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By Bo-Hyun Kim, Steve Berry, Mark K. Larson

"When coal pillar bursts occur, they involve the sudden expulsion of coal and rock into the mine opening. These events occur when relatively high stresses in a coal pillar, which serve as support in underground workings, exceed the critical capability of the pillar. This causes the pillar to rupture without warning. A cleating is a type of naturally occurring joint system unique to coal seams. However, cleatings are pervasive throughout the coal seam and are much more tightly spaced than joints occurring in other rock types. In this study, researchers from the National Institute for Occupational Safety and Health (NIOSH) investigated 84 coal samples obtained from a Utah coal mine by conducting both unconfined and triaxial compressive tests in order to examine the response to stress as a function of both confinement and orientation between cleat and loading direction. The results of the study show that the confining pressure dictated not only the peak compressive strength but also the brittleness as a function of the major-to-minor principal stress ratio. The spalling limit of the coal estimated by the s-shaped brittle failure criterion appears to be very dependent on the angle between the cleat and the loading direction, particularly at lower confinement.INTRODUCTIONCoal pillar bursts, or bumps, remain a most critical safety concern in underground coal mining, despite significant advancements in engineering controls. As shown in Figure 1, between 1983 and 2014, there were nearly 400 cases of reportable dynamic failure accidents in coal and nonmetal mines, resulting in 20 fatalities, 155 lost-time accidents, and an estimated 48,000 lost man hours (MSHA, 2015). These types of events have been documented for well over 100 years within the American underground coal mining industry. During this time, mining practices and support technologies have evolved considerably, resulting in an overall decrease in the rate of dynamic failure-related injuries and fatalities. Although techniques and practices of mining are highly advanced, coal pillar bursts, or bumps, continue to occur (Mark and Gauna, 2015). If coal is expected to fail in a brittle manner, behavior changes associated with relatively low tensile strength, such as the transition from tensile to shear failure, have to be considered and reflected in the adopted failure criteria. Rock failure in tension takes place at low confinement around excavations due to tensile failure in heterogeneous rocks (Diederichs, 2007). The prospect of tensile-dominant brittle failure diminishes as the confinement increases away from the excavation boundary. Therefore, it is expected that the transition from tensile to shear mode in the failure mechanism occurs as the confinement level changes and conditions for tensile failure are prevented or strongly diminished. In addition, many uncertainties remain because of the highly anisotropic characteristics of coal seams associated with geologic structure and the mining-induced spatial redistribution of stress in coal pillars. Thus, to prevent fatalities in underground coal mining, continuous effort is required to better understand the catastrophic failure mechanisms in coal mines."

Jan 1, 2018

By Stephen C. Tadolini, Anand Bhagwat

"Intrinsic supports installed in mines with limited seam or mining heights have always proven to be difficult when the required support length is longer. To date, three solutions have been used to ensure that the correct length bolt is installed to achieve the required anchorage or to maintain immediate roof material: combination bolts, milled-notch bolts, and hot-notch bolts. The combination bolts use threaded sections and couples, which can be screwed together to form the required length. The primary bolt of choice, to minimize costs and achieve the desired support, has been milled- or hot-notch rebar bolts. These notches can be placed at strategic (required) locations along the bolt axis, so the bolt can be successfully bent and straightened during the installation process. However, these notches can reduce the strength of the bolting system; special reductions are detailed in ASTM F432-13 (2013). Technical advances in steel rolling and forged heading processes have led to the development of a deformed rebar bolt termed, the BORE bolt, which stands for Bendable Oval REbar. The BORE bolt can be bent at any location and straightened during the bolt installation cycle. This support exceeds the minimum designed capacities described in ASTM F432-13 and meets the yield and tensile strength requirements of a full diameter no. 6 rebar bolt. The profile and shape also reduces the resin annulus during the installation process to enhance mixing and shreds the cartridge film to help eliminate glove fingering.INTRODUCTIONThe height or depth of competent strata material, in the immediate mine roof, can be greater than the developed mine openings heights. This is especially true in coal mine openings where 50% of the minable seams in the US are less than 50 inches (132 cm). When this scenario occurs, it can be difficult to insert steel bolts in a solid single length. Figure 1 illustrates what difficulties can be experienced when a roof bolter operator is working and supporting a thin seam coal mine. There are three viable options currently available. Lengths of the bolt can be threaded and a steel coupler can be used between the sections to reach the required depth of anchorage. These systems, called “coupled bolts,” usually require a larger diameter hole to facilitate the larger coupler diameter and can be anchored with an expansion shell, resin, or a combination of both. To ensure that the coupler will develop the tensile capacity of the bolting system, required by ASTM 432-13 Standard Specifications for Roof and Rock Bolts and Accessories (2013), the threaded bars should be completely rotated and engaged into the coupler. Premature failures can occur when the threads “strip out” if not inserted completely."

