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By Martin C. Kuhn, Lawrence G. Stevens, J. Mark Richardson

The majority of the world's known and anticipated reserves of nickel ores are lateritic deposits At present, there are several different process alternatives for recovering nickel and other desired metal values from laterite ores An overview is given of the selective reduction roast/ammoniacal leach process, one of the processes in commercial operation.

Jan 1, 1981

By Alexander E. McHugh

A study of some variables affecting the recovery of molybdenum from oxidized molybdenum minerals was made. The effects of variations in contact time, temperature, pulp density, particle size, and solvent concentration on the recovery of molybdenum using acid and alkaline solvents were investigated. Chemical analyses, screen analyses and microscopic examinations were made and the presence of important elements determined. A hydrometallurgical method was developed to recover the molybdenum from the oxidized molybdenum ore by simultaneous dissolution of the molybdenum with sulfuric acid, adsorption on activated charcoal and flotation of the charcoal. The molybdenum was extracted from the activated charcoal with sodium hydroxide, precipitated with calcium hydroxide as calcium molybdate and this regenerated sodium hydroxide for additional stripping. A solvent extraction method, using mines, was also studied to recover molybdenum from both alkaline and acid molybdenum solutions. Satisfactory results were obtained.

Jan 1, 1967

The NorthMet deposit of PolyMet Mining contains a large reserve of copper-nickel-cobaltprecious metal mineralization. Historically it has been difficult to recover separate copper and nickel concentrates of a quality broadly acceptable to custom smelters using conventional mineral processing techniques. Recently, improved flotation methods have made it possible to produce separate and saleable concentrates which will be produced during initial startup of the NorthMet deposit and construction of the PLATSOL TM hydrometallurgical plant. After the initial transition period, it is planned that production of a bulk flotation concentrate will be followed by PLATSOL TM hydrometallurgical treatment to produce copper, nickel/cobalt and precious metal products. The PLATSOL TM process uses chloride-assisted total pressure oxidation treatment of a bulk concentrate to digest the base and precious metal content of the bulk concentrate into an autoclave solution. Following solid-liquid separation, the solution is then treated sequentially for precious metal recovery by reductive precipitation, copper solvent extraction and electro winning and finally, nickel and cobalt recovery by precipitation. During the pilot plant testing of the PLA TSOUM process for NorthMct bulk concentrate treatment, two variations on nickel and cobalt recovery were tested including mixed hydroxide precipitation with magnesia and separate nickel and cobalt hydroxide precipitation following solvent extraction separation with Cyanex 272. The results of the pilot plant testing of the nickel and cobalt recovery circuits for NorthMet arc presented.

Jan 1, 2009

By Angove JE, Johnson GD

Sulphide concentrates are produced in a wide range of mineral processing operations. In the case of base metals they are further treated by smelting and/or hydrometallurgical techniques. In the case of refractory gold ores, sulphide concentrates containing gold are roasted or oxidised by various hydrometallurgical processes. This paper discusses the development of the Activox(R) Process which can provide an alternative method for treating sulphide concentrates. This process consists of fine or ultra-fine milling, followed by mild pressure oxidation to produce solutions of base metals or to liberate refractory gold. Base metal solutions (eg of nickel and cobalt) can be further processed by solvent extraction and electrowinning, while refractory gold can be recovered by cyanidation. The economics of the Activox(R) Process are often better than alternative oxidation techniques.

Jan 1, 1993

By Hao Rang

The article Is specially to Introduce the recovery process of Nb, Mn, P from slags and reutilization techniques of the remaining slags. Nb-Mn-Fe alloys which contain 14-16% Nb, 40-60% Mn can be obtained after using fire refining process of " Blast Furnace-Revolving Furnace-Electric Furnace -Electric Furnace, "meanwhile, phosphorous slags which the content of phosphorus pentoxide Is 29.21%. Further going through wet metallurgical processing and electrowinning technics to recover grading 99°o metallic manganese. Finally, phosphorous slags are grinded finely to be as phosporous fertilizers to be directly used for farmlands. Small particle size (0-8mm) rich slags are mixed with slight solid additions for balling directly, and taken Into blast furnace with blocky slags all together. The remaining tall slags are fully used as constructional materials, slags cement as well as a serial of wall materials, road materials, ground ornamental materials without burning operation and steam curing. The compressive strength range of those materials is up to 15-30 MPa. There is no more solid waste slags effluent In the full technical process.

