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By T. A. Wright

THE SAMPLING of waste materials containing copper, lead and tin has taken on a new significance within recent years, and is of increasing importance, on account of the entry of some of the copper refineries into this field and the tendency towards consolidations among the smelters. In the past, such material was sold on the basis either of a rough or grab sample or of casual inspection of the lot. The two principal sources of this waste are the processes of fabrication and the scrapping of automobile, electrical, plumbing, railroad and other equipment. In both cases much material can be sorted, salvaged and sold by classification and grade to be melted "as is" or with the addition of new or virgin metal. The increasing care in classifying exercised in the large manufacturing plants tends to increase the proportion of this usable scrap; on the contrary, the increasing use of more complex and special non-ferrous alloys tends to make sorting and grading more difficult in turnings, borings, clippings, etc., and consequently is making sampling of more moment. In tin, lead, aluminum, and zinc furnace by-products such as ashes, drosses, skimmings, etc., the trade practice of evaluation has not varied much. For instance, tin drosses, battery lead, and aluminum skimmings are ordinarily sold on what is called' "metallic button" or "metallic recovery" or "metallic yield." Copper and lead waste, however, is being sold more and more on actual copper and lead contents based on sample and chemical analyses. This paper will be devoted chiefly to the sampling of materials suitable only for refining and recovery of one or more of the metals present. The methods described are often based on the fact that but one metal, such as copper, is to be paid for. The loss of volatile constituents in melting does not appreciably affect the final copper result, as the "after-melting" assay can be readily calculated to the "before-melting" basis, but if smelter efficiency continues to increase, so that in time zinc, for instance, may be recovered on a paying basis, the method of sampling dirty brass, for example, will have to be changed also.

Jan 1, 1928

By Chas. A. Dobbel

The Chief Consolidated properties are situated in the Tintic mining district of Utah, being included in Juab and Utah Counties, about 70 miles south of Salt Lake City. The drilling referred to in this paper has been carried on in the Chief Consolidated No. Pmine, the Gemini mine, and the Eureka Hill mine at Eureka, the Grand Central mine, above Mammoth, and the Apex Standard mine just south of Dividend. The occurrence of ore through the district is typically in lenses on bedding or breaks, and in connecting pipes, without accurate geometrical trend so that the ratio between the total length of development in drifts and crosscuts to the square feet of ore exposed is very high. A section of one of the western channels (Fig. 1) through the Tintic district, shows, roughly, the irregularities calling for excessive development work, so typical of the Tintic mines. This irregular trend of the ore constantly calls for excessive cutting up of the whole area for offshoots and the connecting links always present. In the planning of prospecting and development, no regular system may be laid out without overlooking some orebodies or segments of orebodies. It was with these facts in mind that Cecil Fitch, the general manager of the Chief Consolidated Mining Co., decided that some method of drilling was the more feasible way of successfully cutting up and bringing about the discovery of ore believed to exist; yet this drilling had to compete with a great margin of cost over crosscutting, as a crosscut in the Tintic district, even though barren, brings out geologic facts that, in many cases, are worth the price of the working itself. Diamond drills met with the success that might be expected in a district that has a reputation for the Gear and loss of diamonds; after 6 or 8 months, this method was discarded on account of the high costs. Simultaneously, a small churn drill outfit was set up underground but a few shifts proved its inapplicability. At the same time, the ordinary hammer drill was used to drive prospect holes with jointed steel. Just what equipment to develop for this work with the hammer drill involved much experimenting, as nothing was found in the literature