Jan 1, 2018

By Dezhong Kong, Jaichen Wang, Shengli Yang

"The stability control of longwall coal face is the key technology of large-cutting-height mining method. Therefore, a systematic study of the factors that affect coal face stability and its control technology is required in the development of large-cutting-height mining method in China. After the practical field observation and years of study, it was found that the more than 95% of failure in coal face are shear failure. The shear failure analysis model of coal face has been established, that can perform systematic study among factors such as mining height, coal mass strength, roof load, support resistance, and face flipper protecting plate horizontal force. Meanwhile, sensitivity analysis of factors influencing coal face stability showed that improving support capacity, cohesion of coal mass and decreasing roof load of coal face are the key to improve coal face stability. Numerical simulation of the factors affecting coal face stability has been performed using UDEC software and the results are consistent with the theoretical analysis. The coal face reinforcement technology of largecutting- height mining method using the grouting combined with coir rope is presented. Laboratory tests have been carried out to verify its reinforcement effect and practical application has been implemented in several coal mines with good results. It has now become the main technology to reduce longwall coal face failure of large-cutting-height mining method.INTRODUCTIONLarge-cutting-height mining method of more than 3.5m mining height, is one of the main methods for thick-seam mining in China. Currently, the largest mining height is 7.3m and the support capacity is 1,800mt. Shendong and Jincheng mining area have the best use of large-cutting-height mining technology (Wang, 2009). Generally, the mining height ranges from 6.5m to 7m and the support capacity ranges from 1,000mt to 2,000mt, annual output of the panel is between 10 and 12 million tons and the highest are 15 million tons. The main goal of large-cutting-height mining technology is to achieve high yield and high efficiency by increasing the advancing speed of longwall face (Yuan, 2011- 2012). However, in certain geologic conditions, rib spalling and roof falls in the unsupported area often occur. Because of the large tonnage of large-cutting-height mining equipment, once the rib spalling and roof falls cause the equipment to be out of alignment, the face advancing speed could be affected seriously that in turn affects the output (Yang, 2012). Therefore, studying the failure mechanism and control technique of coal face in large-cuttingheight mining method is of great significance."

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 Anton Sroka, Antoni Tajdus, Axel Preusse, Krzysztof Tajdus

