Search Documents

By Craig Collins

The objective was to develop a Real Time Drilling Display System for Rotary Roof Bolting using Analyzing Software to display mapped holes as they are being drilled in real time. The Analyzing Software was developed to analyze and interpret drill sensor data produced during a rotary drilled roof bolt hole by using algorithms and drill sensor data trends to find and locate roof separations or voids. After this software was validated through lab and field testing, its application needed to be tested in a production environment. However, the current method used to record the drill sensor data proved to be cumbersome for the mine to use on a continuous production basis. With different more practical mediums being explored and considered, a Drilling Display System was developed that will both map holes in real time and record the data for later review. With the Analyzing Software's current level of accuracy and with the convenience offered by the real time Drilling Display, the Drill Display System was implemented along with a test mine's current roof control plan and tested as a roof separation "Early Warning System". After installing the Drilling Display System onto a machine, input and suggestions from the mine environment were considered and design changes were made. The real world application provided a design direction that made the information produced by the Analyzing Software both convenient and significant for the test mine. Also, using the Drilling Display System increased the number of drilled holes that were both mapped and recorded. The results from studying these recorded data files lead to refinements in the current algorithms used by the Analyzing Software which increased the overall accuracy and capability of the system A fairly accurate representation of void or separation locations in the mine roof can be determined fromsensor data recorded during production bolting cycle. The immediate feedback on the local roof condition that the Drilling Display System provides, amounts to a quicker and safer alternative to the scratch hole and the video bore-scope examinations that arc performed after the bolting is completed. This project has shown that making the alternative method of examining the roof condition practical for the mine to use is just as important as the accuracy of its results. Based on the results of this field test, this or other methods of reviewing the analyzed drill hole data will be explored.

Jan 1, 2004

By Nicholas Chlumecky

One of the oldest mine supports in use today is timber cribbing. In many mining operations material, installation and maintenance costs for timber cribbings are substantial. Their performance characteristics also add to mining costs. During the 1970'6, the U.S. Bureau of Mines recognized many of the deficiencies of timber cribbing and very successfully initiated the development of steel-fiber- reinforced (SFR) concrete blocks. Since the early 1980's this SFR concrete cribbing has been manufactured as a high quality product which has found increasing acceptance by mine operators. The rapid acceptance of the SFR concrete block cribbing has proven that its use is eliminating many of the deficiencies of timber cribbing while providing superior load carrying capability. The use of SPR concrete blocks to build the cribbing is based on the same principle as wood. That is, stacking longitudinal, rectangular blocks in layers in a crisscross manner. The blocks are positioned at right angles to the blocks of the layer below. With this situation, the possibility of misalignment exists when the concrete blocks at the base are not placed exactly parallel or when the blocks of the base are subject to unequal settlement. Such misalignments creates localized, high stress areas which reduces the load capacity of the cribbing made from SFR concrete blocks. Recognizing the problem of misalignment, the Underground Supports Division of Commercial Shearing, Inc., conceived a circular, one-piece, steel-fiber-reinforced concrete cribbing which is being marketed under the name - Donut Cribbing. This concept eliminates the possibility of misalignment and provides 100% bearing between layers of stacked cribbing as well as other advantages. The purpose of this paper is to discuss the performance characteristics, installation speed and economical factors of each of the three cribbing types -- timber cribbing, SFR concrete block cribbing and SFR concrete Donut Cribbing.

