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By L. Sahu, S. Bhattacharya, S. Dey

"An ideal procedure to determine the limits of flotation-based separation, as is the case with washability analysis for density separation, does not exist. Therefore, a procedure called release analysis, considered to be the counterpart of the washability process in density separation, was developed to evaluate the flotation characteristics of coal fines. Since flotation is a complex process it is necessary to know the maximum recoverable yield that could be achieved at a target ash level and what could be obtained by release analysis. Due to a number of limitations in the initial procedures, modifications were recommended by various researchers to generate an optimum separation curve. This paper presents a review of, and a discussion on, the chronological evolution of release analysis methodology over the decades. Notwithstanding the availability of a standardized release analysis procedure, such as BS 7530, the paper highlights the critical parameters involved in the procedures, such as coal type, feed characteristics, choice of reagents and dosage, operator bias, and design features of the flotation cell IntroductionCoal is a heterogeneous substance consisting of various combustible constituents (macerals) and associated mineral matter. The kind and abundance of the various macerals present in the coal, such as vitrinite, semivitrinite, exinite, and inertinite, determine the type of coal. As macerals are formed by biological degradation of different parts of the plant material, they have different properties based on their original composition. Coal cleaning processes are dependent on the differences in physical properties between the coal components (the macerals) and also between the macerals and the mineral matter. Thus the cleaning characteristics of the coal might vary from one type to the next. The cleaning processes influence the distribution and the concentration of macerals in the various size and density fractions of the cleaned product."

Jun 1, 2019

By P. N. Anyalebechi

Keywords: Aluminum alloy 3004, Solidification rate, Homogenization, Second phase particles The effects of the fineness of as-solidified cast microstructure and temperature on the response of aluminum alloy 3004 to homogenization heat treatment (a.k.a. preheat) have been investigated. The study was conducted at 538oC and 595oC on directionally solidified laboratory-size ingots with a range of solidification rate of 0.10oC/s to 29.4oC/s. Optical and electron microscopy, quantitative image analysis, electrical conductivity, and electron probe microanalysis were used to quantify the effects of the homogenization heat treatment on the microstructure of aluminum alloy 3004. In general, increase in fineness of as-cast microstructure and homogenization temperature enhances the a-transformation reaction, dissolution of Mg2Si, elimination of dendritic microsegregation, and dispersoid formation. The homogenization heat treatment reduces the average size of second phase particles in castings solidified at <1oC/s. However, in castings solidified at above 1oC /s, homogenization causes the coarsening of the second phase particles. Mn in solid solution is independent of homogenization heat treatment temperature.

Jan 1, 2005

By Khaled Morsy

In the past two decades, high horizontal stress has been attributed to many ground control failures. Most believe that entries parallel to the high horizontal stress is most stable and least stable when perpendicular to it. Entry stability decreases when the angle between entry and horizontal stress increases. The concept of 2-D stress shadow has been proposed for selection of proper panel orientation, often with mixed results. In this paper 3-D finite element modeling was used to analyze the stress distribution at the headgate of a longwall panel subject to high horizontal stress at various angles. It was found that the directions of induced principal stresses at the longwall face and headgate are independent of the panel orientation. The induced principal stresses concentrate or relief at headgate based on the panel orientation with respect to the directions of the induced principal stresses. Accordingly, only 3-D modeling can fully account for the effects of high horizontal stress on panel orientation. The paper illustrates in details the changes in orientation and magnitude of all three principal stresses around the longwall face and T-junction area as panel orientation changes.

