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By D. G. Chedgy

Careful consideration is given to the type of cleaning equipment and control sys- teas to be used when selecting circuits for wash-ing fine coal (28M x 0 size fraction). However, the desliming, declaying and sizing procedures required to ensure efficient operation normally are not evaluated in sufficient detail resulting in disappointing system performance. This paper details desliming and dewatering circuit design which is extremely effective based on recent test data. The circuit maintains control of the 28M x 0 (declayed) size fraction with minimal misplaced material and maximized circuit efficiency. A discussion of circuit design parameters and control systems, as well as results from the testing program, are included in this text.

Jan 1, 1986

This one day field trip is sponsored by the Denver Region Exploration Geologists? Society, May 14th 2009. The trip highlights major geologic features and some notable uranium occurrences along the north-south mountain-basin boundary southwest of Denver and continuing north to Fort Collins, Colorado. Uranium deposits visited include sandstone?hosted roll front type and polymetallic veins hosted in metamorphic rocks. A replacement ?episyenite? type occurrence is considered via description and hand samples but is not visited owing to logistical constraints. Recovery methods addressed include underground mining and in-situ leach recovery. Samples of Precambrian basement rocks, deposit host rocks, uranium ore samples, and ore polished sections viewed by reflected light microscopy will be available. A road map showing the locations of the stops serves as the front cover of this field guide. The Robinson episyenite-hosted occurrence (not visited) is labeled as Stop 6. A stratigraphic column is provided which shows the position of the host rocks at each stop (Figure 1-1).

Jan 1, 2009

By A. V. Castelli

Since the record year of 1981 through 1985 the tonnage of weighting materials (barite and other heavy minerals) used in drilling fluids has decreased an estimated 49%. As in the past, barite still makes up over 90% of the minerals used for weighting materials in drilling fluid. Domestic production of barite has decreased 69% since 1981 and imported barite has made up over 50% of the total barite consumed since 1982. See Figure 1 for usage of barite. The number of operating grinding plants has decreased from 69 in 1982 to 39 at the end of 1985. With the decrease in demand of weighting materials for drilling fluids and an over capacity of grinding existing, prices for weighting materials have decreased dramatically. Although the American Petroleum Institute has not changed any of the specifications for barite, some users have initiated their own new specifications and have proposed to have them included in the API specifications. The exploration for and development of oil and gas dictates the level of weighting material used in the drilling fluids industry. Most forecasters do not foresee (or predict) increases in the level of drilling until the 90's. The primary or unground barite produced in the U.S. mines has a difficult time competing with imported barite due to the location of the mines. These mines are located in North Central Nevada, with small tonnages being produced in California, Missouri, Illinois, Montana and Washington. The major reason, is the difference in transportation costs from the mines versus ocean freight rates from the foreign ports to the major area of use, the U.S. Gulf. Since 1981, the production of primary barite has decreased from 2,849,000/ST (2,584,596/MT) in 1981 to an estimated 890,000/ ST (807,403/MT) in 1985(1). If it were not for the use of unit trains by some producers the production of primary barite in the U.S. would have probably declined another 20% in 1985. During the same period, of the total amount of barite consumed in the U.S., imported barite ore increased from 417, in 1981 to 65% in 1985. The major exporting countries to the U.S. during this period were China 46.27, Morocco 15.27., Chile 10%, Peru 8.4%, India 7.9%, Mexico 5.4%, Thailand 3.9% and Ireland 2.3%.

Jan 1, 1986

By A. W. Khair

This paper deals with the causes and mechanisms of pillar failure in a southern West Virginia coal mine. The conclusions are based on in-mine observations and analysis of mining methods, eyewitness account of the sequence of events, topographic and geologic conditions, and mechanical properties of mine roof, coal, and floor. The study indicated that massive pillar failure was caused by the collapse of the mine roof. The collapse of the mine roof was due to the combined effects of high lateral compressive stress and high extraction ratio causing excessive bending of the roof beyond its limit. It was assumed that the higher lateral compressive stress under such a shallow cover was probably due to the drastic change in the structural geology of the area and possible residual stresses. Eased on the analytical study it was concluded that the sudden release of elastic energy (shock) due to the rupture of the roof caused the squeezing of approximately eighty mine pillars.