Jan 1, 1994

By S Maritsa, E K. Chiwandika, S M. Masuka

The iron and steel industry is experiencing an increase in the use of iron and steel commodities within societies, decarbonisation pressures, depletion of high iron ore grades and high energy consumption. This necessitates scrap metal recycling, low-grade ore utilisation using the induction or electric arc furnace to achieve the global target of net zero CO2 emission. The available highquality scrap metal cannot sustain the steel industry but the world has unused hematite-goethite ores. It is the aim of this research that this resource be utilised using the induction furnace to sustain the steel industry in an environmentally friendly manner. Indurated pellets prepared from the hematite-goethite ore, 5 wt per cent limestone and 5 wt per cent coal addition were smelted in the induction furnace at 1350 ± 50°C with 25 mins holding time. The effects of basicity and carbon content were also investigated to optimise the smelting parameters. The results showed that the careful control of the basicity, and iron-to-carbon ratio maximised total metal recovery. The optimal basicity ratio was determined to be 1.3 and a carbon to iron ratio of 1.4 achieved a total metal recovery of 73 per cent. The metal recovery was closely related to the estimated slag melting temperature. The basicity ratio of 1.3 achieved the lowest slag melting temperature estimated to be around 1380°C based on the CaO-SiO2-Al2O3 phase diagram. The recovered metal had a comparable chemical composition and microstructure to that products from an ITmk3 furnace.

Jun 19, 2024

By S. C. Dhara

The recovery of Platinum Group Elements from various industrial scraps and spent materials in the form of high and low grade, such as, impure metals, matte, catalysts and other ceramic substrates, etc., has .been described by using high pressure reaction method. Nickel, Stainless Steel, Titanium and glass lined reaction vessels? are used in the reaction process. Primarily, the method involves either direct leaching of the substrates from the metal value or dissolving the precious metal from the substrate. Both acids and alkalies are used along with other reagents in the reaction process. The temperature and pressure of the leaching process are the main criteria for the reaction kinetics.

Jan 1, 1989

By G. A. Bezuidenhout, S. Bradshaw, W. Malan

"The purpose of the study was to investigate the feasibility of SiC reduction of low-grade concentrates from Lonmin’s Rowland and Easterns operations with respect to metal fall and PGM recovery. These concentrates are rich in SiO2 and MgO with low concentrations of chalcopyrite and Cr2O3. Pd is the most abundant of the PGMs. SiC reduction of samples was conducted at 1600? with 2.5-3.5 kg SiC per 100 kg concentrate.PGM recoveries for Easterns concentrate were better than for Rowland. More than 85% of the Ir and Pd and almost 60% of the Pt were recovered with 3.5 kg SiC per 100 kg concentrate. SEM of slag samples showed little entrainment of metallic prills compared to Rowland samples. This was attributed to the relatively higher melt viscosities of the Rowland concentrate. In order to decrease slag viscosity and to enhance PGM recovery, the FeO content of the Easterns concentrate was increased with the addition of 10 kg converter slag per 100 kg concentrate. Ir and Pd recoveries were increased to about 95%, while Pt recovery was around 70%. On the basis of these results an optimum feed ratio between Easterns and Rowland concentrates and converter slag is proposed. Carbothermic reduction of the optimum charge was also compared to SiC reduction. Carbothermic reduction yielded a marginally higher metal fall; however, the calculated gas emissions and energy requirements were higher than for SiC reduction."