Jan 1, 1925

By C. P. Fiske

The pottery of the New Jersey Zinc Co. (of Pa.) is equipped to make three classes of refractories; namely, spelter vessels, spelter condensers, and high-grade fire-brick. The most important of these are the spelter vessels, on account of the extreme severity of the service which they must perform, necessitating great care in the selection of materials and in the whole process of manufacture. Next to the spelter vessels in importance come the condensers, as large numbers are used in the manufacture of spelter, and improvements in quality result in material decreases in working cost. While the fire-brick which form the laboratory, or high-temperature portion of the spelter furnaces, are placed at the foot of the list, they are nevertheless of much importance, since high quality will decrease the frequency of the always expensive shutdowns for repairs. To promote simplicity, the manufacture of these three products will be described separately in this paper, except in so far as materials and the crushing and handling system common to all are conccrned. Materials Bond Clay.—The bond clay used entirely at the Palmerton pottery is a St. Louis clay furnished by the Grand View Fire Clay Co. It is stored in a system of covered bins, and no attempt is made to weather it in the accepted sense of the term " weathering." Some years ago our clay was weathered before pugging, but we did not find the benefit sufficient to offset the additional handling and the loss of clay incurred. Furthermore, we believe that the rotting of the pugged mix (described later) accomplishes the same results to better advantage. That there is some action taking place in the clay while stored in the covered bins, however, is evidenced by an increase in volume which at times has caused damage to the retaining walls of the bins. We have at present a stock of bond clay weathering in the open, and when this is put into service an opportunity will be afforded for making comparisons with the clay as usually handled.

Jan 1, 1918

By S. J. Laskowski

In potash ore flotation, carried out in NaCI-KCI saturated brine (about 6-7 molar KC1 and NaCI solution), long-chain primary amines are used as flotation collectors. Long-chain primary amines are poorly soluble in water and their solubility further decreases at high electrolyte concentration. For ionic surfactants there is a characteristic temperature (Krafft Point, KP) below which their solubility is too low for micellization. In a previous study, the effect of pH on the KP of dodecylammonium chloride in water was determined [Q. Dai and J.S. Laskowski, Langmuir, 7, 1361-1365 (1991)]. This paper deals with the effect of electrolyte concentration on the KP of dodecylammonium chloride. The derived theoretical relationship between the KP and the critical micelle concentration (CMC) for dodecylamine can be used to calculate the KP for electrolyte concentrations of up to 0.5 mol/L (about 8% NaCI-KCI brine). A simple experimental procedure for the determination of the KP for higher electrolyte concentrations is also described. The results show that the Krafft point of dodecylamine increases from about 17-18°C (in distilled water) to about 80°C at 16% brine saturation.

Jan 1, 2006

By H. C. Counsell, R. Webber

"The mill at Denison Mines' Elliot Lake Operations started up in 1957, with a capacity of 6,000 tons per day. Over the years various plant modifications expanded th.is to 7 500 tons per day. In 1976, management decided to carry out a major expansion of the milling capacity, and during 1980 the grinding capacity was increased to 15,000 tons per day.The original comminution system consisted of a conventional crushing plant feeding minus S/s-inch material to rod and pebble mills, in closed circuit with mechanical classifiers. The key decision that had to be made was whether the expansion should include conventional or autogenous grinding flow sheets.During 1977, extensive pilot plant studies at Lakefield Research showed that two-stage semi-autogenous grinding, with some steel media in the primary mill and with pebble media in the second stage, produced stable grinding conditions. This type of circuit was selected by Denison for the expansion, and went into production in 1980. Primary millThe flow sheet of the semi-autogenous grinding circuit is shown in Figure 1. The primary grinding mill is a Koppers 28 foot by 10 foot mill, driven by a 4,000 h.p. synchronous motor. It is fed with a minus 8-inch run-of-mine ore, and 4-inch grinding balls are added as supplementary grinding media. There are two Koppers 15.5 foot by 22 foot secondary pebble mills, driven with 1500 h.p. synchronous motors. Each grinding stage operates in closed circuit with cyclones, and the primary cyclone overflow is the feed to the secondary mills. The circulating load in the primary stage can be controlled by adjusting the cyclone apexes."

