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By Xianxin Shi

In surface electrical exploration the high resolution earth resistivity method (HRRM) is a very effective method for detecting abandoned drift mines workings. When the abandoned mines are more than 500 ft (150 m) deep, its detection capability reduces greatly and requires more effort to implement. In Yan Quan coal mine, Shanxi province, we tried to adapt this method for underground application. Two survey lines were designed with the spacing of current and potential poles 20 m and 10 m, respectively and measuring points at 2m. The measurement radius of I line and I1 line were 140 m and 60 m, respectively, and the infinitely far pole was 1200~2000 m from the survey line. (Note the I and II lines are located on the north and south ribs of main tunnel, respectively). The survey results showed that abandoned workings were located at 25-70m, Survey Line I and at 80-1 10m, Survey line 11. Based on this finding, the longwall panel gateroads and set up entry were properly located thereby providing safe mining of the No. 15 Coal seam. Key Words: Abandoned coal mine, High resolution earth resistivity, Underground electrical exploration

Jan 1, 2005

By Jisheng Han

Interaction effect exists when two or more neighboring seams are mined. The mining activity in one seam causes stress redistribution and strata movement in both the roof and floor. When the thickness of interburden between two neighboring seams is small, some hazardous situations like roof cutters, roof falls, floor heave and rib spalling in one seam may be encountered due to the effect of the mining activity in the other seam. An investigation about roof cutters, roof falls and floor heave was conducted in a West Virginia mine to study the interaction of multi-seam mining. The mine is mining the Sewickley Seam, which is about 90 ft above the Pittsburgh Seam which was mined in the 1960's using room and pillar method. The roof falls and other stability anomalies of entries observed underground in the Sewickley Seam were found closely related to the location of barrier pillars and gas well pillars left in the Pittsburgh Seam. Then numerical modeling was carried out to check the interaction between these two seams of mining. It was found through the modeling results that in the areas where pillaring operation was conducted in the Pittsburgh Seam, the stress concentrated less in the Sewickley Seam while near the edge of the gas well pillars and barrier pillars, high stress concentration was observed. It was concluded that under the geological conditions of the study mine, the entries stability would be affected by the lower seam mining activity, especially above or around the edges of pillars.

Jan 1, 2005

By Stephen C. Tadolini, John P. McDonnell

"High-strength cable bolts have been used for ground support in underground coal mines for many years. Cable bolts are routinely used as secondary ground support in numerous longwall gateroad support applications, areas that possibly experience the most severe conditions in current underground coal mining. Additionally, when mining ground conditions deteriorate, engineers oftentimes recommend cable supports as a principal solution because of the established credibility for successful ground control. Nevertheless, there is still a reluctance to utilize non-tensioned cable supports as an effective primary roof support alternative.This paper provides historical and recently documented examples of non-tensioned cable supports utilized as an effective primary ground support method in a western U.S. underground coal mine, and it will continue the technical discussion between the differences of various degrees of tension in cable supports with respect to ground behavior and reactions.INTRODUCTIONResin-grouted cable supports are widely used to support difficult ground conditions throughout the world. Through the use of innovative applications and configurations they have proven to be a major paradigm shift in roof support and ground control technology in the U.S. (Tadolini and Koch, 1993; McDonnell et al., 1995; Tadolini and Gallagher, 1994; Tadolini and Trackemas, 1994)Today, the basic cable bolt support consists of a high-strength steel cable installed in bore holes that range in diameter from 25 to 50 mm (1 to 2 in) depending on the strand size and drilling equipment. Cables are offered in both bright and galvanized, for corrosion protection, and range in capacity from about 30 to 60 tons. An example of a resin-grouted cable bolt and the key components is shown in Figure 1. The steel cables are flexible and the support length is not limited by the height of the opening. Lengths of up to 10 m (32 ft) have been supported with high pattern densities when standing supports were not operationally feasible. The most common lengths are 3 to 5 m (10 to 16 ft). An important physical characteristic is that PC-strand has an ASTM minimum elongation requirement of 4%. For comparison purposes, the ASTM 432-08 Standard Specifications for Roof and Rock Bolts and Accessories specifies elongation requirements for traditional roof bolts are a minimum of 12% for grade 40, 9% for grade 60, and 8% for grade 75 materials, respectively. Traditional roof bolts can stretch and elongate twice the length that cable bolts can, which makes design considerations challenging."

