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By H. Razavizadeh

The recovery of copper from oxide ores containing from 2% to 4.1% acid-soluble copper from the Sarcheshmeh mine in Kerman province of Iran was investigated by agitation leaching in sulfuric acid media. Quartz, as a major constituent, and malachite, as the major copper mineral, were identified in the ore by X-ray diffraction (XRD) and microscopic studies. A series of leaching experiments on the ore was performed over a concentration range of 10 to 20 g/L and temperatures from 15° to 50°C for a maximum of 3 hours. An initial high dissolution of copper was followed by a very slow reaction. By increasing acid concentration and temperature, the dissolution rate of copper was accelerated. However, by increasing acid concentration, the acid consumption of ores was also increased. Surface morphology of solid residues at various leaching times was examined with a scanning electron microscope (SEM). All SEM micrographs of the partially reacted particles displayed a needle structure. The kinetic studies showed that the malachite dissolution rate could be modeled in two stages. About 85% of the malachite dissolution occurred in Stage 1, while the remaining 15% occurred in Stage 2. The data obtained for the leaching kinetics indicate that the dissolution for Stage 1 was a diffusion-controlled reaction with an activation energy of about 20.6 kJ/mole, and the dissolution for the Stag 2 was a chemical-controlled reaction with an activation energy of about 87.6 kJ/mole.

Jan 1, 2008

By S. Jolivet, A. Klöfverskjöld, P. Latostenmaa, B. P. Wilson, M. Lundström, P. Halli

Doré slag is secondary product of copper electrore fining anode slime treatment. The slag consists mainly of different slag formers, such as borax and silica based minerals together with soda and traces of sealing materials. However, Doré slag also contains metals present in the anode slime, such as Ag, Cu, Bi, Sb and Te. At some plants, slag cleaning has been conducted utilizing conventional grinding and flotation circuits in order to enrich the content of precious metals in the slag [1]. Nevertheless, the routine consideration of the slag as a secondary raw material has so far not been adopted on an industrial scale. The current study presents the preliminary results of the nitric acid and chloride leaching of antimony from Doré slag. Chemical and mineralogical analyses (SEM-EDS, AAS, XRD) were performed for the raw material, leaching residue, and leaching solution in order to ascertain the metal recovery and kinetics during leaching.

Mar 1, 2017

By Fathi Habashi

SELENIUM AND tellurium are by-products of electrolytic copper refining. Copper anodes contain 0.01-0.2 % Se and up to 0.04 % Te; during electrolysis these deposit as a slime at the bottom of the cell. Such slimes contain 2-15% Se and 0.5-8 % Te, as selenides and tellurides of gold, silver, and copper. The following compounds have been identified in the slimes: (Ag, Au)Te2, Ag2Se, Ag2Te, CuAgSe, Cu2Se, Cu2Te. Anodic slimes from copper electro¬lytic refining are a grayish black powder, minus 200 mesh. When this complex mixture is decomposed in an acid medium, elemental selenium and tellurium can be obtained, while in an alkaline medium, both metals go into solution. Acid leaching The slime is charged in acid digesters of 800-900 lb capacity, heated to 260°C with 78 % H2SO4, for 8-12 hr. During digestion the following reactions take place:

Jan 1, 1970

By Fathi Habashi

CLAYS CONTAIN the mineral kaolinite, a hydrated aluminum silicate. It is an important potential raw material for aluminum, especially for those countries possessing no bauxite deposits. Therefore, there is no surprise to find a large volume of research done for the recovery of aluminum from clays. Kaolinite (Al203?2 SiO2?H20) contains 39.8% A1203; 46.3 % SiO2, and 12.9 % H2O. The analysis of a good-quality dried clay of an Eastern Canadian siliceous shale is shown in Table 8-1.

Jan 1, 1970

By Jaroslav Briancin, Katarína Gáborová, María Florencia Márquez-Zavalía, Marcela Achimovičová, Ol’ga Šestinová, Pavel Škácha

Naumannite (Ag,Se), a silver selenide mineral, was first discovered in 1828. This study investigates the structure, morphology, chemical composition, and leaching kinetics of naumannite. Mechanical activation in a planetary ball mill increased its specific surface area, enhancing the leaching rate of silver in nitric acid (HNO3). The study found that mechanical activation significantly improves silver recovery.