Jan 1, 1983

By H. G. Naidu, Raj K. Singhai, S. P. Singh

Introduction The economics of open-pit mining de¬pend heavily on the cost of overburden removal. It is necessary to maintain a pit slope as steep as possible. Steeper slopes lower the volume of overburden to be handled, thus lower the cost of material handling. There is, however, a limit to which the pit slopes in any given situation can be steepened. Steepening beyond this point could jeopardize the pitwall stability, ren¬dering it unsafe. It might even result in curtailment of pit production. Several factors affect the stability of pitwalls. Some are natural, such as topogra¬phy and regional and local geology. Others are man made, such as pit geometry, ground water management, equipment, drilling and blasting, and the type of excavation. Greater control can be exercised in the latter with proper planning and by following a set of procedures based on field experiences. This article deals with one such control¬lable factor - drilling and blasting. The eco¬nomics of steeper pit slopes is discussed, followed by a record of drilling and blasting practices in western Canadian surface mines. Microprocessor-based instrumentation and its application to blasthole drilling is also described. Economics of steeper pit slopes The economics of steeper pit slopes is not a new issue. It should be reinforced, though, that pit slopes must be as steep as possible to ensure maximum recovery of minerals at a minimum stripping ratio. For operating, economic, and safety reasons, different pit slopes are maintained during different phases of pit development. Figure 1 shows a hypothetical ore body, cylindrical shape, with a radius of 225 m (740 ft) and located 30 m (100 ft) beneath the surface. The mining operations are planned at two different slope angles of 25° and 40°. The ultimate pit depth was established on the basis of the criterion of "break even pit." The variation of the stripping ratio at differ¬ent mine levels is shown in Fig. 2. Table I summarizes the eco¬nomics involved for two pit slopes, 25° and 40°. Figure 3 presents the profits during the life of the mine for these two pit slopes. This study shows that steeper slopes for an open pit operation are far more advanta¬geous than flatter slopes. But, what is the optimum slope at any one point in the life of a moving pit? And how can the stable slopes be maintained over the short- and long-term? Canadian blasting practices A survey of blasting practices in western Canadian surface coal mines revealed that, at most mines, hard consolidated overburden is drilled and blasted. Blastholes are drilled with rotary crawler-¬mounted rigs using bits in the 220 to 310 mm-diam (8.5 to 12 in.-diam) range. Typi¬cal hole depths are 12 to 18 m (40 to 60 ft) to accommodate the pit bench heights of 12 to 15 m (40 to 50 ft). For a 12-m (40-ft) high bench, a typical hole will be 14.6 m (48 ft) deep including the subgrade drilling of 2.6 m (8.5 ft). Half of this hole will typically be a powder column and half will be stemmed. For a 15-m (50-ft) bench, a typical hole will be up to 18 m (66 ft) deep including the subgrade drilling. Drilling patterns are usually 9 m (29.5 ft) spacing and 9 m (29.5 ft) burden. De¬pending on local geology and hole diameter selected, drilling patterns of 8 x 8 m, 8.5 x 7.3 m, and staggered patterns of 11 x 5.5 m

Jan 1, 1989

By James M. Peay

This paper discusses the recently completed organizational development research project, which was addressed to the underground mining industry. The three specific purposes of the project were as follows: (a) to determine whether Organizational Development approaches are feasible in the mining industry, (b) to provide a demonstration that Organizational Development methods can successfully contribute to the safety of underground mines, and (c) to compare the efficacy of classical Organizational Development methods, which are highly trainer- intensive, with trainer-nonintensive management development approaches. The term "Organizational Development" (OD) is here broadly defined as a management intervention for the purpose of assisting an organization with the solution of problems, which are seen by the organization as important. More specifically, we can define OD as a planned effort to improve the effectiveness of an organization through (a) team building and problem solving, which concentrates on the issues of immediate concern to the organization, and (b) efforts to improve organizational communication at all levels of the organization.