Jan 1, 1984

By John Wulff, David A. Stevenson

The rates of solution of copper, nickel, and three copper-nickel alloys in liquid lead were studied at 527° and 727°C under dynamic conditions. The relative velocity at the solid-liquid interface was varied from 8.65 to 125 cm per sec. Analysis of the data for the pure metals indicates that the solution rate is transport corztrolled at lower velocities, but a transition to mixed control occurs at higher velocities, higher temperatures, and higher percent saturation. 1 HE present work concerns the effect of relative velocity at a solid-liquid interface on the dissolution kinetics of pure solid materials in liquid metals. This general problem is of interest in electrode kinetics, in electrolytic processes, and more recently in the field of liquid metal corrosion. Considerable work has been done on dissolution kinetics in aqueous environments because of the obvious importance to electrolytic processes.1,2 In purely metallic systems investigation has been more limited. There are, of course, extensive data from loop tests which are used in an engineering evaluation of liq- uid metal corrosion. Here, however, the dissolution process is complicated by several interdependent processes. In general, one of two factors is considered to be rate limiting in the dissolution process: the transport of material from the solid-liquid interface into the bulk liquid, or an activation process directly at the interface. A number of studies of the dissolution rates of solid metals in liquid metals under static conditions have been published.3- Since the attendant hydrodynamic conditions imposed by thermal gradient convection and concentration gradient convection are difficult to analyze, it is doubtful if a clear interpretation of the experimental findings of these papers in terms of the mechanism involved in the dissolution process can be made. A clearer picture of the dissolution process is provided by

Jan 1, 1962

By Van Dyne Howbert, Fred E. Gray

The Presidio mine of The American Metal Co. of Texas is situated 45 miles south of Marfa, Texas, a town on the Southern Pacific R. R., and lies approximately 20 miles north of the Mexican border. Communication with the railroad is maintained over a state highway. All supplies for mine and mill are hauled in by truck from Marfa and precipitates and concentrates are freighted to the railroad by the same means and shipped to Selby and El Paso, respectively. The mill is at the town of Shafter, 1 mile from the mine and at approximately 350 ft. lower elevation. The ore is conveyed to it by an aerial tramway which requires practically no power on account of the difference in elevation. The mill is a countercurrent decantation cyanide and gravity concentrating plant of 200 tons daily capacity. Coarse and fine crushing are done at the mill. Two neal-by shallow wells in the bed of Cibolo Creek furnish an abundance of water. Power is generated for the mill and town by a Diesel engine plant. Ore Treated The ore is oxidized and siliceous. The greater part contains approximately 16 oz. of silver per ton, but mixed with this are many particles of ore of much higher grade, assaying in many instances 100 oz. or more. In the 16-oz. material there is an almost constant amount of silver which is difficult to recover by cyanidation. The high-grade material contains a much greater proportion of silver which is easily amenable to cyanide treatment. Thus, as the grade of ore becomes higher, the percentage of recovery of silver increases rapidly, under equal milling conditions. The silver minerals contained in the ore are principally argentite and cerargyrite, the former predominating The lead minerals, all of which are argentiferous, are chiefly cerussite and galena, with occasionally a little anglesite. The gold is free; most of the ore contains merely a trace

Jan 1, 1935

By R. Hammen, J. Lifton

"The purification of rare earth elements (REE) at mining operations is traditionally accomplished by multiple chemical phase-transfer processes including precipitation, solvent extraction, and ion exchange. We describe REE purification results with Solid Phase Extraction (SPE) columns that were developed to improve the speed and economy of REE purification. The SPE columns have the properties of enhanced equilibration kinetics, so that residence time of a few seconds will accomplish selective chemical transfers from solution phase to solid phase. We describe the separation of REE’s as a class from leach solutions and the further purification of REE salts. INTRODUCTIONHistorically the rare earths were first chemically separated from each other and purified by fractional crystallization. This slow, cumbersome, and laborious process was not suited to commercial production. Solvent Exchange (SX) technology was applied to the problem of separating and purifying the rare earth elements when it was determined in the late 1960s that there might be commercial uses for them individually. At the same time the very slow and expensive technique of ion exchange was applied either for ultra-purification of the desired “light” rare earths, or for the recovery and purification of the tiny amounts of higher atomic numbered, “heavy,” rare earths. Solvent Exchange (SX) has today become the standard for separating the rare earths across their atomic number spectrum. Whereas the separation of the light rare earths commercially may take 30 “cascades’ in a modern Chinese plant, the total separation of all of them from each other may take up to 80 such “cascades,” and is only even “practical’ with preconcentrated feed stocks, such as those produced in China by heap-leaching techniques performed in situ on very low grade radio-nuclide free ionic absorption clays.. Ultra-pure materials of all types are produced today by ion-exchange processing of ionic species already separated by solvent exchange.Hard rock deposits and refractory residues of tin mining and residues from uranium mining have been identified outside of China with very high ratios of dysprosium and other heavy rare earths to the majority lighter ones. Several issues face those who wish to process these materials so as to separate the heavy rare earths commercially:"

