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By Donald C. Violetta

During 1963 more than 200 million tons of coal was consumed by the public utility industry. Ash residual, from the burning of this coal has caused disposal problems. In order to prevent air pollution, dust collection systems have been installed to suppress the fine ash emitted from boilers before it is discharged into the atmosphere. The collected ash, called fly ash, must be removed from the dust collection facilities and disposed in some manner, usually at a substantial cost to the utility. Many years ago fly ash disposal was not a serious problem because the demand for power was relatively low and coarse coal was burned on stoker type traveling grate machines. The low burning rates utilized on these grates did not produce large quantities of air-borne coal ash while the mechanics of this system of burning resulted in a clinkered residue of cinders. Both of these conditions resulted in the production of only small quantities of fly ash. As the demand for electricity increased, a method of burning pulverized coal was developed and was introduced to the public utility industry in 1918 at Milwaukee, Wisconsin.1 This method of firing was considered a significant advance in steam electric generation as it (1) permitted the use of coals that could not be economically fired by other methods and (2) has generally resulted in improved efficiency, greater generating capacity and lower costs.2 The utilization of larger generating units using pulverized coal substantially increased the quantity of fly ash which resulted in a' more severe disposal problem.

Jan 1, 1966

By J. E. Murphy, L. A. Walters, F. H. Nehl, G. B. Atkinson

The US Bureau of Mines investigated the selective electrowinning of Ag and Au from cyanide solutions contaminated with Cu with the goal of decreasing the amount of Cu codeposited. Decreasing Cu codeposited will reduce refinery costs. Direct current was applied to the cell in pulsed and square wave voltages at 0.7 A h/150 mL of solution. Times tested for each cycle ranged from 1 to 10,000 msecs. Graphite, lead and stainless steel were evaluated as electrode materials. Square wave electrowinning at 70°C (158° F) gave the best separation of Ag and Au from Cu. Applying 2.5 V during the duty cycle, 0.0 V the rest of the period, a duty cycle of 10% to 30%, and a period of 100 msecs to 1000 msecs gave the following results: Ag concentrations decreased from 34 µg/mL to 0.1 µg/mL. Au concentrations decreased from 54 µg/mL to 0.2 µg/mL and Cu concentrations remained almost constant at 560 µg/mL.

Jan 1, 1995

By Y. Lv, H. Ye, D. Du

Electrolytic manganese residue (EMR) is a kind of industrial solid waste with high silicon content that contains harmful metal elements such as chromium (Cr), copper (Cu), lead (Pb), cadmium (Cd), nickel (Ni) and cobalt (Co). Developing a highly effective and environmentally friendly process to treat EMR is important, both for the reutilization of metal resources and for environmental sustainability. In this study, the bioleaching of available silicon from EMR using silicate bacteria is investigated, with flask experiments conducted to optimize the parameters in the bioleaching system. For metal ions, the concentrations of manganese (Mn), Co, Ni and Pb in solution were measured to determine the environmental safety effects before and after bioleaching. The results show that the bioleaching rates of available silicon in the slag were mainly affected by the temperature, pH and pulp density. Under the optimal conditions of temperature of 30 °C, pH of 7.0, pulp density of 5 percent, shake flask rotational speed of approximately 210 rpm, particle size of approximately 0.150 mm and leaching time of 16 d, the concentrations of available silicon in the bioleaching system rose to approximately 110 to 140 mg/L. The metals remaining in the residues were mainly bonded in stable fractions, with concentrations lower in the leachate after bioleaching. This study demonstrates that silicate bacteria can be used to increase the contents of available silicon from EMR, and that bioleaching technology can realize the harmless resource utilization of EMR.

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

By M. S. Prasad

This paper presents the process methods of beneficiating various copper-cobalt and copper-zinc ores mined by Gecamines in Zaire. Geology and mineralization of both sedimentary and vein type deposits are described with special emphasis on the production of oxide, dolomitic and sulfide concentrates. Description of different types of concentrators and their treatment methods are discussed for the production of sulfide, oxide and dolomitic concentrates. Production of concentrates, their consumption and stocks are briefly reviewed. Modernization of the concentrators with substantial investments on grinding, classification, flotation methods and tailings treatments are also discussed to improve the metallurgical efficiency of the concentrators with a view to improve the production of concentrates both qualitatively and quantitatively.

Jan 1, 1990

By N. A. Abdel-Khalek

An experimental study was carried out to compare the results of laboratory column flotation versus conventional flotation for the cleaning of fine coals of different rank and their oxidized samples as well as discrete size fractions. Surface oxidation caused a significant reduction in the flotation yields achieved with both methods. However, the column flotation produced concentrates? of lower ash content. The recovery -size curve achieved with column was similar in shape to that obtained with conventional cell, with superior results in the region of fine fractions. The influence of some column operating parameters was also discussed.