Jan 1, 1983

By Albert W. Deurbrouck

The coal reserves in the United States are estimated to be in excess of 300 billion recoverable tons, and current projections of energy requirements indicate our increased dependence upon these reserves. The Department of Energy's Energy Outlook for 1984 projects that coal consumption in the United States will increase from 737 million tons in 1983 to 958 million tons in 1990, and then to 1,110 million tons in 1995 (Table 1). As shown in this table, the future utilization of coal in the United States is closely tied to the electric utility market, a market that consumes about 85% of the coal produced. Be- cause of this dependence on the utility market, the uncertainties of new environmental regulations, particularly those concerning sulfur emissions, could have a significant adverse impact on future coal demands in the United States. SULFUR IN U.S. COALS Most of the Nation's coal reserves contain too much sulfur to be fired directly and meet existing emission control standards. In Table 2, the sulfur contents of the coals of the eastern United States are summarized by sulfur levels according to regions. While the coals of some of the eastern regions are predominantly high in sulfur content, it is important to recognize that large tonnages of coal with relatively lower sulfur content are available in these regions. Furthermore, an often significant portion of the sulfur occurs in the form of iron sulfide (Fes2 or pyrite). Because pyrite occurs as discrete particles of high specific gravity, it can often be removed by simple, inexpensive physical coal beneficiation techniques after being liberated by stage crushing. The potential for physical removal of pyrite by conventional specific gravity processes combined with stage crushing is summarized in Table 3 on a regional basis.

Jan 1, 1986

By ZHONGLING ZHAO, WANTING SONG, GUOZHONG LIU, LEIXIN SHI, ZUI WANG, YUANYUAN LIU, KE GAO, DONGHONG JIANG, WANG LUO, PEICHAO LU, Jianwei Cheng, WANCHENG ZHENG, Gang Zhao, YUNGANG WANG, JIANBING ZHAO

Coal spontaneous combustion is one of the major disasters in coal mines, and it can have devastating effects on the normal mining and production activities of mines. This paper presents the field investigation, field measurement, data analysis, and formulation of fire-extinguishing measures conducted to extinguish the fire from a coal spontaneous combustion incident that occurred in a high-gas coal mine in China.

By Todd Hollingsworth

The Boddington gold mine is located near Perth in Western Australia in the Saddleback Greenstone Belt, a fault-bounded sliver of Archaean volcanic and shallow level intrusive rocks, surrounded by granitic and gneissic rocks. The mine began operations in 1987, producing about 7.7 t/a (250,000 oz/year) until mining ceased in 2001. Built on the footprint of the original mine, the Boddington gold mine expansion project involved opencut mining from two large pits. Expected gold production is around 31 t/a (1 million oz/year) for the first five years. It will be the largest gold mine in Australia. The mine has proven and probable reserves of more than 622 kt (20 million oz) of gold and has an expected life of more than 24 years. The mine is currently solely owned by Newmont Gold Corp., the world?s second-largest gold mining company, employing around 28,000 people worldwide. The conveyor and belt feeder contract for the Boddington gold mine expansion project was awarded by Aker Kvaerner Clough Murray Roberts (AKCMR) to FLSmidth Conveyor Engineering Inc. in August 2006. The contract involved the design and manufacture of 15 conveyors and 17 belt feeders with belt widths ranging from 450 mm to 2.4 m (18 in. to 96 in.) and conveyor lengths from 16 m to 2,172 m (52 ft to 7,125 ft). The total motor power added up to more than 16,700 kW (22,400 hp). The total length of all belting was almost 11.5 km (7.1 miles), and the total fabricated steel weight was approximately 3.7 kt (4,100 st).

Jan 1, 2010

By Alden B. Carpenter

Oil-field brines have been collected and analyzed from Jurassic and Cretaceous rocks in the Mississippi Salt Dome Basin, from Jurassic and Lower Cretaceous rocks, southwestern Arkansas, and from Ordovician and Silurian rocks, southern Michigan. All of these brines are genetically related to evaporites and appear to be brines expelled from halite-bearing rocks, presumably during compaction. The brines in rocks closest to the halite-bearing rocks are uniformly low in lead and zinc suggesting that the metals in metal-rich brines are not acquired within evaporite sequences. The metals in metal- rich brines in Mississippi and in Russia appear to be derived from red beds and may have been originally associated with ferric iron oxyhydroxides. Reduction of these compounds by migrating hydrocarbons appears to release lead and zinc to the associated brines accompanied by the fomation of authigenic ankerite.