Jan 1, 2015

By G. A. Bezuidenhout, S. Bradshaw, W. Malan

"Currently, the matted-based collection process is most widely used for platinum group metal (PGM) recovery. However, because PGM-containing concentrates are becoming more enriched with UG2 (Upper Group 2) low-grade concentrates, this is expected to be integrated or replaced with an alloy collection process.The reduction LG (low-grade) concentrates with silicon carbide (SiC) was investigated. The purpose of the study was to ascertain the feasibility of using SiC as reductant with regard to metal fall, PGM grade in the alloy, slag composition, and overall PGM recovery, and if possible, the integration of such a process in the current matte-based collection process. Rowland and Easterns LG concentrate samples from Lonmin were analysed by Xray diffraction (XRD), X-ray fluorescence (XRF), and inductively coupled plasma mass spectrometry (ICP-MS). It was found that SiO2 and MgO are the most abundant oxides and Pd is the most abundant of the PGMs. Sulphide-bearing minerals such as chalcopyrite were detected in low concentrations (below 1%), and Cr2O3 concentrations are between 2-4%. SiC reduction of Rowland and Easterns concentrate were conducted at 1600°C with reductant to concentrate ratios ranging from 2.5 to 3.5 kg SiC per 100 kg concentrate.The reduction experiments revealed that PGM recoveries from Rowland concentrate were poor. SEM images showed entrained metal prills in the slag phase. The slag viscosity at the end of the melt was calculated by using FactSage, and the results indicated that the viscosity was higher than 4 poise, which is above that of Easterns slag. PGM recoveries for Easterns concentrate were significantly better. More than 85% of the Ir and Pd and almost 60% of the Pt were recovered in a test with a reductant to concentrate ratio of 3.5 kg SiC per 100 kg Easterns concentrate. At this reductant to concentrate ratio, Fe recovery was also the highest, at 66%. SEM imaging of a slag sample revealed few or no metal prills entrained. The FeO content in the initial concentrate charge was then increased by adding Peirce-Smith converter slag in order to decrease the slag viscosity and further increase PGM recovery. A test was conducted with the addition of 10 kg converter slag per 100 kg Easterns concentrate. The reductant to concentrate ratio was kept at 3.5 kg SiC per 100 kg Easterns concentrate. The Ir and Pd recoveries were more than 95%, while Pt recovery was almost 70%. Fe recovery increased to 76%. On the basis of these results an optimum feed ratio between Easterns and Rowland concentrates and Peirce-Smith converter slag were calculated. FactSage predicted that the concentrate charge should consist of 70-80% Easterns concentrate with 20-30 kg converter slag per 100 kg LG concentrate. SiC reduction of this optimum LG concentrate charge is expected to recover more than 90% of all PGMs. Cr and Si concentrations in the alloy will be below 1% in total.The effectiveness of SiC as reductant was also compared to carbon reduction. Carbon reduction of an optimum concentrate charge at the same reductant to concentrate ratio resulted in a marginally higher metal fall than SiC reduction. However, gas emissions and energy requirements are higher for carbon reduction.The study has shown that base metals and PGMs can be recovered in an iron alloy by SiC reduction of LG concentrate with converter slag additions. Integration of such a process into the matte-based collection process could be considered as a future alternative for smelting UG2 LG concentrates."

Jan 1, 2014

By E. Schoeman

Ion exchange resins were investigated for the effective recovery of platinum, palladium, and gold from a pregnant cyanide heap leach solution. Screening results indicated that strong-base anion exchange resins proved to be most effective in the recovery of these metal cyanides. Gold was selectively recovered with the use of Minix, while effective recovery of gold, platinum, and palladium over the base metals, copper, and iron, was achieved with Amberlite PWA-5. Successful elution of these metal cyanides was possible from both Minix and Amberlite PWA-5, using a 2M potassium thiocyanate solution at ambient temperature. Acidic thiourea was unable to elute platinum from either Minix or Amberlite PWA-5. Effective elution of palladium and gold were, however, achieved with use of acidic thiourea. Using these results, two conceptual flow diagrams for the recovery of these metals from cyanide solutions are proposed.

