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By E. S. Chalpin

On 25 August 1971, the U.S. Bureau of Mines contracted with the Aerojet Liquid Rocket Company to perform an engineering research program to be completed on 25 May 1972. The program objective was to evaluate the advantages and disadvantages of the underground use of several power sources including but not limited to a closed cycle diesel, electric cable/motor, rechargeable battery/ motor and exhausting diesel for mine shuttle cars, with emphasis on health and safety. This report, Research in Advanced Power Systems for Mining Health and Safety consists of four (4) main parts: (1) Hazards and Safety Analysis of Mine Power Systems, (2) Evaluation of Closed Cycle Diesel Power System for Mine Applications, (3) Closed Cycle Diesel Power System Consumables Compatibility, and (4) Concluding Remarks and Recommendations. The hazards introduced to a mine by shuttle car power system failure modes were evaluated and reported. A "fault tree" analysis was generated for each power system and the probability of death or injury during an eight (8) hour shift and for a week of operation (105 hours) were estimated. Advantages and disadvantages were also reported for each system studied. Aerojet's assessment, based on a "forced decisions analysis," of the four shuttle car power systems as to their relative health and safety is that the exhausting or open cycle diesel is preferred, followed by the closed cycle diesel, the electric cable/motor and the rechargeable battery/motor. The assessment was based on ten parameters pertinent to the health and safety aspects of shuttle cars. Drawing layouts of two typical mine shuttle cars incorporating closed cycle diesel powerplant concepts are included in this report. The performance parameters established for the two concepts are tabulated. A study was made of the logistics involved in transporting and replenishing the consumables needed to operate a shuttle car equipped with a closed cycle diesel. The closed cycle diesel powerplant consumables -oxygen, potassium hydroxide, and diesel fuel - were analyzed as to their compatibility with materials found normally in a gassy coal mine. Spill tests simulating inadvertent leakage of liquid oxygen in a coal mine were performed. A critical evaluation of existing oxygen sensitivity data and the relationship to the use of oxygen on a shuttle car in a coal mine environment is presented.

Jan 1, 1972

By Michael Li, Arie Moerman

"Paste backfill as a means of supporting underground mining is gaining popularity. Despite this trend, engineering and design of mine paste backfill systems are still largely based on empirical methods rather than established scientific principles. Because of a lack of sound understanding of mineral paste flow, some degree of trial and error is often involved during commissioning and operating of new paste fill systems. New operations often have to be built with flexibilities and contingencies, escalating the overall project costs.Disciplines that can lend some science to mineral paste flow include slurry transport, fresh concrete rheology, and soil mechanics. This paper reviews the applicability of some scientific and engineering principles found in those related disciplines to mineral paste flows. Examples are given to illustrate important points. The phenomenon of mineral paste flow is examined from new perspectives. Gaps in our understanding of mineral paste rheology are identified and from those some future research directions naturally emerge. Status and problems of the current engineering and design practice are also briefly discussed.INTRODUCTIONPaste backfill as a means of supporting underground mining is gaining popularity. Despite this trend, engineering and design of paste fill production and distribution systems today are still largely based on empirical methods rather than established science. One of the weaknesses of such empirical methods is the lack of scalability and transferability of rheological data from pre-engineering test phases to production phases. It is not uncommon that flow fiction data generated from expensive loop pumping tests turn out to be uncharacteristic of the paste in fullscale production. In such a case the operators have to re-climb the learning curve, sometimes with expensive disruptions to the mining operation."

Jan 1, 2002

By N. L. Piret

Primary and secondary copper smelting dusts are being generated in varying quantities with very dissimilar characteristics due to the different compositions of feed materials and the type of process utilized. After categorizing the multitude of dust types, the criteria for on-site and off-site recycling of smelter dust in terms of type and concentrations of the accompanying elements are established. Thereby, consideration is given to the important aspects of: ? optimization of copper recovery ? quality of copper product ? valorization of accompanying elements ? elimination and disposal of deleterious impurities ? minimization of impact on operation ? effects on operating costs ? environmental impact of dust treatment on overall process Existing commercial, emerging and proposed copper smelting dusts, pyrometallurgical as well as hydrometallurgical, treatment schemes are reviewed and evaluated technically, economically and environmentally in an attempt to streamline the vast number of proposed options. Thereby, the high selectivity achievable in hydrometallurgical dust processing due to it being complementary to the smelting process is expected to be given increased consideration.