Jan 1, 2010

By Kot F. v. Unrug

In spite of the extensive research and accumulated experience from mining operations, an appropriate pillar design remains an open issue. Due to the shrinking reserve base, present mining operations move to more challenging conditions. There are usually thinner seams and geologic conditions are often unfavorable. For instance, mine roof and floor are weak or deteriorate rapidly when exposed to moisture fluctuations in mine atmosphere. It is common that pillars of current operations are more- squat than several decades ago when the bulk of pillar research have been carried out in thicker seams. The methods of pillars calculations were developed regionally (in particular countries) to suit design needs for the most important seams, from which the majority of coal production was coming (e.g. Pittsburg or Pond Creek seams in the eastern of US). Most of the pillar formulae use similar approach by assigning weights (using constants) to the influence of coal (seam) strength and pillar widthheight ratio on its final strength. That mechanistic approach has sound theoretical base and good data supporting its validity. However, what seems to be missing is a qualification system of the mine conditions, which are outside of the data base or do not comply with the theoretical model applied in pillar strength formulae. This paper attempts to address that issue in two major categories, first when only a drill core data are available, and second during the mine development observations can be made directly on a coal seam. The composition of a coal seam is also addressed, (which often changes rapidly across the reserve), and discussed its influence on pillar's structural performance. In summary a set of practical design rules is presented.

Jan 1, 2005

By Luis Giraldo

Rock bolt installation characteristics near roof falls have been identified as contributing to failure. One documented and regularly occurring failure mechanism is loss of anchorage between the grout and the rock wall of the bolt hole. Key contributors to the integrity of the grout interlocking with the rock mass are the diameter of the hole relative to the diameter of the bolt, resin versus cement type grouts, rock type, and condition of the hole. Smooth bolt holes consistently exhibit less rock bolt load bearing capacity than rough walled holes. The goal of this study was to quantify the benefits of borehole conditioning and develop a systematic approach that would allow improvement of roof bolt performance. A new rock drilling technology, the Helical Drag Bit (HDB), was used in the early stages of this project to condition the borehole. The HDB cuts a prescribed helical groove in the borehole wall. When a bolt is grouted into a hole that has been conditioned using the HDB, the grout fills the helical groove and provides a stronger, more reliable mechanical lock between the rock and the grout. This paper presents results of a series of tests conducted in the laboratory and in the field. Short encapsulation pull out tests have demonstrated considerable increase in pull out strength and more consistent bolt performance when hole geometry is modified using the HDB. The results have also shown that greater load bearing capacity improvements are obtained in mines with weak roof rock, where it is most needed. Additionally, bolts recovered for examination suggest that hole conditioning during roof bolting reduces the effects of finger gloving. Therefore, mines with stronger roof rock could also benefit from hole conditioning. A new roof bolt that uses the HDB principle has been designed and tested. The new system works with existing bolting equipment requiring minimal modifications to the traditional procedure used for installation of fully grouted rebar bolts. The use of these higher anchorage capacity bolts will allow mines to reduce severe injuries and fatalities resulting from ground failures while maintaining or increasing productivity.

