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By Mikhail P. Nesterov

The paper presents the results of the research that can forecast the development of subsidences and deformations of the earth surface with time in the permanent marginal areas of the movement troughs arising when mining sylvinite beds in the Upper Kama deposit of potash ores where the rock formation support is made by means of interchamber (rib) pillars of various yieldability.

Jan 1, 1997

By Mark K. Larson

Researchers at the National Institute for Occupational Safety and Health?s (NIOSH) Spokane Research Laboratory are developing a borehole profilometer to dynamically measure the roughness pro-file of boreholes used to install roof bolts fully encapsulated with resin. Numerical measures of roughness were calculated from profiles of borehole surface roughness to determine potential roof bolt anchorage capacity. The concept of dynamically profiling was proven through the following series of laboratory tests: a sensitiv-ity test using various grits of sandpaper, a known-profile test along a section of a concrete hole, and a simulated borehole test in a pipe with machined grooves. Two profilometer prototypes were devel-oped that measure slight changes in the radius of a borehole using either strain-gauged, spring-steel arms or rigid, steel arms attached to a spring-loaded cone and a linear potentiometer. Both instruments performed reasonably well in the laboratory and were subsequently tested in boreholes drilled in two concrete blocks using different drilling techniques to obtain three different borehole surface roughnesses. A Z2 roughness parameter was calculated with data from the first profilometer, and an average radial increment was calculated from the test results of the second profilometer. Although both instruments were able to readily distinguish the difference between smooth and rough boreholes, the correlation coefficient (r2) for a linear relationship between Z2 and average radial increment was 0.577. Pull tests were conducted with resin bolts installed in the concrete blocks, but maximum pull-out strength could not be correlated to borehole roughness because in each of these tests, failure occurred at the resin-bolt interface. A more sophisticated prototype has been designed to measure surface roughness and also accurately determine the location and tilt of the instrument in the borehole. However, this instrument has not yet been fabricated or tested.

Jan 1, 2007

By Murali M. Gadde

Welded wire mesh is the most commonly used skin control technique in the United States underground coal mines. Currently, selection of the mesh is based on trial and error rather than any rigorous design methodology. Several factors affect a wire mesh?s performance including the amount of rock material to be supported and how the load is distributed, wire gauge, aperture size, grade of steel, weld strength, bolting pattern and sheet size. Although the best way to choose the mesh is to conduct full-scale physical load tests, because of the limitations on the number of variables that can be considered in such tests and the requirement of an elaborate test set-up, it may not always be possible to use this option. An alternative would be to use numerical modeling to study the load-deformation behavior of a wire mesh. In addition to providing a very detailed stress, strain distribution in the entire mesh, the advantage of this approach is the ease with which parametric studies can be conducted. In this paper it is shown that the load-deformation curves generated by nonlinear numerical models match closely with laboratory derived ones. Then, based on parametric studies, design guidelines are developed to choose the mesh appropriate for different loading conditions. The modeling studies also bring forth the limitations of the current lab studies. For example, under some conditions it was noticed that the small sheet size considered in lab tests would produce higher mesh deformations than a full-size one used in actual practice.

Jan 1, 2006

By Gary B. Petersen

The author has developed a seven-step process called the Applied Rock Mechanics (ARM) Protocol that the author has used to solve numerous roof stability problems in mines throughout North America since 1975. It is a practical evaluation process using trained observation, innovative data collection instrumentation, video borehole cameras, analytical processes, and controlled testing methods that have become highly effective in solving ground control problems in every type of underground mine. This paper will describe each of the steps of the protocol and how the rock mechanics engineer applies them to resolve a typical ground control problem.

