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By J. B. Gwiazda

Introduction Thick seams constitute a considerable portion of Polish coal reserves. In producing mines and those under construction, the reserves where the height is more than 4.5 m (15 ft) total about 24% of all the hard coal reserves. The reserves that are under urban and industrial areas can be extracted using stowing methods for surface protection. About 50% of thick seam reserves can be extracted by caving methods. Thick seam mining methods Thick seams are usually extracted by multilift methods. There are some methods of thick seam mining that generally can be divided into caving and stowing methods. The stowing method is more expensive, but must be used when mining to protect surface features. The more known methods of thick seam mining include • extraction of thick seam in ascending order with stowing ; • thick seam mining in descending order with consolidated stowing; • thick seam mining with gob consolidation; • thick seam mining in descending order with roof caving ; • sublevel caving thick seam mining system; and • thick seam mining system using an artificial roof. Thick seam mining with introduction of an artificial roof The most advantageous method of thick seam mining is the method using an artificial roof. The artificial roof usually consists of wire mesh. Such wire mesh must be durable and must prevent small rock fragments from passing through. The mesh is usually made of steel properly protected against corrosion. The artificial roof is placed on the coal floor of the top slice (Fig. 1). Caving rock falls on the artificial roof and is thus separated from the coal. The next slice is extracted under the artificial roof. The producing face of the next slice must progress within 10 m (33 ft) behind the front of the first slice. As the second producing face progresses, the wire mesh moves down on the coal floor and may be used as an artificial roof for the next slice. Thus, one artificial roof allows the extraction of several consecutive slices. Application of the artificial roof allows for clean extraction and reduces coal losses. It also eliminates the hazard of spontaneous combustion. Introduction of the artificial roof has been limited due to a lack of artificial roof laying equipment. Manual laying of wire mesh is a labor-intensive operation, and it delays face progress, resulting in poor productivity. To eliminate these problems, artificial roof laying equipment was designed at the Polish Central Mining Institute. The support and the method of laying operation is covered by a patent. Artificial roof laying powered support To extract thick seams using an artificial roof, four-leg shield supports with lemniscate linkage were developed (Fig. 2). This shield is used in the first top

Jan 3, 1987

By Z. Hyder

Simulation and modeling is powerful tool to analyze real time data and can be used to analyze different scenarios using appropriate software that may reduce need of infield data collection process and can predict reasonably accurate results. In this paper the real time data of a mine is used to simulate and model the stresses on the pillars and evaluate the pillar strength and safety factor. The material properties assigned to the model are obtained by testing the representative samples taken from the selected locations of the mine. The mine under investigation has 50 ft wide pillars and 50 ft wide rooms under an average overburden of 650 feet. The effect of reduced pillar dimensions and increased roof span on overall stability of mine and extraction ratio is analyzed using LaModel [1] software. The predicted results of the reducing pillar size are analyzed to compares the current mining practices with the optimized room and pillar size to get better extraction ratio without comprising the safety.

Jan 1, 2011

By E. J. Garrison

A.C. Noble presents a clear and concise summary of the factors pertinent to a competent reserve evaluation of a mineral occurrence. However, I believe that he confuses the difference between a resource and a reserve. By definition, a reserve estimate includes all known economic, legal, mining, metallurgical and environmental factors impacting the recovery of the product(s) found in the deposit. A resource estimate, however, is an estimate of what is there - what is available for exploitation and should not include economic and other considerations. Only after defining the resource, its extent, location, physical and chemical properties, can an intelligent selection process begin of the extraction and processing technologies most appropriate to the circumstances encountered. The early and almost certainly inappropriate application of these often cookbook choices has probably lead many companies to walk away from good properties. (How many mines have been found in others rejects?) An example of the process is the early selection of metallurgical samples, after encouraging values are found in a few holes or surface samples, in an attempt to determine the economics of the prospect. Unfortunately, the sampler does not know how representative of the occurrence the sample is because he does not know what the deposit looks like. As a consequence, the property becomes saddled with a bad metallurgical sample that is often inappropriately used in evaluating the deposit which it does not represent.

