crown pillar between open-pit and underground mining."" In general, they both play the similar role in mining. Since the primary purpose of a surface crown pillar is to protect surface land users, the mine, and those working in it from inflows of water, soil, and rock; it is vital that the surface crown pillars remain stable throughout their life. When a crown pillar could potentially remain between open-pit and underground mining it is specially proposed to prevent water entering from the open-pit floor into the stope, as well as to reduce open-pit wall and floor caving."

Jan 1, 2010

By Xingkai Wang, Wenbing Xie

"To control the large deformation of soft rock around the roadway subjected to high stress, it is important to improve the stability of support structure, in addition to increasing support strength. In this study, the causes for the destabilization of U-steel frame and its loading characteristics were analyzed by theoretical analysis, laboratory experiments and field tests. The technology of constant support resistance, the principle of support structure compensation and the compensation technique are proposed. The results indicated that the main reasons for frame structure destabilization are caused by low strength of connector, irrational structure of connector, poor interaction between the surrounding rocks and support frame and structure, and unstable support structure. Thus, the corresponding technological measures, such as connector with high strength and rational structure are adopted. The bearing capacity of frame is unstable after slipping. In order to maintain the high support resistance when the frame is performing the support function, the frame connector should be retightened after slipping. Though the working resistance may be improved greatly by using the backfilling or back grouting of U-steel, the frame remains unstable because of the unstable frame structure. Therefore, according to the deformation characteristics of rocks surrounding the roadway and the structure instability of U- steel frame, the stable structure device is used to carry out the structural compensation of U-steel frame, and the whole bearing capacity of the frame and its structural stability are improved greatly.INTRODUCTIONThroughout the development of the roadway supports for coal mines in China, it can be seen that the supporting strength and supporting resistance increase continuously with increase in mining depth and supporting difficulties [1]. With respect to the ground control mechanism and supporting technology of rocks surrounding the soft rock roadways subjected to high stress, considerable research has been done both at home and abroad. The backfilling technology behind the high-resistance and yieldable U-steel wall, the new type of grout/bolt support technology and the bolting with high-strength mesh support technology were developed, and have been used widely in coal mines with notable economic benefits [2-7]. However, a large number of engineering practices show that in many cases the existing rock control technology of roadway may not control the large deformation of the soft rock roadways subjected to high stress. The reason that the large deformation of soft rock roadway occurs is mainly due to poor structure stability of existing frame bearing structure, not insufficient support strength."

Jan 1, 2015

By Khaled M. Mohamed, Michael M. Murphy, Ted M. Klemetti, Heather E. Lawson

"The design of rib support systems in U.S. coal mines is based primarily on local practices and experience. A better understanding of current rib support practices in U.S. coal mines is crucial for developing a sound engineering rib support design tool.The objective of this paper is to analyze the current practices of rib control in US coal mines. Twenty (20) underground coal mines were studied representing various coal basins, coal seams, geology, loading conditions, and rib control strategies.The key findings are: (1) any rib design guideline or tool should take into account external rib support as well as internal bolting; (2) rib bolts on their own cannot contain rib spall, especially in soft ribs subjected to significant load—external rib control devices such as mesh are required in such cases to contain rib sloughing; (3) the majority of the studied mines follow the overburden depth and entry height thresholds recommended by the Program Information Bulletin (PIB) 11-29; (4) potential rib instability occurred when certain geological features prevailed—these include draw slate and/or bone coal near the rib/roof line, claystone partings, and soft coal bench overlain by rock strata; (5) 47% of the studied rib spall was classified as blocky—this could indicate a high potential of rib hazards; and (6) rib injury rates of the studied mines for the last three years emphasize the need for more rib control management for mines operating at overburden depths between 500 ft and 1,000 ft.IntroductionEfforts to improve the stability of underground mine ribs have continued for decades. These efforts by mining professionals around the world persisted throughout the 1990s and 2000s as well, though much of this work focused on rib support methodology and selection of proper support for a specific location. Despite these efforts, it is unfortunate that there has been a continual occurrence of rib-related fatalities in underground coal mines. While overall improvements in rib safety have been achieved, the average fatality rate is still about 1.3 fatalities per year over the last 18 years (Jones, 2014).To manage and control the rib hazard, rib support is used in U.S. coal mines. The current rib support methods for U.S. mines fall into two main categories: rib control based on intrinsic support systems and those based on external support systems. Intrinsic or rib-bolt support requires a roof bolting machine that can install a bolt fixture into the rib. Furthermore, the appropriate level of support needed to control the ribs with an intrinsic-based support system as well as the design parameters for those support systems for various mine conditions have not been established in U.S. coal mines. Furthermore, surface control systems are often an integral part of these support systems and must be considered in the system design."

