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By R. J. Roman

The results of column leaching tests on 3 different mixed oxide/sulfide ores from a Chilean deposit are presented. The 3 samples contained different ratios of oxide copper minerals to chalcocite and chalcopyrite. The results of the column leach tests are interpreted in light of the shrinking core model for coarse particle leaching. The copper recovery from the oxide minerals and the sulfide minerals each conforms to the shrinking core model with the total recovery equal to the weighted average of the individual minerals. The use of column test results to predict field heap leaching results using the multi-mineral shrinking core model is discussed.

Jan 1, 1993

By S. P. Fleming

An acid mine drainage sludge containing 0.26% nickel, 0.21 % cobalt and 4.5% manganese was leached in a one-liter reactor with sulfuric acid in the absence and presence of sodium sulfite. The effects of temperature (25°C and 52°C) and sodium sulfite concentration (up to 0.03 molar) on the leaching rate were particularly investigated. The leaching rates of nickel and cobalt with sulfuric acid have been determined to be controlled by diffusion through product layer. It has been found that effective leaching conditions to give more than 90% extraction of nickel, cobalt and manganese while keeping the iron extraction below 10% come with an elevated temperature and a low sodium sulfite concentration at pH 3; for example, 52°C and 0.01 molar sodium sulfite.

Jan 1, 1992

By D. L. Leach

This report covers laboratory experiments simulating in situ copper recovery from primary sulfide ores in sulfuric acid systems pressurized with oxygen. Copper extraction and acid consumption data are correlated with temperature in the range 50-150°C and quenched pH in the range 0.25-2. Adverse changes in the chemical composition of the leach solution and the formation of gangue-mineral alteration products and secondary sulfides can severely limit the extraction of copper above 90°C. Keeping the pH below the self-buffered value of about 2 and the temperature below about 110°C produces the maximum extraction rates. Forcing the pH too low not only results in excessively high acid consumption but may also limit the extraction of copper at temperatures above 100°C.

Jan 1, 1975

By J. E. Murphy, Y. Topkaya, R. E. Johnson, B. R. Palmer, D. K. Dysinger, R. B. Bhappu, A. E. Clark

The U.S. Bureau of Mines utilized a central composite experimental design to investigate leaching kinetics of pure chalcopyrite in the sulfuric acid-oxygen system. The six factors studied were oxygen pressures of 150, 300 and 450 PSI, leach temperatures of 35O, 65O, and 95O C, oxygen-to-chalcopyrite mass ratios of 0.5, 1.0 and 1.5, leach times of 8, 16 and 24 hours, initial acid strengths of 0.6, 1.2 and 1.8 g H,SO,/g chalcopyrite and pyrite to chalcopyrite mass ratios of 0, 1, and 2. Chemical analysis and oxygen consumption measurements indicatedthat leaching occurs through three consecutive reactions. Temperature was the primary determinant of leach rate and solution acidity was the primary determinant of oxygen consumption, sulfur oxidation, and activation energy.

Jan 1, 1993

By J. G. van der Watt

Due to increasing concerns of global rare earth element shortfalls in the near future, possible alternative sources of rare earth elements have recently become of economic interest. One such alternative is decanting acid mine water originating primarily from abandoned old mines in the Witwatersrand region of the Republic of South Africa. In this study, a novel way of rare earth element removal from the acid mine drainage was employed, making use of bentonite clay, which has very good adsorbent properties, as a rare earth element carrier material. The process can be economically viable only, if the elements can be selectively removed from the bentonite clay carrier material so as to yield reusable clay. Acid leaching was proposed to liberate the adsorbed rare earth elements from the bentonite clay. Accordingly, acid leaching experiments were performed to study the desorption of three commonly-found rare earth elements, namely neodymium, samarium, and dysprosium, from bentonite clay in the presence of sulphuric and hydrochloric acid. It was established that the three rare earth element species could be selectively removed as a group from iron, magnesium, and manganese metals through the careful manipulation of the pH. An investigation into the kinetic aspects of the rare earth element desorption process from the bentonite clay was also undertaken. The applicability of various kinetic models such as zero-order, pseudo first-order, pseudo second-order, elovich, parabolic diffusion, and power function were tested to describe the time-dependent desorption of rare earth elements from bentonite clay. It was determined that the pseudo second-order kinetic model represented the dissolution processes for neodymium, samarium, and dysprosium from bentonite clay the most accurately.