Jan 1, 1994

By G. Deschênes

In Part I of the work, gold and silver were shown to be efficiently extracted from Kupol samples using lead nitrate in a cyanide leach system. In Part II, a comparison of the technical and economic merits of a new technology called CELP (CANMET Enhanced Leaching Process) was made with the cyanidation/AVR(Acidification-Volatilization-Recycling) process using 8 metallurgical variability samples as well as a large composite sample which represented initial two years of Kupol mill feed. The composite sample had 33.4 g/t Au and 289 g/t Ag. The metallurgical demonstrated that the CELP produced very similar gold and silver leach extractions compared to the standard Cyanidation /AVR leach conditions at 1/2 the cyanide concentration (1000 ppm NaCN in CELP vs. 2000 ppm for cyanidation/AVR leach conditions). It was estimated that the total cost per ton, were approximately the same for the two technologies (total cost of the CELP with the use of Merrill-Crowe at $46.48/ton, with cyanidation/AVR at $46.13/ton). However, the implementation of the CELP would eliminate a capital expenditure of approximately $5 million. Eight variability samples used to confirm the previous results showed that on average the gold and silver leach extractions with the CELP and the feasibility leach conditions were very similar for gold extractions, but 4.6% higher for silver extractions compared to the conventional cyanidation. The decision was made to remove AVR cyanide recovery from the Kupol flowsheet and implement the CELP at Kupol. The ability to use conventional leach equipment with the CELP and simplify the flowsheet was also a factor, given the remote location of the Kupol Mine.

Jan 1, 2009

By L. Hoffman, F. Seymour, S. Hoerger

Newmont has a large integrated operation in Northern Nevada with numerous mines, stockpiles, and processing facilities. There are over ninety defined metallurgical ore types and over sixty defined gold recovery process options. To take better advantage of the available synergies, a complete Nevada wide mine plan and process model along with a mixed Integer and Linear Programming optimization tool has been developed and implemented. This system is maintained and operated by a Nevada Planning Group. Standard in-house planning software tools and methods have been implemented at all mine sites. While increasing the effectiveness of planning in Nevada, these changes have allowed a significant reduction in overall Mine Engineering staff.

Jan 1, 1999

By Dermot M. Winters, Brent T. Wahlquist, C. Y. Chen

The enactment of the Surface Mining Control and Reclamation Act (SMCRA) in August 1977 addressed subsidence resulting from underground mines, in addition to creating the requirements for reclamation of surface coal mines. Subsidence control regulations pursuant to SMCRA were initially promulgated in March 1979. Several revisions have since been made. In addition, due to court challenges some regulations have been remanded and others are still on appeal. In the meantime the Secretary has agreed to initiate rulemaking on some issues. It is yet to be determined how the prohibition of certain areas from surface coal mining operations under Section 522(e) of the Act applies to subsidence. This presentation discusses the status of the regulations and issues related to or affecting the subsidence regulations.

Jan 1, 1986

By N. Arbiter

The theoretical analysis of tumbling mill energetics and performance has largely neglected mill dimensionality, and, in particular, the importance of the length/diameter ratio. This is in spite of the fact that practice varies substantially in this area: geographically, for semi-autogenous and autogenous mills, as between North America and South Africa/Scandinavia; historically, for overflow ball mills, for which the L/D ratio has increased significantly from the earliest small mills to/the largest mills currently. The present study is concerned primarily with the influence of the L/D ratio on the design and operation of overflow ball mills; on the occurrence of the overload phenomenon; and on the limits, if any, it may impose on mill capacities because of critical pulp axial velocity limits. It will be shown that the shape face or is of major importance in this area; and that its adjustment to the extent that this is practical should remove the diameter limitations previously postulated for this mill type.

Jan 1, 1990

By A. Karami

It is well established that an ordered sequencing of charge initiations in a blast pattern is an important design consideration that affects blasting performance. Delay times must be chosen such that damage to the rock mass is controlled. The accuracy of delay detonators is also of major concern. A large scatter could lead to an incorrect blast sequencing that would lead to damage to the rock mass causing increased over-break, degraded fragmentation, increased fly-rock and poor vibration control. In this study, the effect of using accurate delay detonators on the explosive energy generated is investigated. For this purpose, power spectral density of the blast vibration for delayed shots are calculated and compared with the case where holes are shot at the same time (no delay). At the end the results are presented and discussed.