Jan 1, 1986

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

By Gift Makusha

Pre-feasibility work at Goedehoop Colliery indicated that stooping of the undersized pillars could be undertaken safely and economically. In light of the findings a decision was taken in may 2003 to undertake the project in the 600 area where three panels (628, 630 and 632) were identified for pilot test sites. The findings of the pre-feasibility studies were presented in a paper to the 22nd International Conference on Ground Control in Mining held at Morgantown in 2003 (1). This paper is a progress report for the project 12 months down the line. Work commenced to prepare the sites for mining including, acquisition of equipment, equipping the panels, selecting and training people to undertake the project. Three monitoring sites were set up distributed evenly along panel 628. Each site comprised of a surface extensometer station in the overburden strata adjacent to a pillar underground containing hard inclusion vibrating wire extensometer cells. This combination would be used to record pillar loading ahead of stooping as well as the depth of caving and strata deformation in the goal zone behind the stooping face. In July 2003, support work began to secure the roof and rib sides. The support was required to be at least three-pillar center distances ahead of the mining front to ensure that the effective stooping zone is secured at all times. The NEVID method of pillar extraction developed at Sasol and described in detail by Ismet and Moolman (2) in a Coaltech 20-20 report was adopted for the pilot project. As part of training the selected stooping team was send for a week to Sasol's Bosmanskrans Colliery to familiarize with the NEVID method of stooping. Further training sessions were carried out on site especially on the use of Mobile Roof supports. The first cut was taken on 20 August 2003 by implementing a form of chequerbord mining taking every second row of pillars. Three rows were mined this way which period was considered training to get workers familiar with the new equipment and mining method. On the fourth row the full design span equivalent to mining 4 pillars in a row was implemented taking out all the pillars in successive rows. Production figures gradually increased from 300 to more than 1,000 tons run of mine coal per shift. Panel 628 was mined out at the end of December 2003 and the second of three panels (630) is being mined. Substantial data from the monitoring set up has been recorded since the start of the project. Monitoring will continue to the end of the pilot project and the data will be analysed and fed back into the design tools to improve the design in future undertakings. The pilot project has been a success so far and other old workings are being evaluated for stooping.

Jan 1, 2004

By P. Palmer

Using 3D numerical modeling as a tool to predict rock behavior during mining extraction is still an uncommon practice at many underground operations. This paper will discuss how 3D numerical models were used to identify potential underground instability problems in the later stages of mining. Detailed modeling was then used to make modifications to the mine design in order to minimize the impact of the potential instability problems. This paper will also discusses how the back analysis of a previous underground collapse event at the same mine in a room and pillar area was used to identify the instability problems for a new orebody being developed deeper in the mine. A 3D numerical model was first created in the MAP3D software to determine the conditions of the pillars and the magnitude of the vertical displacement above the mining zone prior to the underground collapse. The results from the ground collapse model were then used as a threshold limit for predicting potential instabilities and developing a model for the new mining zone. sing the results from this analysis, mine management has been able to implement a much higher percentage recovery of the reserve in a new, high-grade area of the orebody while minimizing potential instabilities in the future.

Jan 1, 2005

By F. Ziaie

Line electrical resistivity method using physically infinite distance between current line electrodes is proposed to determine the location of mine workings. In this method at least three line electrodes are used. One of the current electrodes is used as a sinkhole electrode while the other two are used for different activiations of line electrodes when the survey area is between one of the line electrodes and outside the other one. The resistivity profiles perpendicular to the current line electrodes for different activation of electrodes is plotted versus the distance of the midpoint of the measuring potential electrodes. The anomaly due to presence of any locally lateral resistivity variation (e.g. mining cavity) will represent itself as peak or trough depending on resistivity contrast between the source of the anomaly and surrounding medium. Physical simulation of the proposed model was conducted by using a tank model in the laboratory. The results of experiments appear to be consistent with the theoretical model. Experiments for different depths and different spacial locations of the simulated cavities appeared to be successful for locating the depth and size of the cavity.

Jan 1, 1989

By Andrew P. Jarosz

The present computerized mine planning systems use CAD-like and Data Base Management System (DBMS) environments for the handling of numerical, textural, and graphical data. These two environments are usually integrated, that the data provided by the DBMS are linked to any graphical element rendered by the CAD system. The subsidence monitoring, analysis and prediction process should be integrated with the mine design process. It should allow for the assessment of impacts of mining on surface, surface utilities, and structures. It should also provide for the measures to control mining, that will create only acceptable levels of negative impacts on the surface. The common data base, and graphical capabilities of the CAD interface, may be used with great advantage for both systems. The findings of the subsidence engineering modulus should be directly incorporated into the mine design system, and vice versa. In this paper the author suggests guidelines for the development of the comprehensive computer system, able not only to predict the vertical and horizontal components of the ground movement, but also assess the induced surface damages, and costs related to subsidence prevention and compensation.