Jul 1, 2023

By Katarína Gáborová, Marcela Achimovičová, María Florencia Márquez‑Zavalía, Jaroslav Briančin, Oľga Šestinová, Pavel Škácha

In this paper, the structure, morphology, chemical composition and leaching of naumannite ( Ag2Se) were studied. Mechanical activation of Ag2Se in a planetary ball mill resulted in increasing its specific surface area. The influence of milling time on the particle size was studied by particle size distribution analysis and scanning electron microscopy. The leaching kinetics of non-activated and mechanically activated samples in 3.5–7.9 mol/L HNO3 solutions were examined. Subsequently, the effect of leaching temperature in the range of 25–50°C on the recovery of silver from mechanically activated naumannite was documented. The maximum yield of silver 94% was achieved after 120 min of leaching in 7.9 mol/L HNO3 at 50°C for the sample mechanically activated for 30 min. Based on calculated activation energy values from the Arrhenius equation, a mixed leaching mechanism consisting of diffusion and surface chemical reaction as the rate-controlling steps of the solidliquid reaction was assessed.

Mar 4, 2023

Sulphuric acid leaching at atmospheric pressure of a number of Australian nickeliferous laterites was examined in detail. The results obtained indicate that nickel can be extracted from these ores, although the conditions required to achieve about 90 per cent extraction vary widely. These differences in reactivity can be related to the types of nickel-bearing minerals present.Nickel is more readily extracted from nickel-bearing clays such as chlorite, than it is from nickel-containing goethite.The results clearly show the problems associated with the processing of complex ores such as the nickeliferous laterites, and they indicate the need for complete characterization of each deposit.

Jan 1, 1978

By Tomás Vargas

Treatment of copper sulphides minerals via leaching-solvent extraction-electrowining (lix-sx-ew) have been for years the process scheme proposed from hydrometallurgy as an alternative to the conventional flotation-smelting route. Many different hydrometallurgical technologies of this type have been proposed, investigated and tested over the years, but until today none of them has come close to endanger the predominance of the flotation-smelting route in the copper industry. However, bioleaching introduced an important breakthrough in this context as this technology enabled the alternative hydrometallurgical treatment of secondary copper sulfide ores and low grade ores by applying heap bioleaching processes. Extension of its application to primary copper ores have been limited until now by the nature of chalcopyrite which is very refractory to leaching. However, important improvements are expected by using of improved designs that enable to operate at high temperature ranges where copper is leached faster under the catalytic action of thermophile microorganisms. The present work analyzes future scenarios for the application of this technology either in heaps or in-situ, which can become very attractive as copper can be produced reducing drastically the energy consumption associated to the mining, crushing and grinding stages.

Jan 1, 2014

By Fathi Habashi

A LARGE number of metals are found in nature, in rich deposits, in the form of sulfides as shown in Table 4.1. Treatment of sulfide ores is done mostly by pyrometallurgical operations; it is only recently that hydrometallurgical processing started to compete with pyrometallurgical methods.

Jan 1, 1970

By B G. Lottermoser

The placement of sulfidic waste below the groundwater table ensures limited interaction with the hydrosphere and suppression of sulfide oxidation. However, if sulfidic waste is placed above the groundwater table and remains uncovered, the backfill becomes part of the unsaturated zone and is exposed to atmospheric oxygen and leaching. This study aims to establish the leaching behaviour of sulfidic waste placed above the groundwater table and the impact of such leachate on the local aquifer at the Thalanga base metal mine. Mining of the Thalanga copper-lead-zinc deposit resulted in a large final mining void (600 m + 150 m + 70 m) and extensive underground workings. The underground workings were partly filled with tailings and the open pit was partly backfilled with acid producing sulfidic waste rock. In addition, the pit serves as a sink for acidic run-off from adjacent waste rock piles and mine workings. To date, the backfill of sulfidic waste rock placed into the pit has not been capped with benign materials and for most of the dry season, the surface of the backfill is covered by melanterite-type efflorescences. Results of kinetic column leach experiments conducted on the sulfidic waste indicate that Cd, Cu, Zn and SO4 rich waters migrate from the backfilled sulfidic waste into the local unconfined aquifer. However, the seepage of alkaline (pH 7.3 - 8.0), high conductivity (>10 000 ¦S/cm) tailings waters into the remaining pit void clearly shows that the acid leachate originating from the sulfidic waste rock does not impact beyond the waste repository and its immediate environment. Geochemical modelling implies that minimal or no mixing occurs between the acid waste rock leachate and the alkaline tailings waters.

Jan 1, 2003

Fluid bed sulphation roasting followed by leaching at low pH was developed as a means of separating copper and bismuth from a mix- ture of flotation concentrates and flash fur- nace flue dust. A 65 tonne/day plant was established and typically achieves 89 percent copper extraction with negligible bismuth and iron extraction. Acid-chloride leaching of bismuth from copper leach residue was then established, with bismuth oxychloride being precipitated from solution by dilution with water. A small scale batch plant presently in operation typic- ally achieves 80 percent recovery of bismuth to bismuth oxychloride.