Jan 1, 1995

By James H. Cobbs

The utilization of large diameter drilled holes for mine shafts seems to have had its beginning in Europe about 1850 (28) where the mining industries of Germany and The Netherlands first used drilled shafts. Recent shafts in softer formations of diameters as large as 25 feet and as deep as 1, 700 feet have been drilled in both The Netherlands and Germany. The first shaft drillers utilized cable tool or churn drilling techniques which limit both hole diameters and drilling speeds. Though some churn drilling of shafts is still performed, rotary drilling has been the dominant technique for many years.

Jan 1, 1973

By S. B. Levin

The Bureau of Mines has been investigating deposits of critical and essential minerals in the United States and Alaska since 1939. During the fall of 1942, the Bureau of Mines put down nine diamond-drill holes at the Cape Rosier zinc-copper-lead mine in Hancock County, Maine, that aggregated 2,883 feet of bore. These holes supplemented 13 holes, aggregating. 5,501 feet, put down on the property about 2 years earlier by the St. Joseph Lead Co. Veins of mixed sulfides were intersected in 15 of the holes drilled. Beneficiation tests were made on samples of ore obtained from old dumps. Zinc-copper-lead sulfides occur at the Cape Rosier mine as lenses in a zone of altered agglomerate beneath and between a series of diorite sills. The mineralized zone strikes F. 17° E. and dips southeast. Between July 1881 and September 1883 the mine produced about 10,000 tons of ore reported to have contained 20 percent zinc, 2.8 percent copper, and some lead. The crude ore was hand-sorted, and some. 3,000 tons were reported shipped.

Jan 1, 1948

By James P. May

Operating mine haulage trucks near the crest of stockpiles and waste dumps is a potentially hazardous practice often resulting in a localized slope failure and a dump point accident. The U.S. Bureau of Mines has developed a computer code (INSLOPE3) based on the kinematic method of limit analysis which determines an admissible truck weight (upper bound value) for varying distances from the slope edge. The program considers the material strength parameters, slope geometry, and inertial forces induced by vehicle braking. The program can be utilized to assist in the determination of safe operating distances for a haulage truck from a slope edge, the development of safe vehicle operating procedures, and the admissible weight for a static concentrated loading near the crest of an otherwise stable slope. This report provides information on the use of INSLOPE3, interpretation of the results for mining applications, and instructions for acquiring a copy of the software.

Jan 1, 1991

By S. A. Scott

Mintek developed the MinataurTM leach-solvent extraction-precipitation technology as a cost effective, easily operated alternative to conventional gold refining processes. The technology was first commercialized at the Harmony Gold Refinery in 1997 to produce 99.99 per cent gold product and has subsequently led to development of various value-addition steps including the casting of various investment bars, gold wire drawing and gold potassium cyanide powder production. In addition to refining gold to high purity, if an appreciable amount of silver is present in the feed material it is also refined to 99.99 per cent purity. The major advantage that the Minataur? technology has over other gold refining processes is the relatively low gold lock-up and reduced operating costs. The operating cost1as at April 2001 for the expanded Harmony refinery (24-ton per annum gold production) was approximately half that of conventional smelting and toll refining. The extractant employed for solvent extraction is relatively cheap and highly selective for gold, making the technology attractive in situations where significant quantities of platinum group metals or base metals are present.

Jan 1, 2004

By Marvin S. Seppanen

Computer programers are often required to make modifications to unfamiliar Fortran routines. This Bureau of Mines report describes a computer program designed to aid the programer in such a situation by reorganizing: Fortran routines. This reorganization includes a sequential renumbering of the routine's statement numbers, a sequential renumbering, and relocation of format statements, an alphanumeric reordering of dimensioned and typed variables, a uniform pattern of text spacing, and a sequential numbering of the records in the final Fortran routine. The computer program has been extensively tested by the author and has proved to be a valuable tool for reorganizing Fortran routines developed under contract and later utilized by the Bureau, and for preparing routines for publication.

Jan 1, 1975

By T Atkinson, A Cobb, R AJIington

Risk management is concerned with the whole spectrum of financial, geological, engineering, legal, governmental, social and occasionally ethnic issues, concerned with mining projects. The inherent risks in mining operations are well publicised, but new issues and hazards arrive on the scene and lake precedence, but the older risks still remain, while insurance cover may not he ‘full cover’ in the future. It has long been recognised that quantitative approaches to risk analysis can play a major role, but although these methods are capable of considerable theoretical precision, the lack of sound risk input data and experienced personnel limits the usefulness of theoretical analysis. It can, however, greatly assist in comparing cases, in identifying areas of risk and help in the investment decision process. Its major contribution to a mining project is that it imposes an investigative discipline that also assists in improving the quantitative approach to risk analysis and hence risk management. Risk analysis and risk management also requires sound common sense and major inputs from experienced geologists, mining, mechanical, electrical and mineral processing engineers. The Engineering Council, in 1992, issued a ten point code of professional practice entitled ‘Engineers and Risk Issues’ (The Engineering Council. 1992), covering our professional responsibilities, the law, our conduct, the need for a systematic approach to risk issues, the use of professional judgement, communications, risk management, evaluation of risk, our professional development in this field and the encouragement of public awareness of risk issues. This is a specific area of training perhaps long neglected by our industry and academic institutions (Atkinson. 1983).