Jan 1, 1995

By Tingkan Lu

To understand the reinforcing mechanisms of the fully grouted resin bolt and clarify its performance in weak roof strata, the field monitoring investigation using instrumented bolts has been conducted in a coal mine. The total time of monitoring was around ten months. In the first 8 month period, the rock bolt performance relates to the initial development of the gateroad and subsequent time dependent strata response. After 8 months, the rock bolt performance relates to the effect of the approaching longwall mining operation, up to the point where the measuring station was destroyed due to the mining operation.

Jan 1, 1998

By Naj Aziz, Peter Craig

"Double shearing tests were carried out on the 28 mm (1 in) hollow strand Jennmar TG cable bolt to gain a better understanding of the shearing behavior of the cable. The first test was limited to a 50 mm (2 in) travel on the testing machine and produced a shear load of 900 kN (92 t) at the maximum 50 mm (2 in) displacement. The axial load generated on the cable bolt was 238 kN (24.3 t). In the second test, the machine travel was increased to 75 mm (3 in), and first strand failure of the cable, due to shear loading, occurred at 1,354 kN (138 t) vertical load at a vertical displacement of 59 mm (2.3 in). The cable axial load was in the order of 385 kN (39.3 t). Analysis of the failure mode and loads achieved showed that the cable strands bent and the concrete crushed along the shear plane; the shear loading across the concrete and grouted cable then reached the tensile strength of the steel wires.INTRODUCTIONCable bolts were introduced to the mining industry around 1970, initially to surface mining and underground metalliferous mining and then to coal mining in the early 1980s, mainly for roadway reinforcement as a secondary means of support. Cable bolts have since been used as both primary and secondary supports. As primary support, Fuller et al. (1994) described the application of cable bolts, known as FLEXIBOLT, for strata reinforcement in both Angus Place and Ellalong Collieries, in New South Wales (NSW), Australia. As secondary support, cable bolts have also been used as cable trusses, which act to support the immediate roof in a sling-like manner (Fabjanczyk and Tarrant, 1988; Fuller, et al., 1991; O'Grady and Fuller, 1992), and for reinforcement at higher stratification and beyond the rebar bolt length, mainly for anchoring lower strata layers immediately above the coal seam to the higher, competent bedding formation above. Initially cable bolt anchorage was by cementatious grouting and, since the 1990s, by chemical resin. The dominant type of cable bolts used for secondary support in Australian underground coal mines are 588 kN (60 t) capacity cables which are point anchored, pre-tensioned and post-grouted. Point anchored refers to a cable resin anchor. The cables are tensioned on the resin anchor; then the same cable is grouted by cementatious grout pumped to ·fin the annular gap around the cable from the collar of the hole to the resin anchor. Bulbs are located within the resin anchor section for mixing of the resin chemical components and efficient anchorage. The installation of these typically 8 m long cable bolts involves a lot of manual handling of the cables, lifting of heavy hydraulic tensioners, and exposure to resin grouts. Table 1 shows the specification of various cable bolts currently marketed and installed in Australian underground coal mines."

Jan 1, 2010

By Takashi Sasaoka, Pinyo Meechumna, Hiroshi Takamoto, Pipat Laowattanabandit, Akihiro Hamanaka, Kikuo Matsui, Hideki Shimada

