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By C. Arthur Lefler

The Smoky Valley Mine is jointly owned by Louisiana Land and Exploration Company, Felmont Oil Corporation and Case Pomeroy Company. Copper Range Company, a subsidiary of L.L. & E. is the operator of the mine. The mine is located about 60 miles (100 km) north of Tonopah, Nevada on Highway 376 (formerly 8A) in the Big Smoky Valley on the western flank of the Toquima Mountain Range In the shadow of Mount Jefferson. The mine is at an elevation of about 6100 ft. (1860m) and is situated on the southwestern flank of Round Mountain, the hill for which the town of Round Mountain is named. The ore is a rhyolitic iglmbrite with coarse gold occurring along fractures and with fine microscopic gold disseminated in the rock mass. Minor silver is associated nth gold. Annual mining is planned at 5.8 million tons (5.3 million tonnes) per year with 2.2 million tons (2 million tonnes) as ore (7,000 tons or 6,400 tonnes per day).

Jan 1, 1981

By Zhigao Xu, Ming Wu, Jun Qu, Rong Zhang, Wenrui Nie, Ruan Chi, Zhenyue Zhang, Zhengyan He

The leaching processes and mechanisms of weathered crust elution-deposited rare earth (RE) ores with ammonium sulfate were investigated for their high efficiency and low consumption exploitation. The leaching process is a typical reversible heterogeneous ion exchange reaction in liquid–solid systems. The raw ore and leached tailings were analyzed by SEM and FTIR. Results indicated that no dissolution or structural changes of the ores occurred during the leaching process. The migrations of RE3+, Al3+, and NHþ 4 in the leaching process were studied under various liquid–solid ratios. With increasing liquid–solid ratio, the content of NHþ 4 adsorbed on clay mineral by physical effect decreased, NHþ 4 adsorbed by ion-exchange increased, and the content of exchanged RE and Al increased until reaching equilibrium. The adsorption capacity of the RE ores for NHþ 4 remained about 0.61–0.63 mg/g. It was difficult to desorb NHþ 4 from RE ores with washing water alone. The leaching mechanism of Al was the same as RE, but the leaching of Al slightly lagged behind RE. At the optimal liquid–solid ratio of 0.4:1, the leaching efficiencies of RE and Al were 95.51% and 90.99%, respectively, and 56.65% of NHþ 4 remained in the leachate. After impurity removal and RE precipitation, the leachate can be reused by replenishing fresh leaching agent. The aim of this study is to enhance the leaching process of RE, restrain the leaching of impurities, and decrease the consumption of leaching agents.

By N Guresin, R Dunne, S Blanco

This paper presents the results of research work conducted to assess if aqueous solutions of sulfur dioxide (SO2) can be used to treat roaster calcine leach residues such that recovery of gold in this material can be achieved with subsequent application of the thiocyanate (SCN-) leaching system. The project is mainly aimed at assessing an alternative use for sulfur dioxide gas generated during flotation concentrate roasting. The calcine leach residue sample used in the study contained 2.56 g/t gold. The reason for this high level of residual gold is thought to be due to gold surface passivation by thin iron oxide films, presence of gold in solid solution with refractory iron oxides and encapsulation of gold in the structure of these iron oxides.The study was conducted in three stages. The first stage was the characterisation of a sample from a roaster calcine leach residue. The second stage in the project was a set of leaching tests to assess the feasibility of using aqueous solutions of sulfur dioxide to reductively dissolve refractory iron oxide mineral phases contained in the calcine. The final objective was to assess the efficiency of thiocyanate, a gold complexing ligand stable under acidic and oxidative conditions, to extract gold liberated during sulfur dioxide leaching.The study showed that the reductive sulfur dioxide treatment of the calcine leach residues followed by oxidative thiocyanate leaching to extract gold was not a particularly efficient process as only around 20 per cent gold could be extracted. Leaching the calcine residues by any of these systems separately also extracted the same amount of gold (12 - 14 per cent) implying that irrespective of the solution potential, the acidity helps dissolution of gold encapsulating minerals, mostly hematite, but at a very slow rate. The findings on the topic in the literature and during this study show the requirement for further technical investigations and economical analyses on two cases. The first case is to roast gold concentrates to produce magnetite rather than hematite followed by sulfur dioxide and thiocyanate leaching of magnetite containing calcines. The second case is to apply a magnetic concentration process to conventional roast-cyanidation calcine residues to produce a gold-rich magnetite concentrate that can be further treated with sulfur dioxide followed by thiocyanate leaching.For both cases, the best evaluation of the sulfur dioxide leach system in extracting residual gold from calcine leach residues will require a detailed economic analysis to find out the balance between the value of the gold recovered and the loss of revenue from acid production. There are a number of roasters that do not produce acid. In these cases, a portion of the unwanted sulfur dioxide gas can be used to recover gold in magnetite containing calcines as an economic incentive.CITATION:Guresin, N, Blanco, S and Dunne, R, 2013. Leaching roaster calcine leach residues with aqueous solutions of sulfur dioxide and thiocyanate, in Proceedings World Gold 2013 , pp 129-136 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013