Jan 5, 1996

By Sidney Katell

An estimated capital investment of $76,499,800 will be required for a plant designed and built to produce 90 million standard cubic feet per day (std.c.f.d.) of synthetic pipeline gas through Lurgi gasification of bituminous char. This investment figure does not include the pretreatment plant, the cost of which is assumed to be reflected in the cost of the char fed to the gasification system. Shift conversion, C02 removal, and methanation are required after gasification in the Lurgi unit to make a gas suitable for pipeline transmission. Capital investment data are based upon equipment costs, January 1959. Based on a 90-percent or 330-day operating year, the operating cost before profit and Federal taxes with the char price as a variable will amount to the following: [ ]

Jan 1, 1959

By C. C. Jobes

Operating large face equipment, such as continuous mining machines, is one of the most basic and most dangerous elements of an underground coal mining operation. The operator usually performs cutting and tramming tasks in a restricted workspace with reduced visibility. During the execution of these tasks, operators often cannot see all of the locations needed to safely operate the continuous mining machine (CM ), which has sometimes resulted in accidents. In an effort to prevent these accidents, researchers at the National Institute for Occupational Safety and Health (NIOSH), Pittsburgh Research Laboratory, examined the continuous mining machine operator?s current information cues, or visual attention locations, so that risk of injury relative to the task, equipment and workplace environment can be evaluated in the future. The eventual goal is to develop interventions to enhance the information cues, and thereby increase the safety and health of machine operators. This paper details the operator interview process, determination of visual attention locations (VALs) and simulation of various mining scenarios.

Jan 1, 2011

By B. C. Alberts

Planning the use of South Africa's energy resources, especially coal, is of the utmost importance. The State and the mining industry require reliable official estimates of the economically mineable coal reserves for planning purposes. An overview of al the previous reserve estimates is given, with specific attention to the two most recent estimates-those of Petrick and De Jager. The differences between these two estimates are highlighted. The following important aspects are discussed: . the need to update the information regarding South Africa's mineable coal reserves, . the necessity of understanding and agreement between the coal-mining industry and the State department responsible for updating the coal-reserve data, and . the importance of the specific assumptions used in the updating of the data. A downward adjustment of mineable reserve figures is proposed, and the factors that influence the meaningful exploitation of the coal reserves are discussed. A recommendation is made as to how the Waterberg coal reserve should be exploited, and it is concluded that the Soutpansberg Coalfield cannot be exploited economically in the foreseeable future. An evaluation of the status of the coal-export trade is made and compared with the assumptions and targets set at the start of the coal-exportation scheme. The comparison shows the extent to which the exploitation of the coal reserves is being negatively influenced by the current market situation. Recommendations are made regarding planning for the coal-export market, and the aspects that have to be taken into account to keep South Africa in the export market are identified. Reference is made to the fact that the present unrestrained wage and salary adjustments, which show a total disregard of South Africa's poor productivity record, require serious thought and speedy rectification if the country is to survive in the export market. A salary and wage strategy must be developed to suit the specific needs of the Republic of South Africa.

Jan 1, 1987

By M. Rakoma

Following the blasting fatality on 17 December 2009, Landau Colliery was forced to change from the blasting practices that had been utilized since the start-up of the mine. Since then, attempts have been made to optimize blasting results. The challenge is rooted to the mining environment - during the early life of the mine, mining took place in virgin areas, whereas currently mining is taking place in areas overlying old underground bord and pillar workings. The underground bord and pillar workings aggravate phenomena such as spontaneous combustion, hot holes, and sinkholes. The 2009 fatality was attributed to hot holes causing premature detonation. Landau has been struggling with poor results from their overburden and coal blasts; leading to poor fragmentation of overburden and the PI parting that lies between the No.1 and No.2 coal seams. An extensive literature survey was completed in order to identify possible alternative solutions that could be evaluated for this unique mining environment, which demands a high degree of caution during mining to ensure factors such as dragline stability and personnel safety. The literature survey confirmed that the previous blasting technique, which was mainly characterized by air decking, was indeed the more efficient, although it was deemed unsafe because of the possibility of hot holes and the high sensitivity of detonators. Potential modifications to the blasting technique were modelled in a Microsoft Excel simulation that was available from the mine, and pit visits were conducted in order to determine whether implementation would be a problem. From the model it is evident that results can be improved by changing the current burden and spacing ratio. Visiting the pit revealed that the tie-in pattern and drilling inaccuracies are contributing to problem areas.

