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By W. R. Smith

Introduction No commercial deposits of silica sand are known to occur in British Columbia, so all requirements for foundry and other uses must be imported. During the war, nearly 5,000 tons per year were used, all originating in Illinois and delivered to local consumers for about $12.00 per ton. Present requirements are somewhat less, but the cost is still of the same order. The relatively high cost of silica sand in British Columbia is not only a serious item for present consumers, but might be a deterrent to the establishment of a glass industry, should this be contemplated at any time. The present investigation was undertaken in an attempt to provide a cheaper source of silica sand within the Province. In 1936, a large number of beach and bank sand deposits in the lower coastal region of British Columbia were examined by the British Columbia Department of Mines, and samples were analyzed chemically and mineralogically. The deposits over the whole area proved remarkably uniform in mineral composition, consisting essentially of quartz, feldspar, ferromagnesian minerals, rock fragments, and various accessory minerals, and differing only in the relative proportions of the various constituents. In no case did the quartz content of any sand exceed about 40 percent and the average was of the order of 20 to 30 percent.

Jan 1, 1947

By C. V. G. K. Murty

Sillimanite, chemically named Aluminum silicate, 3Al2O3SiO2 is one of the three forms of alumino silicate polymorphs and commonly encountered in aluminous metamorphic, plutonic, and volcanic rocks and their weathering products. Most of the sillimanite deposits are granular and are of high grade available in coastal areas as associated mineral along with Titanium bearing minerals namely ilmenite, rutile etc. Sillimanite is one of the widely accepted raw materials for the manufacture of high alumina refractories (55-60 % alumina) or bricks. Sillimanite based refractory bricks are used for the lining of Blast furnaces, Arc furnaces, Soaking pits, Reheating furnaces of iron and steel, rotary kilns for cement manufacturing and general kilns for lime production. Considering the vast potential of unexploited sillimanite reserves in India, M/s Trimex Sands Pvt. Ltd., (TSPL)has set up a Mineral Sand Processing plant in 2010 at Andhra Pradesh with a capacity to produce about 50,000 tons of sillimanite per annum besides the production of various heavy minerals such as ilmenite, rutile, zircon and garnet. Traditionally in India, sillimanite is upgraded by froth flotation technique using Oleic acid as a collector cum frother, sodium silicate as depressant and soda ash as pH regulator. At TSPL, Conventional mechanical float cells are installed to float sillimanite. In addition to the usual problems associated with conventional cells, poor flow ability of the sillimanite concentrate is experienced in dry circuit due to the presence of sodium oleate. Fine garnet in the sillimanite concentrate contributes to high iron content in the final product, making sillimanite separation a nightmare for operators (lower recoveries coupled with inferior grades). Various efforts were made to improve the performance of the circuit such as introduction of additional frother like MIBC, optimizing the operating parameters such as pulp density, collector dosage, depressant dosage, air flow rate, pH etc., use of collector aids, stage wise addition of reagents etc. The challenges faced to produce a sillimanite product with stringent quality norms and the efforts made to improve the grade and recovery are described in this paper. Keywords: sillimanite, froth flotation, conventional mechanical cell

Sep 1, 2012

By B. Pesic

The recovery of silver and gallium from flue dust with acidified thiourea solutions was examined. The studied parameters were temperature, particle size, thiourea concentration, acid concentration, and the presence of oxidants. The study of the extraction of these two metals was undertaken in sulfuric acid medium. Silver dissolution was dependent on all examined parameters, whereas gallium extraction was only dependent on the initial acid concentration. The experimental results suggest a new very efficient method for recovery of both silver and gallium. Almost complete recovery of silver and gallium can be accomplished within 90 minutes. A technological scheme for recovery of both metals is also proposed.