Jan 1, 2000

By J. A. Lombardi

Acid-Minesite-Solutions (AMS) are defined as copper-containing pit lake, tailing pond, acid-mine-drainage, and heap and waste dump irrigation solutions. All of these solutions are characterized by copper, acid and water contents that have value as individual commodities. HW Process Technologies, Inc. (HWPT) has developed nano- and hyper-filtration membrane methods for the separation of copper and other base metals, acid and water from AMS. These treatments are collectively called HWPT?s Engineered Membrane Separation® (EMS®) systems. The EMS® method of copper recovery produces small volumes of higher-concentration copper solution from large-volume, copper-dilute AMS. A case study of the 4,000 gpm Cananea EMS®-AMS plant for the recovery of copper from an acid-pit-lake resource is described. The EMS® method of copper removal from AMD and tails pond waters has been bench-scale demonstrated and are described, complete with projected budgetary capital and operating costs. Additional treatments for the recovery of sulfuric acid and high-quality water from the permeate (membrane passing) waters for the above-described AMS treatments are presented, with budgetary capital and operating cost estimates. [ ]

Jan 1, 2009

By H. Do

In flotation, hydrophobic particles are selectively collected by air bubbles, and the bubble-particle aggregates enter the froth phase. However, less hydrophobic particles drop off from the froth phase, while more hydrophobic particles are recovered. Thus, the froth phase provides an important cleaning mechanism. In the present communication, the froth-phase recovery process has been modeled by considering particle retention times and the carrying capacities of bubbles. The model predictions are compared with the experimental data reported by Seaman et al. (2006). The results show that the froth-phase recovery varies with particle size, particle hydrophobicity, and bubble coarsening.

Jan 1, 2008

By Massimo Concilia, Remo Grandori, Paolo Nardone

Paul Y. – SELI Joint Venture had been contracted to build the longest section of the Strategic Sewage Disposal Scheme (SSDS) – Stage I. This involved two tunnels under urban and marine areas excavated by high performance hard rock open type TBMs operated from deep shafts. A specific ground treatment program was developed using advance probing and grouting to control water inflow into the excavations. The article describes: the technologies adopted, the experienced problems, the encountered productions and their interrelation with the water inflow amounts, the possible improvement in the technologies for the excavation of the future stages of the same project as well of any other project under similar conditions.

Jan 1, 2001

By Ali Ahmadi, Andreas Rausch, Angela Binder, Stefan von der Mark, Dhananjeyan Jeyaraj, Jana Thomas, Nils Nöther, Stefan Wittek, Thomas Ulrich, Eleni Gerolymatou, Mario Kolling, Helmut Mischo, Michael D. Martin, Tugba Sen, Jens-André Paffenholz, Erik Farys, Oliver Langefeld, Wilfried Ließmann

This paper outlines the MOVIE project, which is dedicated to the development of a digital twin for an underground laboratory at the research and education mine “Forschungs- und Lehrbergwerk Reiche Zeche” in Freiberg (Germany). Data from induced rock deformations and temperature fluctuations is collected using fiber-optic sensors. The objective of the project is to establish a virtual laboratory for real-time analysis of the interactions between geology, geomechanics, geometry and ventilation. This will be achieved by integrating models from these disciplines with artificial intelligence, creating a sophisticated hybrid model. The insights gained will enhance safety and efficiency in mining through improved digital planning options, ultimately contributing to more effective resource management and operational decision-making in different mining environments.

Feb 1, 2025

By Y. J. Wang

The selection of a shaft size for ventilation purposes is a classic problem of concern to mining engineers. This presentation is concerned with the sizing of conventionally-sunk circular shafts with concrete or shotcrete linings. Costs of sinking shafts of three different depths (from 500 feet to 1000 feet) and six different diameters (from 14 feet to 24 feet) were obtained through questionnaires to shaft contractors. The primary objective of the economic model presented is the balancing of detailed annual capital and operating costs to achieve the minimum overall cost of the shaft installation over its life. The economic model is similar to previous analyses but treats several of the individual cost categories in an updated fashion. Included in the paper is a detailed analysis of the effects of varying power costs on optimal shaft size. Also included are a statistical analysis of current shaft costs and comments on the choice between shotcrete and concrete linings.