Jan 1, 2012

W. D. BRENNAN,* Cheyenne, Wyo. (written discussion?).-My experience has been that, where possible, it is preferable for each foreman to employ his own men, rather than to have them handled through an employment bureau. The fact that each foreman comes into direct contact with the man before he is put to work gives him an opportunity to choose, at certain times, the kind and class of men that he wishes. Where possible, after a man is employed he is sent to a central office where a complete history of each individual is kept on a card. If a man has previously been in the employ of the Company, his history can be obtained from this card, and if for any reason he is not wanted he can be refused work at the time he reports to, the central office. In connection with Workman's Compensation costs, we in Wyoming feel that although the rates of compensation paid may in some cases be rather low our law works satisfactorily. As the State and the number of employees are small, compensation cases are handled through Judges of the various District Courts. In the State Treasurer's Department, a few clerks were added to look after the handling and issuing of warrants and compiling of statistics. Most of the large employers have an arrangement by which a joint claim and assent, specifying the amount of compensation due, is made tip by the employer and signed by the workman. This is forwarded, to the Judges so that personal investigation of individual cases is not necessary. In this manner, compensation is promptly received and' overhead expenses of operation of the law is kept at the minimum. The law requires each employer to pay into the State Treasury, for the benefit of the industrial fund, a sum equal to 1.5 per cent. of the money earned by each employee, engaged during each calendar month of such employment. Each employer contributes monthly, unless the sum heretofore contributed, after deducting all payments made on account of injuries to his employees, shall equal a full 1.5 per cent. of his annual pay roll, computed by multiplying his current monthly pay roll by twelve, and shall likewise be not less than ($5000). It has been found during the years that this law has been in effect, that a number of the larger employers have been exempt from payment for from five to eight months per year, so that in some cases, even in coal nines, industrial insurance has cost less than 0.75 per cent.

Jan 3, 1919

By Wilbert G. McBride

Two Also 3 1/2-ton motor trucks were used by Young Bros. while operating at the Mammoth Collins mine at Shultz, Ariz. One was equipped with an oil tank holding 1,075 gal. and was used for the transportation of "tops." The other was fitted with a stake body and used to carry machinery, wood, rails, pipe and all classes of miscellaneous supplies. The bodies were made of oak with maple flooring and were attached to the frame of the chassis by U-bolts, to avoid drilling the main members of the frame. Most of the hauling was done from Tucson, a distance of 47 1/2 miles. During the first 3 months, part of the road was in very bad condition and the tire cost was excessive. After this part was repaired the road was in fair condition, but never good. There were no excessive grades or bad sand, but wagon ruts, too narrow for the truck wheels and of a different gage, caused heavy tire loss; while chuck holes, sharp curves and stones, both imbedded and loose, were objectionable features. During wet weather the trucks could not get sufficient traction to climb some of the hills and were likely to stick in the mud in certain places, so that no attempt was made to run them unless they were on the' road when the rain started. This lost time amounted to about 5 per cent. of the total, but, whenever possible, it was utilized in making minor repairs. The price of gasoline was from 17 to 21 c. per gal. Rubber tires were used throughout. Drivers were said $4.50 to $5 per shift, and a return trip to Tucson was counted as two shifts even when made in one day. Drivers were provided with a room in Tucson and were paid for a11 time lost due to causes beyond their control. Trucks were loaded one way only. Speedometers were placed on both trucks but the excessive vibration soon caused them to fail. For this reason, and because no account was taken of the distance covered in picking up a miscellaneous load or in other minor ways, the mileage given is under the actual distance traveled. Some of the weights had to be estimated, but care was taken to have the number of ton-miles low rather than high, to avoid underestimating the costs. The cost of hauling from Tucson to the mines was $12 per ton