Jan 1, 1986

Cyanidation tests by bottle agitation were conducted on an oxidized lead-silver ore containing approximately 12% Pb and 500 g/t Ag. Over 60% of the silver was leached when the ore was ground to 100% passing 0.15mm, whereas 50-60% of the silver was leached when the ore was only ground to 100% passing 0.5mm. Slurry handling and dewatering at the fine grind size were very difficult due to the high clay content of the ore, however this problem was greatly reduced at the coar- ser grind size. The silver recovery at the fine grind size was very dependent on cyanide solution strength with stronger cyanide sol- utions (containing more than 0.1% NaCN) re- sulting in lower solution silver contents. This was thought to be due to the precipi- tation of leached silver by zinc, which was also leached from the ore by the higher cyanide levels. There was a critical zinc level in solution of approximately 1000ppm, above which the silver recovery fell markedly. Cyanide and lime consumptions for this ore were high (typically 5 kg/t and 40 kg/t respectively), hence the processing of this particular ore by cyanide leaching is un- economic at current silver prices.

Jan 1, 1987

By T. G. Carnahan

The U.S. Department of the Interior, Bureau of Mines, performed a laboratory-scale investigation of technique for recovering silver from pregnant solutions generated in the hydrometallurgical treatment of complex base-metal sulfide concentrates with oxidizing chloride media. This provides a means for recovery and simplified reduction of accessory silver to metal from ions. The method consists of adding NaI or KI to the solution to precipitate AgI, which is then contacted with Na2S to produce Ag2S and to regenerate the iodide salt solution. Silver recoveries range from 99 pct with 20 pct excess iodide precipitant, to-92 pct when the theoretical stoichiometric amount of precipitant is employed. The Ag2S product can be reduced to silver metal by contact with aluminum chips in NaOH solution followed by purification with conventional fire refining techniques.

Jan 1, 1980

By Charles A. Rhoades

The Bureau of Mines is conducting research on a dual leaching method that offers the potential for the economic recovery of the silver as well as the manganese contained in some of the domestic manganiferous silver deposits. The method is an initial leach of the manganese with aqueous SO2 followed by a neutralization rinse, and then by a second leach for the extraction of silver with cyanide solution. Refractory silver in these ores is either bound by the high-valence manganese, chemically resistent to cyanide leaching, or is rendered inaccessible to cyanide leach solution by the surrounding manganese mineralization. Less than 10 pct of the silver was extracted when -1 in +1/2 in (-2.5 cm + 1.3 cm) ore pieces were leached directly with cyanide. However, extractions of 25 to 82 pct for silver and 65 to 96 pct for manganese were attained in laboratory column leach tests using the dual method on the same size ore pieces from six domestic deposits.

Jan 1, 1984

By C. Flores

It is well known that manganiferous silver ores containing manganese intimately associated with silver are not amenable to conventional extraction methods. The segregation process has been applied primarily for recovering copper from oxidized and oxide-sulfide ores. A new alternative process combining segregation roasting followed by flotation to recover silver from refractory manganiferous silver ores, was studied using a 23 factorial experimental design. Laboratory investigations have been conducted on a manganiferous silver ore analyzing 0.02% copper, 0.09% lead, 0.33% zinc, 12.55% manganese and 187 g/ton silver to study the effects of additives salt, coke and copper oxide on silver recovery. Results indicate that 79.3% silver recovery can be attained in the absence of copper oxide and 88.4% when the copper oxide content is 1.5% at '750°C, 2% salt and 1% coke and 30 minutes of roasting time. It was concluded that doping with copper oxide improves silver recovery.

Jan 1, 1994

By C. S. Ek

As a continuation of previous work, further tests have been carried out to recover silver from the residues obtained in a zinc jarosite process. Selective flotation has been performed with different reagents, and ?cleaning steps have been introduced to upgrade the silver concentrate. Leaching of crude residues, as well as flotation tailings, has been carried out with cyanide and thiourea. Finally, the different beneficiation methods are compared, and the industrial aspects? of their application are reviewed.

Jan 1, 1990

By W. Kunda

Silver nitrate solution is a very common product in hydrometallurgical processes involving the treatment of various silver bearing materials. Silver is recovered by electrolysis. The alternative method for the recovery of silver from aqueous solutions is the precipitation by hydrogen at elevated temperature and pressure. A previous study by the authors revealed the feasibility of this technique for the silver ammine sulfate system. Further study showed different behavior for silver in the silver ammine nitrate system and a systematic study was initiated to establish the optimum conditions for the process. The study showed that the process can be controlled by (1) composition of the feed solution, (2) temperature and (3) hydrogen pressure. The recovered silver product was in the form of very fine powder.

