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By Boqiang Wu, Ganggang Xu, Xiaodong Wang, Hai Wang, Shibin Zhu

"Using a paste-like slurry as a filling from the ground surface to the goaf, artificial pillars can be built in the cavity goaf to support the goaf roof to prevent roof collapse and ground subsidence. If the accumulation laws of paste-like slurry are known well, the filling technology can be optimized. The laws can establish the foundation for finding out the best methods for building artificial pillars and filling technology. In order to understand the accumulation laws of paste-like slurry in the cavity goaf, Aeolian sand with wide distribution in northwestern China was selected as aggregate, and Portland cement was used as cementitious material to prepare a paste-like slurry. A test model of the cavity goaf was built to simulate the filling process by pump from the ground surface to the goaf to study accumulation laws of the paste-like slurry. The results show that the whole process of expansion and accumulation of the paste-like slurry could be divided into stages: free expansion, fill tightening expansion, and stable expansion. The angle of repose of the accumulation body of the paste-like slurry is constant in the stage of stable expansion, and it is the final angle of repose of the accumulation body. The angle of repose of the accumulation body is highly relevant to the slump of the paste-like slurry, and it is unrelated to the distance between the floor and roof. Under certain conditions of the cavity goaf, reducing the slump of paste-like slurry is one of the effective ways to obtain cost-effective pillars.INTRODUCTIONA large number of cavity goafs of coal mines were left in northwestern China; they were formed in the mining process by room-pillar or strip mining. Safety coal pillars were used to support the roof to keep long-term stability in the goaf, so the ground subsidence was not obvious (Fu, Deng, and Zhang, 2011; Xie, 2014). With gradual weathering, destruction of coal pillars, or additional load from ground surface buildings and structures, there could be a risk of instantaneous collapse of the roof in the goaf. This would affect the safety of existing structures on the ground and adjacent mines (Qu et al., 2017; Li, Xiang, and Jia, 2011; Tan, Guo, and Zhao, 2016; Tan et al., 2017). For disaster prevention of the falling cavity goaf, the current common practice is full filling (Hu et al., 2008; Liu, Zhu, and Huang, 2015; Pan and Wu, 2010). However, for buildings and structures with lower deformation requirements, full filling is excessive. For example, it is acceptable for goaf treatment under general roads and industrial sites to control the maximum deformation and keep the deformation continuous (Liu, 2011; Liu, 2012)."

Jan 1, 2018

By Christopher J. Brown

"Proper ground stability in gate roads has been a critical aspect of success in longwall mining. In particular, the tailgate has proven to be one of the more active mine areas that needs to be managed (Barczak, Dolinar, and Mucho, 1999). As longwall tailgate support technologies continue to be developed and improved, Strata Worldwide has worked to design and manufacture a revised Propsetter to provide additional support capacity and enable applications at greater mine heights. The support is targeted for tailgate applications; however, it can also be used for roof support in various types of mining environments. This paper describes the design changes for enhancing the support capacity of the Propsetter to more closely align its performance with other support technologies currently being employed in longwall tailgates. Balancing changes to each design aspect of the Propsetter was key to successfully improving the roof support load capacity while maintaining controlled deformation. Increasing the load capacity would allow a mine to improve support efficiency with a smaller, less intrusive support compared to support alternatives and to reduce the cost per foot of supported entry. The target support capacity was 80 to 100 tons at two inches of deformation to more closely match the performance of competing tailgate supports, such as the 24-inch-diameter CAN®. To achieve this goal, changes were made to all design aspects of the Propsetter: diameter, wedge cut depth and angle, pod size, and confinement rings. Test results from the NIOSH Mine Roof Simulator verified that these design changes were successful in increasing the capacity of the Propsetter to the 80- to 100-ton range at two inches of deformation. INTRODUCTION Roof support is one of the most fundamental elements of the underground mining process. As materials are removed from the earth, geological conditions will almost always require some form of roof support to keep the extraction area open and stable. Coal mining environments often experience difficult geological conditions that require both primary and secondary forms of roof support to keep the mining environment safe and economical. The secondary support system serves as additional support in the case the primary support system is inadequate. If designed properly, the secondary support system will also assist in maintaining integrity of the immediate roof beam (Barczak, 2000)."

