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By Raymond M. Stateham

The Bureau of Mines has developed a portable, lightweight device for determining the integrity of resin-grouted roof bolts. The instrument functions by sending a known pulse of ulstrasonic energy into the bolt and comparing the amount of energy reflected hack with the original measured pulse. Where the bolt is properly bonded or coupled with the rock, the energy passes through the bolt-resin-rock interface and dissipates in the rock. Where the bolt is not properly coupled, the energy is reflected back along the bolt to the instrument. The ratio of reflected energy to the pulse approximates the fraction of the bolt's surface that is improperly bonded. The instrument is easy to use. The operator holds the sensor head of the device firmly over the head of the bolt to be tested, causing the pulse of ultrasonic energy to be imparted to the bolt. The device then listens for reflected energy. This cycle is repeated about 200 times per second until an evaluation of the bolt installation is made and indicated by colored lights in the display window. Bolts having 50% or less of their surface bonded are considered had and cause a red light to he illuminated. Bolts with 50% to 75% of their surface bonded cause a yellow light to be lit and are considered marginal. Bolts having more than 75% of their surface bonded are considered good and cause a green light to be lit. The indicator lamp remains on for about 4 seconds and then goes off, indicating that the instrument is ready for another test.

Jan 1, 1982

By Filippo Vallianatos, Michael Verigakis, Vasili Saltas, Kosta Kaklis, Stelio Mavrigiannakis, Zach Agioutantis

"Development of failure in brittle materials is associated with microcracks, which release energy in the form of elastic waves called acoustic emissions. This paper presents results from acoustic emission measurements obtained during three-point bending tests on Nestos marble under laboratory conditions. Acoustic emission activity was monitored using piezoelectric acoustic emission sensors, and the potential for accurate prediction of rock damage based on acoustic emission data was investigated. Damage localization determined based on acoustic emissions evident in the critically stressed region in the form of scattered events or stresses below and close to the strength of the material.IntroductionNon-destructive testing (NDT) has gained popularity in recent years due to commonly available sensors, devices and software tools. Although the majority of such testing still occurs under laboratory conditions, the ultimate goal is a direct field application, especially in the case of underground mining. NDT, if properly applied, may be used to directly provide material property values and characteristics that can be used to determine rock mass or soil mass behavior, with a direct impact to mine design (Iannacchione et al., 2003).AE is generally defined as the high-frequency transient elastic waves originating from the sudden release of energy at localized points within a loaded material (Nomikos et al., 2010). The field of Acoustic Emissions (AE) due to compressive, tensile or other types of loads is a promising research area due to its wide application. The accurate identification of failure precursors stemming from AE is still an open research topic (Lei et al., 2004).Rock is a typically inhomogeneous and anisotropic material that contains several natural defects with various scales such as grain boundaries, microcracks, pores and joint inclusions (Jing, 2003; Read, 2004). Damage induced by microcracks is an essential mechanism in many brittle rocks subjected to stresses. An applied stress induces microcrack growth and damage is distributed uniformly throughout the sample. Finally, the microcracks coalesce and the damage is localized to a shear band, leading to strain softening and macro-scoping failure (Dragon et al, 2000)."

Jan 1, 2015

By Yongping Wu

The steeply dipping coal seams in this paper refer to those with the angle of inclination ranging from 35° to 55°. The reserves of the steep coal seams account for approximately 15%~20% of the total coal reserves in China. In the past 12 years, the top-coal caving mining method had been used to mine the steeply dipping thick (11.5-16.4 ft or 3.5-5.0m) coal seams. However, the mining process of top coal caving is complex with low recovery rate and nearly 30% of coal was wasted. Those problems can be solved by using the single pass large mining height (11.5-16.4 ft or 3.5-5.0m) method. But the movement of the overlying strata around the mining face area of large mining height is complex and it is hard to control the stability of ?rock-shield support? system. Therefore, field measurement, numerical simulation and physical (similarity material) simulation methods were used simultaneously to analyze the strata movement, roof structure and rock-shield support interaction characteristics. The results indicate that strata movement, which is the same as mining the steeply dipping seams with conventional mining heights, is asymmetrical about the dipping direction of the face; deformation, failures, and movement of the surrounding rocks are more intense; the roof weighting interval decreases; the weighting intensity increases significantly; and face sloughage occurs frequently. In the gob, the space for sliding and rolling of the roof increases; the broken main roof tends to form an anti-dip pile structure; the backfilling and compactness of broken roof increase at the lower portion of the face, thereby inducing significant unbalance in stress distribution between the upper and lower portions of the face. The overlying strata above the face form a multi-level step structure. The characteristics of contact and loading of roof-shield support system are more complex. The range of shield load variation is larger, and interaction between shield supports occurs frequently. Anti-toppling and anti-sliding devices at the face become more difficult to implement. The research results provide a theoretical basis for the mining practices in the steeply dipping thick seams with a large mining height. The practice in panel 25221 of 2130 coal mine indicates that the accident rates have been greatly reduced and good technical and economic benefits have bee realized.