Jan 1, 1977

By Malcolm D. Flavel

Much has been said and written about the world's, and particularly the united states of America's, energy crisis. The relative changes between cost components of production have altered because the days of cheap energy have now gone. Decisions that made capital investment reduction attractive to treat large tonnage lower grade ore bodies appear to be no longer valid because average energy costs for comminution that were less than 1 cent per kilowatt hour in 1972, are now more than 3 cents and will shortly become 6 cents according to some projections. Increased capital investment to improve processing energy efficiencies are certainly going to be pursued. Energy today is the largest cost component in the average mineral production facility. It takes more than 12000 kilowatt hours to produce one ton of finished steel and more than 60000 kilowatt hours for each ton of copper and aluminum.

Jan 1, 1980

By Jack T. Work

Spiraling costs in every segment of the open-pit mining cycle is a fact of life that every operator must reckon with. The introduction of various shock absorbing devices for rotary drilling is one recent innovation that may significantly reduce overall drilling cost, increase drill availability, and lower maintenance requirements. In a recent all-inclusive test in taconite ores, shock absorbers on Bucyrus-Erie BE 60R and Gardner¬Denver GD 120 rotary drills were directly responsible for overall savings of up to $41.06 per hour. What is It? A shock absorber is simply a device installed between the rotary drive head and drill steel that cushions the two. It functions in the same manner as springs and shocks on an automobile. Some benefits associated with use of the tool are readily apparent on a daily basis; other benefits are realized over longer time intervals and through continued use. A properly designed shock absorber will: •protect the drill from severe vi¬brations, improve rock bit performance, allow application of the best weight-on-bit and rotating speed under all drilling conditions, •function effectively over a wide range of applied bit weights, • have a spring rate soft enough to perform its primary function of absorbing both rotary and vertical vibrational forces generated by rough drilling, •be reliable and subject to as little maintenance as possible, •be easy to install and remove, and • be of such an economic benefit so as to offset the added expense of its application. A shock absorbing tool developed by BJ-Hughes Inc. was used in a recent test. The "Smoothdrill" is just one of the products currently available. This unit absorbs both rotary and vertical shock loads and has no mechanical parts so lubrication is unnecessary. A rebuild of the unit consists of removing the worn rubber element and inserting a new element, an operation taking less than one hour. Minimizing Vibrations Severe vibrations on a blasthole drill operating in broken and discontinuous formations have visible effects. These vibrations often cause structural damage when allowed to' continue, resulting in fatigue cracks in the mast and related structural components in the drill. When rough running formations are encountered, drilling must be interrupted and possibly shut down. Usually, the drill operator resumes drilling using less down pressures and various rotational speeds until firmer drilling is encountered. An effective shock absorber can substantially reduce these vibrational forces so more down pressure can be applied in bad ground. It is generally accepted that longitudinal vibrations and cyclic forces are more effectively reduced if a tool is designed to incorporate a relatively "soft" shock absorber. The softness of any shock absorbing tool is best judged by its spring rate-the weight required to produce a given amount of deflection, commonly expressed in units of kg per mm or pounds per inch of deflection. All shock tools use a resilient member such as rubber springs, steel springs, fluid, or gas springs to absorb

Jan 4, 1980

By Peter Shuttleworth, Chris Fesq, Eddie Woods

Precast concrete segmental linings for tunnels have traditionally been designed with conventional steel bar reinforcement. In many cases, this has been satisfactory in service, but in others, major problems have occurred with the durability of the linings. Commonly this has resulted from corrosion of the reinforcement, mainly from chloridesin ground water, causing spalling of the concrete and leading to very expensive and disruptive repair programs. This paper discusses the relative merits of designing tunnel linings with steel fiber reinforcement to provide long-term durability, improved impact resistance and toughness for greater load capacity across joints, enhanced robustness during transportation and erection, plus improved fire resistance with the addition of polypropylene fibers. Guidance is also given in the selection of the correct fiber type, mix design and testing methods to establish the optimum dosage required and production techniques to ensure even distribution and compaction.