Jan 1, 2017

By Mark Van Dyke, Gamal Rashed, Khaled Mohamed, Timothy Barton, Morgan Sears

"The unconfined compressive strength (UCS) of coal is an important parameter that controls the stability of coal ribs in underground mines. Estimation of the UCS of coal units composing the rib is required for conducting the rib stability assessment technique. This technique requires conducting large numbers of conventional UCS tests in a laboratory, which are expensive and time consuming. The research work outlined in this paper examines the use of indirect strength estimation methods—the point load test (PLT) and the in-situ Schmidt hammer test (ISHT)—for evaluating the intact strength of coal. The conventional UCS test was conducted on 80 coal specimens from 10 different coal seams for eastern coals. For each coal seam, the PLT and the ISHT were compared with the UCS test results to establish a correlation between the strength indices from the PLT or ISHT and the UCS. Moreover, to simplify the estimation of the UCS, a correlation between the megascopic coal lithotype and UCS was established.INTRODUCTIONBased on Mine Safety and Health Administration (MSHA) reports, rib failures have resulted in 17 fatalities, representing 52% of the ground-fall fatalities in underground coal mines in the United States over the past decade (MSHA, 2018). In an attempt to control rib failures in underground coal mines, the National Institute for Occupational Safety and Health (NIOSH) is working on developing a coal rib stability rating technique to provide a quantitative means to characterize the stability of coal ribs and to provide guidelines for early selections of rib support. The unconfined compressive strength (UCS) of coal units composing the ribs is one of the key input parameters required to estimate coal rib stability. The UCS of coal is an important parameter that controls the resistance of coal ribs to fracturing and slabbing. Direct determination of the UCS is time-consuming and expensive. Moreover, the UCS may be limited to testing only the strong coal samples since the weak or highly cleated samples might not survive the transportation and the preparation process (cutting and grinding). Therefore, alternative methods have been tested to indirectly find the intact strength of coal, such as the point load test (PLT) and the in-situ Schmidt hammer test (ISHT). The acronym ISHT refers to Schmidt hammer tests that were conducted in-situ on coal pillar ribs, while the acronym SHT refers to tests that were conducted on small scale blocks of coal."

Jan 1, 2018

By Uwe Wyink, Guenther Thiele, Isabel Gollnick, Klaus Thyrock

"The application of a thin spray on liner was implemented in longwall T-junction sections in 2009 for the first time. The development of the material and the application is based on a technical cooperation between RAG Deutsche Steinkohle AG and BASF Corporation. The report gives an overview of the first trials. The underground test occurred in a seam with a height of 1.5 m (5 ft) in a depth of 1,100 m (3,609 ft). The roadway is dual-use entry and the roof strata is composed of shale and sandy clay. In some sections geological faults impact the surrounding strata. The poor conditions induce separating and loosening of roof rock. Several remedial actions have to be implemented for decreasing deformations at the face entrance.The application of thin spray on liner (TSL) provides simplification and cost reduction of supporting work at longwall T-junction. The installation of TSL to avoid roof and coal fall operations for lagging are omitted.INTRODUCTIONFor safe operation of face retreat in dual-use roadways, the processes of securing and stabilizing the face end area are very important in order to minimize deformation and stress. Avoiding roof and coal fall are the main concerns for safe operation. For that reason different methods of reinforcement are tested in underground operation. In this paper the application of TSL is presented.The specifications of TSL are described, followed by the case study results. The underground study is located in one mine of RAG Deutsche Steinkohle. The extraction of the hard coal in this mine is characterized by great depth and high stress. There are poor geological and geotechnical conditions. The roadway used for the underground application was headed with the support system combination type A, which will be described later in the paper. After the first face retreat, deformation of steel arch support and bagging destroyed the roadway, so for the second face retreat remedial actions had to be undertaken. For this a new system of reinforcement, simplifying the supporting work at the longwall T-junction was tested. The reinforcement consisted of different parts that are described in the following report. Furthermore two different versions were applied. One method is the operation with TSL and the other one is the use of concrete for sealing. Both methods are compared in the text."