Jan 1, 2015

By Eric Zahl

Researchers at the Spokane Research Laboratory of the National Institute for Occupational Safety and Health, Spokane, WA, have collaborated with three Western longwall coal mines in an ongoing effort to develop technologies that will aid in providing safe and stable working areas. The goal of the research described here is to develop a stress monitoring system that will provide immediate information to mine managers for making daily safety decisions as areas of poor ground are mined through. Initial work has focused on answering preliminary questions regarding the reliability and use of stress change patterns. Research is concentrated on monitoring horizontal stress because horizontal stress is transmitted over long distances through stiff strata, thus allowing an extended length of entry to be monitored. This paper presents an explanation of the concept, key results from field tests at two mine sites, and a proposed process for imple¬menting a monitoring system. System design layouts, instrument use, data collection and interpretation methods, and processes to present findings to mine staff are described. Additional validation and correlation with actual failure mechanisms are required before this approach can be recommended at ongoing operations. How¬ever, initial results indicate that this approach shows promise in mines prone to bumps and roof falls associated with large stress changes. In particular, analyses of measured changes in horizontal stresses appear useful.

Jan 1, 2002

By Rafal Misa

In order to prevent the formation of mining damage or to reduce its reach, it is advisable to look for technical solutions which would enable effective protection of construction sites from the impacts of underground mining. So far, other authors have mainly focused on developing methods of securing construction sites in order to minimize the damage resulting from underground mining; the issue of geotechnical methods for protecting buildings against mining damage has not been thoroughly covered in those articles. It should be noted that there are relatively few publications directly related to this issue, nor are there practical examples of geotechnical protection. In this paper, a geotechnical solution available for securing building structures is discussed. Trenches presented in this research reduce the size and range of mining deformation. The results of numerical modelling for sample protective solutions, aimed at reducing the impact of mining activity on the terrain surface under various conditions, are presented. The calculations were performed with the aid of the Abaqus 6.10-2 software, based on the Finite Element Method.

Jan 1, 2014

By Xicai Gao

With the intensification and expansion of deep coal mining, the surrounding rock stress environment becomes further deteriorated because of the complicated geological environment and powerful mining effect, so the major deformation of roadway appeared frequently. The proportions of floor heave are increasing and , the most serious problem is the floor heave of preparatory workings. The coal seam of Binchang coal area in Shaanxi province is at a depth of 621.6 ~ 817.0 m, the average thickness of the coal seam is 25 m, and the overburden and coal measure strata are the main regional aquifers and are highly permeable. The properties of water-weakening are obvious in the floor strata. The main preparatory workings or chambers were driven in the coal seam. The severe floor heave of the water chamber appeared during early mining and the amount was 180 ~ 910mm, and seriously affected the normal installation and use of the water chamber. Theoretical study and numerical simulation as well as the in-situ measurements were carried out to investigate the deformation and failure characteristic of the surrounding rock with non-support in the floor and the main reasons which caused the floor heave were determined. A new combined support technology was proposed which consists of overcutting, grouting, backfilling and invertedarch and bolt-grouting. The field monitoring showed that the cumulative convergence of the water chamber was 20mm, the maximum convergence speed is 1.2 ~ 2mm/d, the accumulative convergence of two ribs was 36mm and the maximum convergence speed is 2mm/d. The combined support technology will enhance the support intensity between the roof and floor, the two sides of the entry and improve the loading capacity of the floor effectively and the floor heave can be controlled effectively.