Jan 1, 1917

By M. Tanaka, T. Oishi, K. Koyama, S. Alam

"A novel hydrometallurgical process for copper recovery from waste printed circuit boards has been proposed, which consists of selective leaching of copper metal as Cu(I) by Cu(II) in ammoniacal alkaline solution, purification, and electrowinning. The distinct feature of this process is much lower power consumption during the electrowinning than that during the conventional one from acidic copper sulfate media, because the deposition of copper metal from Cu(I) at the cathode and the oxidation of Cu(I) to Cu(II) at the anode take place in the present method. In the laboratory-scale tests using waste printed circuit boards, high purity copper which contains 1.1ppm of impurities was obtained in a chloride system. INTRODUCTIONThe recovery of various metals from waste materials is an important issue in order to conserve the environment and metal resources. Copper is used for cables, electric appliances, etc., and is currently a very important material for industries and will continue to be in the future. Among the wastes that contain copper, one of the most complex recycling materials is electronic scrap like printed circuit boards (PCB), which generally contain 10-30% copper as well as plastics, fibreglass and other metals. There are two ways to recover copper, either pyrometallurgical or hydrometallurgical. Hydrometallurgical treatments generally have more flexibilities in the process scale and control. The separation processes tend to become complicated because of the poor selectivity during the leaching in the sulphuric and hydrochloric acid media. On the other hand, the use of ammonia as a complexing agent, gave a higher selectivity for copper leaching (Nigo, Majima, Hirato, Awakura, & Iwai, 1993; Majima, Nigo, Hirato, Awakura, & Iwai, 1993; Zhou, Shinme, & Anezaki, 1995; Koyama, Tanaka, & Lee, 2006) .In this study, recycling process containing leaching, copper deposition at a cathode and oxidation of copper(I) at an anode from ammoniacal alkaline solutions containing Cu(I) or Cu(II) is introduced. Furthermore, copper recovery from a waste printed circuit boards and impurity behaviour was examined in ammonia-ammonium sulfate (sulfate system) and ammonia-ammonium chloride (chloride system) systems."

Jan 1, 2012

By A. T. Iannacchione

Recent fatal injuries occurring during pillar retreat coal mining call for better identification of the hazards and recognition of risks. Ground control hazards associated with room-and-pillar retreat mining tend to intensify with depth, requiring extra precautions in the form of additional controls. In many locations these activities are further complicated by local complexities often associated with unique or novel circumstances, such as over-mining or under-mining. One technique recently studied by NIOSH, Major Hazard Risk Assessment (MHRA), may help mine operators to mitigate the risks associated with pillar recovery operations. The approach was trialed at two underground coal mines in southern West Virginia that are currently practicing pillar recovery. The first step in the MHRA process involved reviewing the segments or parts of the pillar extraction mining system and identifying associated hazards and threats to the operation. Some of the high risk hazards at these two sites included: ? Rock fall during pillar extraction covers equipment and injury ? occurs during recovery operations, Rock fall above the roof bolt horizon, potentially injuring ? workers and/or requiring recovery action,Rib fall under deep cover, potentially injuring workers, and; ? Strata instabilities associated with subsidence of the ? interburden due to simultaneously mining two seams in close vertical proximity to one another. The mines? risk assessment teams then considered each threat individually in order to systematically identify potential unwanted events. The top unwanted events were examined individually using structured risk analysis tools. The output from the process includes a list of priority existing controls for monitoring and auditing, and a second list of potential new controls. These controls consisted of: ? Examples of best practices, ? Enhanced communication, ? Implementing standard operating procedures (SOP), ? Protocols for emergency response actions, ? Effective layouts, ? Efficient monitoring, and; ? Successful audits. This paper documents the process as it unfolded at the two mines, and analyzes the strengths and weaknesses of the MHRA technique as it applies to pillar recovery operations. This work was part of a larger NIOSH research project. NIOSH was responsible for: 1) facilitating the risk assessment process known as Major Hazard Risk Assessment (MHRA), 2) documenting the risk assessment process, and 3) providing mine management with a written draft summary of the risk assessment process. The results of the risk assessment represent the thoughts and opinions of the individual risk assessment teams and should in no way be construed as an endorsement of the risk assessment output by NIOSH or the University of Pittsburgh.