Jan 1, 2012

By Erik Westman

Few studies have been conducted regarding the location, movement, and relative magnitude of the rear abutment of a longwall coal mine. The rear abutment, or the abutment pressure arch in the gob area, is controlled primarily by the quality of pack of the gob; the more compacted the gob is, the larger the stress would be theoretically. While there is no definitive location for the rear abutment, early studies showed that it could be located as far back in the gob as 300 m (1,000 ft). Like the location of the rear abutment, there has been little research conducted and very few answers as to the exact magnitude of the rear abutment load. Due to the fact that the rear abutment has been studied considerably less than the front or side abutment, the need for further studies is evident. From a ground control standpoint, the location and magnitude of a rear abutment could reveal information on the caving process associated with longwall mining, such as when the initial cave could be expected to occur in future panels. Besides the possible importance of the rear abutment from a ground control perspective, evidence related to the longwall rear abutment could be vital from a ventilation standpoint. The peak rear abutment location could serve as an area for harmful gases, such as methane, to accumulate because forcing air through repacked rock is more difficult than through loose rock. The objective of this study is to determine whether passive seismic tomography could image the location, movement, and relative magnitude of the rear abutment as the longwall face retreated, in addition to the forward abutment and gob. The results of a passive seismic imaging study at a western US longwall show evidence for the location, movement, and relative magnitude of a rear abutment. A clear zone of highlystressed strata is shown where the forward abutment is expected and a low-velocity, damaged zone is shown behind the face where the gob is located. These two zones retreat with the face over a three-week period. In addition, the images show a return to stress levels behind the gob that are slightly higher than overburden stress levels, which is presumably the rear abutment. This occurs at a distance behind the face that is approximately the same as the face width for this mine. We can conclude that passive seismic imaging can be used to identify the location, movement, and relative magnitude of the rear abutment of a longwall mine. This conclusion is dependent on the use of appropriate data collection and processing methods. The information provided can benefit the both ground control and ventilation engineers as they strive to improve the safety and efficiency of longwall mines. The knowledge associated with the rear abutment of a longwall could not only help increase production through proper initial planning and reduce down time, but it could also make positive strides in improving the overall safety surrounding a longwall mining operation.

Jan 1, 2012

By Thomas Z. Jones

Deep mining within the Sewell Coal member of the New River Formation in southern West Virginia has exposed several small-scale thrust faults and related features. The faults were en- countered in Westmoreland Coal Company's Eccles No. 6 Mine located near Beckley, West Virginia (Figure 1). The purpose of this paper is to: 1) describe the nature of these faults and the effect they have on mining, 2) discuss the roof support methods used in the fault zones, 3) establish a mechanism for the origin of the faults. and 4) demonstrate how the mechanism was used to forecast fault trends across the mine property.

Jan 1, 1986

By Haile Manteqi

In longwall coal mines, estimation of roof strata condition, sequence of loading, and supporting requirements are of a high degree of importance, and they are directly related to production and safety. Also, a prediction of the initial caving distance is rather complicated to calculate, because there are many parameters that effect caving process?roof and floor strata conditions, thickness of immediate roof, depth of cover, excavation geometry, and magnitude, and direction of principal stresses. This paper presents the results of 3D numerical modeling of the initial caving mechanism using the finite difference code FLAC3D at the E1 longwall panel of the Parvade 1 underground mine located in the Tabas area in the central part of Iran. The initial caving distance obtained was 10.4 m. Also, the results of numerical modeling have been compared to the analytical methods, such as Peng?s method. The results show good agreement with each other.

Jan 1, 2012

By Guoqi He

Common offset seismic data collected over two active longwall panels at different stages of mining reveal that significant arrival time delays are produced in reflections from different levels within the mine cover. Mining-induced arrival time delays are also observed beyond the edges of the mine over unmined areas. Reflection arrival time delays increase with reflector depth and are laterally variable. The arrival time delays are produced by reductions of seismic velocity within the overburden. Reflection arrival time delays from a major strong unit located neutrally within the cover are proportional to subsidence, however, delays in other events are not.

Jan 1, 1989

By Chris Mark

Successful longwall mining requires a stable tailgate entry. Gate entry performance is influenced by a number of geotechnical and design factors, including: - Pillar size and pillar loading; - Roof quality; - Floor quality; - Entry width; and, - Artificial support (primary and secondary). This paper describes a comprehensive, practical, design methodology, based on statistical analysis of a nationwide data base of longwall ground control experience. Geotechnical surveys were conducted at 44 U.S. longwall mines, and underground observations of site geology, entry conditions, and support design were recorded at each mine. The observations were combined with discussions with mine personnel to identify 69 longwall gate entry designs as satisfactory, unsatisfactory, or borderline. Only conventional longwall designs, in which the pillars are expected to carry the full abutment loads, were included in the data base. Designs which employed yield pillars only were excluded. The case histories were characterized using five descriptive parameters. Pillar design was described by the Analysis of Longwall Pillar Stability Factor (ALPS SF). A major new contribution is the Coal Mine Roof Rating (CMRR), a rock mass classification system that quantifies the structural competence of bolted mine roof. Other quantitative measures were developed for primary support, secondary support, and entry width. Multivariate statistical analyses indicated that in 84% of the case histories the tailgate performance could be correctly predicted using just ALPS and the CMRR. Most of the misclassified cases fell within a very narrow borderline region. The analyses also confirmed that primary support and gate entry width are essential elements in successful gate entry design. The relative importance of the floor and of secondary support could not be determined from the data. Based on these results, a simple equation was developed to guide the design of longwall pillars and gate entries: ALPS SF,, = 1.76 - 0.014 CMRR Where: ALPS SF, = ALPS SF suggested for design. Guidelines for entry width and primary support density, as related to the CMRR, are also provided.