Jan 1, 2014

By Madan M. Singh

Currently longwall coal mining operations require at least 3 entries on both the headgate and tailgate ends because of ventilation and safety requirements. Driving of entries, however, is a slow and costly process. Often it is a bottleneck in longwall development. Single entry gateroads are permitted in Europe and commonly used there. This study entailed ascertaining the feasibility of using a single entry in the United States, without sacrificing safety. This design was conducted for an operating longwall mine in the Black Warrior coalfield in Alabama, An understanding of rock loads and consequent support requirements is prerequisite to the design and successful demonstration of the single entry system. Efforts in this regard were directed at determining center (packwall) support requirements for various single entry widths. Critical to the analysis is knowledge of the nature and extent of side abutment stresses. In collecting geological and geotechnical information relating to the stability of the single entry, Engineers International, Inc. engaged in the following: ? Examination and inspection of the single entry demonstration at the Sunnyside Mine in Utah. Severe ground conditions and unacceptable levels of entry closure were experienced there prior to the tailgate phase. Observations at the mine presented a clear perspective on possible ground control problems for future single entry designs. ? Reconnaissance inspection and classification of roof rock conditions. ? Detailed mapping of roof structure in the areas to be instrumented. ? Installation of convergence stations in the longwall headgate to monitor roof movement during mining. ? Measurement of side abutment stresses in a yield pillar in the area of the convergence measurements. ? Measurements of coal fracture distribution by the detail area method. ? Rock properties determinations of roof, coal, and floor rocks; samples obtained by core drilling. ? Determination of floor bearing capacity by in-situ testing.

Jan 1, 1982

By D. L. Price

This paper presents a method of measuring field power consumption of shearers and deriving specific energy consumption versus shearer haulage speed performance curves. A description of a computer program which simulates the cutting action of a shearer drum is presented. A comparison between the field specific energy curves and the computer simulation results is made. The simulation results compare well with the field data where all the coal seam is being cut by the shearer. The computer simulation results give hack-calculated field pick force cutting characteristics, which allow the performance of the shearer to be predicted for other operating conditions, e.g. effect of changes to pick lacing, effect of changing the drum rotational speed, allows the calculation of power requirements for a particular production rate.