"The EGAT Mae Moh coal mine is an open-cut coal mine in Thailand that produces about 16 million tons of lignite annually to generate 2,400 mw of electricity. The surface mine pit covers is 4 by 7 km (2 by 4 mi) with mining depths ranging up to 260 m (853 ft). A large pit slope is formed with the progression of the mining operation. However, massive slides in the pit slope often occur due to weak planes, such as faults and bedding, and the weak mechanical properties of the rocks. Hence, a 200-300 m (656-984 ft) boundaries of rock block including coal seams is left in front of faults in order to prevent slides and maintain the stability of the pit slope. As a result, there are significant coal reserves beneath the abandoned area along the pit slope.The focus of this paper is on the applicability and design of the highwall mining for recovering much of the coal left along the pit slope.INTRODUCTIONThe EGAT Mae Moh Mine is located about 630 km (186 mi) north of Bangkok, Thailand (see Figure 1). It is the largest coal mine in Thailand and produces about 16 million tons of lignite annually from open-cut mining (70 % of the total coal produced in Thailand). The coal is used to generate 2,400 mw of electricity at the Mae Moh power plant, which is situated near the mining area. The mine pit covers an area of 4 by 7 km (2 by 4 mi) at varied depths up to 260 m (583 ft) as shown in Figure 2. As the mining operation progresses, the pit becomes to be deeper and deeper. In the year 2035, the bottom of the pit is expected to reach depths of 500-600 m and the open-cut mining operation will be finished. The possibility of an underground mining operation is being investigated.A large pit slope is formed with the progression of open-cut mining. However, massive slides often occur in the pit slope along weak planes including faults and bedding planes due to the weak nature of the rock. As a result, 200 to 300 m (656 to 984 ft) boundaries of rock including coal seams are left in place in front of faults to prevent slides and to maintain the stability of pit slope. Moreover, the dip of the slope is less than that of the planned design. As a result, there significant coal reserves abandoned along the pit slope."

Jan 1, 2010

By Esterhuizen S. Gabriel, Su W. H. Daniel, Tulu B. Ihsan, Klemetti M. Ted, Mohammed M. Khaled, Gearhart F. Dave

"A numerical-model-based approach was recently developed for estimating the changes in both the horizontal and vertical loading conditions induced by an approaching longwall face. In this approach, a systematic procedure is used to estimate the model inputs. Shearing along the bedding planes is modeled with ubiquitous joint elements and interface elements. Coal is modeled with a newly developed coal mass model. The response of the gob is calibrated with back analysis of subsidence data and the results of previously published laboratory tests on rock fragments. The model results have been verified with the subsidence and stress data recently collected from a longwall mine in the eastern United States. INTRODUCTION In 2015, there were 40 longwall mines operating in the United States, each producing an average of 4.5 million tons of coal per year, and they supplied 60% of the U.S. underground coal production. This represents a substantial increase from 50% three years earlier (Sears, Tulu, and Esterhuizen, 2017). During this period, reportable roof fall rates in U.S. longwall mines also increased. Large roof falls that can block the gateroads are not only a ground-fall hazard, they can disrupt the ventilation system, block the escapeways, and increase the potential for elevated methane levels in the gob. To address these hazards, the National Institute for Occupational Safety and Health (NIOSH) Pittsburgh Mining Research Division (PMRD) is conducting research to improve the design of ground control systems in longwall gateroads. Gateroad stability is primarily determined by the longwall pillar design. Generally, the required dimensions of the pillars around a longwall panel are determined first, which then dictates the location of the gate entries relative to the mined panel. The Analysis of Longwall Pillar Stability (ALPS) method is the most accepted gateroad pillar design procedure in the United States (Mark, 1992). The ALPS method accounts for local roof geology in the gateroad stability assessment by including the Coal Mine Roof Rating as an input parameter (Mark, 1992). The key assumption in the ALPS method is that unstable pillars result in unstable gate entries. However, experience provides examples of mines where pillar stability and gateroad stability are only loosely correlated (Su, Draskovich, and Thomas, 2003)."