Jan 1, 2013

By W J. Dyck

Past attempts to establish an integrated steel industry on the West Coast of Canada have been unsuccessful for at least three reasons: (a) insufficient reserves of a suit-able iron ore supply; (b) a relatively small local market for iron and steel; and (c) the non-availability of an economical process to produce primary iron on a small scale. Today, the reserves of economically mineable ore deposits are ample to support a steel production of 100,000 to 200,-000 tons per annum. The local market has increased so as to be able to absorb most of this capacity. With-in the next few years, a small integrated steel plant, including the "H-Iron" direct iron ore reduction process for electric furnace charge stock production, will be able to show a profitable operation. Introduction

Jan 1, 1963

By A. K. Ratte, C. M. Brown

"IntroductionIn autogenous mills, grinding of the ore feed is accomplished by the direct action of the variously sized rocks making up the feed on each other and the mill interior. Size reduction is achieved without the addition of grinding media foreign to the mill feed, differing in this respect from rod, ball, and pebble milling. The term ""semi-autogenous"" has entered common usage in the industry to describe autogenous mills to which supplementary grinding ball additions are made to increase the rate at which size reduction of the mill feed occurs, beyond that achieved solely by the action of rock on rock. In a semi-autogenous grinding (SAG) mill, 50Jo to 10% of the mill volume may be occupied by grinding balls, whereas a range of 25% to 45% is more usual in ball milling.The comminution of mineral ores using ""conventional"" techniques has been described in previous papers. In conventional operations, size reduction of mined ore, and liberation of the valuable mineral content, is typically accomplished using flowsheets employing two to four stages of crushing. Crushing is followed by one to three stages of grinding and classification using various combinations of rod, ball, and pebble mills. This paper describes operations employing autogenous or semiautogenous grinding as the principal stage in the comminution process. Primary crusher discharge, or in some instances run of mine ore, is fed directly to an autogenous or semi-autogenous grinding circuit. Final product sizing may be accomplished in a single grinding stage, the mills operating in closed circuit with classification equipment, or in several stages using autogenous and semi-autogenous mills for primary grinding, and ball or pebble mills for secondary grinding. In several operations a portion of the primary mill discharge is crushed in an intermediate comminution step, these circuits being called autogenousball- crusher circuits (ABC). Many different flowsheets exist, but common and important to most are crushed ore stockpiles, reclaim feeder and mill feed conveyor systems, pumps, screens, and classifiers. Autogenous grinding is employed in both wet and dry processes, although today wet predominate."

Jan 1, 1989

By Johnny Zhang, Yang Anguo, Li Weifeng, Xueyi Guo

"SKSTM oxygen bottom blowing smelting technology was developed by SKS-ENFI at SKS Smelter, Hunan in the 1990s to replace the conventional sintering process and thus to meet the high national EHS standard for lead.Since then, two major milestones have been achieved:1. Since 2002, the SKS oxygen bottom blowing smelting-blast furnace reduction has been in operation at SKS, YGL, Chizhou smelter etc. Now in China, about 50% of the lead (4.2 millions t/year) is produced using SKS process; and2. Since 2008, the blast furnace has been replaced by direct reduction of molten rich lead slag at YGJ smelter, which is more energy-efficient since no slag cooling is required. This two stage process (SKS Oxidation- Direct Reduction) is named as the YGL process.Since 2008, numerous technological innovations have been implemented in the furnace design, process control, and maintenance at YGL. The plant has been successfully demonstrated its unique excellent performances on all aspects from processing capacity, product quality, energy reduction, cost reduction and emission reduction. The purpose of this report is to update all the improvements with proven tracks.Meanwhile, a preliminary comparison on the following aspects between the YGL process and other similar technologies has been discussed:1. Plant design and operating conditions;2. Mass and heat balances, metal (Pb, Ag) recovery;3. Consumables, oxygen gun and refractory lining life;4. Energy consumption and operating costs;5. Sulfur capture and emission reduction etc.It is concluded that the modified SKSTM process is a clean SO2-free emission process. The process is characteristic of other features such as no extra fuel addition, low capital with a small foot-print design, low OPEX, high throughput, easy operation, long refractory lining life and high productivity (95%+ on line time annually). It offers an alternative cost-effective technology for lead smelting industry."