Jan 1, 2010

By Long Fan, Shimin Liu, Guijie Sang

"Weak shale roof falls have long been the main cause for fatalities in underground coal mines. Moisture-induced swelling is one of the root causes for the degradation of roof shales. In this study, slake durability tests are conducted on shale samples collected from Illinois basin coal mines, and different levels of durability are analyzed according to Gamble slake durability classification. There are two main mechanisms influencing the shape of retained fragments after two cycles of wetting and drying: mechanical abrasion and swelling pressure. A prototype optically based shale swelling apparatus is designed and manufactured for the shale free-swelling measurement. The experimental results suggest that swelling pressure tends to deteriorate shale laminations. Moisture-induced swelling of the No. 6 roof shale from Bear Run Mine was measured under 100% relative humidity condition. The measured swelling strain normal to beddings is ~5 to 7 times greater than the swelling strain parallel to beddings. This suggests that interlayer expansion plays a key role in moisture-induced swelling. Cracks induced by swelling pressure tend to occur along the bedding plane and lead to shale deterioration. INTRODUCTION Background Ground falls have long been the cause for nearly 50% of all fatalities in underground coal mines and continue to be one of the greatest safety hazards in underground spaces. Among the various ground fall hazards, roof failure takes up 23% of total fatalities in underground coal mines, of which 13% are due to skin failure and 10% are due to massive roof falls (Mark, Pappas, and Barczak, 2011). Roof failures associated with weak shales are summarized in previous studies (Molinda and Mark, 2010; Aughenbaugh and Bruzewski, 1973; Bajpayee, Pappas, and Ellenberger, 2014; Murphy, 2016). For mine roof instability and failure, two aspects can explain the failure mechanism of weak roofs. One is the elevated stress concentration in lateral direction due to the mining-induced excavation and its stress redistribution around the mine opening (Song and Stankus, 1997; Fan and Liu, 2017; Hill, 1986). The other mechanism is the passive stress-relief due to the strength weakening induced by the ambient moisture and temperature fluctuations during a period of time. Shale is moisture sensitive and presents a lower durability if it is exposed to a highly humid environment for a long time, thus leading to roof deterioration and instability in underground coal mines (Aughenbaugh and Bruzewski, 1976; Stateham and Radcliffe, 1978; Chugh and Missavage, 1981; Klemetti, Oyler, and Molinda, 2009). It is well known that shale is a clay-bearing sedimentary rock and that clay minerals are water sensitive. The moisture uptake for shale will induce the micro-deformation of a shale matrix, thus the structure and strength of shale is a function of time and intensity of moisture exposure. All mine engineers notice moisture-induced shale deterioration; however, the underlying mechanism(s) are still not fully understood due to the complexity of geophysical-chemical water-shale interactions (Huang, Aughenbaugh, and Rockaway, 1986; Hensen and Smit, 2002; Diaz-Perez, Cortés-Monroy, and Roegiers, 2007). The coal mining industry is still in the process of finding a technically sound, cost-effective technology to prevent this unexpected and unpredictable shale roof failure. Therefore, understanding how the moisture-sensitive shale responses to various ambient moisture and what role exposure duration plays in shale roof failure is key to prevent unexpected roof incidents, which will lay the foundation for the long-term roof support design and ground control management."

Jan 1, 2017

By Nikzad Oraee, Parham Khajehpour, Kazam Oraee, Arash Goodarzi

"Rock fall related accidents continue to occur in coal mines, although artificial support mechanisms have been used extensively [1]. Roof stability is primarily determined in many underground mines by a limited number of methods that often resort to subjective criteria. It is argued in this paper that stability conditions of mine roof strata, as a key factor in coal mines, must be determined by a survey which proactively investigates fundamental aspects.Failure of rock around the opening happens as a result of both high rock stress conditions and the presence of structural discontinuities. The properties of such discontinuities affect the engineering behavior of rock masses causing wedges or blocks to fall from the roof or slide out of the walls [2]. A practical rule based approach to assess the risk of a roof fall is proposed in the paper. The method is based on the analysis of structural data and the geometry and stability of wedges in underground coal mines. In this regard, an accident causing a huge collapse in a coal mine leading to four fatalities is illustrated by way of a case study. A comprehensive investigation of the hanging wall has been gathered through a systematic collection of evidence. The investigation results are then analyzed, and an interpretation of the evidence gathered is provided.For this purpose, horizontal and vertical profiles are prepared by geophysical methods to define the falling zone and its boundaries. The collapse is then modeled by the use of sophisticated computer programs in order to identify the causes of the accident and hence recommend corrective actions to prevent or minimize reoccurrence probability of similar accidents. The corrective measures are placed into a rule-based framework for the benefit of the mine operators.INTRODUCTIONRoof rock failure has generally been the most dangerous hazard for underground coal mining in the world. Failure of rock around shallow coal mine openings often results from loosening of blocks of rock on the weakened planes under the influence of gravity. This condition is not typically seen in large deep coal mines. There are a number of differences between large mechanized mines and small mines related to ground control [3]."