Jan 1, 2012

By D. E. Deegan, G. Holcroft, T. P. Johnson

"The rising cost of extracting metals from existing traditional deposits, increasingly stringent environmental regulations and a greater awareness of corporate social responsibility issues are all factors in stimulating the examination of mine tailings and small-scale alluvial deposits as a potential source of new raw materials. Furthermore, despite tougher economic times the demand for commodities remains strong, in no small part as a result of the growth of the world’s emerging economies.This combination of social and economic factors looks set to continue in the long term which will encourage greater recovery of materials from previously ignored sources such as mine tailings and ores from alluvial deposits.An expansion in the recovery of useful materials from mining-related activities is likely to require the use of robust and innovative techniques that have a built-in degree of future-proofing; DC plasma technology is one such technology.As a result of its ability to separate the valuable metal and minerals from the waste material while destroying hazardous compounds, Tetronics’ DC arc technology offers an alternative approach to resource recovery from a wide range of mine ores. Harmful cyanide compounds and the vast quantities of water used in conventional hydrometallurgical processing are avoided when plasma processing is adopted. DC arc technology also brings additional benefits, including: the extremely effective destruction of organic materials, the conversion of the inorganic material into a re-useable aggregate and the high tolerance of the process to fine and dusty materials. This paper presents details of Tetronics’ DC arc technology as one of the key pyro-metallurgical solutions available for the recovery of valuable materials from traditional mine ores and tailings, as well as giving specific examples of how the technology has been successfully employed in the recovery of enriched slags, ferroalloys and other metals."

Jan 1, 2016

By David R. Mitchell

THE mineral pyrite (or marcasite) occurs in coal beds as balls, lenses, veinlets and bands. Several million tons are wasted annually on the refuse dumps from coal mining and coal-preparation activities. Sporadic attempts have been made in the United States to recover this pyrite for manufacture of sulphuric acid but only a few plants have succeeded in recovering it on a commercial scale. Several chemical companies in the East became concerned about supplies of pyrite early in 1943, but by late summer of that year the situation had eased. If for no other reason than national security and self-sufficiency during war-time, methods of recovering pyrite from coal-mine refuse should be worked out and made available to coal-mine operators who are potential producers of pyrite. Eastern sulphuric-acid plants are particularly likely to have their sources of raw material blocked by war conditions, so that special studies to determine where supplies of pyrite can be obtained for these plants from inland sources are desirable. Davis's1 summarization of conditions during World War I gives the following facts: During those years, the shortage of pyrite became acute. Shipments of hand-picked pyrite were made from a number of coal mines in Illinois, Indiana, western Kentucky, Ohio and Pennsylvania. A mechanical concentrating plant for separating pyrite from picking-table refuse was built at Danville, Ill., and operated for several years, and a small pyrite-concentrating plant treating washery refuse near Gillespie, Ill., operated for several months in 1918. The Consolidation Coal Co. of Fairmont, W. Va., constructed and operated a sulphuric-acid plant using pyrite recovered from coal-mine refuse. The U. S. Bureau of Mines during 1917-1919 made a comprehensive survey of pyrite-production possibilities from coal mines and estimated that there was a potential yearly production of 435,000 tons from Ohio and Pennsylvania and 928,000 tons from Kansas, Missouri, Iowa, Illinois, and Indiana mines. This estimate included only mines that could produce one per cent or more of their coal production as pyrite containing a minimum of 40 per cent sulphur. PYRITE PRODUCTION According to the Minerals Yearbook,2 approximately 660,000 long tons of pyrite was produced in the United States in 1941, with an average sulphur content of 42 per cent. The average value f.o.b. mines was $3.09 per long ton. Spanish pyrite, 48 per cent sulphur minimum, was quoted at 12 cents a unit c.i.f. U. S. ports, equal to $5.76 a ton for shipments of minimum sulphur content. Most of the pyrite produced in the United States is obtained as a by-product in the dressing of copper and zinc ores. Metal mines in the East making pyrite concentrate for the market or use in company-operated sulphuric-acid plants