Jan 1, 2009

By Henry Hebley

IN recent years, especially in the last decade, great interest has been shown and many advances have been made in the preparation and clean-ing of coal. In the major coal-producing countries, the percentage of tonnage cleaned has been increasing ever since 1924. As separating media, both liquids and gases are used, these generally being water and air respectively. Some systems, instead of obtaining separation by the velocity of the medium in which the coal is cleaned, depend on the density of water mixed with sand, or coal dust and shale, or with chemicals such as calcium chloride. In studying the results of coal-cleaning equipment, it will be observed that the efficiency of the cleaning apparatus lowers as the size of the coal decreases. This inability to obtain any marked improvement in the fines has been partly responsible for the installation of dedusting equipment, to remove the fine sizes from the raw coal prior to cleaning. When considering the possibility of installing dedusting equipment, many factors must be taken into account. Very often, but not always, both the ash and sulfur content of the coal increase as the coal size decreases. When bituminous coal, containing soft powdery fusain, is subjected to a wet washing process, the porous fines of the fusain remain in the small coal or in the wash water and delay the dewatering of the washed coal and increase the difficulty of clarification in settling tanks or thickeners. When a dedusted coal is cleaned by a wet process, the sludge recovered from the settling tank often runs so high in ash that the sludge can be discarded without further treatment and still the loss is small. Where treatment of the sludge is desired, however, froth flotation may be used to advantage if the size of the feed is below 1/10 in. Dust from the deduster can be treated in a similar manner. Fusain, which is usually found in the fine sizes, has often a much larger ash content than other coal.

Jan 1, 1933

An abstract of a paper from the Canadian Mining and Metallurgical Bulletin of August 1978 by E. Stamboliadis, J. McHardy, T. M. Salman, Department of Mining and Metallurgical Engineering, McGill University, Montreal, Quebec, states that: “The recovery of gold, silver and other heavy metal cyanides from mill solutions is technically possible using the ion exchange technique. However, the costs involved are greater than the costs of the zinc precipitation technique used for gold recovery. There are two main reasons for this: one, the fact that in gold ore cyanidation metal cyanides other than gold and silver occupy most of the resin capacity and hence reduce the adsorption efficiency of the resin for the Au and Ag cyanides; secondly, the high cost of the eluting agents.” Resin, therefore, cannot in most cases be efficiently used for recovery of precious metals. With copper in solution, only 1.47% of the resin capacity is used for gold and 0. 8970 for silver; the rest is occupied by unwanted metal cyanides, such as copper occupying 73.7%. On the other hand, although activated carbon will adsorb copper cyanides, they will be preferentially displaced by gold and silver cyanides. Ion exchange is by one or two orders of magnitude more expensive than the conventional method of zinc precipitation. The cost of removing impurities, such as copper from the resin, is greater than the cost of recovering gold. With an oxide ore amenable to cyanidation and with a low content of other metals, such as may occur with some of the Soviet gold deposits, ion exchange would be an efficient as Carbon-In-Pulp.

Jan 1, 1981

By M. G. Corson

UNTIL recently, the characteristics of cast copper have seldom been studied and such study as has been made has been inadequate. The prevailing idea seems to be that since pure copper is infrequently used in the cast state, there is no reason for trying to establish precise data regard-ing the properties of the metal in that condition. On the other hand, wrought copper either in the shape of sheets or wire rarely fails in service, except as a result of the action of reducing gases at the annealing tempera-tures, so there is little incentive to study the properties of the cast shapes -cakes, bars, ingots and billets-for-the sake of tracing the causes of failures of the wrought metal to the cast stock. However, about a decade ago, wirebars of American origin often cracked badly while being hot-rolled in mills in Germany. This led to four investigations: by E. Seidl and E. Schiebold,1 by P. Siebe,2 by O. Bauer and G. Sachs,3 and by G. Welter.4 These authors came to the conclusion that failures in hot-rolling were due essentially to the presence .of a wide zone of columnar crystals between the thin layer of fine crystals near the surfaces of the bars and the large but more nearly equiaxed grains of the central core. This macrostructural defect was explained in turn by the probable upward deviation in the temperature of the metal while it was being poured into the molds of the casting machine. More recently, cast shapes of copper were investigated by N. P. Allen5 and by C. Blazey.6 Allen examined vertically cast wirebars and hori-zontally cast wedge cakes of electrolytic tough-pitch copper, but only for the distribution of densities; i.e., porosity. Blazey was interested mainly in the mechanism of the variations in electrical conductivity clue to cold-work and anneal, but gave introductively an account of the composi-tion and average densities of the tough-pitch horizontally cast wirebars used. As far as the present writer can tell, no investigation was under-taken for the sake of presenting a picture of the distribution of mechanical characteristics in the wirebars of copper, tough-pitch or otherwise.