Jan 1, 2017

By Gennaro G. Marino

As part of the development of a shopping center in West Virginia, a significant bench had to be cut into a mountainside that was over 700 ft. high. As a result, a steeply stepped, over 400-ft.- high wall was created. The high wall exposed old abandoned room-and-pillar coal workings at the toe of the cut. Because of exposure to the elements, the mine openings began to severely deteriorate with significant potential of undermining of the rock cut face and causing a large rock fall(s). Exacerbating this wall problem was a set of fairly planar sub-parallel and sub-vertical joints creating, in essence, thin rock slabs. Also present were stress relief fractures that were observed in the exposed pillars. Given the proximity of the project building to this high rock face, mitigation of the rock fall potential was deemed necessary. In order to prevent continued damage and deterioration of these exposed mine entries, the rubblized openings were filled with bulkhead grout and pressure injected with grout. The total amount of grout installed in these openings was about 3,945 CY. Also, a rock drape was attached to the entire face to mitigate small rock falls from natural causes. There also existed another level of mining below the ground surface and under an existing building at a depth of about 160 ft. Under a portion of the building, the pillars in these old works were second mined, and the building seemed to have been already affected by limited subsidence above this area. Also, calculated pillar safety factors in this high extraction section of the mine were too low. Consequently, these old works were stabilized by saturation grouting. An indication of overall roof distress was the significant fracturing encountered in the bedrock overburden during the drilling of the injection holes and the volume of grout placed in these holes. The total grout installed in the high extraction section of the lower mine was about 6,660 CY. This paper describes the investigation and the grouting techniques employed at the rock cut face and under the existing building to mitigate continued damage to avoid future subsidence problems at the face and the building premises.

Jan 1, 2013

By Robert J. Coutts, Dan A. Payne

"The Broadmeadow mine is a punch longwall mine owned and operated by BHP Billiton Mitsubishi Alliance (BMA) in the Bowen Basin of Queensland, Australia. In 1999, an exploratory adit was driven as part of the feasibility study for the establishment of the mine. Subsequent development of the mine in 2004 started some distance away from the initial test adit. Ten years later, the trial adit would be used as the headgate of the 5'h longwall panel. However, multiple roof falls had occurred in a heavily jointed area over a distance of 1, 000 ft (300 m) during the past 10 years. Due to the immediate coal roof, low stress environment and time, some of these falls had progressed to 38 ft (11.5 m) above the mining floor. As a result of the extensive work involved and safety implications of trying to recover the roof falls conventionally, a strategy was developed to fill these roof falls with a cement based product from the surface, down boreholes, and redrive the roadways through the consolidated roof fall material. This paper describes the process of rehabilitating these roadways, the results when the roadway was redriven and the headgate performance as the longwall mined through the area.INTRODUCTIONBMA operates 7 open cut and 2 underground mines in the Bowen Basin of Queensland, Australia producing over 50 mt per year of high quality coking coal. As a result, BMA has established over 150 miles (250 km) of open cut highwall.Broadmeadow Mine is a punch longwall mine located 22 miles (35 km) north of Moranbah in Queensland, Australia on the western limb of the Bowen Basin Coal Measures (Figure 1). Broadmeadow Mine commenced development in 2004 and began longwall operations in 2005. Broadmeadow is currently completing the fifth longwall panel and produces over 4 Mt per year of high quality coking coal. Mining is undertaken in the Goonyella Middle Seam, which is 21 to 25 ft (6.5 to 7.5 m) thick in the current mining area.The mine layout required the use of an old exploratory gateroad as the headgate of its 5th longwall. This gateroad had experienced numerous large roof falls and required extensive rehabilitation."