"In Europe, the magnitude and range of mining deformation predicted on the surface have been, in most cases, determined using methods based on Gauss' influence function (e.g., Knothe's theory, Ruhrkohle method, Stochastic method). The increase in the application of numerical methods (finite element method FEM, finite difference method FDM, etc.) provides an opportunity to offer a different approach in describing mining deformations. However, this phenomenon is very complex, and due to numerical limitations, a two dimensional analysis was made (in plane strain) in most cases. Such 2D models could not accurately describe the mining geometry and geological conditions. The authors of this paper carried out the calculations in 3D models which describe a mining situation and the character of the strata in one Upper Silesian Polish mine. Furthermore, the authors include some tips for three-dimensional modeling.INTRODUCTIONThrough the years, many numerical models have been used to describe a phenomenon of underground mining exploitation influence on the ground surface. However, such rock masses are highly disturbed (cracks, fractures, changes in water conditions, etc.), and it is very difficult to suitably describe, in a mathematical manner, such materials. Therefore, many of the numerical models created have been corrected in various degrees.At first, models widely used to describe the behavior of strata in the region of exploitation appeared as elastic models, where scientests assumed that a whole strata behaved in a linearly elastic manner, according to Hook's law. Those models have been modifed to received suitable VALUES (of settlement and range of subsidence troughs. Scientests created this amongst others: elastic models with contact between the rocks layers (Shippam, 1975; Salamon, Chugh and Yang, 1993), elastic models with no tension (Chrzanowska-Szostak, 1988), and elastic models with a contact on the subsidence limit plane (Tajdus K., Tajdus A., 2005). However, in most cases, those models did not bring the expected results and the differences between shapes of measured and the calculated subsidence trough were substantial."

Jan 1, 2010

By Michael Gauna, Christopher Mark

"Coal bursts involve the sudden, violent ejection of coal or rock into the mine workings. They are almost always accompanied by a loud noise, like an explosion, and ground vibration. Bursts are a particular hazard for miners because they typically occur without warning. Despite decades of research, the sources and mechanics of these events are not well understood, and therefore they are difficult to predict and control. Experience has shown, however, that certain geologic and mining factors are associated with an increased likelihood of a coal burst. A coal burst risk assessment consists of evaluating the degree to which these risk factors are present, and then identifying appropriate control measures to mitigate the hazard. This paper summarizes the U.S. and international experience with coal bursts, and describes the known risk factors in detail. It includes a framework that can be used to guide the risk assessment process.BACKGROUNDCoal bursts involve the sudden, violent ejection of coal or rock into the mine workings. They are almost always accompanied by a loud noise, like an explosion, and ground vibration. Bursts are a particular hazard for miners because they typically occur without warning. During the years 2012-2014, serious coal bursts occurred at three different U.S. room and pillar mines. These events resulted in three fatalities and two permanently disabling injuries (MSHA, 2014; MSHA, 2013). In all three instances, the events occurred during pillar recovery at depths exceeding 1,000 feet (300 m). None of these three mines had previously reported a burst to MSHA. Coal bursts also occurred at three longwall mines during this same time period.Despite decades of research, the sources and mechanics of bursts are not well understood, and therefore these events are difficult to predict and control. Experience has shown, however, that certain risk factors are associated with an increased likelihood of a coal burst. A coal burst risk assessment consists of evaluating the degree to which these risk factors are present. In addition, some control techniques are effective in reducing the likelihood of an event or protecting miners from their effects."

Jan 1, 2015

By Rob G. Jeffrey, Ken W. Mills, Jesse W. Puller

"A coal mine in New South Wales is longwall mining 300m wide panels at a depth of 160-180m directly below a 16-20 m thick conglomerate strata. As part of a strategy to use hydraulic fracturing to manage potential windblast and periodic caving hazards associated with this conglomerate strata, the in situ stresses in the conglomerate were measured using ANZI strain cells and the overcoring method of stress relief. Changes in stress associated with abutment loading and placement of hydraulic fractures were also measured using ANZI strain cells installed from the surface and from underground. This paper presents the results of in situ stress measurements and stress change monitoring undertaken as part of the hydraulic fracturing program.Overcore stress measurements have indicated the vertical stress is the lowest principal stress so that hydraulic fractures placed ahead of mining form horizontally and so provide effective pre-conditioning to promote caving of the conglomerate strata. Monitoring of the full-scale hydraulic fractures has confirmed the hydraulic fractures grow horizontally.Stress changes monitored during hydraulic fracturing formed part of a broader program to investigate fracture geometry and growth rates to confirm the characteristics of the hydraulic fractures placed. This monitoring showed that vertical stress in the conglomerate strata was being increased by up to 1 MPa in close proximity to the hydraulic fracture and had the effect of tending to promote asymmetrical growth of subsequent fractures about the injection borehole.Monitoring of stress changes in the overburden strata during longwall retreat was undertaken at two different locations at the mine. The monitoring indicated stress changes were evident 150 m ahead of the longwall face and abutment loading reached a maximum increase of about 7.5 MPa. The stresses ahead of mining change gradually with distance to the approaching longwall and in a direction consistent with the horizontal in situ stresses. There was no evidence in the stress change monitoring results to indicate significant cyclical forward abutment loading ahead of the face consistent with the observations of face conditions underground adjacent to both measurement locations. The forward abutment load determined from the stress change monitoring is consistent with the weight of overburden strata overhanging the goaf indicated by subsidence monitoring."