Jan 1, 2006

The Research on Refined Gold Bullions Surface Upgrading

Sep 13, 2010

By M Passmore, N Poznik

Olympic Dam is a world-class deposit and is unquestionably one of the most geologically, structurally and texturally complex orebodies in the world. As development advances into the Southern Mine Area (SMA) the geological understanding is rapidly evolving as underground drill rigs are prioritised to provide the additional data necessary to convert Indicated and Inferred Resources into Measured Resources. The SMA represents approximately 70 per cent of Olympic Dam’s total resource and is significantly more complex when compared to the Northern Mine Area with respect to the diverse nature and origin of lithotypes, rock mass characteristics and the extent of major structures controlling primary mineralisation trends.Recent mine development into the central SMA has been impacted by a major structural gouge zone. Sound geological interpretation allowed mine geologists to successfully integrate their local understanding of the distribution of lithotypes with critical geotechnical data, chiefly rock quality designation (RQD), to delineate and better predict this major structural gouge zone. As a direct result, development rates improved whilst also increasing awareness for the importance of sound geological interpretation utilising all available data. In order to maximise predictability across all systems in place through the mining geology value chain, the integration of geological and geotechnical data in a geotechnical RQD block model provides an additional tool to assist in preventative solutions for complex problems to achieve a proactive mining environment.This paper details the methods used to build the RQD block model, the geostatistical estimation parameters, 3D geotechnical domain wireframes, challenges regarding legacy approaches, and collection of geotechnical information and incorporating/linking geotechnical data with geological features, including the use of grade and lithotype domains. Although the geotechnical block model proved difficult and complex to construct, the resultant product has seen successful integration of geological and geotechnical data into mine planning providing higher resolution of critical parameters for local mining areas.CITATION:Clark, J M, Passmore, M and Poznik, N, 2017. Olympic Dam rock quality designation model – an integrated approach, in Proceedings Tenth International Mining Geology Conference 2017, pp 27–36 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 20, 2017

By Kelsall D. F, Trahar W. J

The size by size batch flotation behaviour of copper activated sphalerite from synthetic mixtures with quartz was determined. In terms of recovery of sphalerite, the size fractions could be grouped into coarse, intermediate, and fine ranges. The behaviour of the coarse and fine ranges was significantly poorer than that of the intermediate range in both the extent and rate of flotation.Each size fraction had a fast and a slow floating component.The main effect of changing the copper sulphate addition was to alter the proportion of any size fraction which was fast floating. No correlation could be found, between the amount of copper on the spha:lerite andeither the flotation rate of the mineral or the amount of copper added to the system. It was found that fine particles did not appear to consume greatly disproportionate amounts of copper sulphate during conditioning. A modified conditioning procedure was found to lead to a higher rate of flotation and recovery of coarse sphalerite and to make poss1ble substantial savings of copper sulphate.

Jan 1, 1975

By Dean Bray

During the past three years, a new rear-dumping dragline bucket technology has been patented, developed and tested at North Dakota coal mines. Field testing has shown that this new concept in dragline bucket technology can improve strip-mining efficiencies by reducing cycletimes and increasing dumping radius. This article details the process, participants and the results for the development of a rear-dumping dragline bucket. Participants in the process In November 1995, the board of directors of Dakota Coal authorized the execution of a field test and option agreement and a license agreement with the partnership of D.F. Schmidt, electrical/maintenance supervisor, and Bill Lengenfelder, president of Modern Industrial. The board also authorized entering into a dragline bucket field-testing agreement with the Falkirk Mining Co.

Jan 1, 1999

By J. Mason

In order that the use of continuous mining machines be most economic, the equipment must be operated at rated capacity as much of the time as possible and downtime for repairs kept at a minimum. A large factor influencing any proposals for maintenance programs is the complexity of the machines—often the intricacies of the precision-built equipment requires engineering knowledge byond the scope of mine maintenance personnel. A training program is part of the answer; there are others. In addition to investigating a joint company-manufacturer equipment-design improvement program, Mountaineer Coal Co. has launched a detailed preventive maintenance program to combat extensive downtime often encountered when continuous miners are used. Everyone in the coal industry knows the importance of operating continuous miners at their rated capacity as much of the time as possible. The potential capacity of the miner is such that l-min. delay means from 3 to 8 tons of coal lost. Today&apos;s miner is a well designed, precision-built machine, complicated with hydraulic and electrical equipment and complex mechanical assemblies with all possible combinations of gears, levers, and rods. As the machine becomes more complex, the problem of maintenance becomes more acute. The transition from conventional to continuous mining has taken place so rapidly that the mine mechanic has been unable to keep pace. When there is a machine failure, the maintenance man usually employs either the trial-and-error or the play -it-by-ear method of repair, but this represents an expensive proposition. Mine mechanics, maintenance superintendents, and mine managers complain about the complexity of the machines and how difficult it is to keep them running. There is no shortage of skill in the design offices. Machines are designed by good engineers, but for engineers. What&apos;s the answer? Better trained maintenance men—better maintenance instruction—better field service. These are all good corrective measures, but the real cure is to provide for the men at the do it level. Designers have got to cooperate with men to simplify maintenance on the increasingly complex machine. The designer must be aware of how his machine can be maintained. Sometimes a simple design change in a bearing, oil seal, or a bushing will correct recurrent breakdowns. Parts such as filters, hydraulic cylinders, hoses, and pivot pins can be better arranged on the machine to make them accessible so that the mine mechanic can easily keep the machine producing. Therefore, to get our maintenance off on the right foot, the program must begin back in the design office, before the machine is delivered to the mine property. Realizing the importance of this aspect, our company set out on a joint company-manufacturer equipment-design improvement program. This is a formal program in which quarterly reports are written covering machine defects, design weakness, and general problems relating to machine maintenance and performance. The manufacturer in turn comments on each and every item appearing in these reports, indicating the action he proposes to take. In addition to the written reply, meetings are scheduled with manufacturer representatives to discuss further the outlined problems. The benefits realized from this organized program are both substantial and far reaching. Good maintenance begins with a good design. The subject of preventive maintenance is not a new one. Volumes have already been written on the subject. It consists primarily of doing the maintenance work when you want to rather than when the machine demands it. However, a preventive maintenance program can be overdone. Theoretically, if carried to ridiculous extreme, all working time could be spent inspecting and repairing the machines and no time would be left for production output. Our goal in maintenance is to maximize running time for all machines, and to prevent or at least minimize production downtime. INSPECTION AND REPORTING The center of any preventive maintenance schedule is regular and systematic inspection, but it must be