Jan 1, 1979

By Clarence O. Babcock

Model pillars of limestone, marble, sandstone, and granite, with length-to-diameter ratios of 3, 2, 1, 0.5, and 0.25 (0.286 for granite), were broken in axial compression to determine to what extent an increase in end constraint increased compressive strength. Radial end constraints of 13 to 23 percent of the average axial stress in the pillar, produced by solid steel rings bonded with epoxy to the ends of dogbone-shaped specimens, increased compressive strength somewhat above that of cylindrical pillars without ring constraint. However, when the results were compared with those obtained by other investigators for straight specimens of several rock types taken collectively, with L/D ratios greater than 0.5, the resulting strengths were not significantly different. Thus, the amount of end constraint produced by the solid steel rings was about the same as that produced by the friction from the steel end plates. In other tests a radial prestress of 3,000 psi or 5,000 psi was applied prior to axial loading by adjustable hardened steel rings to increase the constraint above that obtained for the solid rings. The average radial constraint stress, expressed as a percentage of the average axial pillar stress at failure for the 3,000 psi prestress, was 54.3 percent for limestone, 40.3 percent for marble, 44.7 percent for sandstone, and 23.4 percent for granite. The average radial constraint stress, expressed as a percentage of the average axial pillar stress at failure for the 5,000 psi prestress, was 74.2 percent for limestone, 51.2 percent for marble, 61.6 percent for sandstone, and 29.7 percent for granite. These constraints increased the compressive strength significantly above the strength of straight specimens and solid-ring constrained specimens. These results suggest that large horizontal stresses in orebodies mined by the room and pillar method should increase the strength of the pillars and allow an increase in ore recovery by a reduction of pillar size when major structural defects are absent.

Jan 1, 1969

By Edmund C. Bitzer

The unit costs of dry crushing, within the efficient reduction ranges of equipment now available, are considerably less than those for the reduction of the same material in the same size range by wet mills using grinding media. This is a commonly recognized fact which cannot be disputed; still we find considerable imbalance in costs between the crushing and grinding sections of many operating mills. In other words, there are many instances where more reduction could be accomplished in the lower cost area of dry crushing. In addition, equipment is now available which enlarges this area and makes it possible to consider much finer crushing than has been heretofore economically feasible. There has been a tendency for the past two decades to utilize wet rod-¬milling in the reduction step. This actually amounts to a fourth stage of crushing in a size range formerly served by roll crushers which have fallen from favor on account of high operating costs, Some investigations have indicated that there is not sufficient difference in the operating costs of rolls and rod-mills to give a clear cut advantage to either method, There has for some time been a need and an opportunity for a dry crushing process to improve costs in the range now served by these two systems.

Jan 1, 1965

By R. S. Conti, M. R. Krump, C. P. Lazzara Ph. D., R. L. Derick, L. L. Chasko

A mine fire is one of the most challenging safety issues facing a mine operator and can occur at any location underground. The National Institute for Occupational Safety and Health (NIOSH) and the Twentymile Coal Mine, CO, conducted in-mine experiments to determine the capability of high-expansion foam for addressing underground fires. The tests were conducted in sloping entries with high-capacity diesel-powered foam generators. The following results were observed: a well-designed, maintained and properly operated high-expansion foam generator can propagate a foam plug hundreds of feet in steep upward-sloping (20%) multiple entries against a ventilation pressure; stoppings and partitions designed to contain the foam plug in upward-sloping entries must be substantially constructed; and the predicted quantity of foam concentrate can be significantly less than the actual amount used, due to foam losses, such as bubble breakage on dry surfaces. This paper describes the production of high-expansion firefighting foam and discusses the in-mine experimental tests and results.

Jan 1, 2010

By Robin J. Hickson

The Ortiz Mine is operated by Gold Fields Mining Corporation, a subsidiary of Consolidated Gold Fields Ltd., one of the world's major gold producers. The Ortiz Mine is Gold Fields only heap leaching facility as well as being their first operating mine in the U.S.A. The hydrometallurgical processes utilized for recovery of gold and silver at Ortiz are discussed along with the several practical changes in operation adopted since start-up. Included are heap leaching by cyanidation, concentration of values on activated carbon, desorption, and the final electrowinning steps prior to shipment. Construction costs and pertinent technical data are shown.