Jan 1, 2005

By Taghi Sherizadeh, Pinnaduwa H. S. W. Kulatilake

"This paper examines the effect of different geological and mining factors on roof stability in underground coal mines by combining the field observations, laboratory testing, and numerical modeling. Observations at an underground coal mine that uses the room-and-pillar mining method showed several failures of the immediate roof, and is used as a case study. Three-dimensional distinct element analyses were performed to study the intrinsic roof failure mechanisms and to evaluate specifically the effect of in-situ stress magnitudes and directions, discontinuity mechanical properties, and variation of pillar to room size on roof stability.INTRODUCTIONThe lithology of the immediate roof with respect to the underground openings plays an important role in stability of the roof in underground excavations. Laminated, or thinly bedded shale, is one of the most frequently seen overlaying strata above the coal deposits. The bedding planes and interfaces have very low or close to zero tensile strength in the direction perpendicular to the bedding planes, and their shear strength is much lower than of the rock layers. As a result, the bedding planes are much weaker than the rock layers. Therefore, due to the stress concentrations caused by excavations, slippage and separation can easily happen along the bedding planes before initiation of failure within the rock layers.In the past, several attempts have been made to understand the mechanism behind the roof failure in underground coal mines. Most of these earlier attempts were based on empirical techniques and continuum based numerical methods that have limited capability in simulating the response of discontinuum rock mass behavior to excavation. Different factors such as the magnitude and direction of the horizontal in-situ stresses, type and mechanical properties of roof rock, surface topography, geological anomalies, gas pressure, slippage and separation of bedding planes at the roof and floor, entry width and excavation sequences are identified as key factors that could significantly impact the stability of the roof. Among all the aforementioned factors, in spite of its importance, the impact of slip and separation of bedding planes on roof stability has not received enough attention from researchers. In a few available numerical modeling studies, the bedding planes are modeled using the continuum based numerical codes such as ABAQUS (Chen, 1999) and FLAC3D (Ray, 2009). However, since the bedding planes among the roof layers become discontinuous during the numerical analysis, this behavior is most likely not be accurately captured by the continuum based numerical models. Chen (1999) conducted finite element analysis to study the impacts of the roof and coal interface slip and the bedding planes slip in the roof on the stress distribution in the roof strata by using the finite element analysis software named ABAQUS. To avoid the complications associated with the finite sliding between two deformable bodies, he limited his study to the two-dimensional finite element analysis (plane strain). He concluded that the slip and separation of interfaces and bedding planes have significant impact on the behavior of the roof strata and without considering the interfaces, the roof bolt functions cannot be simulated properly by using the numerical analysis methods. Gadde and Peng (2005) performed three-dimensional finite difference analysis using FLAC3D software with strain softening material behavior to simulate the cutter roof in an underground coal mine. Their numerical model was fully continuum and no bedding plane was included in the model. They suggested that the strain softening material model can help to simulate the cutter roof failure up to a satisfactory level (Gadde and Peng, 2005))."

Jan 1, 2015

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 David Miller

Longwall gateroad support can be supplied by a variety of methods. The system of support also needs to be flexible as conditions change. This is especially true in the complex geological conditions of the west Kentucky # 13 seam. Methods of gateroad support used in this seam will be presented along with the successes and failures of the systems as conditions changed Lodestar Energy's gateroad support progression through the seven longwall panels retreated to date will provide an insight into planning for future panels

Jan 1, 1998

By Terry Medhurst

Geophysical logging is routinely undertaken as part of most Australian coal mining exploration programs. Typically, the main application is the determination of coal seam depth and the qualitative estimation of coal quality, lithology and rock strength. For rock strength estimation, sonic transit times relate to the elastic properties of the rock and provide a general correlation with uniaxial compressive strength. However, experience has shown that site-specific relationships are necessary to account for different geological environments. In this paper we describe results from a series of recent studies that have provided a more comprehensive method of geotechnical strata characterisation using detailed lithological interpretation and rock property estimation from borehole logs. This has resulted in an improved method of estimating rock strength via the determination of the proportions of quartz and clay plus the porosity. We also propose a rock classification scheme based on geophysical log analysis. This builds on the existing acceptance of the sonic log as an indicator of rock strength and adds scores according to the moisture, cohesion, porosity and quartz content inferred from the geophysical logs. The new scheme can be related to existing mapping based schemes such as CMRR and the Q-system.

Jan 1, 2005

By Dion Pastars

Resin based grouts are the main form of rock bolt anchorage in the underground coal industry in Australia and New Zealand. To be effective, the system requires the mixing of the catalyst and mastic components of the resin, as well as shredding of the laminate cartridge that contains the resin. An unknown measure is the load transfer characteristics of a bolt where the resin is well mixed but remains encased in the cartridge (gloved). Laboratory and in situ field investigations have been undertaken to quantify the performance loss due to mixed gloved bolts. This work showed repeatable results, indicating serious performance loss of the gloved and mixed system, with load transfer approximately 10- 15% of a non-gloved system (MacGregor, 2004). The in situ testing has demonstrated the relationship between the adhesion qualities of the resin and the mechanical interlock generated by radial confinement with progressively increasing tensile load. Effective load transfer is defined by the ability of the system to sustain shear stress on the bolt hole wall.