Jan 1, 2005

By Paul B. Queneau

Formulation and commercial utilization of solid-waste incinerator slag by-product is the focus of this paper. Particular attention is given to iron aluminosilicate melts having compositions similar to that of nature's basalt The earth?s crust can be viewed as an environmental laboratory for predicting stability of various silicates; small-scale laboratory experiments are necessary to confirm findings there from. An ideal slag consumes flux additives at acceptable cost, is compatible with furnace refractory, flows readily from the furnace, and has physical properties suitable for its commercial utilization -all within the critical constraint that the slag is environmentally stable. Examples of successful solid waste slagging practice, including tonnage, utilization and pricing, are provided from metallurgical and waste treatment operations of North America. Western Europe and Japan.

Jan 1, 1991

By B. Boberg

Annual requirements of uranium to fuel the 441 currently operating nuclear reactors in the world is 170 million pounds U3O8 (65,385 tonnes U). 2005 global production was 108 million pounds (41,595 t U), of which only 2.7 million pounds U3O8 (1,039 t U), was produced from US uranium deposits. US nuclear utilities represent 26% of the global nuclear capacity but only obtain 2.5% of their supply from US uranium mines. Since 2002 annual global production increased from 93 million pounds U3O8, a 15% increase and the US production increased from 2.3 million pounds to 2.7 million pounds, a 17% increase. Reasonably Assured Resources (RAR) of 12,332 million pounds U3O8 (4.7 million t U) are reported from a wide variety of countries and types of uranium deposits. This RAR suggests that there are significant global uranium resources potentially producible at a price of $50 per pound U3O8 ($130 per kg U). Table 1 shows the distribution of global uranium resources. This table, prepared from historic data maintained by the World Nuclear Association, does not include some areas of the world that are currently considered to contain significant resources of uranium such as Mongolia. Table 1 suggests that the United States contains the fourth largest uranium resource in the world with 890 million pounds U3O8, behind Australia, Kazakhstan and Canada in resources that are producible at a price of $50 per pound U3O8 ($130 per kg U). Within the US, the states that demonstrate the greatest amount of uranium resource are Wyoming, New Mexico, Texas and the Four Corners Area of Arizona, Colorado and Utah.

Jan 1, 2007

By Harold J. Schock

The cases of no auxiliary ventilation and ventilation with tubing were mathematically modeled, using a turbulent dispersion model. The scooptram that is used to load is modeled as a variable-speed, variable-strength source of pollutants moving in and out of the dead-ended drift. Model predictions are compared to experimental measurements. A parametric analysis was performed to determine the effect of exhaust pipe location and time spent out of the drift, as well as to determine if a relationship existed between CO2 and NO levels. It was concluded from this analysis that when modeling a load-haul-dump cycle, it is necessary to use a model that has the capability of predicting changes in source strength as well as source location. It was also determined that four relevant factors include effective heading height, apparent longitudinal diffusion coefficient, transverse concentration gradient at the face, and average dispersion width of the exhaust jet. As these factors change with heading geometry and ventilation configuration, they were determined experimentally, Carbon dioxide was used as a tracer to calibrate the model; since, knowing the fuel flow rate, it is possible to calculate CO2 production. The above four factors in the computer model were calculated using a regression analysis, comparing predicted local concentrations to measured local concentrations. The model was then verified after this calibration process using additional experimental results. With the calibrated model, it is possible to predict pollutant concentrations at any location in the heading using an engine map and the previously calculated flow field. A parametric analysis shows that the exhaust pipe location can have a significant impact on operator exposure to exhaust gases. It was also shown that NO concentration varies directly with C02 levels. Predictions using the dynamic model are contrasted with constant source strength (pollutant flow rate) results. An analysis of these predictions for computing time-weighted average NO shows the dynamic model predictions of operator exposure to this pollutant to be less than one-half of the value predicted using a constant source model.

Jan 1, 1981

Rising energy costs are a growing concern among mine operators. At sur’ace operations, the size of a dragline, production requirements, pit geometry and layout, and operator skill are all factors in potential energy savings. The following tips are offered as a practical means to conserve energy demand, and can be applied to diesel fuel draglines as well.

Jan 7, 1980