By Luísa Maria Abrantes

Environmental control is becoming more and more stringent. As conventional processes for zinc production are pollutant routes, alternative methods have become an important subject of research. In this regard, the leaching processes have aroused a considerable interest, namely the electrochemically induced approach. The behaviour of zinc concentrate ore in the leaching processesses is influenced by several parameters. Chloride concentration and leaching solution pH, time and temperature of reaction were studied in order to evaluate their influence on the zinc recovery. The efficiency of the leaching process for zinc extraction is performed by factorial design. The experiments were carried out by 23 and 22 factorial designs of the factors. Results show that all factors, except for the time of reaction have a significant influence on the optimisation process, within each experimental region.

Jan 1, 2001

By W. A. McKinney

The U.S. Bureau of Mines, Kennecott Copper Corp., and New Mexico Institute of Mining and Technology are cooperating on a research project to improve extraction of copper from previously leached waste dumps that are in part compacted and impervious to leach solutions. The field work is being done at the "J" leach dump, Chino Mines Division, Kennecott Copper Corp., Santa Rita, N. Mex. The laboratory percolation leaching tests were made at the U.S. Bureau of Mines Salt Lake City Metallurgy Research Center on drill core samples from the "J" dump to determine how much copper might be leached if permeability could be restored by explosive heaving or other means. The core samples from nine drill holes on "J" dump contained from 0.06 to 0.14 percent copper. Copper extractions ranging from 15 to 3 percent were obtained in 75 days by percolation leaching of minus-1-inch material with dilute sulfuric acid. The tests results indicate that about 2'7 percent of the copper remaining in the "J" dump could be extracted under ideal leaching conditions.

Jan 1, 1970

By Renfeng Song, Shaoyan Wang, Zhiqiang Zhang, Ke Guo

Leaching titaniferous magnetite concentrate with alkali solution of high concentration under high temperature and high pressure was utilized to improve the grade of iron in iron concentrate and the grade of TiO2 in titanium tailings. The titaniferous magnetite concentrate in use contained 12.67% TiO2 and 54.01% Fe. The thermodynamics of the possible reactions and the kinetics of leaching process were analyzed. It was found that decomposing FeTiO3 with NaOH aqueous solution could be carried out spontaneously and the reaction rate was mainly controlled by internal diffusion. The effects of water usage, alkali concentration, reaction time, and temperature on the leaching procedure were inspected, and the products were characterized by X-ray diffraction, scanning electron microscope, and energy dispersive spectroscopy, respectively. After NaOH leaching and magnetic separation, the concentrate, with Fe purity of 65.98% and Fe recovery of 82.46%, and the tailings, with TiO2 purity of 32.09% and TiO2 recovery of 80.79%, were obtained, respectively.

Jun 16, 2021

By Corale L. Brierley

An unidentified, high-temperature microbe, which oxidizes reduced sulfur and iron in an acid medium between 45° and 75°C, is studied to determine its ability to leach copper sulfide minerals and molybdenite. Copper is leached from low-grade ore as well as a chalcopyrite concentrate, and preliminary data indicate a tolerance of 1.0 gpl copper by the microbe. A molybdenite concentrate is also biogenically leached, and the microbe is able to continue growth when the soluble molybdenum concentration is 750 ppm. Molybdenum can not be biogenically extracted from waste material using the present technique. Direct observation using SEM illustrates colonial growth and developmental characteristics of the microbe on molybdenite fines. The organism responds to hexavalent molybdenum by reducing this metal to the pentavalent state in aerobic and nitrogen gas atmospheres.