Jan 1, 1944

By Roger Payne, John R. Goode, Feng Gao, James Brown

The Eco Ridge deposit is located in the Elliot Lake mining camp in Northern Ontario, Canada. Between 1955 and 1996, the area produced more than 138,000 t (300 million pounds) of U3O8 and very significant amounts of thorium and heavy rare earths. Based on the NI 43-101 Preliminary Economic Assessment by Roscoe Postle Associates (RPA) dated June 2012, and the economic review by RPA as announced in Pele Mountains’ Press release dated September 12, 2013, the Eco Ridge Mine Project would extract and process 46 Mt of material to produce about 19,000 t (42 million pounds) of U3O8 and about 65,000 t of rare earth oxide (REO) over a 14-year life. This paper discusses process optimization studies including beneficiation and hydrometallurgical testwork. The proposed modified flowsheet includes the flotation and processing of 1) a pyrite concentrate, 2) a dominantly light rare earth elements (LREE) concentrate, 3) a concentrate containing most of the balance of the uranium and heavy rare earths (HREE), and 4) a relatively inert tailings.

Jan 1, 2014

By John Blackshaw

NEW sources of supply of tungsten concentrates became necessary when the flow from the Far East was blocked in 1942. In Canada, one source has been noted in the workings of the Hollinger mine at Timmins, Ontario. For many years, a small amount of scheelite had been removed with the gold ore, and an appreciable tonnage of scheelite still remained in the mine. At the request of the Dominion Government, a plant was built to recover this scheelite in the form of a concentrate that could be used by the alloy-steel industry without further preparation. The plant began operating early in March, 1942, and a small tonnage of ore, carrying scheelite and gold values, has been treated daily since then. The possibility of recovering the scheelite had been considered on several previous occasions. Each time the proposal had been given up because the total profits from the enterprise were estimated to be insufficient to repay the capital cost of building the recovery plant. At the time the plant was built, additional information had been gathered about the tonnage and grade of the scheelite ore in the mine, and several thousand tons of ore grading higher than 0. 25 per cent WO3 (tungstic trioxide) were known to be available.

Jan 1, 1943

By Zuoren Huang

The flowsheet of the hyd.rametallurgical zinc plant which is in use at Zhuzhou Smelters is described. The principal features of the process are fluid bed roasting, continuous calcine leaching, solution purification, zinc electrolysis, electrolyte cooling, and finally, the melting and casting of zinc forms. The Zhuzhou process for the flotation of silver from the leach residue collected on the Moor filters is described. The technical details of the process are outlined and the operating data are analyzed. The flotation facilities are described in detail and concentrate grades and recoveries are presented. The waelz process is used to concentrate In, Ge and Ga remaining in the leach residue after silver flotation. These elements subsequently are recovered by selective leaching in sulphuric acid, and details of this operation also are presented.

Jan 1, 1990

By R. W. Ruddle

The advent of new methods of roasting, “flash roasting and fluidisation”, which are now under development, combined with the use of oxygen-enriched air, may markedly increase the SO2 content of roaster smokes. Skogmo suggests that the production of smokes containing as much as 50 per cent SO2 should be feasible. The use of flash smelting processes may well enable a similar increase in the SO2 content of smelter smokes to be made. In this event the recovery of sulphur would be greatly cheapened and the process might well be much more economic, even if allowance is made for the cost of generating the oxygen

Jan 1, 1953

By F. Botha, J. S. Adams, W. T. Ruhmer

The report describes the Amuran project in the Welkom area, which is being conducted by six members of the Anglo American Group. The project comprises three plants for the flotation of pyrite, a twin-stream uranium plant in which 'high-grade' slimes and pyrite can be leached separately, a pyriteburning sulphuric acid plant, and a plant for the recovery of gold from calcines. Details of these plants including capital costs and estimated production are given. Mention is made of the adsorption of gold onto activated charcoal and wet high-intensity magnetic separation as possible suitable processes for these residues.