Jan 1, 2013

The U.S. Bureau of Mines (USBM) has developed the coal Mine Roof Rating (CMRR) to aid in ground control. This paper describes the application of the CMRR at four coal mines operated by a major U.S. coal company. The mines are widely separated geographically, and cover a wide variety of geologic conditions and mining methods. The study found that the CMRR was a useful predictor of roof performance. In mines where the CMRR varied, and other parameters held constant, roof falls were associated with low CMRR values. Important similarities in roof performance were found in mines with similar CMRR values, despite radically different geologic environments. The analyses also indicated that statistics can be used to evaluate ground control performance. For these mines, a "roof fall probability" can be estimated from the CMRR and the intersection span. A spreadsheet program has been developed to facilitate treatment of CMRR data, and it is briefly describe

Jan 1, 1994

By Francis S. Kendorski

High-extraction mining of coal and other stratified minerals has a predictable effect on the surface and subsurface. Impacts on ground waters in overlying strata and on surface waters, that are associated with overlying sequential zones of collapse, fracturing, aquiclude, constrained strata, and surface disturbances can be anticipated. Since first quantified around 1979 in Bureau of Mines sponsored contract research work and others (and often adopted by state regulators), with zone extents based upon multiples of mined seam height of 24 to over 60 for the fractured zone, much new data have become available. In examining the sources of data and the methods and intents of the researchers of over 65 case histories, it became apparent that the strata behaviors were being confused with overlapping vertical extents reported for the fractured zones and aquiclude zones depending on whether the researcher was interested in water intrusion into the mine or in water loss from surface or ground waters. These more recent data, and critical examination of existing data, have led to the realization that the former "Aquiclude Zone" defined for its ability to prevent or minimize the intrusion of ground or surface waters into mines has another important character in increasing storage of surface and shallow ground waters in response to mining with no permanent loss of waters. This zone is here named the "Dilated Zone." The existence of the dilated zone explains the ratios of up to 60 times the mined thickness. Surface and ground waters can drain into this zone, but seldom into the mine, and can eventually be recovered through closing of dilations by mine subsidence progression away from the area, or filling of the additional void space created, or both. A revised model has been developed which accommodates the available data, by modifying the zones as follows: ? collapse and disaggregation extending 6 to 10 times the mined thickness above the panel ? continuous fracturing extending approximately 24 times the mined thickness above the panel, allowing temporary drainage of intersected surface and ground waters ? development of a zone of dilated, increased storativity, and leaky strata with little enhanced vertical permeability from 24 to 60 times the mined thickness above the panel above the continuous fracturing zone, and below the constrained or surface effects zones, whose thickness or existence is dictated by the zones above and below ? maintenance of a constrained but leaky zone above the dilated zone and below the surface effects zone, whose thickness or existence is dictated by the zones above and below ? limited surface fracturing in areas of extension extending up to 50 ft or so beneath the ground surface

Jan 1, 1993

By William Blake

Mining induced rock bursts have been a severe operational hazard associated with the deep silver mines of the Coeur d'Alene mining district for over 30 years. Recent studies during the past fm years have led to a better understanding of the different types of rook bursts and their causative mechanisms. This has resulted in ground control measures that more effectively minimize burst occurrences as well as contain burst damage. Case histories of burst incidences and examples of control measures are presented.