Jan 1, 2015

By Venkata S. Vutukuri

Compressive strength of cylindrical specimens of a model material and sandstone with various diameter/height (d/h) ratios was determined in unconfined compression tests. It was observed that the compressive strength increases with increase in d/h ratio. Triaxial compression tests were also conducted on the specimens of the same materials with a d/h ratio of about 0.47. It was observed that the axial stress at failure increases with increase in confining pressure. A very interesting correlation was found between the effect of d/h ratio on compressive strength in unconfined compression tests and the effect of confining pressure on axial stress at failure in triaxial tests. The relationships between d/h ratio and confining pressure, s3 in MPa are:- For model material - s3 = -0.95 + 1.7957*(d/h) 0.7883. For sandstone - s- (-0.4554 + 0.9888* (d/h) 1.0355. The correlation coefficients for these equations are 0.9997 and 1. These equations are valid for d/h ratios of up to 10.

Jan 1, 1993

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 Nan Zhou, Jixiong Zhang, Qiang Zhang

"Mining activities under hard roof could easily induce rock burst. Based on the geological conditions of widespread hard roof in China, this paper discusses the type, mechanism of occurrence and prevention of rock burst caused by hard roof from energy evolution point of view. It analyzed the interaction between the solid backfilled body and roof under different backfilling ratios. And by establishing and applying the mechanical model, the deformation of hard roof and energy evolution of the panel under different backfilling ratios were determined, revealing the control effect of solid backfilling on disaster-causing energy. As a result, an innovative method using the solid backfilling method was proposed as a solution to the hard-roof-induced rock burst hazards. Moreover, mine trial was conducted in Panel 6304-1 of Jisan Mine. The result shows that the stress concentration factor of the front abutment stress was only 1.44; the coal dusts in drill holes didn’t exceed the limits and microseism of large energy didn’t occur during mining mining, indicating that solid backfilling mining could be applied as an effective method for preventing rock burst caused by hard roof.INTRODUCTIONSHard roof refers to the thick strata above the coal seam or thin immediate roof. It is strong with poor joint development. After mining, instead of immediate caving, it will overhang for a large area in the gob, resulting in high stress in front of the face and difficulty in gas dispersion. As a result, the face spalls, the roof exposes and the gas accumulates in the gob. With the continuing advance of the face and increase in area of overhang, the stress in the hard roof will eventually exceed its ultimate strength and cause a large area of caving, followed by quick release of energy concentrated in the roof and coal seam, which may cause rock bursts and other dynamic disasters in the mine as well as serious equipment damage and casualties. Therefore, the hard roof becomes one of the main factors causing rock bursts at the face.Hard roofs in China varies in conditions. It ranges from dozens of meters to hundreds of meters in thickness. However, coal reserve under hard roofs accounts for about 1/3 of the total reserve; moreover, nearly 40% of the fully-mechanized coal mining panels are those with hard roofs and more than 50% of the mining areas are associated with problems of hard roofs. Aiming at this problem, researchers at home and abroad have proposed many solutions, e.g. coal pillar support, strength weakening and forced caving, to prevent rock bursts caused by hard roof. These solutions have been employed in some cases with good results. However, due to the variability in geological conditions of mining and hard roofs, the application of the above solutions is limited."