Jan 1, 2010

By Charles D. Albright

FMC built its first angle bolter in 1969. This was a single mast unit which could drill holes for conventional bolts or angle holes for installing roof trusses. From this drill we learned very quickie that the mast design was an important factor to the success or failure of any angle bolter. The first mast installed on this machine (EXHIBIT 1) was fixed in length with endless chain and hydraulic motor feed. It took less than a shift underground to determine that the fixed length was not going to work. The undulating top created conditions where the mast was too long to be positioned at a 450 angle. Several mast designs were tried and discarded before we felt we had the answer to our problems. (EXHIBIT 2} We developed a mast that would extend or retract without moving the position of the drill unit. This allowed the operator to "wedge" the mast between the floor and roof before starting his drill feed. This arrangement worked exceptionally well, but unfortunately the demand, as perceived by our Sales Department, for angle bolters did not materialize. This unit became an "orphan" and was used by many coal mines until it died a natural death. FMC did not decide to re-enter this market again until 1979. At this time many coal producers were contacted to determine what features should he incorporated into a dual mast angle bolter. From this information we were able to arrive at three main areas of concern that we should consider when designing a new machine. 1. Eliminate overturning moment on drill unit. 2. Provide some sort of steel support near roof line to aid in starting a 15° hole. 3. Dual mast units need a method of raising center of mast rotation to facilitate various mining heights and still maintain a 45° angle and proper distance from rib. I would now like to discuss each of these individually. EXHIBIT 3 indicates the overturning moment resulting from the feed cylinder thrust and the reaction of the drill hit against the roof. To eliminate this moment a mast was designed using two feed cylinders mounted in the plane of the drill center line, thus removing all overturning moments on the drill unit carriage. This mast has been used extensively in vertical drilling with no downtime attributed to wear on mast guides and drill unit supports. The cylinders are designed to provide uniform thrust and feed speed throughout the total stroke. For manufacturing and economic reasons, the final mast design on the first drill was not used for this new drill. A drill steel support mounted on the mast, and positioned by a hydraulic cylinder provides the needed guidance near the roof for starting a hole at a 45° angle. (EXHIBIT 4) This support does not surround the drill steel but simple cradles the drill steel for "collaring" the hole. Once the hole is started, the support can he lowered. The support plate is hinged to allow the drill mast to pass. The advantage of having vertical motion of the center of rotation on a dual mast drill is shown in EXHIBIT 5. Following the criteria recommended by Birmingham Bolt Company, the truss bolts should he at a 45° angle and one fifth of the width of the entry from the rib. Therefore, in an entry 20 ft. wide, the hole should be drilled 4 ft. from the rib and at the 45° angle previously mentioned. With the limitations inherent to dual boom drills, that is, the masts cannot pass each other nor can one mast pass the center line of the machine even if the other is positioned out of the way, it is impossible to accommodate much variation in seam height and maintain the recommended criteria. As shown in EXHIBIT 5, a center of rotation 17 inches above the floor will start a hole 4 ft. and 45°angle at a height of 67 inches. Any height above 67 inches

Jan 1, 1982

By E. Bane Kroeger

For many years, researchers around the world have been investigating coal pillar stability. Many have focused on trying to optimize the size of the pillars by examining stable and failed pillars in underground coal mines. For mines that are already in operation or companies that are opening a new mine adjacent to an existing mine where the pillar dimensions and stability can easily be determined, this is the preferred technique. However, when a new mine is planned for an area where there has been relatively little mining, then samples of the coal need to be drilled and tested for strength. Because of the fractured nature of coal, this lab testing is often problematic and may often yield poor results because of the limited size and number of samples that can be obtained from the core. To remedy this problem, research was conducted in the labs at SIUC to determine if improvements could be made when optimizing the size of pillars for underground coal mining in Illinois. Through a greater understanding of the relationship between dimensions and strengths of coal samples tested in the lab, it was hoped the in situ strength of coal pillars could be determined with greater accuracy. This could allow the extraction ratio to be increased and more of the coal resources in Illinois recovered without impacting the safety of the mine workers. Many different agencies around the world have published recommended minimum numbers of samples for strength determination based on the size of the coal samples. For two inch cubes, uniaxial testing of at least seven samples is recommended and this minimum number decreases to four when testing four inch cubes. However, the laboratory testing of coal from two seams thus far has suggested this number should be almost doubled to accurately determine the coal strength. Testing also showed there is a pronounced decrease in the variation in the strength of the samples tested as the dimensions of the sample were increased. From the testing results thus far, it is recommended that a sample size versus strength curve be constructed for every coal seam for Illinois. Creating such a curve will better determine both the minimum number of samples and minimum sample dimensions for that seam or region of a particular scam. A second set of tests were conducted in the laboratory to determine the increase in strength with reduction in height for samples with the same width and length. In Illinois, typical pillar dimensions are 40 to 80 feet (12.2 to 24.4 m) wide and long, with coal thickness varying from about 4 to 9 feet (1.22 to 2.74 m). This provides width/height ratios ranging from 4.5 to 17.5. Coal samples having width/height ratios ranging from 0.5 to 13 have been tested thus far. As with most pillar design formulae, the strength of the Murphysboro coal versus width/height ratio was linear and closely matched shape factors published by other researchers. The laboratory testing thus far has identified three factors that affect the strength of laboratory samples, the number of discontinuities, the angles of critical discontinuities, and the shape of the samples. The factor that probably had the most pronounced effect on the strength of the coal samples was the angle of critical discontinuities. This testing has confirmed that there is still a gap that needs to be bridged between the strength obtained from laboratory testing and the in situ coal strength predicted.