Jan 1, 1990

By Jian Deng, Hehua Zhu, Lianyang Zhang

Rockburst is defined as a sudden and violent explosion of rock on the surface of a rock excavation. Rockburst has been regarded as the least understood and most feared hazard problem for underground rock excavations. Traditionally, rockbursts were classified into the strain, pillar, and fault-slip categories. However, rockbursts are also known to occur at stress levels well below the rock yield strength, in mines where the rock is far from brittle, or without the presence of geological discontinuities. A plausible explanation of rockbursts is buckling, which has been identified as one of the main mechanisms for rockbursts. It is observed that rockbursts are often induced by dynamic loadings (pulses) from adjacent blasting. Therefore, the present paper studies the effect of pulse shape and amplitude on pillar rockbursts induced by dynamic loadings from blasting using the method of numerical analysis. The governing equation of a pillar’s motion is set up and the mechanism of pillar rockbursts is investigated through a practical example. The results show that it is not the shape of a pulse but the impact energy to decide the occurrence of pillar rockbursts. If all other parameters remain the same, pillar rockbursts are more likely to occur when the pulse duration is longer. The parametric resonance and the perturbation of dynamic loadings are also found to play an important role in the occurrence of rockbursts. These insights would provide invaluable clues about the mechanism and for better control of pillar rockbursts.

Apr 26, 2022

By B. J. Scheiner

The Bureau of Mines, U. S. Department of the Interior, is conducting research on a dewatering technique for fine particles generated during the mineral beneficiation process that will recover a portion of the water now being lost using conventional waste disposal methods and significantly reduce the time required for land reclamation. The technique utilizes a flocculant, polyethylene oxide (PEO), that forms strong stable flocs that can be dewatered rapidly on static and/or rotary screens. In field tests, phosphate clay waste has been dewatered from a nominal 2.5 percent solids to greater than 20 percent solids. In other field tests, coal-clay waste has been dewatered from a nominal 3 percent solids to greater than 55 percent solids, in addition, small-scale continuous tests have shown that bentonitic clays can be dewatered from 12 percent solids to 40 percent solids.

Jan 1, 1983

By J. S. Hall

The performance of a full-scale balling drum circuit is well known to fluctuate, and is difficult to control. The fluctuations stem from causes that are not readily quantifiable. Results from plant tests show that surging of the recycle tonnage is dependent, inter alia, on the method of bentonite addition, small variations in drum or cutter bar speed and the method of drum water addition. This paper presents detailed observations from experiments on full-scale drum circuits, including preliminary results of a computer measurement and control scheme. Size distributions measured at points along an industrial drum are presented, together with the results of a mathematical model fitted to one plant data set. Model predictions are also included emphasizing the importance of screen performance on circuit stability.

Jan 1, 1983

By H. Rush Spedden

A process has been developed and piloted to reject copper, lead, iron, phosphorus, bismuth, antimony, elemental sulfur, and insol from molybdenite concentrate without decomposing the contained molybdenite and rhenium. The impure concentrate is first heated with concentrated sulfuric acid at 260 to 290°C for 1 hour. This bake not only oxidizes the contaminating sulfides, but also activates the concentrate for later rejection of insol by flotation. Two stages of leaching follow the sulfation step. The first is an acidified water leach which extracts the oxidized copper, iron, and phosphorous; the second leach rejects the lead, bismuth, and antimony into a warm concentrated solution of sodium chloride. Final upgrading is achieved by conventional froth flotation. The chemistry of the process as well as results obtained from its implementation in a 5,000 lb/day pilot plant are presented.

Jan 1, 1971

By D. Dolinar, T. P. Mucho

Horizontal stresses are caused by global platetectonic forces. During 1995 alone they were largely responsible for the closing of two longwall mines in the United States. This paper presents six case histories from Pennsylvania, West Virginia, Kentucky and Alabama. In each case, a mine encountered roof falls or difficult ground conditions at the headgate caused by horizontal-stress concentrations. The problems are detailed, and the control measures adopted are described. In most cases, nearby longwall panels without stress concentrations were trouble free. The paper also discusses detailed measurements that were made at two adjacent Pennsylvania longwalls. One headgate was oriented to avoid a horizontal-stress concentration, and the other was not. Eliminating the stress concentration dramatically reduced roof support loads and roof deformation. The paper concludes that proper panel orientation and sequence is the key to maintaining headgate ground control. The optimum orientation is not parallel to the maximum horizontal stress, as previously thought, but rather it is 20o in the stress shadow of the gob. Other stress-control techniques, including artificial support, are briefly discussed