Jan 1, 2015

By Brent A. Slaker, Michael Murphy, John Ellenberger, Erik Westman

"Photogrammetry, as a tool for monitoring underground mine deformation, is an alternative to traditional point measurement devices, and may be capable of accurate measurements in situations where technologies such as laser scanning are unsuited, undesired or cost-prohibitive. An underground limestone mine in Ohio is used as a test case for monitoring of structurally unstable pillars. Seven pillars were photographed during four visits over a 63- day period. Using photogrammetry, point clouds of the mine geometry were obtained and triangulated surfaces were generated to determine volumes of change over time. Pillar spalling in the range of 0.29–4.03 m3 of rock on individual rib faces was detected. Isolated incidents of rock expansion prior to failure, and the isolated failure of a weak shale band, were also observed. Much of the pillars remained unchanged during the monitoring period, which is indicative of proper alignment in the triangulated surfaces. The photographs of some ribs were of either too poor quality or had insufficient overlap, and were not included. However, photogrammetry was successfully applied to multiple ribs in quantifying the pillar geometry change over time.IntroductionAdequately measuring underground rock mass movements is integral to understanding how rock masses behave and how to interact with them safely and efficiently. Many modern measurement techniques employed in underground mining environments rely on point measurements, such as through extensometers or borehole relief methods (Christiansson, 2006). These techniques, while commonplace, do not provide a comprehensive view of how the rock mass is behaving. The dynamic changes in stress states underground, coupled with the mechanical uncertainty of rock masses, near active excavations, creates a need for measurement systems which better capture the behavior of the rock mass. Photogrammetry is a technology capable of producing the wide-area displacement measurements required for a comprehensive understanding of rock mass movements.Digital photogrammetry is a means of obtaining threedimensional point clouds from digital photographs. By finding the same point across multiple images, all viewing an object from different angles, the location of that point in three-dimensional space can be inferred. Close range digital photogrammetry (CRDP) is photogrammetry applied to measuring objects or scenes less than 100 meters away (Atkinson, 1996), and is used for various functions in underground mining environments. These uses include, but are not limited to, mapping fracture networks (Kim, Kim, & Won, 2005)(Somervuori & Lamberg, 2010), characterizing fractures (Styles et al., 2010) and measuring volumes of blast rock (Heal, Hudyma, & Potvin, 2004)."

Jan 1, 2015

By G. P. Neobutov

The near surface inclined gold ore deposit is mined out at freezing temperature. The replacement of ore pillars by artificial freezing rocks during the rock pressures control in this conditions presented no difficulties. At the present, the mining operations are maintained at a depth of more than 300 meters. The exploration of mining conditions showed partial warming of rock mass and unstable ground conditions. Analytical calculations using a mathematical model of heat exchange predicted undesirable duration of freezing of artificial pillars. Practical techniques and recommendations for mining with ice pillars as an artificial support system using a new freezing method are discussed.

Jan 1, 1996

By R. F. Valene

This paper is not available for inclusion in the Proceedings at the time of printing. but it will be included as an insert

Jan 1, 1984

By Anthony (Sam) J. S. Spearing

Backfilling using coal waste is not a novel means for waste disposal throughout the world; yet, backfilling using a high density (paste) backfill is relatively new, except in Germany and Poland. One of the main benefits of a high density backfill, as opposed to a low density (slurry) backfill, is that, if placed at a low permeability and low moisture content, the risk of water table contamination can be relatively low. In addition, if paste backfill can be placed cost effectively, environmental benefits are huge because the surface impoundment footprint can be greatly reduced or even eliminated. This is especially true in room-and-pillar applications. In its current state, backfilling is considerably more costly than surface disposal. This is mainly because paste, as opposed to slurry, significantly increases backfill placement costs; however, this could be offset by increasing the achievable extraction rate. Overall, such a system could potentially improve the public?s image of coal mining in the US for a myriad of reasons. This paper discusses a proposed backfill system for a typical room-and-pillar coal mine in the Illinois Basin of the US and clearly shows the technical advantages, although costs are still higher with fill. The cost differential will likely reduce over time, in favor of the backfill option, as costs of traditional surface disposal increase, due mainly to ever-restrictive disposal regulations.