Jan 1, 1992

By James H. Fletcher

Much data and experience has been accumulated, especially in the last 5 or 6 years, to show that modern roof trusses, both of the Birmingham type and the bolt-and-channel type, perform well to support certain kinds of coal mine roof. For some time now, roof trusses of the Birmingham type have been effective in controlling thick shale roof, which tends to sag and fall over a period of about eight weeks. These falls leave dame-shaped cavities which are reminiscent of the pressure arch theory of roof failure. Lately there has also been significant interest in supporting some longwall panel entries with channel and-bolt systems. These systems utilize two vertical center bolts and two 45 degree end bolts through a heavy gage channel. Both types of truss systems have long bolts, which require 8 to 9 Foot holes on a 45 degree angle, and point anchor resin bolts, which utilize a 2 or 3 foot rebar grouted in a thin annulus of polyester resin. Consequently, many coal mining people have been impressed with the very good anchorage characteristics and tensionability features of these systems. However, the same people have been proportionately dismayed to find that very often the most difficult part of introducing a trussing program to a mine is deciding on the machine specifications which determine the truss installation rate. The purpose of this paper is to indicate aspects of truss bolter design and use which are not shared by roofbolter design and its vertical pattern of bolts. Of the two types of drill units conventionally accepted, arm feed units and mast units, only one 15 judged suitable. While the arm feed unit offers low collapsed height with a long drill feed length, a great attribute for low seam use, it has not lent itself to being rolled over to a 45 degree angle. One reason for this inability is that many items which are conventionally mounted on the arm feed unit cannot be allowed to rotate with the unit. These items include drilling controls, tool Cray, lights, canopy or safety post, panic bar, and stab jack(s). If these items were independent of the arm feed drilling unit, they would have to be out of the way of the boom arms, links, and feed cylinder as the unit moves. To ease this problem, the arm feed unit can be made very short and even mounted across the machine. However, the resulting short feed length is undesirable. Several short drill steel extensions are required to drill a relatively long hole, making this procedure time-consuming as well as increasing the likelihood of drilling a crooked hole. Another reason for abandonment of the arm feed unit is that a very stiff drill guide (or "centralizer") is required to start the hole. Using the mast tilt control, the roof bolter operator will force the bit against the roof and will rotate the bit without feeding to cut sideways into the roof to form a hole face. Omission of this step often results in the skidding of the bit toward the rib. To date, no such sturdy drill guide has been designed for an arm feed drill unit. The mast style unit has been the most successful design for angle drilling. It is easily rolled over to a 45 degree angle or to 90 degrees and, on single head models, often has been made to roll through 180 degrees. The mast style is compact, in plain view and is easily fitted with heavy duty manual or hydraulically operated drill guides. It can be mounted directly on a machine chassis, or onto a swinging boom, or onto a slider arrangement in a broad variety of configurations which will accommodate controls, tool trays, stab and roof support jacks, lights, canopies, and panic bars. The only significant difficulty with mast drilling units occurs in lower mining heights when desirable low mast height is accompanied by undesirable short feed length. While the lead extension in the drill string can be as large as the mining height will allow, a short feed length will permit only the use of shorter middle extensions and drivers, and perhaps will. require the use of two different length wrenches. The associated problem is the relatively short distance from the drillhead to the drillguide. This distance is further shortened when raising the drillhead to get the bit to the roof. One might keep longer lead extensions on the machine; drill a vertical pocket for the bit, starting the angle hole at this pocket; or modify the mast. If the mast is made so that the distance from the drillhead to the drillguide remains constant while the mast extends, after which the drillhead travels up the mast, the shorter mast can be made to start the hole more easily. However, short teed length will still result in longer drilling time. It is obvious from these points that a proper choice of mast is simply the highest one which will be moved easily in the given mining height. This, of course, goes against the grain of most experienced coal mining people who realize that a relatively small decrease in mining height can create a problem correctable only with the purchase and installation of shorter masts. If the mining

Jan 1, 1982

By André Zingano

Pillar collapses have been studied for several years and can be classified into two types: nonviolent squeeze or violent pillar collapse, i.e., controlled or uncontrolled pillar collapse. Underground coal mines in Brazil are considered shallow mines, because the overburden thickness varies between 20 and 400 meters. For this reason, the coal pillars should be designed as permanent pillars to avoid subsidence and groundwater problems. In the last two years, various pillar collapses occurred, one being a violent collapse and the others being squeeze collapses. All these collapses were considered massive pillar collapses because many pillars were involved. In the violent pillar collapse case, the width¬to-height (w/h) ratio was less than three; although, the pillar safety factor (SF) was more than 1.3. There was a collapse of about 100 pillars in less than three hours. The objective of this study is to determine the mechanism of pillar collapse for this violent pillar collapse. Convergence monitoring and numerical modeling were applied to combine field data and theory to simulate the pillar collapse. Convergence field measurements introduce the arc-effect behavior for the roof. The results showed that other factors besides pillar geometry caused the pillar failure. Other parameters, mainly dip of the coal seam and presence of fractures and faults may affect the strength and failure mechanism of a pillar.