Jan 1, 1961

By Radianto Elprama

For the realisation of the underground railway line RandstadRail in the city area of Rotterdam a bored tunnel is under construction. Diaphragm walls up to a depth of about 40 m were used to construct the building pits for the start shaft, the Blijdorp Station and the end shaft. The water tightness of the diaphragm walls was determined by leakage tests that were executed after the construction of the walls, but prior to the excavation of the building pits. The tests showed satisfactory results with a hydraulic resistance of the diaphragm walls of at least 500 days. To determine the water tightness of diaphragm walls it is recommended to follow the specific procedure as described in this paper.

Jan 1, 2006

By Linneas Laage, Thomas Weber, William Pomroy

"Underground mine fires can significantly influence mine ventilation airstreams, in some cases throttling or even reversing airflows. As a result, the performance of a metal and nonmetal mine stench fire warning system, which depends on the ventilation to carry the vital warning signal, under fire conditions is different from performance under nonfire conditions. The safety of underground miners can be jeopardized if the warning signal is delayed. This Bureau of Mines report describes research to investigate fire and stench warning system interactions. A computer model is presented that permits quantitative analysis of stench warning signal delays as a function of fire location and intensity. The results of a case study involving computer simulations of stench distribution in a hypothetical mine network subject to various fire exposures are also discussed. This case study illustrates a technique for identifying the areas within a mine that are subject to unacceptable warning signal delays, thereby enabling preemptive action by mine personnel, such as redeployment of stench injectors. INTRODUCTIONFires are an ever-present threat to the safety of underground miners. Since the smoke and toxic gas produced by a mine fire can be spread rapidly by the mine's ventilation system, mine evacuation must be accomplished as quickly as possible in the event of fire. In metal and nonmetal mines, the most common means of passing the fire warning signal to each miner is the stench system. The typical stench system utilizes ethyl mercaptan, a highly odoriferous organic compound, injected on the surface into the compressed and/or ventilation airstreams. Upon smelling the stench, workers evacuate the mine according to an emergency preplan.Although the stench system has been used successfully for over 60 yr, it has several serious shortcomings, owing to certain chemical properties of ethyl mercaptan and to certain performance characteristics and limitations of present injection systems.Recent Bureau research succeeded in upgrading the overall safety and effectiveness of the stench system through the use of a superior stench odorant and the development of improved stench injection equipment (1).4 This research has also investigated a specialized computer simulation model capable of calculating the precise concentration of stench in any mine ventilation network branch at any time after stench release (2). An adaptation of the computer model now enables the analysis of stench system and ventilation system interactions under the influence of a mine fire."