Jan 1, 1989

By William Romanoff

IN AN ARTICLE on "Recovering Smelter Dust and Oxide," published in the Engineering and Mining Journal (Vol. 131, No. 2), the authors briefly described some dust-recovery equipment and its operation at a secondary-metals brass-smelting and refining plant. A description of our experience in overcoming certain operating difficulties may prove of interest or value to those who are confronted with some of the same problems or who should equip their furnaces with dust-collecting equipment. The dust loss from reverberatory smelting and refining of brass con¬sists largely of zinc, with smaller percentages of lead, tin, and copper in the order named. At the Kramer plant, the fume discharge from the two stacks amounted to 3 to 4 tons of solids per day, making a dense smoke. With dust equipment in operation, the discharge from the stacks is invisible and approximately 2500 lb. of zinc and other metals is recovered daily thereby from each furnace operating.

Jan 1, 1933

By B. Kruger

After the milling and flotation of the platinum bearing ore, the flotation concentrate is electrically smelted in 6-in-line electric arc furnaces and the resulting furnace matte treated in Peirce Smith converters to remove the iron. The iron-free (less than 1% Fe) converter matte is granulated and sent to the base metals and precious metals refineries for further processing. During the smelting and converting processes two distinct gas-strength gas streams are produced: ?The Peirce Smith converter off-gas contains 4%?9% SO2which is treated in a conventional sulphuric acid plant to produce 94% H2SO4. ?The electric furnace off-gas contains 0.9% SO2which was (until September 2002) vented through a 91- metre high stack to atmosphere. In October 2002 Impala Platinum commissioned a SULFACID® plant at its smelter in Rustenburg, which recovers the SO2from the electric furnace off-gasses into a weak sulphuric acid solution. The SULFACID® technology utilizes wet activated carbon to oxidize the SO2toSO3. The SO3reacts with the water on the catalyst and is washed off the catalyst bed as a sulphuric acid solution with up to 13 wt. H2SO4. The plant operates at 60ºC and at atmospheric pressure. The SULFACID® technology is widely used in the pigment industry. This is, however, the first application of the technology on electric furnace off-gas. Keywords: SO2, sulphuric acid, electric furnace, off-gas, SULFACID®

Jan 1, 2004

By David T. Hobbs

Large quantities of highly alkaline, radioactive, liquid wastes are stored in underground tanks throughout the Department of Energy complex. These wastes will be pretreated to separate radionuclides and the remaining decontaminated liquid wastes incorporated into solid waste forms for permanent disposal. Significant savings in disposal costs could result by recovering .the sodium as sodium hydroxide for reuse or disposal as nonradioactive waste. Recent testing demonstrated an electrochemical separation process for the recovery of sodium hydroxide from alkaline waste solutions; The separation process uses ion-selective membranes to separate sodium from the bulk of the waste constituents under an applied electrical field. The sodium is recovered as a 10-15 wt % solution of sodium hydroxide. Results of bench-sc3Ie and pilot-scale tests with simulated and radioactive waste solutions are presented.

Jan 1, 1999

By J. W. Spyker, D. H. Filson, W. H. W. Husband

"Large deposits of sodium sulphate exist in southern Saskatchewan in the brines or muddy crystal beds of alkali Jakes. Currently, more than 300,000 tons per year of sodium sulphate are produced from these lakes and sold as salt cake to the paper industry. Most of the production is obtained from the brines of a limited number of Jakes.The Saskatchewan Research Council was requested to develop a process for recovering sodium sulphate from the muddy crystal beds so that this type of deposit could be exploited in the future. A flowsheet was developed with the aid of phase diagram data and the process was studied on a pilot-plant scale using raw material from three different lakes. Satisfactory yields and product purities were obtained in all cases, and an economic analysis indicated that a 100,000-ton-per-year sodium sulphate plant could produce salt cake at a cost of $12.26 per ton for a rate of return upon investment of 16 per cent at current market prices."