Jan 1, 1999

By J. A. Marts, R. C. Gilmore, Jr. Bogin, S. Saki, J. F. Brune, J. W. Grubb

"Researchers at the Colorado School of Mines have studied the influence of headgate side ventilation controls near the longwall start-up room on the formation of explosive gas zones (EGZs) in underground coal longwall bleeder ventilated gobs. In a project funded by the National Institute for Occupational Safety and Health (NIOSH) researchers developed a Computational Fluid Dynamics (CFD) model to study the formation of methane-air mixtures in the gob, start-up room, and bleeder entries. The relative change in size and location of EGZs are examined in response to ventilation controls in the headgate side bleeder entries near the start-up room. Modeling suggests that adjustments to the ventilation controls can be made to minimize the size of the EGZ. However the EGZ may form in or around active working areas regardless of the ventilation control settings. Research found regulators on the crosscuts into start-up room and first entry inby can force air into to crosscuts outby the start-up room causing more air to flow through the gob. INTRODUCTION Underground longwall mine ventilation systems may operate with either a bleeder ventilated gob or a progressively sealed gob. Bleeder ventilated gobs must meet the standards for safe operation of the bleeder system by examining bleeder system effectiveness. As defined in the United States Mine Safety and Health Administration (MSHA) Program Policy Letter (PPL) No. P13-V-12 (MSHA 2013), the pillared area or gob is “any area where a reduction in pillar size has occurred during retreat mining” and includes the “caved and broken material, the interstices within the caved material, and the open spaces surrounding the caved material.” The PPL further states that the intended purpose of ventilating the gob is “to protect active workings from the hazards of methane accumulations …” The requirements for operating a bleeder system as stated in The United States Code of Federal Regulations (CFR 2013) Title 30 §75.334(b) is to pass ventilation air though the gob “and to continuously dilute and move methane-air mixtures [from the gob] away from active workings and into a return air course …” where the concentration cannot exceed 2 percent methane (CFR 30 § 75.323). Further clarification of the bleeder system as described in the PPL states the following: “A bleeder system includes the area from which pillars are wholly or partially recovered, bleeder entries, bleeder connectors, and all associated ventilation control devices that control the air movement through the area. Bleeder entries are air courses designed and maintained as part of the bleeder system.”"

Jan 1, 2015

By F. F. Peng

For a given flotation system, the rate-controlling step in fine coal flotation can be the attachment of bubble to the fine coal particles. The probability of attachment depends upon the surface characteristics and the collector adsorption on the coal particle surfaces. Thus, to further improve the flotation rate and selectivity is to alter the distribution of hydrophobicity on coal particles efficiently by applying the effective means of collector dispersion. In order to enhance the combustible recovery and selectivity of fine coal flotation, and lower the amount of collector used in the fine coal flotation processes, four modes collector dispersion/addition techniques are tested. They are direct liquid addition in pulp (mechanical agitation), ultrasonic energy dispersed collector into pulp, atomized collector into pulp and gasified collector transported by air and formation of collector-air bubbles in pulp. The effect of four collector dispersion/addition methods on the floatability of -600 µ m (-30 mesh US sieve) coal samples from four ranks of coals were studied. The ultrasonic energy dispersed collector and atomized collector flotation processes enhance flotation yield and most significantly improve the kinetics of fine coal flotation. The gasified flotation process can increase the selectivity of separation and decrease the collector consumption due to starvation flotation is achieved in this process. The flotation efficiency of gasified collector process is as high as that of conventional flotation. It is because that the effect of the decrease in the combustible recovery is offset by the increase of selectivity. In addition to flotation experiments, induction times of bubble attachment to the bed of fine coal particles for four ranks of coals are measured.

Jan 1, 1992

By G Luxbacher

Computer simulation of mine ventilation systems is rapidly becoming an essential tool in mine planning and design. This paper examines the utilization of ventilation simulation techniques to evaluate the number, shape, size and layout of mine entries in a typical coal mining system. Since the mains and sub-mains serve mainly for transportation and ventilation and the ventilation system is critically important to the health and safety of the workers, there exists a great need for optimal designs in this area. Evaluation criteria, could include, among others, minimization of power costs, shaft requirements and leakage. Constraints of this optimization problem include health and safety regulations and production and mine infrastructure requirements.

Jan 1, 1977

By B. Anameric

Considerable effort has been directed towards understanding the carbothermic reduction mechanisms of self-reducing, fluxing dried green balls to produce pig iron nuggets. Given the geometry of the situation, some investigators believed that the shrinking core model was applicable. Hence, this study involved investigation of the applicability of the shrinking core model to pig iron nugget production. The experiments involved heat treatment of dried green balls utilizing either a laboratory-scale resistance box furnace or a gas-fired muffle furnace at various furnace temperatures and residence times. The products were analyzed for the following: (i) preferred reaction pattern, (ii) distribution of the metallized areas, (iii) % iron content variation in the pig iron nuggets from surface to center, (iv) carburization pattern and (v) physical and chemical properties of the pig iron nuggets. It was determined that the reduction of iron oxide occurred simultaneously throughout the sample and not beginning at the surface and propagating inward; thus the shrinking-core model was not applicable. Heat transfer and gas diffusion from surface to center were not the rate-limiting steps for reduction and carburization reactions.

