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By Mahesh C. Jha

Arsenical gold ores occur in many parts of the world, including numerous mines in Canada and the western United States. Free gold is often finely disseminated in the grains of sulfide minerals, arsenopyrite, pyrite, orpiment and realgar. Generally a sulfide concentrate is produced by flotation. Arsenical ores or concentrates are not amenable to direct cyanidation and therefore are considered refractory. A wide variety of roasting equipment and conditions have been used by the mills to remove arsenic and sulfur from these concentrates. The resulting calcine is easy to cyanide and gold is recovered from the cyanide solution by zinc dust cementation. Room-temperature potential-pH diagrams and high-temperature phasestability diagrams are presented to discuss the chemistry of the gold cyanidation, cementation, and roasting steps in the presence of arsenic. The environmental aspects of arsenic discharge are also included.

Jan 1, 1984

By Kenzo Murao, Sei Nakao

HISTORY OF ASHIO SMELTER The Ashio copper mine, which formed the main part of Furukawa mining industry, is located inland of Kanto area, and is known as the oldest copper mine of Japan. The mining started even since 1610, and Mr. Furukawa began management in 1877. Afterwards, within a short period, the mine became to have the highest productivity in the Far East. The total quantity of mined copper reached approximately more than 800 thousand tons. The mother rock of the ore deposit in Ashio was rhyolite, and was formed of sissurefulling and irregular clod that was called Kajika deposit. The lode was nzrrow of 6 to 7cm. The mineral was formed mainly of chalcopyrite, including also pyrite, pyrrhotite, and arsenopyrite that were not favorable. After the founding of the copper mine, the Ashio smelter started the operation of Mabuki process as the beginning in 1611, and subsequently adopted the processes of the blast furnace, reverberatory furnace, and flash smelting furnace of the present use. As indicated above, the copper ore of Ashio mine included arsenic. For the purpose to remove it, the operation of the arsenic trioxide production using MacDougall's multiple hearth roaster was started in 1918, and it maintained the status of having the most qualified technique within the country. However, as the completeness in the disposition of anti-pollution and safety had been requested in those years, we studied and developed a unique equipment in 1970. It excludes the heavy metals especially arsenic and copper from the waste water of sulfuric acid process of Flash Smelter's exhaust gas.

Jan 1, 1976

By James E. Hoffmann, John A. Bonucci, Peter D. Parker

A hydrometallurgical process has been developed for the recovery of high-purity silver from sulfated residues. The process eliminates the high temperature operations ordinarily associated with silver recovery and refining. The crux of the process is a metathetical leach which maintains the sulfate ion concentration at a level which both allows silver nitrate to be freely soluble and simultaneously depresses the solubility of lead nitrate. The subsequent hydrolytic purification of the silver nitrate solution, the recovery of high-purity silver by electrowinning in a diaphragm cell, and the final product purity are discussed.

Jan 1, 1981

By H. S. Hanna, C. Rampacek

The problems and limitations encountered in the Physical beneficiation of the Appalachian Red Iron Ores are discussed. Two major problems are identified - the high phosphorus content and the fine dissemination or encapsulation of mineral impurities in the iron-bearing minerals. A process that combined extreme fine grinding techniques with magnetizing roasting or flotation-separation methods is suggested. By these methods, concentrates assaying 60-61 % Fe, exceeding the widely accepted upper limit of 56% Fe set for these ores, were produced with 75-90% iron recovery. Moreover, it was shown that the phosphorus content of these concentrates could be reduced from 0.25% to 0.03% phosphorus by acid leaching. A brief discussion on production costs based on current and future developments in iron ore processing is given.

Jan 1, 1983

By T. Tanaka, H. Kiuchi

The combination of the following two exothermic reactions was studied with the purpose of recovering H2 from H2S formed in acid leaching of sulfide ores, direct reduction of metal sulfides with H2 or desulfurization of crude oils: In the kinetic study of the recovery of H2, the reaction rate was measured by the two methods of single-stage and circulation blowing of H2S through molten Pb in the temperature range of 550 to 750°C. The experimental results showed that the rate determining step is in the chemical reaction itself and also that this reaction is accelerated markedly by the addition of a trace of Ni, which may be responsible for the catalytic action of Ni dissolved in molten Pb. In the experiments where Pb metal was produced, PbS was oxidized in a stream of atmospheric N2 containing 0.5 to 20 vol-% 02 or in the same 02 partial pressures as in N2 though under reduced pressures in the temperature range of 400 to 800°C. It was found from these experiments that the formation of PbSO4, basic sulfates and PbO are suppressable to yield Pb metal alone above 750°C and for 02 concentrations of less than 1 vol-%. Vapor of PbS was thought to take an important part in the Pb formation mechanism, and the probability thereof was demonstrated thermodynamically.