Jan 1, 1981

By James R. Hindman

Vermiculite is the mineralogical name generally applied to a group of hydrated, trioctahedral ferromagnesian aluminum platty or micaceous silicate minerals. In microscopic particles the mineral as mineral is differentiated from montmorillonite by a higher Si:AI ratio of approximately 3:l and a higher layer charge and higher cation exchange capacities. Commercial vermiculite, the vermiculite which is mined, beneficiated, and discussed in this chapter, is an important industrial mineral which is formed by the alteration of macroscopic particles of biotite and iron-bearing phlogopite. In natural form it has the size and shape of mica, while possessing the ion exchangeable properties of zeolites and some clays. Within the internal structure of vermiculite lie water molecules, which when rapidly heated to high temperature, transform into steam causing the vermiculite particles to increase in volume. This process of thermal exfoliation produces a lightweight product which finds use in various construction products, agriculture, horticulture, and other industrial applications. In exfoliated form it has the low density and biological inertness of expanded perlite, while maintaining a chemically active surface. Most of the uses of vermiculite are for the thermally exfoliated material. In addition, the cation exchange potential of vermiculite and its layer-silicate structure are properties which can be used as the basis for numerous products including intumescent coatings and gaskets, inorganic films, and the treatment of nuclear and other toxic waste. When chemically and physically modified, it possesses some properties common to organoclays and synthetic compounds. In high purity concentrates and manufactured products, these properties can be altered and optimized in applications to produce superior performance. GEOLOGY Mineralogy Vermiculite was first described as a mineral by Thomas H. Webb in 1824. Webb believed vermiculite to be a variety of talc and described its diagnostic property of thermal exfoliation thusly: "If subjected to the flame of a blowpipe, or that of a common lamp, it expands and shoots out into a variety of fanciful forms, resembling most generally small worms having the vermicular motion exact." Observing that with one exception, all commercial vermiculite has formed by the alteration of iron-bearing phlogopite, a simple definition of vermiculite might be the alteration product of trioctahedral micas, macroscopic particles of which display marked exfoliation when heated rapidly to temperatures above 374°C (the critical temperature of water). The first vermiculite described was found near Worcester, MA, but it was later pointed out by Brush (1866) that vermiculite had been known for some time in Japan as a children's novelty. Because the techniques for mineral analysis were somewhat primitive com- pared to those routinely available today, and because the structure and crystal chemistry of the micas and related minerals were not widely understood, the early literature generated at least 20 "new" minerals which were nothing more than variations in the color and chemistry displayed by vermiculite. Trade names such as Dugarlite, Mandolitem, Strong-LiteTM, and Zonolitem are sometimes used to denote commercial vermiculite, thermally exfoliated vermiculite and their products. Vermiculite particles resemble mica and vary in color from colorless through tan and green to black, depending on the chemical composition. Most physical and optical properties are close to those of biotite, with perfect basal cleavage, hardness between 1.5 and 2.5 Mohs, and specific gravity between 2.2 to 2.8. Fine-sized particles will feel talcose, especially when wet. However, for commercial vermiculite the diagnostic test is that of thermal exfoliation. Vermiculite can display a wide range of compositions, depending on the composition of the original mica and the change of chemistry during weathering and ion exchange. Chemical data for some commercial vermiculites are given in [Table 1]. There are numerous analyses for vermiculite given in the literature, and the analyses of vermiculites from commercial deposits will fall within the range indicated by the average analysis in the table. Many analyses separate the adsorbed water (H2OP) from crystalline water and water as hydroxyl (H20+). The basis of separation of adsorbed water is generally considered to be 100°C. However, hydrate water within the vermiculite interlayer begins to be lost at 75°C and lower. Therefore, measured water content and loss on ignition are combined in the table as H20[total].

Jan 1, 1994

By Allan M. Harvey

Asbestos production in the United States is no match for the vast operations of Canada and the U.S.S.R. The majority of U.S. production is limited to the states of Vermont, California and Arizona, and the type of asbestos is limited to the variety of serpentine known as chrysotile. What little production of other varieties of asbestos such as tremolite and anthophyllite that has occurred over the past 100 years since asbestos became a useful commodity is of practically no significance. However, new developments in the field of environmental and occupational health hazards has startled the mining industry in this country and has already given a few particular companies reason to be concerned about the profitability and probability of their continued operation in the business of mineral processing. At the heart of these developments is the question and controversy over just what constitutes asbestos and what exactly is the definition of a fiber. To the mineralogist or, in fact, to practically anyone engaged in the process of mining and separating minerals there is little or no question about asbestos. It is not a distinct mineral but rather a commercial product serving specific applications for which a number of fibrous mineral types can qualify. To the hygienist pr health authority this definition is too general. They demand something they can detect, measure and control. To suit their purposes and simplify their task of detection, they have labelled the minerals chrysotile, crocidolite, amosite, tremolite, anthoph-yllite and actinolite as asbestos, and any particle of these minerals with an aspect ratio of 3:1 as an asbestos fiber. It does not