Jan 1, 1978

By K Gottliebsen

Many hydrometallurgical and chemical processes produce large amounts of acid waste. This is illustrated by the copper S/X-E/W process used at the Mount Gordon mine in North Queensland, which produces a tank house bleed in the form of a concentrated sulphuric acid stream. The bleed stream is necessary to remove impurities such as iron that decrease the efficiency of the electrowinning process. However, this bleed stream comes at a high financial cost due to large quantities of make up sulphuric acid required as well as the need to neutralise the bleed stream prior to disposal. The present work describes the approach taken to develop a solvent extraction process for the selective recovery of sulphuric acid from copper tank house bleed liquor. The proposed extraction system was investigated using a branched long chain aliphatic tertiary amine tris(2-ethylhexyl)amine (TEHA) as the extractant, Shellsol 2046 as the diluent and octanol as a modifier. Computer simulations were used to determine the optimum operating conditions for the recovery process. It was found that the process was able to selectively recover up to 90 per cent of the sulphuric acid from the copper electrolyte. This is recovered as a pure aqueous acid stream at a concentration of up to 130 g/L that can be recycled back into the tank house circuit or other areas of the process. The raffinate acid concentration was reduced from 180 g/L to as low as 18 g/L, significantly reducing waste disposal and the environmental impact of the copper process. The equilibrium data and simulation results were also compared with an alternative extractant, CYANEXR 923. Finally the efficiency of the TEHA extraction system was investigated on a pilot-scale in a pulsed perforated plate column. This information was used in order to scale the process for an industrial application. The same methodology applied in this case study could be used to provide innovative and cost-effective solutions to waste streams encountered in other hydrometallurgical or chemical processes.

Jan 1, 2000

By Samantha A. Beeson, Colin Poole

"A selective flocculation-floe flotation technique is already in use in the Canadian potash industry for the removal of water-insoluble slimes, but the process is currently limited in its application and reagent consumption is high. The work described in this paper details investigations into the separation of sylvite and halite from water-insoluble slimes by selective flocculation utilising other commercially available polymers.Preliminary flocculation performance tests were undertaken on pure sylvite, halite and water-insoluble minerals to establish the importance of a number of experimental parameters. The data from these single mineral tests was then used to establish optimum parameters for the separation of sylvite and halite from a synthetic mixture similar in composition to a typical potash ore. Under optimum conditions, separation was achieved with high recoveries using dispersant and flocculant dosages significantly lower than those currently used in plant practice.IntroductionAlthough solar evaporation and solution mining are used in some operations, most potash is extracted by conventional underground techniques and is then processed by froth flotation, in order to separate the valuable sylvite (KCI) from halite (NaCl) and water-insoluble gangue minerals. Flotation is usually preceded by a desliming stage to remove water-insoluble fines, but this also results in high potash losses and affects process brine requirements. To overcome these problems, one plant operated by Cominco in Saskatchewan uses a selective flocculation -floe flotation process (Banks, 1979/1980), in which a polyacrylamide flocculant is added to produce floes of water-insoluble material with minimal incorporation of halides. A collector is then added to render the floes hydrophobic and suitable for removal by froth flotation.A number of workers have also investigated the use of selective flocculation in potash processing, but these too have been concerned with developments in the removal of water-insoluble slimes (Williams, 1960; Brogiotti et al, 1974; Chan et al, 1986; Thompson et al, 1979; etc.). This paper provides results of novel investigations into the selective flocculation of sylvite and halite from water-insoluble minerals and discusses the possible mechanisms involved."

Jan 1, 1995

By K:-I. Rhee

The recovery of tellurium from Te-cement obtained during the processing of copper anode slimes was carried out by a series of hydrometallurgical processes, i.e. leaching Te-cement in NaOH solution, precipitation of impurites with the addition of Na2S followed by direct electrowinning of tellurium in NaOH solution. The optimum conditions for each process were determined and discussed in terms of various parameters associated with each step to maximize the recovery and purity of tellurium. The final purity of tellurium from electrowinning in alkaline solution was found to be more than 99.9 %.

Jan 1, 1997