Jan 1, 1978

By S. K. Bose, P. K. Sahoo, S. C. Sircar, P. K. Jena

High grade manganese ores are utilized in metallurgical, chemical, and battery industries, whereas the low grade ores are not accepted for these purposes and hence get accumulated at the minehead. In the present study an attempt was made to solubilize the manganese value as its sulfate by roasting the low grade ore with a suitable sulfatizing agent followed by water leaching. The different parameters were determined in the laboratory scale. Based on these, kilogram scale experiments were conducted where an appreciable amount of manganese goes into solution during leaching.

Jan 1, 1981

By E. B. Thornhill

(San Francisco Meeting, September, 1915) IN this paper on the recovery of mercury as sulphide, from the residues from the amalgamation and cyanide treatment of high-grade ores and concentrates, I will not discuss the many reactions, chemical and other- Fig. 1. wise, that take place in the general process, but confine myself more particularly to the methods of recovering the mercury. I submit charts of two amalgamation-barrel charges, Figs. 1 and 2, which are self-explanatory. The flow sheet of the high-grade and the mercury extraction plants is shown in Fig. 3.

Jan 8, 1915

By E. B. Thornhill

Discussion of the paper of E. B. THORNHILL, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 104, August, 1915, pp. 1653 to 1670. D. B. HUNTLEY, Oakland, Cal.-About 15 years ago it chanced to be my lot to cyanide some mill tailings, assaying about $5 per toll in gold and, a few cents in silver. It was in southern Idaho, a. desert region where costs are nearly as large as in central Nevada. We knew that we had a lot of tailings that had been amalgamated and contained about 1 lb. of quicksilver per ton. We knew that we would dissolve some of the quicksilver when we cyanided them with say 0.25 per cent. cyanide solution. We figured (optimistically) that the cyanide would dissolve all of the quicksilver, and we thought to capture all of that, 1 lb. per ton; and we figured out profits as partly gold, partly silver, and partly quicksilver from the recovery in working those tailings. Actually, however, we did this: We had a battery of three large, old-fashioned silver-mill retorts. We retorted our zinc-dust precipitate, recovering about ¼ lb. of quicksilver per ton treated. That was only a loss of 75 per cent. of what we originally had, but still it was profitable by virtue of the conditions. I mention it partly to call attention to the crude old way of doing it, and partly to call attention to how chemical methods for recovering quicksilver have developed in the last 15 years.

Jan 12, 1915

CHEMICAL PRECIPITATION U.S. 4,067,789 - Removal of soluble manganous ions from a zinc sulfate solution formed by leaching roasted zinc sulfide ore with spent electrolyte. An excess of ammonium persulfate or like oxidant is added to the leach solution to convert manganese to insoluble, filterable gamma-manganese dioxide, and the precipitate separated from the solution by filtration. L. Harris and A. K. Hanson, Jr., assigned to Newmont Exploration Ltd. Jan. 10, 1978. 11 pp. (204-119). Filed: 5-3-76 (682.442). (Drawing on page 67.)

Jan 1, 1979

US 4,152,143-Continuous precipitation of cement copper from an aqueous ore leach solution mixed with pieces of iron as the precipitant Copper depositing on the pieces of iron is loosened and removed therefrom by means of constant but regulated oscillating movement of the pieces of iron in a horizontal elongated chamber, and the loosened copper is separated and collected from the mixture E Kausel and R Nissen, assigned to Klockner-Humboldt-Deutz A G May 1, 1979 8 pp (75-100) Filed 1-12-78 (868,981) (Drawings below) [ ] US 4,156,711-Removal of metal impurities from an acidic zinc or cadmium solution prior to electrolysis. Impurities such as manganese, cobalt, nickel, and thallium are selectively precipitated by contacting the solution with persulfuric acid or an ammonium or alkali metal salt thereof while maintaining a pH value of from 1 5 to 4 5 This patent is a continuation-in-