Jan 1, 1976

By William Paul Robinson

I would like to first summarize the relevant health information relating to the workers and general populations around uranium mining as a basis for understanding the hazards associated with these operations and their available remedies. This discussion will summarize the literature on Radon and its daughters, among the most toxic uranium decay products. Second, I will discuss waste handling in the uranium industry: past, present and future, discussing exploration, mining, and milling and the volumes of water and solid and gaseous wastes which are associated with these various phases of uranium operations. This paper will also include a discussion of the various past and present materials handling technologies and their responsiveness to the concern for uranium waste isolation. I will then conclude with a summary of recent modifications in the relevant standards as applied to uranium mining and milling. Joseph Wagoner, a research epidemiologist with the U.S. Public Health Service has summarized the literature on uranium mining and its health effects in his written testimony as a commissioned witness in British Columbia Royal Commission of Inquiry into Uranium Mining in January of 1980, entitled Uranium, The U.S. Experience: A Lesson in History.1 Briefly summarizinq his report, as early as 1546, miners of uranium bearing ores in Central Europe were reported to have unusually hiqh frequencies of fatal lung disease2. As early as 1879, the real nature of pulmonary disorders among these miners was first diagnosed to be malignant neoplasia or lung cancers. Of 655 miners dying between 1875 and 1912, 40% or 276 of these Central Europe miners died of lung cancer.4 By 1939, it was reported that lung cancer mortality rate for miners in these Central European mining districts was several times (9.7 per 1000) higher than that for similar populations in Vienna, Austria (0.34 per 1000). Thus, before uranium mining began in the U.S., there was a considerable body of information linking uranium mining to lung cancers. The U.S. experience can be summarized in Table 1 also from Wagoner's B.C. Testimony.1 This table shows that through 1978, 205 respiratory cancer deaths have been observed among white underground uranium miners in the

Jan 1, 1980

By R. R. Kallu

The current 120 psi seal design and evaluation practices are mostly based on simple analytical solutions. Considering the simplistic assumptions and limitations involved in simple analytical solutions and practical difficulties involved in conducting conventional full-scale explosion tests for evaluating 120 psi seal designs, there is a need for developing new scientific methods and engineering practices for design, analysis and evaluation of mine seals.

Jan 1, 2010

By W. Roggenthen, Z. J. Hladysz, G. D. Callahan, W. G. Pariseau, W. Weinig, C. Randolph-Loar, R. Popielak

A Deep Underground Science and Engineering Laboratory (DUSEL) is being planned at the site of the former Homestake Gold Mine in Lead, SD, under the sponsorship of the U.S. National Science Foundation (NSF) in cooperation with the Department of Energy (DOE). This national facility will host a variety of multidisciplinary science and engineering experiments, including physics experiments that require shielding from cosmogenic radiation. Current plans include construction of large rooms (up to 55 m in width) at depths of 1.48 km (4,850 ft) and 2.25 km (7,400 ft) below the surface. Preparation of the preliminary design required the development of a geotechnical characterization program, including preliminary drilling and coring, core logging, drill hole mapping with a televiewer, drift mapping of the geology and rock structure, in situ stress measurements, and laboratory testing of the mechanical properties of the rocks and discontinuities. Overall rock quality is very good to excellent. Site investigations included geotechnical assessments for the construction of caverns and laboratory modules. The results are favorable and reaffirm selection of the Homestake site for establishment of a Deep Underground Engineering and Science Laboratory.

Jan 1, 2011

By F. H. Nehl

The U.S. 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 co-deposited. Decreasing Cu co-deposited will reduce refinery costs. Direct current was applied to the cell in pulsed and square wave voltages at 0.7 A?h per 150 mL of solution. Times tested for each cycle ranged from 1 to 10,000 ms. Graphite, lead, and stainless steel were evaluated as electrode materials. Square wave e1ectrowinning at 70? C 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 pct, and a period of 100 ms to 1,000 ms gave the following results: Ag concentrations decreased from 34 µ g/mL to 0.1 µ g/mL, Au concentrations decreased from 54 µ g/mL to 0.2 µ g/mL, and Cu concentrations remained almost constant at 560 µ g/mL.

Jan 1, 1993