Jan 1, 1993

By Dennis R. Dolinar

Screen material in the form of welded wire mesh and geogrids are used in underground coal mines to prevent the fall of small pieces of rock from the roof between roof bolts. Further, if the screen is installed during the production cycle, roof fall injuries can be reduced significantly. Therefore, the National Institute for Occupational Safety and Health (NIOSH), because of the safety implications, conducted an evaluation of screen materials commonly used in U.S. coal mines to determine their support characteristics and identify the parameters that could affect their performance with respect to controlling the fall of rock from the roof surface. To evaluate the load and stiffness characteristics of the screen, a test frame was designed and installed in the Mine Roof Simulator (MRS) at the NIOSH Pittsburgh Research Laboratory (PRL). With this test set up, the screen is bolted at four corners of the frame and a center load is applied. The test set up allows for up to 20 in of screen deflection while the load and screen defection are continuously recorded. A series of tests were conducted to evaluate the effects of various parameters such as bolt tension, the type of load bearing surface and the size of bearing plates on welded screen performance. The most common screen material used in U. S. coal mines is an 8-gauge wire welded in a 4-by-4-in spacing or aperture. The peak screen load was normally limited by wire breakage. The conditions at the bearing plate influence the nature of the wire breakage. However, the screen stiffness was controlled by slippage at the bearing plates, and by weld and wire failure. The size of the bearing plate had a significant effect on the performance of the welded screen. Therefore, the design capacity of the 8-gauge screen is evaluated based on plate size. For a 6-by-6-in bearing plate the average peak load is 2,900 lb with a stiffness of 250 lb/in. For an 8-by-8-in bearing plate, the average peak load is 4,500 lb with a stiffness of 430 lb/in. These test results are based on a 4-by-4-ft bolt spacing. A geogrid mesh was also tested.

Jan 1, 2006

By John Ellenberger

The nature of a stone deposit is rarely defined through exploration and engineering design, but more through the actual experience encountered as the mine develops; hence, issues with ground control are often at the forefront for a new mine. Ground control problems account for over 15 percent of lost-time injuries in underground stone mines. The objective of this work is to present actions taken during early development of a stone mine to mitigate the obvious effects of high horizontal stress on ground control. The mine was developed from a box cut with relatively little cover, and high horizontal stresses were not anticipated. Early recognition of possible high horizontal stress resulted in mapping of the mine using the NIOSH-developed Roof Fall Risk Index (RFRI) to provide a baseline to objectively analyze roof conditions. A program of borescoping the mine roof provided valuable roof beam thickness data and found that the roof beam was of minimal thickness, approximately 10 percent of drift width. As the mine developed, it became apparent that ground control problems related to high horizontal stress existed. A concentrated effort was made to improve ground conditions by implemeting a number of mitigation techniques. Actions taken by the mine included systematic roof support, reorientation of mining direction, a lower roof horizon, reduced span widths and reduced mining height in crosscuts to prevent running falls. These strategies resulted in signifcantly improved ground stability.