Jan 1, 2015

By David W. Evans

"Induced hydraulic fracture of strata during roof bolt installation is a potentially prevalent, but masked phenomenon within the underground coal industry. Previously reported resin testing programs (McTyer et al., 2014) examined the relationship between resin mixing effectiveness and varying bore hole diameter. The methodology employed within this earlier test program facilitated a further critical area of research – the measurement of back pressures generated within the bore hole during standard rock bolt installation practices. Experimental data has indicated that fluid resin can be pressurised to levels where it exceeds the compressive strength of the strata, inducing hydraulic fracture within the immediate area of the bolting horizon. The routine cycle of roof bolting serves to propagate this effect, progressively fracturing and delaminating the roof during mine advancement. This masked phenomenon can lead to a perception of difficult ground conditions - mining efficiencies and costs are therefore affected, with increased needs for additional support subsequently required to restabilise the inadvertently damaged roof.Further analysis of the parameters associated with resin bolt installation has now been conducted, assisting in the development of an empirical relationship between bore hole pressure, bore hole diameter and bolt insertion times. This relationship has been analysed for 15:1 ratio resins and 2:1 ratioresins, within 28mm and 30mm boreholes. Further to this, load transfer performance has been comparatively assessed for both 28mm and 30mm boreholes, suggesting that for 2:1 resins, acceptable resin mixing and load transfer can be obtained within a 30mm bore hole. The combination of 2:1 resins, utilised within a 30mm bore hole, may well provide the optimal solution to reduce the risk of hydraulic fracture in weaker strata during resin bolt installation.INTRODUCTIONA growing number of industry papers have previously investigated areas of concern associated with the performance of cartridge style resins in roof bolting, predominantly focussing on the effects of plastic film gloving, inadequate resin mixing and the pressurisation of resin within the bore hole. These three effects have a critical influence on the load transfer of the steel bolt element, through the cured resin and into the surrounding strata.Gloving occurs due to the plastic film of the resin cartridge partially encasing or wrapping around the steel roof bolt element – a known phenomenon over many years (Pettibone, 1987). The plastic film creates regions of discontinuity between the bolt, cured resin and borehole, reducing effective load transfer into the strata. Experimentation conducted into the effects of bolt ends fully encased by plastic film (Pastars and MacGregor, 2005) revealed that load transfer can be reduced by 85 to 90% in worstcase occurrences. It is also known that the aggregate filler size used within the resin can assist in shredding and breaking up the plastic film – this effect was observed in a previous study on resin mixing effectiveness (McTyer et al., 2014)."