Jan 1, 1974

A fundamental component of grade control at Tarmoola Mine has been the monitoring of assaying through quality assurance programs. The programs have concentrated on issues related to accuracy and precision of leachwell and fire assay techniques, sample preparation and grade variability within a sample. An outcome of monitoring has been identification of shortcomings in the application of assaying techniques to Tarmoola ore types that contain coarse particulate gold. Leachwell has shown satisfactory precision since introduction as the primary assay technique in October 1998 but leachwell accuracy and recovery through comparison with fire assay techniques has identified bias. A bias toward higher fire assays in grades above 2 g/t invoked an investigation into the cause of the discrepancy between leachwell and fire assay. Testwork evaluated Leachwell assay recovery by varying leach time duration, which showed that by increasing the leach time improved recovery significantly. Monitoring of leachwell against fire assay had shown that a leachwell (200 g) with a one hour leach has under-called grade when compared with fire assay by up to 18 per cent, on average, for Tarmoola ore types. Alternatively, optimisation of the leachwell assaying technique using a 200 g pulp and a two hour leach has returned, on average, higher assays than fire assay by up to eight per cent. Testwork on screen fire assays, duplicate assays and fire assay to extinction has demonstrated assay variability associated with coarse, particulate gold. The variability and distribution of gold in Tarmoola ore types has emphasised the importance of high quality sample preparation for leachwell assaying to ensure liberation of gold to leaching solutions. Assay quality assurance protocols are vital to assay data integrity and enable confidence in delineation of ore-waste boundaries and provide early recognition of problematic assaying periods and isolation of grade control practices, which depart from performance standards critical to the operation.

Jan 1, 2000

By G. L. Oldright, Henry de Rycker, S. F. Ravitz

The upper limit of richness of concentrates that can be smelted by means of the blast furnace without added diluents is fixed by the operation of sintering. A sinter feed with normal gangue constituents, containing more than about 50 per cent of lead, tends to fuse; as a result, air does not come sufficiently into contact with the sulphides at roasting temperatures, and sulphur elimination is poor. Since most lead concentrates now contain 55 to 80 per cent lead, it is necessary to add diluents to lower the grade of such sinter feed to about 50 per cent or less. This not only entails the expense of providing barren diluents when adequate diluents containing precious metals are not available but also appreciably lowers the capacity of the blast furnace to produce lead and increases slag losses. Attempts to desulphurize rich lead concentrates by methods other than sintering have not proved successful. Flash roasting, for example, converts much of the lead sulphide to lead sulphate rather than to oxide. However, this suggested the possibility that deliberate flash roasting of the concentrates to sulphate, followed by sinter roasting of the resulting product, might prove promising, particularly inasmuch as the elimination of sulphur from lead sulphate requires heat alone and not contact with atmospheric oxygen. The sintering of lead sulphate products has additional interest because important amounts of such products are now treated by lead smelters. Considerable quantities of lead sulphate are present in rich lead flue dusts and other by-products; lead-chamber sulphuric acid plants and other