Jan 1, 1987

By Andrew Hyett

Over the last 30 years, a transition away from end-anchored bolts to fully grouted rock bolt systems has occurred in many underground mines. It seems geomechanically intuitive that the success of such bolts, which interfere with rock movements around underground excavations through a variety of complex interactions, is precisely why they remain challenging to instrument and monitor. Whereas measuring the load profile for an end-anchored mechanical bolt is trivial, for a fully grouted system, the load may vary dramatically due to discrete rock movement at cracks and joints. Despite the challenges, measuring the strain (i.e., load) along fully grouted rock bolts remains very appealing, and a lot of valuable research has been conducted using foil resistance strain gauges, particularly in US coal mines (Signer and Jones, 1990). Despite good intentions, more widespread adoption of this instrumentation remains unlikely for the following reasons: (i) The technology is expensive, partly because Intrinsically Safe (IS) requirements limit the number of suppliers and solutions. (ii) The technology is perceived as challenging to successfully implement except by those who are skilled in the art. (iii) The reliability of the technology is questionable in harsh environments. (iv) The technology relies on discrete pairs of strain gauges that may not capture the complete strain distribution along the bolt. No existing alternative technology can offer a solution to all of these limitations. However, in this paper, we introduce results from a new Distributed Optical Sensing technology that promises significant potential. Since the technology is capable of distributed sensing, our results suggest that this new technology can, for the first time, reveal the detailed strain profile along fully grouted rock bolts.

Jan 1, 2013

By Fernando O. L. Xavier

"The Barro Branco coal mine was located in the southern region of Brazil, in the state of Santa Catarina. In 2002, an incident occurred at the mine, leading to a violent, massive pillar failure, resulting in almost 100 failed pillars in less than 3 hours, totaling nearly 700 collapsed pillars after a period of 6 weeks. The paper outlines the geological and operational conditions before the accident and analyzes the primary factors leading up to the event. This paper also discusses and examines the unusual characteristic of violent, uncontrolled pillar bursts that occur over a longer timespan, in comparison to other equally violent but rapid events, such as those that occurred at the Crandall Canyon Mine, at a Book Cliffs District mine, and at the Solvay Mine. INTRODUCTION The southern region of Brazil holds 100% of the coal reserves for the country, and half of it is located within the state of Santa Catarina. Almost the entire Brazil’s coal production, around 8 million tonnes/year, is consumed for electricity generation. Most of the coal mines in Santa Catarina are underground room-and-pillar mines, and most are shallow mines, with depths of cover ranging from 20 to 400 m. The Barro Branco mine, near the city of Criciuma, experienced a violent mine collapse at the end of March 2003, which was analyzed and discussed by Zingano, Koppe, and Costa (2004). The event consisted of violent failure of 100 pillars in less than 3 hours, and a subsequent failure of approximately 700 additional pillars over the next 6 weeks. After the panel collapsed, the mine was closed due to unsafe conditions as a consequence of the seismic event. A map of the collapsed area is shown in Figure 1. GEOLOGY AND MINING METHOD There are two coal seams mined at Barro Branco mine. The first is the 1.5 m to 2.3 m thick Barro Branco Seam, under about 75 m of cover. The second is the Bonito Seam, where the collapse occurred. The Bonito Seam is 2.5 to 5.5 m thick located at a depth of approximately 80 m, with an average dip of 0° to 15°. Various major faults and inherent joints are present in the area."