Jan 1, 1980

By H. Reinhardt, P. Haglund, J. Rydberg, B. Lunden

As legislation to minimize pollution in effluent streams becomes stricter, it is important to treat used acid pickling baths before discharge. This is usually achieved by neutralization of the acids and precipitation of the dissolved metals in such a way that neither acids nor metals can be re-used. A process has been designed and tested on a pilot plant scale to recover both acids and metals from a stainless steel pickling bath leaving only Na2SOs in the effluent. In this process, H2SOr is added to the spent pickling liquor and the recovery of the pickling acids HN03 and HF is achieved by solvent extraction, using TBP in kerosene as the organic solvent. The extracted acids are stripped with water, giving an acid concentration high enough for direct use in the pickling bath. Molybdenum is also extracted by TBP and is recovered in a separate step. The other metals, Fe, Cr and Ni, are precipitated from the raffinate as hydroxides, dried and recycled to the metal process. In evaluating this process, comparative experiments with mixer-settler and pulsed column have been performed. The process described is being considered for use by a Swedish company pickling 70,000 tons per annum of stainless steel.

Jan 1, 1973

US 4,192,852-Precipitation of iron values as a jarosite from a sulfate solution obtained as a byproduct in the hydrometallurgical processing of zinc sulfide ore The solution is cooled, partially neutralized, and clarified. The clarified solution is heated to a temperature of 80 to 100°C in the presence of sodium, potassium, or ammonium ions to precipitate substantially all of the ferric iron as jarosite The precipitate is thickened, filtered, and washed The residual solution may be processed for zinc or other nonferrous metal values R V. Pammenter and C J Haigh, assigned to Electrolytic Zinc Co of Australasia Ltd. Mar. 11, 1980. 16 pp. (423-42) Filed: 5-8-78 (903,524). Priority- Australia, 5-9-77 (PD0030) (Drawings below.)

Jan 1, 1982

US 4,185,077 - In the recovery of uranium from a solution produced by leaching uranium ore, the leach liquor is contacted with a polymeric crosslinked vinyl benzyl chloride anion exchange resin and eluted with sulfuric acid to recover uranyl sulfate Such a resin is resistant to attrition and thermal degradation during elution. R.L. Albright, assigned to Rohm & Haas Co Jan. 22, 1980. 6 pp (423-7). Filed: 10-31-75 (627,829).

Jan 1, 1982

US 4,185,078-In the recovery of rhenium values from an aqueous sodium molybdate solution contained in the hydrometallurgical processing of roasted molybdenite, improved results are realized by using a described extractant consisting of a quaternary ammonium compound and benzene or other highly aromatic solvent which dissolves a quaternary ammonium-rhenium complex Essentially all rhenium is recovered from the solution and the extractant can be regenerated and recycled. This patent is a continuation-in-part of US Patent No. 4,049,771, dated Sept. 20, 1977. L R. Quatrini and M.B MacInnis, assigned to GTE Sylvania Inc. Jan 22, 1980. 6 pp (423-49). Flied. 7-19-78 (926,167).

Jan 1, 1982

By Minoru Yamada, Eikichi Mohri

INTRODUCTION Most of the copper concentrates treated by the flash smelting furnace at Kosaka smelter come from "black ore" that is produced in the local mines nearby. The copper concentrates contain various impurities such as lead and zinc as shown below. [ ] These impurities are mainly concentrated into the flue dusts in the copper smelting system. The dust production rates of flash smelting furnace and converter at Kosaka smelter are, therefore, fairly high compared with other copper smelters. However, by recovering these valuable metals in the dusts systematically, profit-increase of the by-products must be expected. This also leads to the solution of environmental pollution problems and enables us to establish the technical completion of complex sulfide ore treatment by flash smelting process. HISTORY OF DUST TWATMENT AT KOSAKA SMELTER 1) December in 1967 to January in' 1969 Since the flash smelting furnace started its operation in December 1967, all the dusts from the furnace had been recycled. Only conver-