Jan 1, 1944

By Reed W. Hyde

Some information has been published on the sulphur dioxide concentration of gases from D wight-Lloyd machines sintering lead ores but most of this relates to the customary practice in which the charge is relatively high in lead and is sintered in two passes Over the machine. It is thought, therefore, that the following observations on sintering in one pass a low-lead, low-sulphur charge may be of interest. The data were obtained in sintering lead sulphide notation concentrates mixed with minus 20.mesh oxide ore, the mixture carrying 16,8 per cent Pb and 5.3 per cent S. No returns were used because of the limited amount of sulphur available as fuel, but the mixture pelletized readily. to make a satisfactorily permeable bed on the sintering machine. As there was a possibility that the SO z might be utilized, a series of runs was made to determine the concentration ofSO2 in the gases at different points along the wind box, and the effect on this concentration of changes in the wind-box vacuum. Several runs each were made with vacuums (as measured in inches of water at the wind box) of (1 2 to 1 in., ('1 4 to 3% in. and (3) 6 to 5% in. The averages of the determinations from each series are platted On the graph (Fig, 1). The vertical scale shows the percentage by volume of SOz (as determined by the Reich method), and the horizontal scale shows the time e~pressed as percentage of the total sintering time. Sintering was considered to be complete when no SOz could be detected in the gases. In all cases the depth of bed was 4 1/2 in. and the weight of dry charge per square foot of hearth area was 45.2 lb. Other data are given in the table. The volume of gases corresponding to the several points along the wind box at which SO2 determinations were made could not be measured, but has been calculated from the SO2 Content and sulphur elimination on the assumption that the volume is uniform throughout the sintering period. This is not strictly true, as with a given vacuum the volume usually will drop slightly toward the middle of the period and increase toward the end. However, this made is a reasonable approximation. several interesting features are brought out by these data. In general, the published data on SO 2 concentration in sintering under ordinary conditions show a high SO2 concentration in the early stages followed by a rapid decrease in concentration and with relatively little SOz in the gases from the last half or third of the operation. In other words, the greater part of the sulphur elimination is obtained in the first third or half of the period. In these runs, the 6-in. vacuum, which approximates the usual operating condition, gives a curve that follows this customary pattern. Maximum SO2 concentration is obtained 234 min. after

Jan 1, 1944

By R. A. Perry

A speiss is an artificial arsenide, sometimes an antimonide, formed in lead smelting, smelting of oxide copper ores, and in some lead-refining operations. The production of speiss is closely allied with the amount of arsenic and metallic iron in the blast-furnace charge, while dross is produced mostly from the copper and sulphur, with varying amounts of antimony, arsenic and nickel. Most speiss is essentially an iron arsenide containing some sulphur, copper and nickel, if these elements are present in the furnace. The recent introduction of the soda treatment of lead drosses has produced a speiss high in copper. Speiss is not miscible to any extent with molten copper, lead, matte or slag. Miscibility increases with temperature. This property of the speiss is clarified in the latter part of this paper. Most speisses are produced in conjunction with matte and their sulphur content points to the solubility of matte in speiss. . Some plants have insufficient speiss to warrant separate treatment. When arsenic is high in the blast-furllace charge, there may be sufficient speiss to tap separately from the first settler. Arsenic From the foregoing, it is obvious that a study of the treatment of speiss would not ,be complete without consideration of arsenic. Inasmuch as perhaps go per cent of the blast-furnace feed may be sinter, a definite knowledge of the properties of arsenical ores and compounds that go in to the roaster and sinter feeds is of value. Lcad smelters obtain their arsenic from ores and concentrates, as well as from flue dusts shipped from other plants, particularly copper smelters. The percentage of arsenic in ores is usually low, although occasional lots of high arsenical products are smelted. These crude ores may be sulphides, as arsenopyrite; or oxides like scoiodite, which carry gold or lead minerals. The small amount of arsenic in the original ores received is greatly increased by a circulating load of arsenic in secondary products at are resmelted for their lead or precious metals. More important than the arsenic present directly in the concentrates treated is that in fumes and dusts that may originate in the blast furnace or sinter plant of the lead smelter, or may come from a copper smelter as lead-bearing converter fume or roaster dust. Thus the lead smelter not only must treat its own arsenic but is compelled to handie that from some copper and perhaps other lead plants. Where copper and lead plants exchange products, each plant may be forced to re-treat part of the arsenic it has shipped ,outside its own confines. Speiss from the lead blast furnace contains chiefly iron and arsenic with some lead, copper and sulphur and often appreciable amounts of gold and silver. If its copper content is sufficient this speiss may be sent to the copper smelter where usually it is treated in the converter and the copper is separated as blister. Much of the arsenic and most of the lead are recovered as converter fumes

Jan 1, 1944

By Felix Edgar Wormser

JUST as the ancients used the products of their crude mining endeavors to fashion tools with which to make digging easier, so today mining enterprises are dependent upon the very metals they mine for the efficient conduct of mining and metallurgical operations. In fact, the outlet afforded by the production of minerals in the United States is no mean market for the nonferrous metals. I believe that mine managers frequently lose sight of the opportunity they have to select from competitive products those which will help the market for their own metal output. For example, some years ago a large lead mine used storage-battery locomotives for its haulage, common practice, of course, throughout the United States. These particular locomotives, however, were powered not with the ordinary lead storage battery but by a competitive battery product made from an imported metal. After studying this situation, the manager saw how he could help himself by purchasing a product made from his own mineral output, and in his next order turned tip the use of the lead storage battery.