Jan 1, 1987

Equipment-washing facilities at Freeport-McMo-Ran?s Sierrita Mine in Arizona produce a mixture of oil and water that was difficult to dispose of easily or cheaply. However, after some changes in the facility, waste oil is now used as part of the mining process. Freeport-McMoRan Copper & Gold Inc., head-quartered in Phoenix, AZ, operates copper, gold and molybdenum mines. It is the world?s largest producer of molybdenum. Heavy equipment used at the mine is hosed off periodically in the wash bay. The dirty water and oil mixture drains into the sump next to the bay. Oil was manually collected for disposal. This was a tedious and dirty process Mine managers sought a more efficient method of removing the oil. Instead of spending the time and effort skimming the tank with vacuums or sorbents, they enlisted the help of Abanaki Corp., manufacturer of the Model 4 Oil Grabber. Engineers designed the sump to utilize prevailing winds to blow the oil to one corner. This was the ideal place for the Abanaki skimmer. The skimmer removes the oil and discharges it into a plastic storage container. ?This reduces the time needed to keep the area clean, and the staff can focus on cleaning the equipment rather than worrying about the oil,? according to Jay Salerno, construction manager at the Sierrita Mine. Salerno is considering adding an optional oil concentrator to the skimmer to make the process even more efficient. The concentrator would remove any water picked up by the belt and return it to the sump, leaving nearly pure oil for disposal. Because the mine is located in the Sonoran desert, most of the water evaporates in the hot, dry air and is not reused.

Jan 1, 2009

By George Chung

"The Iron Ore Company of Canada's Carol Lake Project is Located near the western boundary of Labrador, 960 air kilometers northeast of Montreal (Figure 1).The Concentrator operation was started in 1962 with a capacity to produce 7 million tonnes of concentrate per year. The bulk of the concentrate was pelletized at a 6 million tonnes per year pelletizing plant completed in 1963. The remaining concentrate was sold in the international market. The concentrator was first expanded in 1966 to 10 million tonnes per year capacity with the pellet plant also expanded in 1967 to match the increase in concentrate production. In 1973, the concentrator was further expanded to gear with the increase in concentrate sales. In 1977, more than 40 million tonnes of crude ore was processed to produce in excess of 17.4 million tonnes of concentrate and 10.4 million tonnes of pellets."

Jan 1, 1980

By Ryan C. Stephan, Christopher Mark

"Some of the weakest roof in United States coal mines is found in the Illinois Basin. Historically, the Illinois Basin has also suffered from high rates of unplanned roof falls and rock fall injuries. During the past decade, however, roof fall and rock fall injury rates have fallen dramatically, and are now lower than the national average. Several technologies are responsible for the improvement. Roof screen was rarely used ten years ago, but a number of mines now use it on a regular basis to prevent rock falls and rock fall injuries. Other roof skin control fixtures and work practices have contributed as well.The reduction in roof falls is also directly attributed to enhanced intrinsic roof support. Stiffer primary bolting patterns have reduced the magnitude of roof spalling and roof sag. They have also helped create a more intact beam that is able to be suspended from longer cable bolts. Cable bolt design methodology has also contributed to this reduction in roof falls. Some early cable bolt applications were unsuccessful because they employed as few as two cable bolts per intersection. This paper describes a methodology for estimating rock loads for support design.ROOF FALL TRENDS IN THE ILLINOIS BASINAlthough the Illinois Basin has a long history of poor roof conditions and high rates of unplanned roof falls and rock fall injuries, recent data indicates a change in the trend. Roof fall rates are determined by dividing the total number of non-injury roof falls by the total number of hours worked underground. The data is considered for 200,000-man hours. While the Illinois Basin was once a coalfield experiencing roof falls at nearly twice the national average, it is now at or below the national rate. Figure 1 indicates the trend over the past 20 years. In 2008, the Illinois Basin mines experienced 5.06 roof falls every 200,000-man-hours, compared to only 0.75 in 2018.There has also been a significant reduction in the rate of rock fall injuries in the Illinois Basin coal mines, as shown in Figure 2. In 2008 there were 1.48 rock fall injuries every 200,000-man-hours, compared to only 0.43 in 2018.ILLINOIS BASIN COAL ROOF GEOLOGYThe roof rock in the Illinois Basin has been studied in the past to understand the primary causes of roof failures (Molinda et al., 2008; Thompson, 2010; Molinda and Mark, 2010). Rusnak and Mark (2000) evaluated 10,000 laboratory strength tests from 800 rock units and concluded that for shale, siltstone, and sandstone, the strength of the Illinois Basin variety was less than half that of the same rock types in Central Appalachia (Figure 3). A study of Coal Mine Roof Rating (CMRR) measurements in more than 200 US coal mines also found that weak roof was predominant in the Illinois Basin (Mark and Molinda, 2005)."