Jan 1, 2012

By David F. Gearhart

The Burrell Can1 is a thin, steel, tubular shell filled with aerated concrete that is used as a roof support in coal mines. The Can height is always shorter than the mining entry, so it is capped with wooden timbers and wedges when it is installed. This combination of the Can and the wooden header is a standing support system that is designed to yield in a controlled manner and exhibit a constant yield, or elastic-plastic behavior. The Burrell Can is capable of withstanding 20 in or more of vertical convergence and 15 in or more of horizontal displacement. Because of this exceptional ability to sustain load through extended convergence and displacement, it is one of the most robust standing supports in use today. The National Institute for Occupational Safety and Health (NIOSH) conducted testing on seventeen Burrell Cans in 1995, during development of the support where the ratio of the height to the diameter, or the aspect ratio, was between 2:1 and 4:1. Based on the results of that testing, it was suggested that the standard aspect ratio of the Can be less than 5:1. This suggestion is intended to preserve the overall stability of the Can, which will help to maintain the initial stiffness and prevent load shedding. Recently, the necessity of the aspect ratio has been questioned, and this issue will be addressed in this paper. Since the initial development and introduction, more than 130 tests of the Can have been conducted at the NIOSH Mine Roof Simulator (MRS), and 25 of those specimens had aspect ratios greater than 5:1. The 18-in-diameter by 10-ft-tall specimens had the largest aspect ratio, 6.67:1.However, the aspect ratio is not the only parameter that affects the performance characteristics of the Can. Multiple factors, such as the use of different wood species, the use of different header or footer configurations, and the use of pedestals, as well as biaxial loading conditions can also decrease the stiffness and capacity of the Can. Data from full-scale tests that include these other factors are analyzed to show that the initial stiffness is primarily affected by the roof and floor contact configurations. Furthermore, load shedding is more likely to occur or begin earlier when the aspect ratio is greater than 5:1 and one or more of the other factors are present.

Jan 1, 2012

By L. R. Stace

Arguably, one of the most significant technological breakthroughs that had been experienced during the period when British Coal was the major coalmine operating company in the UK, the use of rockbolts as principal support in underground roadways brought with it the need to re-assess the technical knowledge, management controls and legislative frameworks associated with the support of mine workings. Following a major roof fall in a roadway supported principally using rockbolts in which fatalities occurred, the process of the design of rockbolted support and the systems and procedures that had been put in place to control its use in the UK were further reviewed. The paper discusses the outcome of these reviews, and how the incident initiated a major new interest in strata control research in the UK against the background of the fragmentation of the ownership of the mines following the sale of the former British Coal assets at the end of 1994, together with the proposed introduction of new strata control legislation in 1996.

Jan 1, 1996

By Da Zhao Gu

An important problem in rock engineering is to control the properties of artificial rock. A compaction technique for casting artificial rock specimens with controlled properties is described, i.e. using a compression-testing machine to compress the mixture. Uniform density is necessary to achieve uniformity of properties of the specimens, and uniform density can be obtained by using the specific pressing method presented. Rebound and borax content are important to the technique and have been studied as well.

Jan 1, 1992

By Alan A. Campoli

The mining industry has long been searching for an effective structural application that prevents weathering and falls of loose rock, debris, and sediment. J-Crete is a fiber-reinforced, engineered, cementitious composite rock application that is designed, tested, and successfully utilized as such a material. Laboratory evaluation of J-Crete material properties has shown the final compressive strength to be in excess of 7,300 psi (50.34 MPa), with 24 hour strength in excess of 1,500 psi (10.34 MPa). The flexural tensile strength is in excess of 600 psi (4.1 MPa) and is provided through a unique combination of polyvinyl alcohol (PVA) fibers activated by a proprietary liquid additive that prevents them from clumping and enables even distribution throughout the grout. Tensile bond strength in excess of 350 psi (2.4 MPa) at 28 day cure tightens the strata and holds loose material in place. Practically impervious to gas and liquids, J-Crete protects strata against long-term weathering, does not require screening to be effective, and has proven to be a successful waterproofing material. J-Crete has demonstrated the potential to alleviate coal mine roof cutters in soft, wet clay formations. J-Crete reinforces coal mine ribs by bonding directly to the rib, conforms to the strata, strengthens the rib, and eliminates sloughing due to long-term weathering; stabilizes coal mine entries by bonding directly to the brow and roof, and creates a monolithic shell with double reinforced corners.

Jan 1, 2012

By Mingju Liu

China has been suffering from coal and gas outbursts since 1950s. Nowadays, the situation has become worse than ever with the need to mine deeper coal deposits at higher production rates in a more complex underground mining environment. In the year 2006 alone, 37 outbursts occurred in underground coal mines with a death toll of 232. This paper describes a case study of an outburst event, which occurred in July 2006 during the shaft development of the Mengjin Coal Mine, Henan Province, China at a depth of 768m below surface. Detailed analysis of the possible causes that led to the occurrence of the outburst is reported. These include the study of complex geological features, gas storage conditions, possible improper technical management practice as well as incomplete prevention and control measures. Based on the case study and the observation of outburst occurrence in China in the year 2007, suggestions have been put forward to help prevent and control future coal and gas outbursts.