Jan 1, 1935

By T. W. Bingay

The production of both lead and zinc in Canada has steadily increased during recent years until today Canada ranks among the world's leading producers of both of these metals. The high prices which they have commanded have led to an energetic search for new deposits and also to an investigation of abandoned properties which, under the present conditions of better prices and more efficient methods of treatment, offer chances of commercial success. It is confidently expected that during the next few years success in both these lines of activity will further increase Canada's lead and zinc output, giving her a still higher ranking among the world's producers. In 1926, approximately 140,000 tons of metallic lead was produced from ores mined in Canada, and of this, 125,000 tons was refined in Canada at Tadanac, B.C. During the same year the production of metallic zinc from Canadian ores amounted to 68,000 tons, of which 62,000 tons was refined at Tadanac. History Of Lead and Zinc in Canada The production of lead in Canada is much older than that of zinc. The chief source has always been the interior of British Columbia. As earl y as the 'eighteen twenties', Indians and Hudson's Bay Company's trappers had discovered an outcrop of lead ore on the site of the present Blue Bell mine on the east shore of Kootenay lake, but it was not until1865 that active prospecting for lode deposits began. In the early nineties the activity of the silver-lead-zinc region of the Slocan reached its culmination and a steady production has been maintained down to the present time. Up to the end of 1924 the Slocan area produced over 38 million ounces of silver, nearly 300 million pounds of lead, and almost 122 million pounds of zinc.

Jan 1, 1927

By FREDERICK J. ALCOCK

THE high prices which lead and zinc have commanded during recent months have given a great impetus to search for workable deposits of these metals, and there has accordingly risen a demand for information about deposits which have been worked in the past and which might possibly be opened up anew under these more favorable conditions. Production of zinc and lead in Canada has been on the increase during the last few years. In 1924, 175,- 485,499 lb. of lead with a computed value of $14,221,- 345 were mined. Of this, 96 per cent. was produced in British Columbia, 3 per cent. in Ontario, while Quebec and the Mayo district of the Yukon accounted for the remaining one per cent. During the same year the production of zinc amounted to 98,909,077 lb.. valued at $6,274,791. Of this 96,000,069 lb. were produced in British Columbia and the remaining 2,909,008 in Quebec. The large producer of zinc and also of lead is the Sullivan Mine at Kimberley, B. C. In comparison with the Sullivan, the deposits which are known to occur in eastern Canada are relatively small. Two, however, the Kingdoll lead mine at Galetta, Ontario, and the Tetreault zinc property at Notre Dame des Anges, Quebec, are steady producers. The object of this paper is to describe briefly these and other deposits of eastern Canada, to show their relation to geological events and to point out what are the chances of any of those which are idle at present developing into active producers or what are the possibilities of new deposits being discovered.

Jan 1, 1926

By C. A. Summers

The U.S. Bureau of Mines applied process mineralogy to investigate a lead and zinc deposit located near Story Creek, Colville Mining District, Alaska. Mineralization is hosted by the Kayak Shale in veins, veinlets, and breccia. Sphalerite and galena, with supergene anglesite, occur in symmetrical quartz veins and between breccia clasts. The study sample analyzed 2.11 percent Pb, 5.37 percent Zn, 0.014 percent Cu, and 1.2 oz/ton Ag. Analyses of crushed products showed that 33 percent of the sample weight, and more than 50 percent of the Pb and 65 percent of the Zn, are in the minus 850 µ m fraction. Liberated grains of sphalerite and galena were found in crushed products finer than 300 µ m. Beneficiation tests indicate the material is amenable to processing by traditional gravity and flotation techniques.

Jan 1, 1993

By C. A. Summers

The U.S. Bureau of Mines applied process mineralogy to investigate a lead and zinc deposit located near Story Creek, Colville Mining District, Alaska. Mineralization is hosted by the Kayak Shale in veins, veinlets, and breccia. Sphalerite and galena, with supergene anglesite, occur in symmetrical quartz veins and between breccia clasts. The study sample analyzed 2.11 percent Pb, 5.37 percent Zn, 0.014 percent Cu, and 1.2 oz/ton Ag. Analyses of crushed products showed that 33 percent of the sample weight, and more than 50 percent of the Pb and 65 percent of the Zn, are in the minus 850 um fraction. Liberated grains of sphalerite and galena were found in crushed products finer than 300 um. Beneficiation tests indicate the material is amenable to processing by traditional gravity and flotation techniques.