Jan 1, 2004

By A. H. PH. D Wilson

Compared to the rocks encountered in, say, metalliferous mining or tunnel drivage in igneous and metamorphic rocks, the strata associated with the coal measures are relatively soft. If excavations are made at depths greater than one or two hundred metres, zones of fractured or yielded rock develop around the openings. The stress which can be taken by yielded rock at the boundary of an opening is low, but because of the friction present between the fragments or particles, the sustainable stress rises rapidly with distance from the opening until sufficient confinement has been developed to prevent failure of the rock in the first place. Thereafter the laws of elasticity will apply and increasing distance will reduce the stress field until conditions consistent with the cover load are reached. For rocks which maintain their elastic condition right to the edge of an opening, finite element analysis and similar techniques have proved successful in determining the distribution of stress around the opening. This approach, however, is not successful when a yield zone is introduced between the elastic zone and the opening. Nevertheless, some estimate is required if help is to be given in the siting of roadways and in the determination of adequate pillar sizes. This matter is of particular importance in the planning of new mines especially if retreat faces are involved, as extensive roadway drivage must be planned and executed before experience of the effects of face extraction is gained. In 1977, the author published [I] a method based on 'stress balance'. The total aggregate downward load over a large area will remain that of the cover load even after part of the coal seam has been removed. Any rise in stress in ribsides and pillars must be offset by an equivalent stress reduction across roadways and other areas of extraction, and vice versa. Knowledge of one can lead to an estimate of the other, provided the general form of the stress distribution can be postulated. It is the object of this paper to revue the method in more detail and to give further examples of its successful application.

Jan 1, 1981

By Christopher C. Woomer

Cyprus Shoshone Mine uses the longwall method to extract a deep, thick, pitching coal seam in the Hanna Basin of South Central Wyoming. The immediate, and main roof rock consists of weak, thinly-bedded, silty mudstones with weak, interbedded fine-to medium-grained sandstone. Tailgate ground control has been a critical factor impacting productivity at the mine. A gateroad condition mapping program for the 11 left longwall gateroads indicated potentially severe ground control problems for the tailgate. It was predicted that the existing, secondary support pattern of wood cribs would not provide adequate support capacity. As predicted, severe deterioration of the tailgate occurred in the form of a roof cutter, sloughing ribs, and failing wood cribs. Productivity was affected and additional support was required. Shoshone personnel evaluated several options for providing the additional support needed in the tailgate. The additional support system and placement method had to satisfy several requirements with respect to access, materials handling, placement method, set-up time and support capacity. Approximately 6,000-ft. of deteriorating tailgate had to be supported. The initial wood crib support system left a 6-ft. walkway for access and construction of the additional support. This access was further reduced as the roof loaded the wood cribs. This eliminated the possibility of mechanized material handling to the construction locations. All materials would have to be transported by hand. The additional support system had to be placed quickly to stay ahead of, and keep pace with, the required longwall advance rate. It had to incorporate the existing wood cribs and provide at least 500 tons of extra support capacity. And most critically, the system could not affect longwall ventilation. Longwall coordinators and engineers made the decision to use a low density, pumpable cement known to the industry as TeksealTM, to provide the system required. A 200 psi ultimate strength mix was decided on to provide the required load capacity. The existing cribs were formed with 1-in. by 6-in. boards and brattice cloth to provide the containment. To overcome the access limitations, three boreholes were drilled from the surface to the tailgate on 2,000-ft centers. A mobile pumping station was established on the surface and the TeksealTM was pumped 900-ft. down the boreholes through a 1.5-in. steel pipe, then as much as 1,800- ft. along the tailgate entry through 1.25-in. miner spray hose. The materials required for the TeksealTM supports could all be carried into the construction locations by hand. As a direct result of incorporating relatively new methods of pumping high yield, low density, cementitious grouts, the Shoshone Mine reduced downtime due to tailgate ground control problems by approximately 70% in comparison with previous Longwall panels. The longwall set three monthly production records while mining the 11 left longwall under the deepest cover, steepest pitch, and most extreme ground control conditions ever encountered at the mine.