Jan 1, 2004

By C. P. Mangelsdorf

The success of the Birmingham roof truss (Figure 1) in supporting some difficult roof conditions, particularly in the Illinois coal basin, has given impetus to the development of a number of alternate schemes. These schemes attempt to alleviate two major problems with the Birming-ham truss, namely, the time required for installation and the cost of material. Because of the difficulty of drilling angled holes and the number of different pieces of hardware which must be handled manually, most operators have been unable to install Birmingham trusses in the cycle and have had to schedule truss crews during the maintenance shift or whenever they would not impede production. Due to this delay the truss has traditionally been viewed as a supplemental form of roof support. In fact, the truss might qualify as primary support and replace at least two bolts if it could he installed in the cycle (and MSHA would accept it as an alternate in the roof plan.) [ ] Figure 1. The Birmingham truss. Two attempts to deal with the first difficulty have sought to streamline the truss hardware while preserving the essential character of the Birmingham truss. One of these, by Peabody Coal, is reported elsewhere in these proceedings by C. Bollier. Another is a contract issued by the Bureau of Mines to Bendix Corporation for the development of a machine, particularly suited for use in low coal, which will automate many of the operations currently executed manually. This machine is expected to be ready for field testing by fall of 1982. Once in place, a truss installed by either of these schemes will interact with the roof in the same way as a conventional Birmingham truss. An alternate approach has been proposed by several different manufacturers independently. Although the hardware details differ. the essential idea consists of two angle bolts, the heads of which are connected by a horizontal chord tightened to some pre-determined tension. Figure 2 shows the generalized concept and does not depict any one of the four systems proposed. The advantage of this scheme is that the angle bolts can be installed and tightened in the cycle, and as bolts, immediately satisfy the roof control plan. The horizontal chord can be added later to achieve almost any initial desired roof loading, the limits of which are defined below this scheme is referred to here as the angle-bolt truss. Because of a fundamental difference in the way in which load is transferred between chords, this design does not interact with the roof under subsequent loading in the same manner as a Birmingham truss. The purpose of this paper is to examine this fundamental difference and its consequences. [ ]

Jan 1, 1982

Design and selection of support systems for longwall faces call for in-depth knowledge of strata mechanics and in particular of strata-support interactions. The paper presents the results of field investigations and analysis of a difficult-to-cave roof using powered supports under a massive dolerite sill in Central India using a state-of-the-art datalogger system. The mechanised longwall face was equipped with 4/680 tons chockshields operated on the IFS mode using a 300 K W DERD shearer. The datalogger system was capable of storing data simultaneously in 32 channels. The paper describes the monitoring system and the monitoring programme. The system provided useful information on support behaviour, build-up of leg pressures and face convergence during an intense dynamic periodic weighting phenomenon which destroyed the integrity of almost half of the face supports. At this point, the face had advanced by about 197 metres from the barrier. The paper outlines a post-mortem analysis of this dynamic event and provides a critical evaluation of the data collected for understanding strata support interaction and the behaviour of the massive sill in the overburden.

Jan 1, 1992

By Michael Karmis

During the last 25 years, technological advancement and subsidence research have resulted in more accurate and diverse prediction capabilities. The work presented in this paper focuses on the development of integrated prediction capabilities for dynamic deformation indices and for strains over sloping terrain. In order to assist subsidence engineers with accurate prediction methodologies, the research also addresses the development and validation of techniques that can enable model calibration using alternative measured subsidence parameters such as horizontal or ground strain instead of the traditional vertical subsidence values. This paper presents the basic concepts and validation of the above mentioned enhanced prediction and control methodologies, which are necessary for effective assessment and control of mining-induced subsidence, using examples and case studies. In addition, a risk assessment approach for evaluating landscape stability over mined areas is presented. The enhanced prediction methodologies have been incorporated into the Surface Deformation Prediction Software (SDPS) package.

Jan 1, 2008

By Michael Gauna, Christopher Mark

"Underground coal mining in the U.S. is conducted in numerous regions where previous workings exist above and/or below an actively mined seam. Miners know that overlying or underlying fully extracted coal areas, also known as gob regions, can result in abutment stresses that affect the active mining. If there was no full extraction, and the past mining consists entirely of intact pillars, the stresses on the active seam are usually minimal. However, experience has shown that in some situations there has been sufficient yielding in overlying or underlying pillar systems to cause stress transfer to the adjoining larger pillars or barriers, which in turn, transfer significant stresses onto the workings of the active seam. In other words, the overlying or underlying pillar system behaves as a ""pseudo gob."" The presence of a pseudo gob is often unexpected, and the consequences can be severe. This paper presents several case histories, summarized briefly below, that illustrate pseudo gob phenomenon:• Pillar rib degradation at a West Virginia mine at 1100-foot (335 m) depth that contributed to a rib roll fatality.• Pillar rib deterioration at a Western Kentucky mine at 570-foot (175 m) depth that required pillar size adjustment and installation of supplemental bolting.• Roof deterioration at an eastern Kentucky mine at 1300- foot ( 400 m) depth that stopped mine advance and required redirecting the section development.• Coal burst on development at an eastern Kentucky mine at 1700-foot (520 m) depth that had no nearby pillar recovery.• Coal burst on development at a West Virginia mine at the relatively shallow depth of 1100 feet (335 m) that also had no nearby pillar recovery.The paper provides guidance so that when an operation encounters a potential pseudo gob stress interaction the hazard can be mitigated based on an understanding of the mechanism encountered.INTRODUCTIONThe U.S. underground coal mining industry has conducted mining in multiple seam environments, where the active mining has underlying or overlying old workings at varying interburden distances, throughout its history. Often no serious consequences arise from the multiple seam mining. However, sometime mines have been confronted with hazards from underlying or overlying workings that include localized roof and rib failure, pillar system failures through propagating roof falls and floor heave, and also pillar bursts."