Jan 1, 1966

By S. S. Ibrahim

The left behind material after the combusion of fossel fuel in some power stations in Cairo; weighing tens of tons every years, was investigated. The material called "Fly Ashes" containing 7.5% Ni, 10.0% V and 9.50% Cr; could be considered as a promising source of such valiable elements. As well as the small grain size (050 70 "m) of the material, and besides the complexity of their formation history, hydrometallurgical methods appeared to be the effective technique to deal with. Yet, preconcentration is possible by means of magnetic and gravity separation methods. The end product is assaying 16.60% Ni and 20.2S!; V with weight recovery of about 60.15%.

Jan 1, 1995

By J. M. Demarthe

The recovery of minor metals from zinc concentrates is dependent on the process used for their treatment. The behaviour of gennanium and indium in pyrometallurgical processes such as the Imperial Smelting furnace or the cupola furnace is of practical importance. The .Anaconda process for the treatment of zinc leach residues is not well suited to the recovery of valuable minor metals, and new technologies must be studied. In this regard, a solvent extraction process, using Kelex 100, developed by Minenet Recherche for the recovery of gennanium from dilute sulphate leach solutions is described. Also, new hydranetallurgical techniques in chloride media can be applied to the purification and concentration of minor metals depending on the stability of their chloro-complexes and the valency of associated inpurities. A synthesis of methods is proposed to address the problems related to the recovery of valuable minor metals in the zinc industry.

Jan 1, 1990

By Edward L. Fidler, Thomas E. Finch

Thin seam coal mining in the western US nominally equates to recovery of stray seams associated with thicker, major seams. The thin or stray seams encountered are generally recovered, but not in all cases. The alternative to extraction is, of course, spoiling. The majority of thin seams are encountered somewhere in the overburden above the major seam or in the parting between thick seams. Although some are true strays and appear anywhere in the geologic sequence, others might be classified as thin "rider" seams immediately above or below a thick seam and separated by some equally thin parting. For the purpose of this discussion, stray or thin seams are defined as those which are less than 1.2 m thick or represent less than 10% of the expected extractable coal. [Figure 1] is a graphic representation of typical thin, stray, or rider seams which actually appear in a western operation. Operationally, stray seams create irritating scheduling problems and require special handling techniques which do not always lead to efficient equipment utilization. Additionally, coal recovery from a thin seam is considerably less than the 95% usually recovered from thick seams. Higher recovery in thin seams leads to dilution problems, which depreciates Btu quality, and most western coals are not in a position to lose heating value. Thin western seams also undulate and pinch or swell more than thick seams, and this leads to more difficult recovery. Often they are associated with equally dark carbonaceous shales and visual selectivity is difficult at best during daylight hours. Resource Recovery Regulations Thin seam recovery in western mines is encouraged by both federal and state regulations and by the mining industry's concern for total resource recovery. [Table 1] lists the western coal states and their recovery regulations. The federal regulations are found in the Surface Mining Control and Reclamation Act of 1977. They state, "General performance standards shall . . . require the operation as a minimum to … conduct the surface coal mining operations so as to maximize the utilization and conservation of the solid fuel resource being recovered." Economic stimulation for thin seam recovery appears to be marginal or at the least controversial. Scheduling and operational problems often outweigh the nebulous profits created by engineering analysis. The only alternative to extracting thin seams occurring above major production seams is to strip directly through them and lose the resource to the spoil pile. Although this is not palatable to environmental emotions or resource promoters, it is the most simple expedient to operations people.

Jan 1, 1981

By S. Vellinga, G. H. R. Janssen, C. J. N. Buisman, C. F. M. Copini

Non-ferrous metallurgical plants are often confronted with the need to neutralize bleed streams or effluents containing sulfates. The treatment of these streams could produce large amounts of gypsum. With a biological sulfate-reduction process, the production of gypsum is avoided and the metals can be recovered as sulfides. In this paper, the biological treatment of an industrial stream is presented. The bleed solution from a roaster gas scrubber was treated by means of a two-step process. First, the acid was neutralized with calcine. The resulting zinc sulfate solution was converted to ZnS with bacteria using PAQUES-THIOPAQ® technology. The precipitated ZnS was de-watered to 70 wt% solids, and was returned to the roasters. The quality of the effluent stream complied with the environmental regulations.