Jan 1, 2017

By Ihsan Berk Tulu, Keith A. Heasley, Chris Mark

"The Analysis of Retreat l\.fining Pillar Stability (ARJ\1PS) and the Lal\.fodel programs have been used successfully in the U.S. for designing safe pillar recovery operations for many years. However, the recent Crandall Canyon l\.fine collapse showed that further research is required to improve the pillar design for recovery under deep cover. To this end, the National Institute for Occupational Safety and Health (NIOSH) and the West Virginia University (WVU) l\.fining Engineering Department have been working together to improve both the ARJ\1PS and the Lal\.fodel programs.This paper compares and analyzes the overburden loads calculated by ARJ\1PS 2002, the new ARJ\1PS 2010, and the Lal\.fodel program with regard to the retreat pillar line, the gob, and the barrier pillars. The analysis shows a number of distinct differences between the ARJ\1PS and Lal\.fodel loading distributions. Ultimately, the programs, with their distinctive load distributions, are used to analyze a database of 52 deep cover pillar retreat case studies, and the ability of each program to discriminate between successful and unsuccessful retreat pillar plans is evaluated.INTRODUCTIONThe Analysis of Retreat l\.fining Pillar Stability (ARJ\.f PS) and the Lal\.fodel programs have been used successfully in the U.S. for designing safe pillar recovery operations for many years. However, the recent Crandall Canyon l\.fine collapse showed that further research is required to improve the pillar design for recovery under deep cover.Pillar recovery accounts for less than 10% of the coal produced from underground coal mines; however, in the period between 1989 and 1996, it accounted for more than 25% of all ground fatalities (l\.fark et al., 2003). In the past 15 years, the safety of retreat mining operations has greatly improved. l\.fark (2009) explains three steps promoted by l\1SHA and NIOSH for safer pillar recovery:1) Global stability through proper pillar design2) Local stability through proper roof support3) Worker safety through proper section managementThe ARJ\1PS and Lal\.fodel (Heasley, 1998) programs have been playing an important role in guiding mine operators in designing pillars to ensure the global stability of a retreat panel."

Jan 1, 2010

By Jun Lu

This paper presents the evaluation, enhancement and comparison of two hydraulic fracturing techniques which were employed to mitigate the impact of a massive sandstone channel on two successive longwall faces. The hydraulic fracturing technique and Longwall Visual Analysis (LVA) software were initially employed to mitigate the impact of thick, massive sandstoneroof during mining of longwall Panel 21. Improved longwall face conditions and productivity were observed as Panel 21 mined through the sandstone channel. However, moderate sandstone fractures and falls at the face were still present in areas where hydraulic fracture influence was not present. Such adverse conditions caused two to three weeks of production delays due to large face cavities, jammed face conveyor and subsequent face gluing. Based on the mining experience in Panel 21, and site-specific roof geology in Panel 22, a modified and more effective hydraulic fracturing program was employed. Fewer well-placed holes were drilled, and larger volume of water was pumped into each hole in order to propagate the individual horizontal pancake fractures over a larger area, thereby enhancing caving and relieving face pressure caused by overhang of the massive sandstone. Longwall Visual Analysis (LVA) software was also employed at Panel 22 to predict the cavity risk index and guide the longwall face crew in double-pulling shields behind the shearer drum. This helped to reduce the tip-to-face distance at the critical spots along the longwall face, which helped mitigate the formation of the roof cavities and failure of massive sandstone onto the face conveyor. Underground observations, face advance rate and face rock delay analysis confirmed that longwall face conditions and advance rate in Panel 22were significantly improved compared to those of Panel 21. Consequently, safety and productivity were improved.

Jan 1, 2014

By Takashi Sasaoka

A great deal of coal will be left under the final end-walls in Chinese surface coal mines because of lower slope angle in shoveltruck mining systems and larger mining depths. In traditional surface mining systems, the coal under the end-walls will be buried by the inner dumping site, and these coal resources are generally abandoned. Considering these situations, the application of end-wall mining systems was considered in order to recover the residual coal around the end-wall. This mining system can improve economic benefits by exploiting haulage and ventilation roadways from end-walls and utilizing the existing transportation systems. In order to develop the appropriate pillar design methodology for the end-wall mining system, several empirical formulas and methods were discussed based on the formulas developed so far and the mechanics properties of coal and rocks in the Chinese coal mine. From the results of a series of theoretical and numerical analyses, it can be concluded that the coal pillar width calculated by the Mark- Bieniawski formula may serve as an important reference for pillar design in end-wall mining systems, and the end-wall mining system can increase the coal recovery from residual coal resources around highwall and reduce the effect of underground mining operation on slope stability