Jan 1, 2016

By R. Fritschen

In order to investigate locations, source parameters, and source mechanisms of mining induced seismic events, a local network of seismic stations was installed at a coal mine in the eastern part of the Ruhr District (Germany). The recorded events were located with the help of an anisotropic velocity model, which accounts for the fine layering of the carboniferous strata. Most of the events took place about 50 m above the working Luise 76-10. Their locations correlated with the advancing working face, and the number of events increased when pillar edges were undercut. The source mechanism, i.e., the orientation of the fracture plane and the direction of displacement on this plane, could be determined for several seismic events. The majority of them was characterized by a fracture plane parallel to the working face and an almost vertical dislocation on this plane towards the goaf a mechanism resembling the breaking down of the hanging wall. The fracture plane had a dimension of about 100 m. Probably the fracturing of a sandstone layer, located approximately 40 m above seam Luise, was responsible for the observed seismicity

Jan 1, 1999

By J. L. Condon

The Bureau of Mines, through in-house and contract research, monitored mountain bump-prone areas of the Olga #2 Mine, near Welch, WV, using microseismic techniques for 15 months during 1985 and 1986. During this period, two ground failures occurred that were considered "mountain bumps" owing to excessive stress buildups. One failure damaged the roof and 4 pillars around an intersection; the second failure, which released much greater energy, resulted in significant damage to over 100 coal pillars. Analysis of the microseismic data collected provided significant insights into the magnitude and location of each failure. The actual failure zone was clearly delineated prior to the mountain bump occurrences through analysis of source locations of each microseismicevent. Trends in cumulative rate plots indicated the onset of structural instability. Although the data could not be analyzed automatically, the significance of the data interpretation cannot be overemphasized. Microseismic monitoring of bump-prone areas can be successful in providing a means to actually observe the structural stability occurring in underground coal mines. This paper discusses data collection techniques, analysis methods, and implications of the monitoring technique, with special emphasis on actual data from two large mountain bumps in West Virginia.

Jan 1, 1987

By Bruce K. Hebblewhite

Tower Colliery is a longwall mine operated by BHP Coal Illawarra Collieries, Southwest of Sydney, Australia It mines the Bulli Seam at a depth of approximately 450m. The surface topography overlying the mine consists of several steep-sided river gorges, up to 68m in depth, which run at oblique angles across a sloping terrain Apart from some light rural development, the surface land is essentially natural bushland, but is traversed by a major freeway. This crosses one of the gorges on twin, six-span, box-girder bridges, with bridge piers up to 55m in height. Consequently, a major surface subsidence monitoring program has been in place for several years now, including intensive conventional, GPS and E DM surveying, plus real-time monitoring of critical components of the bridge structure. Although the bridge and freeway are outside the conventional angle of draw' subsidence influence criteria, and have seen only negligible vertical deformation as a result of mining, there has been widespread evidence of regional horizontal deformation of the land surface, large distances away from the mining area The effect at the bridge has been well within acceptable tolerances, primarily because of the regional nature of the movements, with very little differential movement observed at the bridge piers Gorge closure and evidence of large headlands of land moving `en masse' have been observed. Horizontal movements of hundreds of millimetres have been recorded in some locations where gorge closure has been monitored. Possible explanations of these movements include one or a combination of mechanisms such as pre-mining stress relaxation. valley notch effects, valley bulging, regional joint patterns, movement toward active goaf areas, vertical gorge shearing' and shear failure of horizontal bedding planes below the surface This paper presents a case study of the regional horizontal ?subsidence displacements? measured during mining. and a discussion of the possible explanations for these phenomena