Jan 1, 2011

By J. L. Huiatt

The U.S. Bureau of Mines and Great Salt Lake Minerals and Chemical Corp. developed a froth flotation process for concentrating potassium salts from Great Salt Lake solar evaporites containing about 5 percent K. The crude evaporites, which were predominately halite (NaCl) with smaller amounts of kainite (KCl.MgSO4.2.7H20) and schoenite (K2SO4MgSO4.6H20), were conditioned in a saturated brine solution, thus converting nonfloatable kainite to the floatable mineral schoenite. Schoenite was then selectively floated with a medium-molecular weight fatty acid. Continuous small-scale pilot plant testing indicated that 80 percent of the potassium could be recovered from the crude evaporites in flotation concentrates assaying about 11 percent K.

Jan 1, 1975

By J. R. Sims

Gold assays from production blastholes in the Mercur Hill Pit were used to delineate known ore zones and to determine how well previous exploration drilling defined these zones. The most important factors causing inadequate ore zone definition were: 1) inappropriate location of drill holes relative to the ore zones, 2) presence of stopes from previous mining, 3) sampling and assaying errors, and 4) smoothing of the ore zone boundaries during ore reserve estimation. Infill drilling strategies were evaluated to determine the efficiency with which a given drill pattern defined the known ore zones. A drilling strategy was developed which improved ore zone delineation when compared to techniques which maintained uniform drill spacing.

Jan 1, 1992

By R. J. Roman

Several research groups have spent considerable time investigating coarse ore leaching. Coarse ore leaching includes both heap and dump leaching. The term "heap leaching" will be used in this paper to mean any type of coarse ore leaching. The results of their work are generally in agreement: the progress of leaching can be described by the shrinking core model and its associated mathematics. The shrinking core model is schematically represented in Figure 1. The particle is contacted by the leach solution on its outside surface and the solution migrates inward by filling the pores on the particle. As the solution filling the pores comes in contact with grains of ore mineral the reagent chemically reacts with the ore mineral to solubilize the contained metal. The solubilized metal diffuses outward to the particle surface and is washed away by the leach solution flowing past the particle. The leaching reagent in the particle interior is replenished by the diffusion of reagent from the surface of the particle to the interior of the particle. If the chemical reaction between the leaching reagent and the ore mineral is relatively fast compared to the diffusion rate of the reagent in the solution in the pores of the particle then leaching proceeds from the outside of the particle inward resulting in a "shrinking core" of unreacted ore surrounded by a "leached shell". The rate controlling step in the reaction can initially be a chemical reaction between the reagent and the ore mineral grain, however, once the reaction has proceeded a short period of time diffusion of the leaching reagent through the leached shell will be the slow step and hence the rate controlling step in the production of metal from an ore particle.

Jan 1, 1992

By Terry I. Mudder

Metal complexed cyanides in wastewaters form as a result of interactions of free cyanide with metals present in the wastewater and exhibit varying degrees of stability, toxicity, and treatability. Thiocyanate, a pollutant commonly found in cyanide bearing wastewaters, is formed through the interaction of free cyanide with sulfur containing species (i. e. pyrrhotite) both present in the wastewater. In certain industrial processes, such as the beneficiation of gold and silver, cyanide is an essential reagent. Since free cyanide, complexed cyanides, and thiocyanate are potentially toxic to humans and aquatic organisms, these compounds must be removed from wastewaters prior to their discharge into surface or ground waters serving as potential potable water sources, marine or fresh water habitats

Jan 1, 1984

By A. W. Khair

This paper presents a strength formula for the design of coal pillars. The formula is derived from results of laboratory experiments, studying the effects of primary factors influencing the strength of the coal specimens/pillars. Five primary factors, significantly affecting the strength characteristics of the coal specimens/pillars were identified and extensive laboratory investigations were carried out. The five primary factors were namely 1) coefficient of friction between loading platens and coal specimens, 2) elastic properties of the loading platens, 3) coal specimen geometry, 4) coal specimen size, and 5) rate of loading. Four to nine levels of variation for each primary factor on the strength of coal were studied. An empirical formula, implementing the variation of the primary factors affecting the strength of coal specimens/pillars, has been developed. A comparative analysis of the results provided by this and some well known formulae is also presented.

Jan 1, 1994