Jan 1, 2013

By I. Sedat Buyuksagis, Jamal Rostami, Ali Naeimipour

"Improving the safety of underground openings and optimizing ground support systems requires reliable ground characterization. This includes understanding the joints, discontinuities, and rock strength. Joint information can be obtained from borehole cameras, but measuring in-situ rock strength is very difficult. Use of borehole probes can be an ideal and efficient way to meet this requirement. This paper discusses the development of the Rock Strength Borehole Probe (RSBP) and presents the preliminary laboratory/field test. This device can be an accurate, flexible, quick, non-disruptive, and cost-effective alternative to estimate the rock strength inside boreholes in underground mines and tunnels. In addition, the scratch testing procedure in the laboratory by means of a newly developed miniature linear cutting machine is explained. Based on the results of these tests, two equations are developed to estimate the uniaxial compressive strength (UCS) and Brazilian tensile strength (BTS) of the sedimentary/metamorphic rocks. These equations, are based on testing 27 different sedimentary, igneous, and metamorphic rocks by full-scale scratch tests, including the cutting tests by a miniature disc. The results show a good correlation between the normal force and the compressive strength of sedimentary rock if the depth of the scratch is known. No significant correlation was observed for igneous rocks, due to the impacts of grain size.INTRODUCTIONEstimated rock mass properties are essential input for geomechanical studies and stability analysis for any structure constructed in or with rock. The availability of geological information is the crucial component of rock mass classification. One of the important parameters in evaluating rock mass properties is the intact rock strength. This parameter is usually measured by testing the core samples, obtained from exploration boreholes. These samples are subsequently tested in geomechanics laboratories. The test results offer limited information about the rock at the few boring locations despite all the efforts and costs. In addition, these results may not necessarily be representative of the behavior of the rock in the field as the test cannot provide the insitu condition of the ground."

Jan 1, 2016

By Dakota Faulkner

Unexpected roof falls often occur in the entries or intersections of active mining sections in underground mines. For large roof falls (greater than 20 ft in height), it becomes dangerous and impractical to clean up the rock debris and re-bolt the newly exposed roof strata. To help mine operators rehabilitate the roof fall areas in a quick, safe, and economic manner, Jennmar Corporation and Keystone Mining Services, LLC (KMS) designed, tested, and developed various impact-resistant (IR) steel set designs, as initially detailed in the proceedings of the 30th International Conference on Ground Control in Mining (ICGCM). Since then, the IR steel sets have been successfully installed in more than 100 roof falls in 43 different coal mines. Significant data has been obtained from these installations, enabling further refinement to and improvement on the design. This paper (1) updates drop test results of the IR lagging panel, (2) presents a case study of the performance of IR arch sets installed 56 months ago at a roof fall rehabilitation site, (3) demonstrates capacity evaluation of the IR steel set, and (4) outlines a preliminary applicability evaluation guideline of the IR steel set for a given roof fall condition. Field evaluation and structural numerical analysis indicate that impact capacity of the system is significantly higher than that of the IR lagging panel, as originally determined. It is concluded that, when evaluating the applicability of an IR steel set, ground control engineers should consider the IR lagging and steel set as an entire system and make a reasonable engineering decision based on actual roof fall geotechnical conditions.

Jan 1, 2014

By Monte R. Hieb

Study of moisture-induced swelling strain in Pennsylvanian Period rocks of the Appalachian Plateau in West Virginia shows a 2:1 biaxial horizontal anisotropy which trends NW-SE. Oriented in-situ slabs of coal mine roof rock were collected across West Virginia, dry-sawed to size, air-dried to stability, and then monitored for moisture changes and deformations in the horizontal plane through one wetting cycle and drying cycle. Biaxial swelling strains were exhibited in nearly all rocks tested, with maximum horizontal deformations occurring normally-incident to the strike of regional geologic fold structures. This relationship persists as fold strike varies across the study area. A connection with Alleghanian fold development and/or post-Alleghanian fold relaxation is suggested. Deformations were largest in shales, but occurred in sandstones and siltstones as well. During testing, moisture-induced fractures (MIF) developed parallel to the regional fold structure and ranged in size from hairline cracks to sample-destroying cracks that were orthogonal to the observed biaxial swelling strain direction. Preferential wetting and swelling of clay minerals along NE-trending fractures, related to a fold-parallel fracture fabric of regional extent across West Virginia, is the most likely reason for the observed systematic NW-oriented biaxial swelling strains. The testing methods used in this study are simple and inexpensive and produce robust results that offer insight into a regional, anisotropic mechanical weakness that imparts a dynamic, asymmetrical horizontal stress to moisture-sensitive coal mine roof rocks over time. Because a similar mechanical fabric has been reported in Devonian-age shales, these findings may have implications for shale gas developers in the Appalachian Basin challenged by rapid aperture closure rates along horizontallydrilled induced fracture networks in Devonian rocks like the Marcellus Shale. The paper concludes with an example of how the concept of a preferred swelling strain direction might be applied to improve the effectiveness of roof trusses.