Jan 1, 2019

By Charles Berkey

CONTENTS PAGE C. P. Berkey-Responsibilites of the Geologist in Engineering Projects (with discussion) 4 Kirk Bryan-Problems Involved in the Geologic Examination of Sites for Dams 10 0. E. Meinzer-Relation of Ground-water Conditions to Leakage of Reservoirs (with discussion) 19 Charles Terzaghi-Effect of Miner Geologic Details on the Safety of Dams (with discussion) 31 G. H. Matthes-More and Better Geologic Data for Use in Hydraulic Projects (with discussion) 47 L. W. Fisher and R. D. Ohrenschall-Geologic Examination of Dam Sites in the Tennessee and Cumberland River Basins 60 C. K. Wentworth-The Geology of Dam Sites (with discussion) 78 L. C. Glenn-The Geology of Dams and Reservoirs 97 H. T. Stearns-Success and Failure of Reservoirs in Basalt (abstract) 111

Jan 1, 1929

A hard lump char containing only 4 to 5 per cent ash, 0&#183;25-0&#183;30 per cent sulphur and a trace of phosphorus has been produced from Latrobe Valley brown coal by slow drying and carbonization of ordinary industrial briquettes at temperatures between 7000 C and 14000 C. The briquette char can be made with a variety of reactivities by varying the maximum temperature of carbonization; the higher the temperature, the lower the reactivity. The strength of the fuel compares favourably with metallurgical coke made from black coal, though it may not be quite so resistant to abrasion. Experiments in a small experimental cupola indicate that it will make a satisfactory fuel for use in iron foundries. It may also find extensive application as a smokeless fuel in other branches of industry and in household heating appliances.INTRODUCTIONVictoria, like South Australia and Western Australia, is apparently without &apos;significant deposits of black coal of coking quality but there are huge reserves of brown coal of high moisture content (60-66 per cent) in the Latrobe Valley (well over 20,000 million tons have been proved) and there are other important deposits in the south central portion of the State (Herman, 1952).For three decades Victorian industries, including a few metallurgical firms, have been using the briquettes made at the Yallourn factory of the State Electricity Commission of Victoria (S.E.C.). These briquettes, having a moisture content of about 15 per cent and a calorific value of approximately 9,500 B.T.U. per pound, are satisfactory for certain grate-type furnaces but they are not suitable for use in cupolas of other shaft furnaces because they disintegrate badly on rapid heating.

Jan 1, 1956

By T. Hudaverdi

Several studies have been conducted on blastability and prediction of fragmentation. The term blastability is used to refer the susceptibility of the rock to blasting and is closely related to fragmentation. Parameters that determine fragmentation by blasting may be divided into two groups. The first group consists of blasting design parameters. The blast design parameters are controllable parameters and determined by the explosive engineers. The burden, spacing between holes, bench height, drillhole diameter, stemming height and explosive density are all design parameters. The second group consists of rock mass parameters that are referred to as uncontrollable parameters. The uncontrollable parameters are physical and mechanical properties of rock and rock mass structure. The strength, modulus of elasticity, density of rock, number of joint sets, joint orientation and joint spacing are among uncontrollable parameters. The rock mass parameters and blast design parameters should be considered together to explain fragmentation process. In this study, a blast database was accumulated to use for multivariate analysis. The blast database contains the blasts performed near Istanbul region quarries and the blasts that were preformed in various mines in the world by the blasting researchers. The cluster analysis was used to separate the blast data into different similarity groups. Group memberships were analyzed and checked by the discriminant analysis. Then, the multivariate regression analysis was used to develop a prediction equation for blast fragmentation for each group. Finally, a validation of the suggested models was performed for various mines. The model comprises blast design parameters, mechanical properties of rock and rock mass structure.