Jan 1, 1986

By W. G. Cornwall

The mining field, inclusive of the treatment plant is located in the foothills, five miles inland from the northern coast of Viti Levu, the largest and most important island of the Fiji Group. Gold and silver occur as tellurides, associated with the sulphides, and as minor occurrences of electrum.

Jan 1, 1975

By W. G. Cornwell, R. J. Hisshion

The development of a process for the treatment of telluride concentrates at Emperor Gold Mining Co. Vatukoula, is discussed. The concentrate is produced by flotation in cyanide solution immediately after milling. Two stages of cleaning are employed and a high-grade concentrate is produced usually containing more than 50% of the precious metal. The process developed in the first instance involved batch roasting followed by batch leaching in strong cyanide solution, recovering gold and silver only. The procedure currently employed is chemical oxidation of the tellurides using hypochlorite as the oxidant in a basic solution of a caustic soda and sodium carbonate. After oxidation the gold and silver are recovered by cyanidation. The residue is then leached with sodium sulfide and elemental tellurium precipitated with sodium sulfite.

Jan 1, 1977

This paper describes experimental investigations into the reductive decomposition of arsenopyrite/pyrite (FeAsS/FeS2), a gold and silver bearing iron/arsenic sulphide using metallic iron powder. The main objective was to establish and experimentally confirm the thermodynamic and kinetic feasibility of reductively decomposing the sulphide matrix for precious metals recovery. In this context, cyanidation leach tests were conducted on the reductive leach residues. The arsenopyrite/pyrite flotation concentrate, mainly studied in the hydrochloric acid medium, decomposed through the non?oxidative chemical dissolution reaction and the reductive decomposition reaction for the arsenopyrite and pyrite components, respectively. Desulphurization levels below 65% were achieved at pH values below 0.15 and iron to concentrate ratios above 1, and the system was characterized by overall very slow kinetics. The reaction system had a direct linear relationship with iron to concentrate ratio and an inverse relationship with pH. An analysis of the pyrite/iron galvanic system showed that pyrite forms a partially inert cathodic surface on which the anodic dissolution of iron occurs, supported by the hydrogen evolution reaction. This phenomenon explained the low desulphurization levels and the mineral decomposition seemed to be restricted to the non?oxidative chemical dissolution reaction for the arsenopyrite component. The cyanide leach of the reductive leach residues showed very little improvement in gold recovery. The reductive leach process for arsenopyrite/pyrite has considerable limitations in terms of both desulphurization and precious metals liberation. Keywords: Pyrite, arsenopyrite reductive leaching, hydrogen sulphide, invisible gold

Jan 1, 2009

By Gale D. C, Giles D. E, James Parker A

A zinc-lead-iron sulphide concentrate from McArthur River can be leached at atmospheric pressure in aqueous solutions with HCI, FeCl3, CuCI2, HCl/02 HC1/02/ CuCI2, Cl2, and CI2/CuCI2 to extract >98 per cent zinc, >90 per cent silver, and <15 per cent iron in 1-7 h at 95°. Galena is more reactive than sphalerite and is converted to lead chloride. Except for leaching with HCI, elemental sulphur is the the main by-product of the leach with little or no sulphate ion detected. The rates and products of these reactions are compared. It is recommended that aqueous slurries of the concentrate (50 per cent w/v) are leached at atmospheric pressure with Cl2 in the presence of copper(II) chloride, which accelerates the extraction of zinc. Hydrochloric acid is a by-product which is consumed bYJeaction with fresh concentrate in air at 60°. Copper is subsequently stripped from the leach solution and recycled by reaction with fresh concentrate at 100°. The separation, recovery, and electrolysis of zinc and lead chloride is outlined since this process would produce Cl2 gas which can be recycled and used directly in the leaching stage.

Jan 1, 1976

By Pelin Altinkaya, Maria Leikola, Jussi Liipo, Mika Haapalainen, MARI LUNDSTROM, Eero Kolehmainen

Chloride leaching has many advantages compared with other alternative processes to cyanidation because it has increased metal solubility, improved redox behavior and high leaching rate properties. Also, chlorination is a nontoxic and environmentally friendly process, which is sought by the mining and metallurgy industry. In addition to these advantages, chloride is an effective reagent for the dissolving of metals from refractory ores and process residues that have trace amounts of valuable metals waiting to be valorized. 