Jan 1, 2005

By S. Mehrdad Heidari

The concentrator plant is located near a very high (105m), large (- 500m) and steep (a, > 45' ) wall. Geologically, the wall was composed of very complicated formations and broad range of soils and rocks type in which a variety of local typical failures models have been occurred on different soil and rocks formations. The construction of concentrator plant structure and other activities have been overshadowed by these local failures. Safety and risk of huge failures were serious concern. For distinguishing the wall conditions a broad campaign of slope stability investigation with detailed geotechnical drilling on berms of wall have been done and the structure of slope was fully identified and further Lab tests and slope simulation models done. Slope stability of the wall was analyzed and the effects of instability factors were described in short term (for construction phase) and long term (for production phase).The results indicated, if a well-appointed drainage system is implemented, the overall stability of the wall in short term will not cause any particular problem but about long term stability, the wall must be supported by rock bolting, consolidation grouting, shotcrete reinforced with fibers. The results have been confirmed by passing one year after this investigation.

Jan 1, 2005

By Shugang Li

The layout and efficiency of the degasification system is heavily dependent on the distribution of and the methane migrating mechanism in the mining induced fractures in the overburden strata In this paper, physical modeling and numerical simulation are employed to study the mining-induced fracture zones and the stress distribution in these zones. The physical models show that both the fracture and compaction zones form a parabolic shape in the over- burden strata along the panel transverse direction. When this is generalized to the panel longitudinal direction, the shape of the mining induced fracture zone over the longwall gob can be represented by a 3-D elliptic paraboloid function. Through the observation of bed separation, fracture density and distribution pattern in the physical models, it is believed that the development of the parabolic fracture zone can be divided into three distinguishable phases. The expansion characteristics and sizes of the fracture zones can be estimated. The numerical stimulation studies can be used to estimate the extent of the stress release zones. The stress release zones can be also divided into a high compressive stress - fracture closure zone, stress relieved - fracture expansion zone, and intact zone. The methane desorption model is then used to analyze the methane migrating pattern in the fracture zone. The developed model has been verified by the results of a longwall gob degasification tests at a coal mine in Yangquan, China with a fair agreement. Such studies show that the model can provide a reasonable accurate tool for designing the gob-well degasification system for longwall mining operations.

Jan 1, 2005

By Dennis Dolinar

"Tailgate entries in U.S. longwalls can be subjected to high loads and deformations. Therefore, tailgate entries are generally supported with varying types and amounts of standing support. However, even with standing support, roof falls and blockages can occur in the tailgate when the support system is not properly designed. Safe and efficient operation of a longwall face requires the selection of the appropriate type and amount of standing support. Because of the importance of maintaining a safe and functional tailgate entry, the National Institute for Occupational Safety and Health (NIOSH) is currently developing a standing support design methodology based on the ground reaction curve. Using the ground reaction curve for design requires that the ground and support interaction be quantified. This can be accomplished by the installation and analysis of data from instrumented field sites. Although necessary, the field results cannot develop the complete ground reaction curve. Numerical models, calibrated to the field results, are used to establish the complete ground reaction curve. In the present study, data was obtained and evaluated to establish points for calibration of the ground reaction curve for two western longwall operations.One test site, Mine A, was located at a depth of 1,550 ft. At this site, the floor consisted of siltstone, while the roof was an interbedded siltstone and sandstone. Because of the depth, a double row of22-in diameter CANs1 were installed to support the tailgate. Both the first and second longwall panels have mined past the instrumented site.The second site in Mine B is at a depth of only 760 ft. The floor at this mine consists of a weak shale/mudstone and the roof consists of thick, top coal overlain by weak shales. At this site, a double row of pumpable cementitious cribs was installed. The second test site has only been subjected to side abutment loading from first panel longwall mining.The measurements indicate that the support at both mines was heavily loaded when subjected to the side abutment loading from first panel mining. At the deeper site, the support began to yield when the face was more than 300 ft outby during second panel mining. In comparison to results from previous test sites in the Pittsburgh Seam and Illinois Basin, the amount of support loading and roof-to-floor convergence was an order of magnitude higher at the western sites. This has resulted in a significant difference in the position and shape of the ground reaction curves."