Jan 1, 2011

By P. J. Bartlett

Premier Diamond Mine exploits a kimberlite orebody using cave mining methods. A drilling level is established at the base of a block of ore to effect undercutting that leads to caving of the overlying ore. A second level is developed some 15 metres below the undercutting level to allow extraction of the ore once it has caved. As much as 45 per cent of the rock is removed on the extraction level creating a series of pillars. Undercutting creates abutment stresses, as high as 3 times the virgin stress, which affect the rock mass to a distance of 20 metres from the undercut face on the undercut level and on the extraction level 15 metres below. If the extraction level is completely developed prior to undercutting, where stress levels exceed 20 percent of the rock mass strength, extensive rock mass and support damage occurs as the undercut passes overhead. Rock mass damage takes the form of movement along existing joints and minor shear failure through intact rock. Support damage takes the form of cracking in rigid concrete linings. Where induced stresses exceed 50 per cent of the rock mass strength it is often impractical to keep excavations on the extraction level open. Rock mass damage takes the form of extensive failure of the rock in shear which evidences itself as propagation of the fracture zone around tunnels and footwall heave. Support damage results in destruction of shotcrete linings and concrete footwalls. As the interbolt support fails, badly fractured rock ravels around the steel tendons. Observation and monitoring at Premier Mine and on other cave mines has shown that induced stress levels and consequent damage are a function of :?the extraction ratio on the undercut and extraction level prior to undercutting ?the shape of the undercut face ?the rate of undercut advance ?the distance of leads and lags between adjacent tunnels on the undercut level. To ensure the stability of excavations, all of the above parameters must be carefully planned and the planned mining sequence rigidly enforced. The minimum of development is completed on the extraction level prior to the undercut passing overhead. Required development is then rapidly completed once the area has been de-stressed. Support is installed in three phases: ?Rockbolts are installed as soon as excavations are developed ?Wire mesh, steel tendon straps and a layer of shotcrete sufficient to protect the mesh and straps from damage by blasting and LHD loading is installed as soon as practical thereafter. Inevitably the shotcrete lining has to be replaced once the undercut has passed overhead.. ?Rigid shotcrete linings and concreted footwalls are completed only after the undercut has passed overhead and the area has been de-stressed.

Jan 1, 2000

By Gene Mattila, Ron Boyd

Introduction Rock bolting is a general term that encompasses pinning, roof bolting, rib bolting, and cable bolting. Rock bolting contributes the highest nonproductive cost and is the most significant safety risk in mining today. Total installed bolt costs are now more than $20 per bolt. Rebolting or repair work doubles this. And bolting accidents account for about 30% of all fatal and nonfatal lost time accidents. It is only logical, then, for the mine operator to install the highest quality anchors possible. Every bolt on the market - tensioned, untensioned, full contact, or point anchor - has designed capabilities in axial, shear, and fluctuating ground conditions. Bolts will provide a designed support if material specifications and installation procedures fall within the manufacturer&apos;s recommendations. This also assumes that there are no unexpected geological or operational changes. Proper material specifications and installation procedures should guarantee the mine operator a quality anchor. To help assure this, the Mine Safety and Health Administration (MSHA), the US Bureau of Mines (USBM), and the AMC Manufacturers group developed stringent ASTM materials standards. These govern mechanical point anchor bolts and bolts using threaded and nonthreaded rebar. There are also minimum standards that suppliers must comply with in the manufacture of frictional bolts. However, most frictional bolt standards are based on the manufacturer&apos;s own specifications. It would be impossible to set uniform regulations on frictional bolts. Each bolt provides support in a slightly different manner and thereby requires different material needs. Quality Control In any case, material quality can generally be assured through tight order specifications and periodic quality control checks. The problem with installing quality anchors does not solely lie with the material. It includes all the events that occur from the time a bolt leaves a manufacturer to the time it is installed. The anchoring ability of any major type bolt can be affected by transportation, storage, mechanical performance of installation equipment, installation demands of the operator, proper drilling accessories, and local geology. Table 1 and Fig. 1 document the negative events that can affect each bolt and the degree to which it can be affected. Drilling a hole of oversized diameter, for example, is an operator controlled factor. It will adversely affect most bolts. An over-sized diameter reduces, or even eliminates, the anchorage capability of a mechanical shell. It will reduce the support capacity of a split set to nearly nothing. And it will cause a partial or incomplete grouting around the resin bolt. Frequently, due to bit wear, holes are drilled undersized in diameter and may not be properly cleaned out. In this case, it will be difficult for the operator to push the mechanical shell into the hole, thereby causing him to drill another hole. In cases where the bolt is obviously not providing anchorage, operators will probably drill another hole and place another bolt. Rock Bolt Regulations It is the situations where a bolt is improperly installed in the hole, but gives the appearance of a good bolt, that are frightening. In fact, some safety officials and engineers believe that only 50% to 60% of installed bolts are working. These bolts cause irregular ground load transfers in other bolts and overload them. This causes hazardous situations and overstressing. It is these bolts that fail prematurely 200 to 255 mm (8 to 10 in.) behind the face and possibly contribute to the high accident rate. The government and several manufacturers recognized the need to check the integrity of rock bolts. These programs are in different stages today. Tight torque regulations have been established for mechanical