Jan 1, 2013

By Houquan Zhang, Ming Li, Guimin Zhang, Bei Zhang

"The support quality of bolts in deep roadways is one of the important factors for efficient and safe coal production. A new kind of non-destructive testing technique and equipment was used to monitor the axial bolt load in real time in deep roadways. Based on the test results, the working status of bolt support in deep roadways was evaluated. The research results show that the variation characteristics of axial bolt load is consistent with the distribution characteristics of rock pressure no matter whether it is in the initial support stage or mining disturbance stage. The axial load of bolts located 20m outby the development face has an overall rising trend within 3-5 days after installation. Within a certain distance in the front abutment pressure zone in front of the face, the axial bolt load increases with the advance of coal extraction. When coal and rock in the abutment pressure zone yield, the axial bolt load is dropped correspondingly. This kind of non-destructive testing technique provides an advanced means of detection and evaluation on bolt support quality.INTRODUCTIONMany coal seams more than 800m deep are being mined in eastern China. Roadways of large cross-section are being developed in coal seams more than 600m deep in western China (Kang et al. 2014). Moreover, these roadways are commonly supported by roof bolts and have encountered many support problems, including soft rock strata, intense mining disturbance, and abnormal tectonic stress. The stability of its surrounding rocks is difficult to maintain, the roof bolt even fails before mining (Shen 2014; Jiang et al. 2015). Therefore, it is very important to develop an effective method to detect the bolt quality in real time and evaluate the bolt's safety status.During installation of roof bolt, the installation quality is heavily affected by installation equipment and tools, field conditions, and operators' skills (Li et al. 2005; Shan 2009). The anchorage length will be shorter than the designed one if the resin cartridges are pushed to the bottom of excessively long drill hole and stay there without being shredded and mixed. The support function of the bolt would be weakened by the shorter anchorage length."

Jan 1, 2016

By Jinwang Zhang, Zhaohui Wang, Jiachen Wang

"Thick coal seams (seam height larger than 3.5 m) account for a large portion of the proven coal reserves of China, and underground thick coal seam mining provides about 50% of the Chinese annual coal production. Such seams are mainly mined using the longwall top-coal caving (LTCC) method. For safe and successful mining, a number of analyses on the stability of surrounding rocks in the LTCC face area have been extensively performed, and certain research achievements have been made. This paper introduces the main findings on rock pressure occurrence in the LTCC face during the last few decades and presents face and roof stability control techniques. INTRODUCTIONUnderground thick coal seams in China provide 1.8 billion tons of coal every year, which accounts for approximately 50% of the total Chinese annual coal production and 25% of the world, indicating that thick coal seams stand an important position in Chinese coal mining industry (Wang, 2005). Before the 1980s, thick coal seams were mined mainly by using the multiple-slice mining method. With the improvement of mining techniques and equipment in last two decades, longwall top-coal caving (LTCC) and high-seam, single-cut longwall are employed for thick coal seam extraction, offering advantages over multiple slicing from entry excavation and production perspectives (Wang, 2013).. The primary author of this paper introduced the most popular thick coal seam mining methods in China at the 30th ICGCM (Wang, 2015). This paper presents the stability of surrounding rocks in face area and ground control techniques for mining thick coal seams. For traditional longwall faces (seam height less than 3.5 m), the well-developed VoussoirVoussoir beam theory has been successfully used to analyze the movement of overlaying strata, which indicates that ground control techniques have become mature for this mining system. In such faces, the four-leg support with loading capacity of 4000–8000 kN is considered adequate for safe and successful mining. These faces normally have an annual yield of 1–2 million tons (Wang, 2007). For thick coal seams, however, the traditional support becomes inadequate and the ground control problem change to be more difficult and complex. Thus, development of large-scale mining equipment and ground control techniques for enlarged mined-out space and entries have been recognized as the most basic technical problems in mining thick coal seams."