Jan 1, 2013

By Chase K. Barnard, Rahul Thareja, Sean Warren, Rajagopala Kallu

"Inflatable rock bolts are commonly utilized for ground support in Nevada underground mines, however, there is limited information regarding what factors influence the bond strength of these bolts. Bond strength is an important parameter in friction bolt support design, and this information is usually not available from manufacturers as a standard. Previous research has investigated the bond strength of friction bolts; however, these studies focused primarily on split set bolts and ground conditions more favorable compared to those commonly encountered in Nevada underground mines. Without site-specific bolt pull-out tests, ground control engineers are required to make assumptions regarding the in-situ bond strength of rock bolts. This is especially a concern for projects in the feasibility and initial development stages before site-specific experience is acquired.The purpose of this paper is to establish confidence in anticipated minimum bond strength for inflatable rock bolts by comparing the bond strength to variable geotechnical conditions using the Rock Mass Rating (RMR) system. To investigate a correlation between these parameters, the minimum bond strength of pull-out tested inflatable rock bolts was compared to the RMR of the rock in which these bolts were placed. Bond strength vs. RMR plots indicate that expected minimum bond strength is positively correlated with RMR, however the correlation is not strong. Cumulative percent graphs indicate that 97% of pull-out tests result in a minimum bond strength of 1 ton/ft and ½ ton/ft in RMR =45 and <45, respectively.Although lower bond strengths are more commonly encountered in low RMR ground, high bond strengths are possible as well, yielding higher variability in bond strengths in low RMR ground. Bond strength of friction bolts relies on contact between the rock bolt and drill hole. Experience in Nevada indicates that RMR is known to affect both the quality and consistency of drill holes which likely affects bond strength. Drilling and bolting in low RMR ground is more sensitive to drilling and bolting practices, and strategies for maximizing bond strength in these conditions are discussed."

Jan 1, 2015

By Zhengdong Liu, Kaizhong Zhang, Zhaonan Jiang, Yuanping Cheng, Biao Hu, Qingquan Liu, Ranpeng Wang

"Major coal mining areas in eastern China have entered into the deep mining state. Gas pre-drainage, which has been successful at shallower depths, is not suitable at greater depths because of complex conditions. Coal permeability, which is closely related to the damage behavior of coal, is the most important parameter in determining gas migration and gas drainage in a coal seam. To achieve an optimal gas drainage design in a deep mining state, the damage and permeability evolutions of coal should be further studied. The progressive failure process of coal has been analyzed by conducting a series of triaxial loading and unloading tests. The development of coal fractures has been successfully quantified, and the corresponding characteristic stress points have been obtained by using Martin’s fracture model and the stress-strain curve. Coal will produce plastic deformation, and its permeability will rapidly increase when the stress level reaches damage expansion stress. INTRODUCTION Coal mine methane (CMM) is a general term for all methane released during and after mining operations. CMM is both a potentially valuable energy and a serious hazard in active coal mines. Degassing coal seams is important for mitigating this hazard and results in the beneficial recovery of a cleanburning, low-carbon fuel resource (Karacan, Ruiz, Cotè, and Phipps, 2011). The permeability of coal is a key parameter in determining CMM production, coalbed methane (CBM) production, CO2-enhanced coalbed methane (CO2-ECBM) production, and CO2 storage in coal seam reservoirs (Connell, Lu, and Pan, 2010; Chen et al., 2011; Perera, Ranjith, and Choi, 2013). Significant research has been condcuted to the study of the evolution of coal permeability. These studies revealed that coal permeability evolution is controlled by the competing influences of effective stress and sorption-induced coal deformation. Chen et al. (2011) investigated the influence of an effective stress coefficient and sorption-induced strain on the evolution of coal permeability. Robertson and his team (Robertson, 2005; Robertson and Christiansen, 2005) measured sorption-induced matrix strain and permeability changes with regard to CO2, CH4, and N2. They found that the compressibility is not constant in a natural coal sample. Karacan (2007) conducted a series of experimental tests to study the swelling-induced volumetric strains internal to a stressed coal associated with CO2 sorption. Liu et al. (2016) investigated the effects of matrix swelling area propagation on the permeability evolution by using both natural and reconstituted samples. Many coal permeability models have been developed to describe the permeability evolution under uniaxial loading, biaxial strain, hydrostatic confining pressure, triaxial strain, and triaxial stress. Recently, some analytical permeability models have been developed to describe the anisotropic permeability behavior (Palmer, 2009; Pan and Connell, 2011; Robertson and Christiansen, 2005; Wang, Zang, Wang, and Zhou, 2014; Chen, Pan, Liu, and Connell, 2012)."