Jan 1, 1995

By Xinzhi Du

Coal strength is essential to analyze the stability of excavation in coal. It is well known that the strength of coal tested in the laboratory is size-dependent, and different size specimens will produce different strength. Many researchers have studied the uniaxial compressive strength of coal and developed a number of empirical formulae to describe the strength variation with specimen size. According to a comprehensive data base that includes most of the laboratory testing results in the last century in U.S. and South Africa, the size effect and shape effect were reviewed and analyzed. The results show that different groups of coal specimens have different size effect on the compressive strength. The critical size for the size effect from the data analysis is about 40 in. The shape (W/H) of the specimens has a large effect on the compressive strength. The pillar strength formulae obtained from the data regression have a larger strength than all other existing formulae. Most pillar strength formulae have their own place in coal pillar design field. In order to analyze the difference of those formulae, twelve design formulae have been selected to compare using the Pittsburgh coal seam. The intention is not to determine the preferred formulae, but only to compare them, and see how do the W/H ratio, width, length, angle, cross-section area and effective width affect different pillar strength formulae?

Jan 1, 2008

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 Daniel W. Su

This paper presents the results of a detailed geotechnical study on the potential multiple seam interaction between the No. 2 Gas (upper) and the Powellton seams, which are separated by an average interburden of 45 ft. The interburden consists of 4 ft of clay shale below the No. 2 Gas seam, 38 ft of massive sandstone, and 3 ft of shale above the Powellton seam. The No. 2 Gas seam was mined about 60 years ago, and many water-pooled areas were projected to be present. The deepest overburden over the No. 2 Gas seam is about 750 ft. To evaluate the potential interaction between the two mine works and to determine the pillar safety factor, Analysis of Multiple Seam Stability (AMSS) software was initially used to determine pillar dimensions to obtain a minimum safety factor of 2.0 for areas under pooled water, as well as a minimum safety factor of 1.5 for areas not under pooled water. Furthermore, to assess the quality of the interburden rock strata, two surface core holes were drilled, and rock cores were tested in the laboratory for their strength and elastic moduli. Using the experimentally determined strengths and moduli as part of the input parameters, the RocScience Phase2 finite element program was employed to model the stress field scenarios and estimate the safety factors of the interburden rock strata before and after Powellton seam mining. The pillar safety factors derived from the Phase2 analysis were in excellent agreement with those obtained from the AMSS analysis. Phase2 results also predicted that the effect of the Powellton seam mining, first mining only, would reach a maximum of 19 ft above the Powellton seam, which led to the proposal of drilling 25-ft deep test holes to detect water, instead of the 45-ft deep test holes mandated by MSHA. The proposed test-hole height was accepted by both Mine Safety and Health Administration (MSHA) District 4 and the state of West Virginia. Mining of the Powellton coal seam commenced in February 2011 and continued for eight months. To date, no 25-ft deep hole has encountered water under the water-pooled zone, which validates the RocScience Phase2 finite element analysis. Also, inmine observation of excellent pillar stability and roof condition confirms the validity of the AMSS analysis.

Jan 1, 2012