Jan 4, 1985

By Christopher Mark

The first International Conference on Ground Control in Mining opened with the topic of pillar design. Two classic papers were presented, one by Bieniwski and the other by Wilson. Unfortunately, the two methods were so radically different from each other that it was nearly impossible to reconcile them. Adding to the confusion were the many other pillar strength formulas (such as the Salamon-Munro, the Holland-Gaddy, and the Obert-Duvall, just to name a few) that were also available. Little wonder that discussions of pillar design in those days often ended with anguished cries of ?but which formula is the right one?? The past 25 years have seen substantial progress in the science of coal pillar design. Indeed, one testament to the improvement is the relative scarcity of papers on the topic at recent Conferences. Two factors have been largely responsible for the progress that has been made. The first has been the collection of large data bases of actual case histories of pillar performance in a variety of settings, from shallow room-and-pillar mines through deep cover longwalls. These have made possible the development of empirical design procedures that are closely linked to real world experience. The second important factor is the development of sophisticated computer models that can accurately simulate pillar behavior and roof/pillar/floor interactions. Together, these two lines of research have led to a new understanding of pillar mechanics that identifies three modes of pillar failure: ? Sudden, massive collapse, accompanied by airblast, for slender pillars (width/height<4) ? Squeezing, or slow, non-violent failure, for most room and pillar applications (4<w/h<10) ? Entry failure or bumps for deep cover and longwall applications (w/h>10) It is particularly satisfying that the insights gained from numerical models broadly support those obtained from the empirical studies. While far less controversial than in the past, pillar design problems continue to arise. One recent example is pillar design for highwall mining. NIOSH has just released a software package, called ARMPS-HWM, which employs a number of modern pillar design concepts. Since highwall mining web pillars are long and slender, the greatest danger is that of a sudden collapse. ARMPS-HWM suggests two possible prevention strategies, one which concentrates on the SF of the webs, and the other which creates a ?pressure arch? using properly sized barrier pillars. The paper will close with a discussion of some current needs in coal pillar design, including: ? Updating older empirical methods, such as ALPS, where changes in technology (new types of roof support, more demanding ventilation requirements, faster retreat rates) may have made some of the original case histories obsolete. ? Methods for determining site-specific coal strengths, focusing on bedding plane strength and other factors that may effect confinement, as input for both empirical and numerical design. ? Improved methods for evaluating coal pillar performance for environmental issues, such as surface subsidence and hydrologic impacts, which consider such factors as depth, w/h ratio, water immersion/drainage, and time dependent seam strength.

Jan 1, 2006

By J. C. Haptonstall

The Bismark Mine is located in northern Chihuahua, Mexico, 150 km (95 miles) southwest of Ciudad Juarez (Fig.1). The ore deposit was discovered in 1979 and was placed in production in 1992. The production rate is about 2.4 kt/d (2650 stpd) from an ore reserve of 8.5 Mt (9.3 million st). Grades currently average about 8% zinc, 0.5% lead, 0.2% copper and 50 g/t (1.46 oz/st) silver. Substantial inflows of water into the underground mine complicate the operation. Complex metallurgy poses another challenge for the operators. Bismark is a joint venture with 90% participation by Minas Penoles, SA de CR of Mexico City and 10% participation by Promociones Industriales Banamex, S A de C V (PIB S A) of Mexico City. Anew company, Minera Bismark, SA de CV was formed to operate the property. Cyprus Minerals was a 40% participant in the joint venture but, in early 1993, sold its share to Minas Penoles. The Bismark Mine is one of the largest new underground mining projects built in Mexico during the last 10 years. It is the country&apos;s fifth largest underground metal mine and provides about 1 % of the world&apos;s zinc supply at about 50 t/a (55,000 stpy). Bismark is a trackless operation with overall productivities of 7.6 t/manshift (8.4 st/manshift). The 2.4-kt/d (2650-stpd) concentrator started up in March 1992.

Jan 1, 1994