Jan 1, 2017

By William J. Conrad, Ryan J. Molka, Erik C. Westman

"In order to study pillar and overburden response to retreat mining, a ground control program was conducted at a Central Appalachian mine. The program consisted of several monitoring methods including a seismic monitoring system, borehole pressure cells in the pillars, and time-lapse photogrammetry of the pillar ribs. Two parallel geophone arrays were installed, one on each side of the panel with the sensors mounted 3m (IO ft) into the roof. A total of fourteen geophones recorded more than 5,000 events during the panel retreat. A MIDAS datalogger was used to record pressure from borehole pressure cells (BPCs) located in two adjacent pillars that were not mined during retreat. A series of photographs were taken of the pillars that had the BPCs as the face approached so that deformation of the entire rib could be monitored using photogrammetry.Results showed that pillar stability and cave development were as expected. The BPCs showed an increase in loading when the face was 115 m (380 ft) inby and a clear onset of the forward abutment at 30 m (100 ft). The photogrammetry results displayed pillar deformation corresponding to the increased loading. The microseismic monitoring results showed the overburden caving inby the face, again as expected. The significance of these results lies in two points, I) we can quantify the safe manner in which this mine is conducting retreating operations, and 2) we can use volumetric technologies (photogrammetry and microseismic) to monitor entire volumes of the mine in addition to the traditional point-location geotechnical measurements (BPCs).INTRODUCTION AND BACKGROUNDExcavation of underground openings is a difficult job in a challenging environment. The rigor of the process is compounded by the lack of a method that allows quantification of changes within the rock mass as excavation progresses. Statistics kept by the Department of Labor, Mine Safety and Health Administration show that 16% of fatalities and lost time incidents in the underground mining industry are due to unexpected rock mass failure (MSHA, 2015). A recent example is the Crandall Canyon, Utah, coal mine accident caused by the vertical collapse of a longwall coal mine on August 6, 2007 (Dreger, Ford, Walter, 2008)."

Jan 1, 2016

By Wojciech Masny, Andrzej Walentek

"The most common extraction method in Polish coal mines utilize pillars between adjacent longwall panels. This is mainly due to the significant depth and difficulties in maintaining stability of the gateroads. This paper presents measurements of entry convergence and range of the fracture zone around a gateroad measured with a borehole camera. The entry where the measurements were taken was located in a parallel entry separated from the gateroad by a 5-m- (16-ft-) wide pillar. The gateroad was located in a hard coal seam 900 m (2,952 ft) deep with a dip of 5°. The extraction method was longwall mining with caving and the mining height was about 3 m (10 ft).Furthermore, the ""in situ"" measurements were compared to the convergence forecast for the gateroad and fracture zone by means of numerical modeling conducted using the Phase2 Version X.X finite element program from Rocscience. The determination of the fractured zone around the openings is of great importance to successful mining practice, especially in support planning and design for longwall gateroads. Moreover, proper determination of the fractured rock zone is one of the basic elements in the process of planning the roof bolting for mine workings.INTRODUCTIONIn the Polish. hard coal mining sector, retreat longwall extraction is most commonly used (Prusek and Masny, 2007). This method is connected with the necessity to carry out a number of development operations during the first stage, including among others, driving of gateroads and a setup room.Currently, most Polish mines eliminate gateroads behind the extraction front, using the ""U"" ventilation system (Szlitzak et al., 2008). In such cases, one of the roads for the next longwall panel is driven along the caving goafs, leaving a protective pillar of some chosen width (Figure 1). The question regarding the width of the coal pillar is especially important to the stability of the gateroads that define the panel for the subsequent longwall. When analyzing the previous experience relating to this type of pillar (Junker, 2006; Rajwa, 2008; Stranzet et al., 1973; Betriebsempfehlung, 1982), one can state that there exist two solutions for the determination of its width."

Jan 1, 2010