Jan 1, 1938

By G. Cockburn

Nkomati Nickel JV exploits the ores of the Uitkomst Complex near Machadodorp in the Waterval Boven district in South Africa?s Mpumalanga Province. Due to factors such as the remote location, stellar growth in production, opencast mining methods, and ore characteristics, a number of innovative processing options were selected. Nkomati has undertaken numerous initiatives over the last few years to improve plant running times, metallurgical performance, and operational profitability. Great emphasis has been placed on effectiveness of management control systems. A number of initiatives such as short interval control and time?in-state metrics have been implemented. A focus on improvement on availability and asset utilization of key items of equipment has been particularly effective. While the ores are remarkably similar to the Merensky and UG2 reefs, the relatively high base metal sulphide content and mineralogical characteristics make metallurgical treatment somewhat different to the ores of the Bushveld Complex. The low head grades and flotation kinetics distinguish Nkomati from other base metal operations. Numerous milling and flotation optimization initiatives have resulted in dramatic improvements in throughput, recoveries, and concentrate grades. This paper discusses the metallurgical, operational, and management challenges and the outcomes obtained.

Jan 1, 2014

Discussion of the paper of LAWRENCE ADDICKS, presented at the Salt Lake meet-ing, August, 1914, and printed in Bulletin No. 91, July, 1914, pp. 1671 to 1674. JAMES H. PAYNE, Baltimore, Md. (communication to the Secretary*). -Mr. Addicks's account of the behavior of the nodulizer at Chrome on blast-furnace flue dust leads me to outline what has been done so far upon the application .of the same apparatus to fine sulphide concentrates. As Mr. Addicks states in his paper, the conditions prevailing in the test kiln at Yorktown in the demonstration I made for him upon Chrome material were strongly oxidizing; so much so that the sulphur was roasted down to less than 1 per cent. It is therefore to be expected that the first test made upon sulphide ore a short time after this, in which Ducktown green ore fines were used, did not show very promising results. Although the Ducktown ore contained only 13.52 per cent. of sulphur its gangue was quite fusible, and the intense heat generated by the oxidation of. the extremely fine sulphide particles slagged the mass, bringing about a condition approximating that of a Brückner roaster in distress. During the spring and summer of 1912, the writer kept working upon different schemes to cut clown this slagging. Two successful methods were finally worked 'out embodying opposite principles, one promoting oxidation of the finer particles and the other retarding. their oxidation. In the early part of the present year the writer succeeded in interesting the Braden Copper Co. in nodulizing, with the result that an extensive series of tests were made at Yorktown upon oil-floated concentrates covering quite a wide range in analysis. These tests led to trials in the Chrome rotary on the part of the U. S. Metals Refining Co. of Minerals Separation concentrates which they had in stock, and later to test runs to check up the Yorktown runs on Braden concentrates. The large-scale tests checked up with the Yorktown tests in every way (except fuel consumption, which was, however, correctly predicted), and led to the Braden company's decision to adopt the process at Braden. The fuel consumption upon oil-floated concentrates is not over 6 gal. per ton and in many cases is less. The action in the furnace is independent of the sulphur content, and there is actually less tendency to form nose rings in the Chrome rotary than upon flue dust. It is believed that the improvements entering into the design of the Braden kilns will cut down the nose-ring trouble to where it will no longer be a drawback to the process.