Jan 1, 2015

By Brian Daniels, James Farroni

"Draw rock (small rock that forms/falls when bolting the roof and/or rib) is the cause of numerous injuries annually. These injuries result in lost work hours and low morale. The report below explains how ""Mine Safe Draw Rock Shields"" (Pans) are used in an underground operation. This report will chronicle the application of these pans in a Southern West Virginia coal mine along with testing data that measures their effectiveness. Since the introduction of these pans in 2002 the amount of reportable accidents relating to draw rock has decreased by 80%. The report below reinforces that perimeter contact is critical to the performance of these pans. In addition the strength of these products at the perimeter will determine their effectiveness. The results of the testing data outlined in this paper, along with the physical records at the study mine, confirm that the use of pans has significantly reduced draw rock falls at this facility.BACKGROUNDSupplemental support is defined as ground control that is used in addition to required support in an underground mine. Some examples of supplemental support are:• Pans• Mesh/Geogrid• Mats• Half Headers• Boards• ChannelA typical roof condition that warrants the use of pans is when loose small rock is dislodged while bolting. If the top is very scaly (loose, thin rock) additional supplemental support can be used in conjunction with pans such as mesh or geogrid (See Figure 1). Pans will provide protection where used and approximately 12 in (250 mm) out from perimeter. This is achieved by contact with a surface which will put scale in tension up to 12 in (250 mm) from perimeter of pan.INTRODUCTIONThe ease of applying a roof support system is critical to effectively operating a deep mine. Figure 2 shows how a pan is installed in conjunction with a roof bolt and bearing plate. The Mine Safety and Health Administration (MSHA) requires that supplemental support cannot decrease the performance of required support. The goal of using all supplemental support is to provide additional safety for the miner which equates to increased productivity.The ability of skin control products to contact a surface is the single most critical characteristic (Figure 3 and Figure 4). If the material does not contact the surface it cannot be effective. Therefore, the flexibility, design and proper installation of the products are critical in determining their effectiveness. Figure 5 shows the proper installation of the roof pan.To reiterate, all skin control must be applied in conjunction with a roof bolt and ASTM approved bearing plate as shown in Figure 4."

Jan 1, 2010

By Fangtian Wang

The fully mechanized longwall mining was employed in a shallow coal seam that was under the gobs of room and pillar mining and overlain by thin bedrock and thick loose sands. Operational practices have demonstrated that severe accidents would occur including roof structure instability, roof steps, damages of shield supports, and face bumps triggered by the large area roof weighting, resulting in serious threats to the safety of underground miners and equipment. This paper uses Shigetai Mine, located in the Ordos area, Inner Mongolia, as the study site. It fully considers the geological conditions of 3-1-2 coal seam and employs theoretical analysis, 3DEC numerical simulation, and field measurements to analyze the overlying strata movement for the shallow seams under an abandoned room and pillar mine without pillar extraction. The research results serve as an excellent guide to mining in similar geological and mining conditions.

Jan 1, 2014

By Kyle A. Perry

Pillar design formulae have existed for decades, and the guidelines for design stability factors are generally accepted by regulatory agencies and broadly used within the coal industry. Nearly all design pillar formulae are based on geometrical parameters of pillars resulting from empirical investigations, with primary focus on the pillar width-to-height (W/H) ratio and associated shaping constants from regression. The most widely used modern pillar design formula in the coal industry, the Mark- Bieniawski equation, also adheres to this principle. Common practice when applying Mark-Bieniawski equation is to use a coal compressive strength of 900 psi. This reduces the design equation to sole dependence upon geometry. However, recent studies have shown that pillar strength is highly dependent on roof and floor properties, particularly the interfaces between the roof-pillar and pillar-floor. This study utilized numerical modeling using FLAC3D to evaluate the relative effects of varying roof and floor interface and mechanical properties. The results confirm that pillar strength is significantly influenced by the properties of the coal pillar interfaces with roof and floor. The mechanical properties of the roof and floor material have only a minor impact on pillar strength when compared with interface effects. These results were used to formulate design recommendations for existing pillar design methods.