Jan 1, 1991

By S. F. Radtke

The properties and characteristics of lead and zinc have made these metals useful to man in many ways since the days of antiquity. Despite their long history of use, however, these metals have proved to be extremely resilient and responsive to modern needs. New alloys and compounds continue to be developed, adding to the growth of these metals. Consumption data is given for 1969 and compared with one and two decades ago. Trends in major markets also are cited and explained. Uses for lead and zinc which have come into prominence within the past few years are described and explained, with emphasis on the role played by the industry's cooperative research program. Current efforts in the research phase are discussed, with details given on major efforts for both metals, including new alloys, process developments, applications research, finishes, coating, paints and pigments, batteries and fundamental research. A forecast is given of promising new areas and a projection made on a longer term of new applications that may emerge from current research.

Jan 1, 1970

By A. Patrick Ryan

To many people, the word cartel implies illegality. However, although it is illegal in the US, cartels are an effective and legal way to conduct business elsewhere in the world. But be-fore we continue, what is a cartel? The American Heritage Dictionary defines a cartel as "a combination of independent business organizations formed to regulate production, pricing, and marketing of goods by the members." However, that definition does not fit many of the "cartels" we are examining today. The International Tin Council is an organization of both consumer and producer nations--no individual companies are represented. The buffer stock is used by the ITC to provide protection for both producers and consumers, and has only been in effect since 1955. (Earlier tin producers groups date back to 1921 when the Federated Malaysia States and the Netherlands East Indies formed the first intergovernmental tin agreement, known as the Bandoeng Pool, which only lasted three years.) The International Bauxite Association only dates back to 1974 when seven nations--Australia, the Dominican Republic, Ghana, Guinea, Guyana, Haiti and Jamaica--agreed to three main objectives: 1) to promote the orderly and rational development of the bauxite industry; 2) to secure a fair and reasonable return on the raw material for the benefit of the nations' citizens; and 3) to safeguard the interests of member countries in relation to the international bauxite industry. 1 Again this organization is composed of nations, not separate firms. This generally follows the pattern that the oil producing countries (OPEC) chose since the nations' mineral product-ion is controlled and often owned by the state. However, the group I am going to discuss fits the more classic mode of a cartel. The informal lead-zinc producers' cartel can be dated as far back as to Roman times when Romans are said to have imposed a quota system on British lead production to safeguard their Spanish lead mines. Also, the lead-zinc producers group is composed only of individual companies, not do not include any lead-zinc consumers. Finally, the activities of the group are not open to public scrutiny, and details are only now available due to a leak in the Australian government. At this point, it is necessary to split the two metals since while many of the members of the cartels produce both lead and zinc they do have different histories and marketing arrangements.

Jan 1, 1978

By Michael G. Mayer

Since its invention in 1859 various designs of lead batteries have evolved to meet commercial needs for traction applications and standby power. In recent years major developments have taken place in lead alloys and the chemistry and mechanical design of industrial lead batteries. These are reviewed in relation to their commercial application together with the new type of maintenance free gas re-combination lead-acid batteries which are opening so many new markets for lead.

Jan 1, 1990

By R. E. Wagner

MISSOURI's famous lead area, in what is known as Southeast Missouri, is locally termed the "Lead Belt." These deposits are in the Bonne Terre dolomite of late Cambrian age which has a thickness of nearly 400 ft. Underlying this is the La Motte sandstone which varies in thickness from 0 to 400 ft. The La Motte beds are deposited directly on the ancient erosion surface of Pre-Cambrian porphyries and granites of probably Algonkian age. The buried igneous peaks extend up into the Bonne Terre in several areas within the Lead Belt. Overlying the Bonne Terre are shales and siliceous dolomites of late Cambrian and early Ordovician age. Within the Lead Belt these have a thickness of from 0 to 500 ft. The district is bounded in general on the southwest by the Sims Mountain fault, on the northwest and north by the Big River fault, and on the

Jan 1, 1947