Jan 11, 1914

By John B. Sabey, Vernon R. Degner

The industrial coal processing trend toward greater fines production, due to mining techniques and customer requirements for reduced ash and sulphur content in the fine coal product, is the basis for ongoing efforts to improve the performance of the flotation machine. WCO has successfully adapted the principles of the beds Column to a full size machine prototype. Recent plant evaluation tests confirmed the initial promise of this concept by achieving a considerably lower ash product while maintaining the combustible recovery level of conventional flotation machinery. The paper will include a review of the fundamental principles of the beds Column concept, together with a summary of the WEMCO R&D Program from concept evaluation and laboratory testing, through pilot, to full size machine tests. The performance of the WEMCO/Leeds flotation column will be compared to conventional flotation concepts using a flotation "kinetics" approach, which quantifies the floatable and non-floatable specie recovery performance using the Klimpel model as the basis for analysis. The variety of distinct features unique to this concept include a series of barrier rods which are used to advantage, often together with countercurrent reflux water flow, to achieve a significant improvement in product grade. The hydrodynamics associated with the mechanics of "gangue stripping" from the rising "air bubble-floatable specie" matrix will be briefly reviewed.

Jan 1, 1988

Located 19 km (12 miles) north of the Carlin gold mine and 43 km (27 miles) north of Carlin, Nevada, is Carlin Gold Mining Company's Bootstrap mine. This deposit is similar in character to the Carlin ore, except it occurs in the upper plate formation of the Roberts thrust fault rather than in the lower plate rocks. The deposit was discovered over 50 years ago and operated on a small scale in the early 1960s using vat cyanide leaching. Newmont Mining Corporation acquired the property in 1967, and from an open pit mined 650,000 mt (717,000 st) of ore which was trucked to the Carlin conventional cyanide plant. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME, 1975. ) With the same grade of ore, the gold recovery by conventional methods is somewhat lower than that of the Carlin ore, even with a finer grind. Leaching tests were first conducted by the United States Bureau of Mines in 1968; and in 1971, ‘72, Carlin conducted extensive cyanidation investigations by bottle testing, simulated heap leaching in tall columns, and small scale pad heap leaches. Gold extractions ranged from 55% to 95% with low reagent consumptions. In 1974, a 545 mt (600 st) heap leach test was made at Carlin on pit-run low-grade material assaying 1.37 g per mt (0.04 oz per st). In a six week period the extraction of gold was 47.57'0. Carlin uses atomic absorption analyses on cyanide leach liquors to determine potential leach ore grade rather than fire assaying because it gives a more practical assessment of leach recovery than does fire assaying, which accounts for the total gold content. Before open-pit mining was started, plans were made to heap leach, the ore at Bootstrap which was too low grade for transport to Carlin. The topsoil outside of the designed pit area overlaid ancient lakebed clays which were found suitable for compaction to form an impervious base on which the dumps to be leached could be placed. By using the cost of truck haulage and milling at the Carlin plant and the price of gold 'at that time, a cutoff grade was established which resulted in the gold content of the first 150,000 mt (165,000 st) of the heap leaching ore being 1.51 g per mt (0.044 oz per st). An additional 500,000 mt (552,000 st) of lower grade material, assaying 0.96 g per mt (0.028 oz per st), is available for leaching as economics permit.

Jan 1, 1981

By Isadora Dias Perez, Denise C. Romano Espinosa, Flávia P. Cianga Silvas, Jorge A. Soares Tenório, Mónica M. Jiménez Correa

In view of the rising generation of wastewater containing metals, sustainable and clean techniques to recover them are being researched. This study investigated the adsorption of aluminum, cobalt, chromium, copper, iron (III), magnesium, manganese, nickel and zinc presents in a synthetic ef fluent by chelating ion exchange resin Dowex XUS43605. The resin was tested in columns experiments and was analyzed the flow rate (0.45, 0.66 and 1.11 BV/h), at pH 1.5 and 20 °C. After loading the column, the elution experiments occurred with 1 M of H2SO4 and 4 M of NH4O. The concentration of the metals in solution was obtained by X-ray fluorescence spectrometry energy dispersive (EDX). The results showed that the breakthrough experimental curves demonstrated that the main metals (aluminum, cobalt, manganese and zinc) were immediately broken through just after the initiation of the feeding, while the breakthrough of copper began later. Therefore, the study showed that the chelating resin could be used to treat contaminated effluents with metals.

Mar 1, 2017