Jan 1, 2012

By M. Walawalkar, C. K. Nichol, G. Azimi

Rare earth elements (REEs) are essential to the production of technologically advanced products, including wind turbines, electric and hybrid cars, automotive catalytic convertors, hand held electronic devices, and military defense systems. Their application in environmentally beneficial technologies puts them at a high priority for ameliorating environmental concerns. Because of their increasingly high demand and limiting resources, it is highly beneficial to find new sources for them. The present study is focused on leaching REEs from a phosphate based fertilizer by-product, known as phosphogypsum (PG). Global estimated world production of PG is about 240 million tonnes that is stored in the form of wet stacks adjacent to fertilizer production facilities, occupying vast amounts of land. PG also contains trace metals, which are detrimental to the environment, and hence it is crucial to monitor these stacks. We obtained our PG sample from Agrium Inc. (a fertilizer manufacturing company in Canada). We performed a comprehensive characterization of the PG sample and found that it predominantly contains Y, La, Ce, Nd, and Sm. For leaching purposes, we used hydrometallurgical acid leaching techniques. We investigated three acids over a range of operating conditions such as acid concentration, temperature, solid-to-liquid ratio, and residence time. We achieved an average leaching of 81% with hydrochloric acid as leachant using operating conditions that can be easily replicated in industrial settings. If our process is implemented, not only will it reduce the size of the existing PG stacks (mitigating their environmental risks), but also enable the development of new innovative routes for extracting REEs, making countries like Canada less vulnerable to supply disruptions.

Jan 1, 2016

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 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

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 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 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 S Moradi

The aim of this study was to investigate the possibility of leaching zinc from the tailing (-2 mm) of Angouran flotation plant using sulphuric acid. The tails were first calcined at 800¦C for one hour. The effect of factors such as stirring rate, temperature, initial acid concentration and solid/liquid ratio, on zinc extraction were studied. The results obtained show that at the best conditions, about 91 per cent of zinc were leached using a stirring rate of 800 rpm, a reaction temperature of 50¦C, a 30 minute reaction time, a 100 g/L initial sulphuric acid concentration and solid:liquid ratio 1:6 g/ml. The P85 of the feed material was about 150 ¦m and acid consumption was 600 kg/t.

Jan 1, 2002

By Kyoungkeun Yoo

A citric acid leaching test after heavy medium separation process was performed to remove the heavy metal contaminat such as Pb, Cu, and Zn from contaminated soil. The lead component was removed from soil by heavy medium separation whereas zinc component shows 0 % - 37.9 % removal ratio with particle size. The citric acid leaching was conducted to remove the remained zinc component. The removal ratio of Zn reached to more than 95 % under the conditions; 10 % in pulp density, 200 rpm in agitation speed, 50 °C, 2 mol/L in citric acid concentration at 2 hrs in leaching time. These results suggest that the heavy medium separation method followed by citric acid leaching is effective for treatment of soil contaminated with various chemical forms of heavy metal. Keywords: metal contaminated soil, citrate, heavy medium separation, zinc

Sep 1, 2012

By E. LaRocque

"Denison Mines Limited has been processing uranium ore in the Elliot Lake area since May of 1957. The ore body is located i11 the Huronion sediments on the north limb of the Quirke Lake syncline. Minerals of metallurgical significance are brannerite, uraninite and monazite. Gangue minerals include quartz, chlorite, sericite, pyrite, pyrrhotite and rutile. Silica content of the ore is greater than 75 % and pyrite-pyrrhotite content around 7 %. Specific gravity of the ore is 2.75.The Denison flowsheet is based on sulphuric acid leaching to extract the uranium values from the ore followed by liquid-solid separation to wash the uranium-bearing solution from the barren ore. The resultant pregnant solution is concentrated and purified by ion exchange followed by two-stage precipitation with ammonia to produce an ammonium yellow-cake.Elliot Lake ores require high acid concentrations and high temperatures to extract the uranium from brannerite and monazite. Uraninite will leach at low acid concentrations as long as oxiclizing conditions exist. The whole ore must be treated so that these conditions do exist.This paper provides a description of the leaching process equipment as originally installed and major changes in that equipment to date. Process control refinements to economically improve extraction are also discussed."

Jan 1, 1979

By C. Arthur Lefler

The hydrometallurgical processes for recovery of gold and silver in use at Round Mountain, Nevada and the practical aspects of the operation are discussed. Included are heap leaching by cyanidation, upgrading of solutions by adsorption of values on activated carbon and desorption (stripping) for electrowinning and subsequent refining steps to produce dore metal.

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