By Pelin Altinkaya, Maria Leikola, Jussi Liipo, Mika Haapalainen, Mari Lundström, Eero Kolehmainen

With decreasing ore grades, the tailings of mining operations are becoming of increasing interest as metal-containing secondary raw materials. The objective of the current work was to investigate chloride leaching of gold, copper, cobalt, nickel, and zinc present in the flotation tailings. In the current study, the effect of cupric ion as an oxidant (0–50 g/L) and NaCl (150–250 g/L) on metals extraction was investigated. The other parameters, such as pH (1.8), temperature (95 °C), solid/liquid ratio (25%), oxygen feed rate (1200 mL/min), leaching time (24 h), and stirring rate (950 rpm), were kept constant. Gold dissolution rate was shown to increase with increase in cupric ion concentrations up to 50 g/L. Also, increase in NaCl concentration up to 250 g/L increased gold extraction. Majority of the copper present in the flotation tailings could be dissolved with all the all solutions investigated. The other base metals, cobalt, nickel, and zinc extractions were also shown to increase with increase in the cupric ion and sodium chloride concentration. However, even in the absence of cupric ion addition (t = 72 h, 250 g/L NaCl), the final extraction of Cu, Ni, Zn, Co, and Fe increased up to 98, 93, 83, 76, and 80%, respectively. This shows the power of inherently originating oxidants present in the tailings. Furthermore, the solid analysis of the leach residue indicated that leaching the flotation tailings in pure NaCl may result in partial gold dissolution. The results demonstrate that significant amount of metals present in the tailings could be extracted even with only NaCl as added chemical in the presence of oxygen feed, oxidizing agents originating directly from the raw material. This can provide an advantageous cyanide-free method for extraction of metals from very low-grade tailings with low chemical consumption.

By Hai-Peng Guo, Lu Jiang, Xiao-Man Yan, Bing Xie, Hong-Yi Li

Traditional, calcium compounds cannot be used in vanadium extraction processing due to the limitation of sodium roasting technology, which results in unsatisfied dephosphorization for vanadium-containing hot metal in converter and thus increasing the dephosphorization burden for subsequent steelmaking. The calcium roasting technology is an alternative for vanadium extraction and allows adding lime into vanadium slag in converter. In the present study, vanadium was leached by soda from roasted vanadium slag with high calcium content. Effects of roasting temperature and time, leaching temperature, Na2CO3 concentration andliquid to solid ratio on vanadium leaching rate were studied. The results show that 94% of vanadium was extracted by roasting the vanadium slag at 950℃ for 2.5h and then leaching in 160g/L Na2CO3 liquid at 95℃ for 1h with liquid to solid ratio of 10:1 ml/g. Impurities of Si and P reacted with calcium and formed insoluble Ca2SiO4 and Ca3(PO4)2 respectively.

Jan 1, 2015

By C. A. Born

AMAX Inc. climax's mine-mill complex at Climax, Colorado is best known for its molybdenite production: 50,000 TPY of molybdenite flotation concentrate (95% MoS2) from mill heads containing 0.35% molybdenite. Less recognized are the byproducts that are concentrated from the tails generated in the rougher flotation circuit (see Figure 1). Climax recovers a heavy mineral concentrate by passing the rougher tailings from molybdenite flotation over Humphrey spirals (Figure 2a). This concentrate is floated using xanthate at pH 4 to remove sulfides, principally pyrite. The tails from the pyrite float are then cleaned of quartz, feldspar and light minerals by tabling (Figure 2b). Amine flotation follows at pH 1.5 and 75°C to float monazite (Figure 2c). The present circuit at Climax dries the monazite flotation tailings and utilizes high intensity cross-belt magnetic separation Figure 2d) to generate a high grade wolframite concentrate and a low-grade cassiterite product, both of which are contaminated with the mineral brannerite, (UO, TiO, UO2)TiO3. The uranium specifications have recently become difficult to meet due to the increasing percentage of brannerite reporting in the mill feed. A plant is being built utilizing hot sulfuric acid and sodium chlorate to leach this uranium from the monazite flotation tailings prior to their being fed to the wolframite- cassiterite magnetic separators. This modification is the subject of our paper.

Jan 1, 1975

By M. Canbazoglu

In this study, optimum sulphuric acid leaching parameters of zinc and other valuable metals from two different waelz clinker samples were researched. Origins of the samples are electric arc furnace dust and original non-sulphite zinc ore waelz clinkers. The main zinc mineral determined as zincite in these samples. Collective dissolution of zinc and other valuable metals in waelz clinkers can easily achieved by sulphuric acid leaching of them. 96% of zinc can be dissolved at 80 ºC, 60 g/L sulphuric acid concentration, 60 min. leaching time and 1/5 solid/liquid ratio. Almost 90% of Cd, Cu and Co in clinker dissolved at these conditions while dissolution of Fe was almost 20%. The consumed acid amount for dissolved per ton of zinc is approximately 1.7 tons at these conditions.

Jan 1, 2014