Jan 1, 2013

By Tom Blevins, Roland Walker, Jinsheng Chen

"The Dywidag-Systems International DSI Tandem Anchor bolt, with an improved expansion shell-washer design, can substantially reduce potential entry roof falls caused by twisted or deformed expansion shells of the traditional mechanical anchor bolts during installation. In addition, the improved anchorage system can also ensure that a consistent and high installed load is achieved, which is critical for effective beam building under thinly laminated or layered roof conditions. This paper discusses the genesis and anchorage performance of the Tandem Anchor bolt, and analyzes the tension/torque relationship of different mechanical shell anchors using data obtained from both lab and field tests. It also discusses roof bolting mechanisms, roof bolting systems, and factors influencing effective beam building. Finally, numerical modeling was used to analyze the effect of bolt pretension magnitude on entry roof stability.INTRODUCTIONThroughout the history of underground mining, keeping the roof from uncontrollable collapse has been the greatest concern of the miner. Tremendous strides have been made in better understanding, predicting, and supporting the mine roof. This conference has spent the last twenty-eight years sharing information and distributing it to the mining community. A paraphrased statement from former Secretary of Defense Donald Rumsfeld adequately describes the challenges to the mining engineer as it pertains to roof control: it is a study of known knowns, known unknowns, and unknown unknowns. It is the hope of the authors to address some of these unknowns.The paper discusses the development of bolt anchorage systems from dry shells to shells with resin-assisted anchorage. It looks at some of the issues faced with the anchorage shells being used when introducing the insertion pressures caused by resin cartridges. The rationale behind the shell design changes was to compensate for this resin induced stresses.An improved expansion shell assembly (a slotted expansion shell base, a tapered cramming plug, and a novel support nut with tapered top) for mine roof bolts makes up the Tandem Anchor (U.S. Pat. No. 6742966); this minimizes the possibility of twisting seen with traditional mechanical shells and thus prevents springy bolts and shell leaves from breaking. A fully grouted Tandem Anchor bolting system has the advantages of both the pure mechanical point-anchor and the non-pretention fully grouted bolting systems. Once installed, the installed load or pretension will be locked in and never bleed off. Compared with the fully-grouted (2-speed resin) tension rebar bolting system, the fully-grouted Tandem Anchor bolting system can (a) reduce bolting cycle time (about 15 seconds per bolt), boosting coal production and productivity; (b) provide a greater installed load or pretention; and (c) achieve a consistent tension/torque ratio, which gives better bolting quality control. Therefore, a fully grouted Tandem Anchor bolting system is the most effective means to reinforce a weak rock mass intrinsically by the theory of beam building. For thinly laminated or layered roof conditions, a higher installed load or pretention (50% or more of the bolt's yield strength) can make a roof bolting system much stiffer, which results in less roof movement. Numerical modeling results show that the magnitude of the roof's tensile stress and relaxation zone can be substantially reduced under a higher installed load or pretention."

Jan 1, 2010

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 Soheil Bahrampour

Development of measurement while drilling (MWD) systems for ground control applications in mining applications has recently been an active research topic. The main goal is to utilize the drilling information for characterizing the ground condition. Such characterization is the key to the accurate and efficient mapping of hazards and proper planning of ground support. However, most of the existing drilling units and measurement while drilling systems are mainly developed for joints and voids detection, and much less research is done to estimate rock strength. This paper focuses on rock classification based on estimated strength from the drilling data. For this purpose, data from J.H. Fletcher roof bolters instrumented with the vibration and acoustic sensors as well as torque, thrust, penetration rate, RPM and position signals is used. All the available data was utilized for rock classification using pattern recognition algorithms. The algorithms were developed based on data from drilling, and into three different rock types. The classification is performed using a simple feature extraction algorithm with a well-known pattern recognition algorithm. The results demonstrate the suitability of the proposed algorithms in identifying the rock types based on their strength properties, which can be adopted in measurement while drilling systems.