Jan 1, 2000

By Richard C. Repsher

The U.S. Bureau of Mines has developed a method and device designed to detect lose rock material in underground mines. The technology is designed to be an aid to mine workers in detecting hazardous roof conditions in underground mines which can complement or replace the traditional roof sounding techniques where the miner relies on experience to determine whether rock conditions are sound. The leading cause of accidents and fatalities in underground mines is falls of lose rock pieces or rock slabs from the mine roof. The device is an electronic roof sounding device that acoustically determines the integrity of mine rock. In previous research the Bureau of Mines found that loose rock, when impacted, vibrates at a much lower frequency than intact rock material. A major problem in determining rock stability using this technique has been the repeatability of the matt signal. This difficulty baa been greatly reduced in the current design by measuring the power spectra contained in two separate frequency bands of the signal produced by striking the rock in question. The ratio of the energy contained in each band is computed. This process minimizes any striking force differences, producing accurate, repeatable results for solid rock as well as loose, drummy material. The prototype has been successfully- tested in a variety of underground environments including coal, uranium, molybdenum, silver and salt. The technology has been investigated by the U.S. Mine Health and Safety Administration and the Department of Energy for use in detecting detached tunnel lining areas in nuclear repositories. The paper will discuss the technique, applicable results, and future applications.

Jan 1, 1990

By Dennis R. Dolinar

Roof falls, even of supported roof, still constitute a major hazard in underground mines. However, associated with any fall or instability is a pattern of roof movement. Therefore, the National Institute for Occupational Safety and Health, Pittsburgh Research Center, has been conducting research into the development of practical extensometer systems that can be used to measure this displacement and to determine if a roof fall may occur, thus increasing miner safety. For extensometers to be used in the detection of potential roof falls in coal mines, an understanding of the movement and displacement rates that will lead to roof failure is required. Because roof falls are unpredictable, it is often very difficult to obtain this information. Therefore, in a coal mine in the lower Kittanning seam in Pennsylvania, a room widening experiment was conducted to intentionally cause a roof fall while roof movement data was collected. In this experiment, two adjacent rooms 5.5 m (18 ft) wide supported with 1.5 m (5 ft) resin-grouted bolts and instrumented with extensometers were widened to 9 m (30 ft). In one of the rooms, 3.6 m (12 ft) cable bolts were installed as supplemental support. In the room with no additional support, a roof fall occurred 72 h after the room was widened. Much of the 3 cm (1.2 in) of movement in this room resulted from the instantaneous and transient response to mining with displacement rates up to 7.6 cm/d (3 in/d) being measured. This was followed by a steady state displacement phase with a rate of 0.43 cm/d (0.17 in/d). Just 12 h before the fall, movement deeper in the roof caused the rate to accelerate to 1.1 cm/d (0.44 in/d). The roof in the room with the supplemental cable support did not fall even though the total roof deformation was over 7.6 cm (3 in). Significant roof movements and strains in the cabled room were detected to a depth of 3 m (10 ft) while maximum cable loads were over 180 kN (40,000 lbf). Essentially, the roof was in postfailure with the cables maintaining the material, while roof stability was indicated by a steady state displacement rate of only 0.025 cm/d (0.01 in/d) at the end of the test.

Jan 1, 1997

By Gheorghe Oncioiu

The phenomena developed on the influence of thick coal seam no.3 mining, by horizontal slices and rocks caving roof control, in the Jiu Valley coal basin, are very complex. By analyzing the influence volume of underground faces and defining the stress and strain zones, around the faces, it is possible to evaluate the suitable powered support performances. Also, Jiu Valley thick coal seams involve about 50% of reserve balance in the main safety pillars of social, industrial and natural objectives, situated on the mining surface perimeters. For a correct design and drawing of main safety pillars complex research was done with the aim of establishing the subsidence parameters. Upper face rocks medium could be assimilated as a granular medium, which is in a permanent movement. For describing the rocks movement, respectively the ground surface displacement, was taken into account the stochastic mechanics principles. The theoretical solutions are in accordance within situ measurements.