Jan 1, 2010

By David A. Newman

In the 1950s, an underground limestone mine in the Camp Nelson Formation was developed from an outcrop exposure. The Camp Nelson Limestone met the requirements for DOT stone and aggregate. The mine ownership changed several times, which resulted in distinct mine designs in Level 1. Based upon the production requirements and an attempt to maximize limestone recovery from a single level, a variety of pillar dimensions and three mining heights (28 feet, 68 feet and 83 feet) were developed. This was sufficient prior to reaching the mineable extent of Level 1. Level 2 introduced previously unknown ground control challenges. These included: ??Pillar spalling and a significant back fall on the Level 1 - Level 2 decline; ??Laminated immediate back and the need to bolt the back on Level 2; ??Delamination of the back adjacent to the Level 1 - Level 2 decline resulting in bolt shearing, and floor heave in isolated areas of Level 2; ??Multiple level interaction, the need to design pillars based upon numerical modeling as columnization between Levels is infeasible: and ??Water inflow as the mine approached river. To meet the challenges, a testing program to determine rock strength and physical properties was implemented along with numerical (LAMODEL and FEM) modeling, and borescope inspections of the immediate back. The efforts to implement rock mechanics were re-focused as Level 2 approached its mineable limit and the initial attempt to drive a decline to Level 3 was unsuccessful. A second decline site was selected based on inter burden stability, a consistent shale parting for the Level 3 back horizon, results from numerical models of the immediate back and multiple level interactions, and the projected low water inflow. The mine is operating on Levels 2 and 3 with long-term planning and ground monitoring for Level 3.

Jan 1, 2014

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 Morgan Sears

The LaModel program (Heasley, 1998) is a displacement discontinuity, boundary element model (BEM) for analyzing stresses and displacements in single and multiple-seam coal mines. After the Crandall Canyon mine collapse, the need for better design guidelines for LaModel was observed and a new calibration method for deep cover pillar retreat was developed (Heasley et al., 2010). However, the scope of this initial calibration was limited to single-seam mines deeper than 750 ft. This limited application of the initial calibration method lead directly to the research presented in this paper, which attempts to define appropriate LaModel calibration guidelines for pillar design where shallow cover and/ or multiple-seam mining scenarios are encountered. As part of this research, a database of shallow cover case histories has been developed, and this database is used to help validate the shallow cover calibration method. By adding the shallow cover database to the previous deep cover database, the LaModel calibration method has now been extended to the complete range of mining depths, as well as to multiple-seam cases. With this shallow cover calibration, the utility of the LaModel program continues to increase.

Jan 1, 2013

By Raj R. Kallu

Underground gold mines in Nevada are typically hosted in geologic zones with a wide range of geotechnical properties ranging from blocky competent rock to faulted and highly altered soil-like material. Geotechnical engineering design in weak rock utilizes rock mass classification systems as inputs to both empirical and numerical methods. The most common rock mass classification system utilized in the gold mines of Nevada is the easy-to-use Rock Mass Rating (RMR) system; however, in practice it is difficult to apply in weak to very weak rock masses. This paper attempts to improve the sensitivity of RMR in weak rock masses by utilizing W-RMR?a modified procedure for calculating RMR?and evaluates its performance against other methods of determining RMR for predicting in situ moduli. In-situ rock mass moduli were determined using a portable plate loading device (PPLD), and these data were used to quantify the improved sensitivity of the modified RMR versus existing classification systems. The PPLD was designed for performing rapid and economical in-situ deformability tests in a mining environment. Tests were performed at eight sites yielding thirteen rock mass moduli measurements. Statistical analyses indicate Detailed-Discontinuity RMR-89 and W-RMR provide the best correlations to the in-situ moduli compared to other methods of calculating RMR.

Jan 1, 2014

By Bruce W. King, David A. Newman

"Underground mines have always been an integral part of the limestone industry. However, beginning in the 1990s, the number of underground limestone mines has increased for many reasons including• Stripping ratios at surface quarries have become uneconomic.• Existing surface quarries have reached the mineral boundaries, and the remaining stone reserves are underground.• Surface access to the entirety of the limestone reserve is limited by development or infrastructure.• The once-rural setting of the surface quarry is now surrounded by subdivisions and development. An underground mine is less intrusive and takes the mining operation out of the public eye.• For larger markets, an underground mine enables year-round production without the sequence of overburden stripping in the winter to expose mineable stone for summer production. This is common at many surface quarries.Multiple-level operations are a natural extension of the decision to mine underground. Most limestone reserves, for example, the Tyrone-Oregon-Camp Nelson, Hermitage-CartersLebanon- Ridley, and Salem Limestones, are sufficiently thick to accommodate multiple levels. Multiple-level mines should be planned by considering each level separately and then integrating the individual levels into a larger mine plan. Failure to plan all the levels of a multiple-level mine in concert can result in ground control (back, pillar, floor, and inter-level sill pillar) problems attributable to stress transfer and interaction between the levels or the necessity to leave a thicker sill pillar between adjacent levels.Multiple-level mine planning begins with geology. Once the upper level back horizon is identified, the geology, rock strength, and physical properties in each level are used to• identify the optimal back horizon• quantify whether back support is required and, if so, the type, spacing, and length of the back support• determine the pillar centers• determine the total mining height, including floor shots"