Jan 1, 2014

By Gennaro G. Marino

Site conditions at several shallow room and pillar mines in Illinois are described and compared with the charac¬teristics of the subsidence profiles at the ground surface. The shape and magnitude of the subsidence profiles were found to be closely related to the thickness of the soil and rock over-burden the percent extraction of the coal, and the shape of mine pillars. The room and pillar mines were located at a shallow depth (from 112 to 220 ft), beneath flat to gently rolling terrain. Thickness of the coal seams ranged from 5 to 9 ft and at all sites was underlain by underclays. This information has been developed in order to improve techniques for evaluating the subsidence potential at sites within the Illinois coal basin. he results presented form a small portion of the work carried out by the University of Illinois with Illinois Abandoned Mined Lands Councils and the U. S. Bureau of Mines at the Minneapolis Research Center in Minnesota (Marino and Mahar, 1985). Presented in the paper are the modes of mine failures observed for Illinois room and pillar mines, an overall summary of the site geologic and mining condi¬tions, case histories of each mode of failure, a comparison of sag profiles, and then an empirical relationship between the site conditions and the maximum sag subsidence. The most ideal case histories are presented to illustrate fundamental behavior. A brief of the Paper is given in a summary and conclusions section.

Jan 1, 1984

By Ximeng Li

The entries that are subjected to dynamic pressure are much more difficult to maintain due to repeated disturbance. A case study of an entry stress distribution was performed using the FLAC3D program. By analyzing the distribution of stress field around the entries, this paper provides guidelines for entry support design in the lower (#16) seam subjected to the repeated disturbance from longwall mining in the immediate overlying (#17) seam, 72ft (22 m) above. The results showed that there was stress superposition inside the barrier pillar in the lower seam, 15ft (5m) to 45ft (15m) deep, during development. The influence zone of the front abutment pressure in the upper (#16) seam was about 100ft (30m) outby the faceline. The stress-relieved zone was from 0 to 100ft (30m) inby the face. The area further inby the face (i.e., 230 ft (70 m) inby the face) was within the gob compaction zone. Along the face advancing direction, the affected zones of the entries in the lower (#17) seam that were subjected to the dynamic pressure from longwall mining of the immediate overlying seam could be divided into four stages: pillar development stress stage, front abutment pressure stage, unstable gob stage, and stable gob stage. Those entries within an area from 100ft (30m) outby to 100ft (30m) to 230ft (70m) inby the faceline should be reinforced with special supports. The numerical modelling matched the monitoring results well.

Jan 1, 2013

By D. F. Howarth

Two innovative ground reinforcement products have been developed for use in the mining industry. One of these products is a high strength polyethylene (HOPE) bar for use as a rib bolt in underground coal mining and as a corrosion proof tendon in open pit mines. The other product is a cable bolt (Ultrastrand) which demonstrates enhanced load transfer and is available in continuous lengths. Ultrastrand's exceptional in¬situ performance and versatility makes it ideal for mechanised installation in both underground and surface mining operations. Details of these products are presented and results from labortaory and field trials are discussed.

Jan 1, 1992

By Richard A. Allwes

A procedure is presented for the design of an arch canopy for the application of rehabilitating a high-roof-fall area. This procedure is specifically developed for evaluating an unbackfllled arch canopy, which Is subjected to an impact load (roof fall). Dynamic line loading of the arch canopy at its crown is only considered and the procedure does not account for an unsymmetrical loading condition. The evaluation of whether an arch canopy is suitable for a particular installation is dependent upon many variables, including in-nine conditions and the engineering properties of the structure. However, a general evaluation can be based upon a comparison of the arch's maximum crown deflection and a prescribed allowable crown deflection. A single spring-mass system is used as an idealized model to represent the dynamic behavior of the arch canopy at its crown. An energy approach is taken to determine the maximum crown deflection of an arch rather than solving a differential equation of notion derived from dynamic equilibrium. The design procedure is based on the concept that when an arch canopy is subjected to impact loading at its crown and deflects from its unloaded state to maximum crown deflection, the structure absorbs strain energy, both elastic and plastic. As a result, this strain energy can be calculated from a static load-displacement diagram for the structure. The significance of this design procedure Is that It gives mine personnel an analytical tool to select an arch canopy to meet the dimensional and functional requirements of a mine entry and a prescribed allowable crown deflection.

Jan 1, 1986

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 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