Jan 1, 2017

By WenFeng Li

A 3D roof bolt model with the consideration of rebar, resin, bearing plate, and resin/rock interface was developed for studying the bolt/rock interactions of the tensioned and fully-grouted resin bolts. For the tensioned bolts, 2 ft compressive zone above the rooflineis generated by the 10 tons pretension. Since the free portion of the tensioned bolts displays a higher vertical stress, the increase of the bearing capacity of the rocks in such range is much larger. For the fully grouted resin bolts, despite no pretension is generated during installation, the roof sag can indirectly generate a compressive zone near the roof line. In addition, the analysis on the differences in bolt/rock interactions when employing the tensioned or fully-grouted resin bolts indicates that the fully-grouted resin bolts are preferred as the primary support when weak and thinly-bedded immediate-roof is present. Meanwhile, it is found that both the tensioned and fully-grouted resin bolts may have very limited effect in preventing and/or control the cutter roof.

Jan 1, 2014

By John C. Stankus

The 3 Right pre-developed longwall recovery entry at Signal Peak Energy?s Bull Mountain No. 1 mine was designed to prevent the abnormal roof conditions encountered in its firsttwo recovery entries. Unique geological conditions, such as the low overburden of approximately 200 feet and prevalence of roof joints in the sandstone overburden, contribute significantly to abnormal conditions during longwall recovery. The 3 Right recovery entry was 42 ft wide and developed in two stages, 21 ft wide in each stage. The entry roof was reinforced with steel wire mesh/non-metallic recovery mesh, high-capacity primary bolts and cables. After all the support was installed, the entry was completely backfilled with specially designed, cuttable, low-density cement. The roof bolting and backfilldesign were based on current practices in this mine and successful practices in other mines in the U.S. and Australia. Extensive instrumentation was installed in the recovery entry for continuous monitoring of load transfer as the face approached and cut into the recovery entries. The support design was completely successful with a smooth cut-through on January 29 and 30, 2014, significantly reducing the shield recovery time and meanwhile protecting the safety of miners. Based on actual observance and practices of the 3 Right recovery entry, recommendations on improvement and modificationsof roof support for 4 Right recovery room are presented.

Jan 1, 2014

By Ihsan Berk Tulu

The complex interaction between multiple seam pillar extraction panels and variable topography can generate unexpected ground control problems. This case study describes stress-related damage in multiple seam workings that resulted in difficult operating conditions and compromised the safety of the support crews. The damage appeared to be caused by horizontal stress but the unexpected location of the damage baffled the authors. A 2D Finite element analysis was conducted of the mine layout and surface topography in an attempt to explain the observed failure. It was discovered that the topography resulted in the rotation of the major stress at the location of the current workings. The stress rotation caused asymmetrical interactions between the workings on the different coal beds, which can explain the unexpected response observed in the mine. Consequently the layout of the next panel was changed by locating the vulnerable entry 30 ft away from its originally planned position. The new location was selected to locate the entry in favorable stress conditions. The change in layout resulted in a significant improvement in conditions when the next panel was mined. At the time of writing this paper, a full panel has been developed using the modified layout without undue ground control problems. The case study demonstrates that numerical models can help engineers to interpret and solve complex multi-seam interaction problems.

Jan 1, 2014