Jan 1, 2007

By Gabriel S. Esterhuizen

This paper addresses the need for a method to compare the effectiveness of different support systems when designing ground support in coal mines. At present, support design methods include empirical methods based on observations of past performance of installed support systems, analytical methods where the roof is typically simulated by elastic beams and numerical model analysis. The approach presented in this paper estimates the relative stability of a support design through geotechnical evaluation of the rock mass and numerical model analysis of the interaction between the rock mass and the support system. Models are used first to simulate the design performance at the expected rock conditions. The rock strength is then reduced until collapse is indicated in the model. The stability factor is then calculated as the ratio of the expected rock mass strength to the rock mass strength at the onset of collapse, and is similar to the well-known factor of safety used in engineering practice. The stability factor can be used to assist in developing a final support design by comparing the effectiveness of various support systems and the stability of excavations under various geological and loading conditions. Examples of the method?s application in three different geological settings are presented.

Jan 1, 2012

By V. M. Shick

For practical purposes of the use of information resources in the field of geomechanics and mine surveying with accepted mining technology in mines, the authors have developed the packaged software programs of the survey-geomechanical information system with database objects of mine field. This system enables to execute the graphical representation of workings layout in preset scale, axonometric projection, the zones of higher ground pressure: program packages for calculations of rock pressure manifestations and for selection of suitable support facilities for the permanent, development and active workings and vertical shafts based on the information resources of the State Research Institute of Mining Geomechanics & Mine Surveying (VNIMI). The program packages are applied at coal mines in Russia and in Kazakhstan. The profitable operation of coal mines can be reached only by the combined optimal engineering decisions in mine working's laying-out and their reliability and safety at the stages of their designing and exploitation, and with minimal harm of the environment as well. This problem is being complicated by the fact that today the mining-practicing engineers and designers are ordinarily facing the problem " of an information barrier", when the complexity of controlling systems exceeds the potentialities of traditional management of production practice. Coal mine is a typical representative of complicated systems, because the information about geological structure, physicomechanical properties and stress-strain state of rock mass are of fragmentary character and insufficiently reliable. A mine has a developed network of interrelated workings, by the way, some of them are productive ones being subjected to control to a certain extent, and some of them are abandoned ones being noncontrollable and hazardous in rushes of water, gas, pulp and fire. The ground pressure forces cause the disintegration of rock mass growing with depth. The progress of mining operations constantly changes the geomechanical situation forming the zones of bearing pressure, stress relieving, coal falling, as well as it leads to the pulp movement on the earth surface, breaks the ground water flow, causes deformations and changes in stability of buildings, constructions and natural objects on the earth surface. One should take into account the existence of hazardous zones within the mine field, i.e. the zones of elevated ground pressure, geological dislocations and those ones under natural water reservoirs, ponds, etc. Major coal seams are prone to rock bursts, sudden outbursts, bleeding, endogenic fire. It is obvious, that in practice it is extremely difficult to account for the mentioned above factors in their interrelation.

Jan 1, 1996

By John Latilla

Exxaro Mining?s Matla Coal situated in Mpumalanga, South Africa operates two shortwall faces on the No 4 and No 2 seam while the No 5 seam has also been longwalled in the past. Severe strata loading events have lead to a number of serious roof falls on the wall face (facebreaks) resulting in safety threats as well as considerable production losses. In 2002 the No 4 seam shortwall suffered a facebreak restricted mainly to the tailgate side of the mining panel due to a thick, near surface dolerite sill not breaking properly. This was partly due to the premature suspension of longwalling in the overlying No 5 seam. The resultant overloading lead to severe pillar damage and subsequent bord collapse. Surface pre-split blasting was successfully used to relieve the pressure as well as to pre-condition the dolerite above the following two panels. The shortwall on the No 2 seam has experienced mid-face loading due to No 4 seam chain pillars lying over this area and creating a secondary cantilever effect. Pre-split blasting of the solid rock above the No 4 seam pillars has resulted in a significant reduction in the number of facebreaks. The panel layout of the No 2 seam was subsequently changed so that the chain pillars on both the No 4 and No 2 seam are superimposed whereas previously they were offset.

Jan 1, 2007

By Dwayne C. Kicker

This paper introduces the concept of design theory and methodology as applied to rock mechanics for more innovative and efficient design. This is a "frontier" research area in rock mechanics because although design is a fundamental foundation to all engineering branches, very little attention has been paid to this aspect in mining. In discussing the theoretical premise of design, this paper identifies the principles which form the basis of design activity. Also, from a practical viewpoint, the findings from a series of interviews with mine design engineers are presented. These findings lead to the identification of needs arising from overlooked design methodology in practice.

Jan 1, 1989