Jan 1, 1976

By L. C. De Jong, J. F. Shirley, M. L. Campbell

The Toquepala concentrator of the Southern Peru Copper Corp., located approximately 720 miles southeast of Lima near the Chilean border, processes 36,500 tons of ore daily containing between 0.50% and 3.00% copper. The molybdenite content in the ore also varies greatly, ranging from essentially nothing to as high as 0.15%. The 1966 yearly averages were approximately 1.20% copper and 0.03 % molybdenite. The three basic problems that plague the opera- tion of any molybdenite recovery plant are all present at Toquepala in large measures. Both the quantity of plant feed and the quantity of molybdenite contained in the plant feed are continuously changing. Even more detrimental to an efficient operation is the fact that the copper mineralization in the feed frequently changes once in every 24 hours and occasionally once a shift. From the point of view of the molybdenite plant metallurgist, there are four distinctive chalcopyrite concentrates and, at least, an equal number of chalcocite concentrates treated in the molybdenite recovery circuit. The various types of concentrates are distinguished mainly by their mineral content and how effectively the copper minerals are depressed by the different reagents used in the circuit.

Jan 9, 1967

By John W. Lane, Richard A. Ronzio, Frederick N. Bender

Climax Molybdenum Co. operated a hydrometallurgical plant at Climax, Colo., from August 1966 to August 1968 to recover molybdenum from an oxidized ore. The feed, tailings from sulfide flotation, was first upgraded by cycloning. The concentration plant handled 6000 tpd of feed and produced 2000 tpd of slime concentrate. The concentrate was leached by a sulfuric-sulfurous acid solution at 65°C. The molybdenum was solubilized in the form of "molybdenum blue." Activated charcoal was added to the leach tailings in a resin-in-pulp type circuit. Molybdenum was adsorbed on the charcoal. The charcoal was screened from the pulp and washed. The washed charcoal was pumped to columns where it was ammoniated and air-blown. The oxidized molybdate then was washed from the charcoal. The eluate was evaporated until ammonium dimolybdate crystallized out. The ammonium dimolybdate was calcined to molybdenum trioxide.

Jan 1, 1973

By R. B. Sudderth, C. R. Merigold

Greater, emphasis is being placed on recovering nickel from laterite ores by ammonia leaching techniques. New methods are required for lower cost, non-polluting techniques to separate more efficiently the metals of interest. Liquid ion exchange techniques meet these requirements with the added advantage of producing salable electrowon nickel cathodes at the plantsite. A complete solvent extraction- electrowinning process is described, including preliminary engineering and cost data for a typical plant producing 50 million pounds of nickel per year.

Jan 1, 1973

By John A. Bonucci

AMAX Base Metals R&D, Inc. has developed technology for extracting and recovering platinum and palladium from spent automobile exhaust emission control catalyst. Our most promising route is a chloride-based hydrometallurgical flowsheet in which catalyst pellets or crushed monoliths are leached in HC1 with an oxidant at 95°C without agitation. Alumina co-dissolution can be limited by thermal pre-treatment of the catalyst. The leach solutions are treated by one of several precious metals recovery techniques to produce a high grade PGM concentrate. For the new generation of three-way catalytic converters which contain rhodium, an H2S04 preleach could selectively extract much of the rhodium while leaving the platinum and palladium behind. Capital and operating costs will be discussed.

Jan 1, 1984

By C. W. Clark

Two papers have been written previously concerning operations at the Montreal East plant of Canadian Copper Refiners Limited. The first one,1 written in 1932, described silver-refinery operations at the time for producing silver and gold bullion from tankhouse anode slimes; the second,2 published in 1933, consisted of a description of the entire plant. Since publication of these papers the refinery has become the world's larg-est producer of refined selenium. The first selenium was produced in August 1934. Tellurium production on a. commercial scale followed in October 1935. This paper will deal with selenium and tellurium produc-tion and also with present operations in the silver refinery. Blister cakes from Hudson' Bay Mining and Smelting Co., Flin Flon, Manitoba, and commercial anodes from Noranda Mines Limited, Noranda, Quebec, are refined at Montreal East. Both these bullions have unusually high gold, selenium and tellurium content. The average assay of raw slimes produced from domestic and Noranda anodes is about 40 per cent Cu, 3600 oz. per ton Ag, 800 oz. per ton Au, 22 per cent Se and 3.7 per cent Te.

Jan 1, 1938