Jan 1, 2014

By T. G. Sieber, T. D. Tiemann

The extraction of quartz from siliceous iron ores by dissolution in sodium hydroxide solution has been described in detail in AIME Transactions. l923 data reported covered the temperature range from 230-416°F, where a continuous increase in the amount of silica extracted with increase in temperature was found. It is the purpose of this paper to describe some unexpected results of more recent work exploring principally the range of 500600°F where higher rates and greater amounts of silica extraction were antici-p ted. In brief, the experimental procedure consists in digesting the ore under pressure in solutions of sodium hydroxide, filtering the suspension and weighing and analyzing the dried residue for iron content. The amount of extraction is calculated from the weight loss or from the iron analysis of the residue and that of the original ore.

Jan 1, 1969

By V. G. Pienaar, B. M. Mbiya, P. T. Slatter

[a constant b constant cc onstant c' constant D pipe diameter (m) Dshear shear diameter (m) E error function fth theoretical friction factor F function F' function K fluid consistency index (Pa sn) kv valve loss coefficient L pipe length (m) n flow behaviour index N number of data points ?P valve pressure drop (Pa) Q flow rate (m3/s) Re Reynolds number Re3 Reynolds number (Slatter) V average velocity (m/s) Vann annulus velocity (m/s) x variable X number of independent variables y variable Y number of fundamental dimensions z variable µeff effective viscosity (Pa s) ? valve degree of opening d partial derivative ? fluid density (kg/m3) [ ] function]

Jan 1, 2007

By C. M. Paiencia

The Bureau of Mines studied the potential availability of primary antimony (Sb) from demonstrated resources in 21 mines and deposits in market economy countries (MEC's). Twelve of these properties were evaluated as producers and nine as non- producers. The 21 studied properties contain 499,600 mt Sb as demonstrated resources, of which 304,000 mt is recoverable as antimony metal; another 973,500 mt Sb is contained in inferred resources. Recoverable antimony from demonstrated U.S. deposits is only 10,700 mt, just 3.5 pct of the MEC total. Domestic production of primary antimony from operating mines has averaged only 561 mt/yr over the 10-yr period 1975-84. This represents just over 5 pct of total U.S. consumption of primary antimony, which averaged 11,081 mt/yr during the same period. The United States will thus continue to rely on imported ores andlor concentrates, antimony oxides, and metal to satisfy domestic industrial requirements. At production costs up to $1.15/lb Sb, 110,000 mt of refined antimony is potentially available from 10 of the 21 studied properties. At $2/lb, 294,000 mt Sb is available, and at $3/lb, the total rises to 304,000 mt Sb available from all evaluated producers of refined antimony.

Jan 1, 1986

By R. Rodgers

The Pinto Valley concentrator of BHP Copper Inc. restarted its sulfide mining operations in September 2007 to produce copper and molybdenum concentrates. During the plant production ramp up, the copper/molybdenum circuit quickly reached the copper grade and recovery targets, but the production and quality of the molybdenum concentrate continued to be a challenge. Even though several improvements in the molybdenum circuit were completed, the final molybdenum concentrate did not meet contractual specifications. After reviewing several options to improve grade and reduce contaminants, the installation of three flotation columns were chosen. The column design and construction were done using in-house craftsmen and engineers. Ten days after commissioning, the columns successfully reached a steady operation and began producing molybdenum concentrate to the required specifications, that is, high molybdenum tenors and low copper and iron content. The Pinto Valley operations, are comprised of the Pinto Valley concentrator, the Pinto Valley dump leaching and a solvent extraction/electrowinning (SX/EW) plant. The concentrator operated from 1972 to 1998 when mining and milling operations were curtailed for economic reasons. The SX/EW plant continued operations for copper cathode production. After a cessation of mining operations of almost 10 years, Pinto Valley operations began a restart project in January 2007, to bring the mine and the concentrator back into production. Copper concentrate production started in October 2007 after refurbishing and upgrading the facilities with a new DCS system and filter plant. The daily targets of the concentrator were to treat 56.7 kt (62,500 st) of ore, produce 762 t (840 st) of copper concentrate containing 27.5 percent copper at recoveries of 86 percent and produce 2.2 t (2.5 st) of molybdenum concentrate containing more than 80 percent MoS2.

Jan 1, 2010