Jan 1, 1999

By Kevin Jinrong Ma

Underground mines often experience roof falls in entries, crosscuts, and intersections of active mining sections, main travel ways, and belt entries. Roof fall heights greater than 20 ft (6 m) make re-bolting of the newly exposed roof dangerous and impractical. To protect mine personnel, belts, moving vehicles, and other facilities, mine operators typically install steel structures such as square or arch sets in the roof fall areas. Wood lagging is usually installed between the steel-sets to enclose the area and protect the entry from recurring falls. Usually the void space above the steel-sets and laggings is backfilled. However, backfilling high roof fall voids is costly, which causes some operators to leave the voids open. In this case, durability of wood over time and the capability of the steel-set and wood lagging to resist falling rocks are unknown. During the last few years, Keystone Mining Services, LLC (Jennmar Corporation affiliate engineering company) designed, tested, and developed an impact-resistant lagging system to protect steel-sets1. Various steel-sets protected with the impact-resistant lagging were approved by MSHA and installed in different underground roof fall rehabilitation projects. This paper presents (1) design, development, and laboratory testing of the patent-pending impact-resistant lagging panel, (2) steel-set design methodology according to the American Institute of Steel Construction (AISC) national standard, and (3) impact-resistant steel-set design method based on Law of Conservation of Energy. Two underground cases are reviewed to demonstrate the successful application of the impact-resistant lagging system.

Jan 1, 2011

By Ashok Kumar, Rajendra Singh, Arun Kr. Singh

"A simple analysis of site conditions and capability of the mechanised depillaring (MD) operation found that this approach is suitable to address the strata control issues in coal pillar extraction under the weak and laminated roof strata of Pinoura mine. A continuous miner (CM) with a ram car and quad-bolter are used for single pass pillar extraction (first time in the country) to achieve a relatively faster rate of depillaring. A total of eleven panels are depillared under weak and laminated roof strata with a fair-rated rock mass rating (CMRI-RMR). It is observed that there is less influence of roof caving over the surrounding mining structures, especially those located out-bye the side of the extraction line. An attempt is made to quantify the goaf influence by field studies, which show that the maximum value of mining-induced vertical stress is 2.83 MPa, and the range of influence remains limited to around 15 m from the line of extraction. However, the problem of roof instability under the influenced zone was experienced due to the inherent nature of the weak roof strata. The chances of trapping of CM under weak and laminated roof strata of the panels is reduced by deriving a threshold value of the roof dilation, on the basis of a simple statistical analysis of the obtained field results. However, few incidences of roof collapses are observed over the CM due to the unpredictable behaviour of the geological structures below the freshly exposed strata, which also caused an obstruction in the speed of the extraction. This paper presents results of the field and laboratory studies, conducted for a successful application of the single-pass pillar extraction method, along with details of some CM trapping cases below the weak and laminated roof strata.INTRODUCTIONCoal is the major natural resource to meet the growing energy demand of India and this trend is likely to continue for the next 20 years in the country. A major part of the Indian coal belongs to Gondwana period, and only a small part belongs to Tertiary period. The estimated coal reserve in India per the exploration work done up to a maximum depth of 1200 m is 315.149 billion tonnes, including the proven coal reserve of 143.058 billion tonnes (CMPDIL, 2017). Even in presence of these huge resources of coal in India, there is a considerable gap between production and the users’ demand of coal. Most of the mines are operated by conventional semi-mechanised Bord and Pillar (B&P) methods; thus, the mining industry has gained a considerable amount of operational experience (Singh, Singh, and Dhar, 1996) and knowledge about strata behaviour (Singh and Singh, 1999; Singh et al., 2008) using this method. However, recently, the Indian coal mining industry has introduced some mechanised depillaring (MD) operations, where the level of production is relatively high. This approach of depillaring improves safety along with the production and productivity of a depillaring operation (Ram et al., 2017). The nature of the overlying strata of these depillaring operations vary greatly: from highly laminated, weak, and easily caveable to massive, strong, and difficult to cave."

Jan 1, 2018

By S. J. Jung

The goal of this research is to demonstrate how optimization methodology can be coupled with the finite element method for greater stability of underground mine openings. As a result of increasing mining depth and intersection of geologic discontinuities, mining environments are becoming more complex. Current design procedures are rapidly becoming inadequate. In this research, the finite element method has been coupled with optimization criteria to design more stable and safe underground openings

Jan 1, 1993