Jan 1, 2000

By D. A. Rice, R. G. Swanton, A. May, O. C. Carter

As part of a cooperative effort between the US Bureau of Mines and the Florida Institute of Phosphate Research (FIPR), with input from the phosphate industry, the conversion of phosphogypsum to sulfur is being investigated. The proposed process incorporates a thermal reduction of phosphogypsum to calcium sulfide (CaS) and a hydrometallurgical treatment to convert calcium sulfide to ammonium bisulfide (NH4SH), which is then oxidized using a catalyst to produce elemental sulfur. The paper summarizes the results of continuing laboratory studies using an activated carbon catalyst which was loaded with sulfur through oxidation of ammonium bisulfide solution. The extraction of the sulfur that forms in situ on the catalyst is a key area in the development of the sulfur recovery process. Two approaches for removing the sulfur are being investigated: (1) thermal treatment to volatilize sulfur. and (2) leaching the sulfur using liquid anhydrous ammonia. The research has shown that it is technically feasible to recover over 90% of the sulfur from the carbon by either volatilization or ammonia leaching. Locked cycle testing has demonstrated that the carbon has the potential to be recycled after ammonia leaching.

Jan 1, 1987

By D. A. Rice

As part of a cooperative effort between the U.S. Bureau of Mines and the Florida Institute of Phosphate Research (FIPR), with input-from the phosphate industry, the conversion of phosphor-gypsum to sulfur is being investigated. The proposed process incorporates a thermal reduction of phosphogypsum to calcium sulfide (CaS) and a hydrometallurgical treatment to convert calcium sulfide to ammonium bisulfide (NH4SH), which is then oxidized using a catalyst to produce elemental sulfur. The paper summarizes the results of continuing laboratory studies using an activated carbon catalyst which was loaded with sulfur through oxidation of ammonium bisulfide solution. The extraction of the sulfur that forms in situ on the catalyst is a key area in the development of the sulfur recovery process. Two approaches for removing the sulfur are being investigated: 1) thermal treatment to volatilize sulfur, and 2) leaching the sulfur using liquid anhydrous ammonia. The research has shown that it is technically feasible to recover over 90% of the sulfur from the carbon by either volatilization or ammonia leaching. Locked-cycle testing has demonstrated that the carbon has the potential to be recycled after ammonia leaching.

Jan 1, 1988

Some design and feasibility studies have been made into possible methods of economic recovery of sulphur from smelter gases emitted at Western Mining Corporation's Nickel smelter at KalgoorIie in Western Australia. The smelter discharges some 330 000 tonne/yr of sulphur dioxide in a stream of approximately 6 per cent by volume of S02.Three possible recovery methods have been considered:1. The manufacture of sulphuric acid using the contact process.2. The concentration of sulphur dioxide to a pure form using dimethyl aniline as an absorbing medium.3. The reduction to elemental sulphur using natural gas as the reducing agent.Design studies were carried out by final year chemical engineering undergraduate students at Royal Melbourne Institute of Technology (RMIT) in 1981 under the supervision of the author and other teaching staff of the chemical engineering department. These studies indicate that sulphuric acid manufacture is the economically preferred recovery route.Markets for sulphuric acid include several potential mineral processing applications and the established fertiliser industry.Sulphuric acid is currently manufactured in a number of plants located at coastal regions of Western Australia using imported sulphur. The sulphuric acid is then used for superphosphate or titanium dioxide pigment manufacture at adjacent plants.The possibility is considered of manufacturing superphosphate at Kalgoorlie from locally produced acid obtained from the smelter gases and from phosphate rock railed to the site from the coast. The advantages of such a scheme include improved scale and technology of operation compared with some existing coastal operations, use of a waste material to replace imported sulphur, and avoidance of costs and hazards associated with railing large quantities of sulphuric acid.

Jan 1, 1983