Jan 1, 2016

By Weihua Song, Hongwei Zhang, Feng Zhu, Sheng Li, Guoshui Tang

"Rock burst is one of the most serious dynamic disasters in coal mining. Rock burst prevention involves two difficult tasks: systematically evaluating the occurrence conditions and effectively forecast for rock burst. This paper systematically studies the effects of natural geological dynamic conditions and mining engineering on and forecast for rock burst. It establishes the evaluation method and index system of rock burst geological dynamic conditions and proposes a method of forecasting rock bursts by monitoring fault activity with anchor cables. Natural geological dynamic conditions are a necessary condition while mining engineering is a sufficient condition for the occurrence of rock bursts. Rock bursts in mine must have suitable geological dynamic conditions, and mining engineering activity is merely an inducing factor. There is a good correlation between anchor cable stress variation and fault activity, when the anchor cable dynamometer indicates that the fault is active. It can forecast a rock burst a few days in advance. This technique has been successfully applied to Panel 83003 of a coal mine in China, which improved the safety production level of the mine and found a new way of rock burst evaluation and prediction.INTRODUCTIONRock burst is one of the most serious dynamic disasters in coal mining. With the increase of mining depth, rock burst disaster is becoming more and more serious. Systematically evaluation and effective forecasting of the occurrence of rock burst in the coal mines are important and difficult work in the control of rock bursts. Chinese researchers have performed numerous researches on rock bursts, Qi et al., established the rock burst model of frictional sliding instability, explained the occurrence mechanism of rock bursts by stick-slip phenomenon in frictional sliding, gave full consideration to the influence of primary rock stress and mining-induced stress on rock burst occurrence, and proposed the stress controlling theory of rock burst. Pan et al., [3] put forward the criterion of perturbation response, explained and analyzed the problem of rock burst in circular openings, put forward the ratio of coal-rock elastic modulus and reduced modulus is the significant parameters for the determination of stabilization; systematically summarized the rock burst distribution, type, mechanism and control in China. They classified rock bursts into three types: coal compression, roof fracture and fault movement. By applying energy transfer principle and conservation of energy, combined with the analysis of the weakening of rock fabric damage, Zhou and Jiang put forward the theory and equation for the occurrence of coal-rock mass impact effect. They analyzed and discussed the formation mechanism of rock burst by combining impact effect theory and energy equation. Wang et al., recognized that the mechanism of rock burst which induced by fault, fold and phase transition is the superposition of tectonic stress and mining stress and that the axes of syncline and anticline and their flanks are the location where rock bursts mostly occur. To prevent stress superposition is the efficient way to control the structural rock burst [7]. Jiang et al., summarized the research achievement of rock burst mechanism and controlling techniques and existing problems in recent years in China, and pointed out the direction to improve the level of prevention of controlling rock bursts in coal mines."

Jan 1, 2015

By Behdeen Oraee-Mirzamani

Tunnel stability has an important role in the production process of an underground coal mine. There are various methods for analysing tunnel stability, such as numerical methods and analytical methods. In this paper, numerical methods (Phase2 software) have been used to determine tunnel wall displacement in a mining tunnel of the Parvade underground coal mine. The Ground Reaction Curve has also been drawn using analytical methods to determine tunnel wall displacement. The comparison of results from the numerical method and the analytical method show a noteworthy difference in the tunnel wall displacement. The displacement calculated by the numerical method shows a lower value than that of the analytical method because the numerical method is more suited to modeling the various coal and rock layers, as well as the shape of the excavations found in coal mines. The methodology used in this paper together with the results obtained from both methods can serve as useful tools for the coal mine design engineer when determining the ground support requirements.

Jan 1, 2013