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In the leaching and recovery of gold from copper containing gold ores using cyanidation, some copper minerals also dissolve which can have a detrimental effect on the process efficiency and economics. There are also environmental concerns with current operations discharging cyanide and copper cyanide species into tailings dams, and strict regulations prohibit this for greenfield plants. Cyanide destruction however is not an economic option for new plants treating copper-gold ores. This paper evaluates a potential recovery process to overcome the copper cyanide problem. Despite gold dissolution occurring in copper cyanide solutions having no free cyanide, gold recovery from ores is lower compared to free cyanide. Thus, cyanide addition has to be optimised to maximise the gold recovery whilst endeavouring to eliminate free cyanide at the completion of leaching (CN:Cu less than 3:1). Subsequently, copper cyanide can be recovered using activated carbon. The loaded copper cyanide is easily recovered from the carbon with a cyanide soak and distilled water wash such that the eluted CN :Cu ratio is less than 4:1. The copper could be recovered efficiently by electrowinning, by decreasing the pH of the eluant to 5. The remaining cyanide solution, after electrowinning, is then neutralised to form NaCN and recycled.

Jan 1, 2005

By J. A. McKeown, B. M. Sanders, A. Dadgar, R. H. Sergent

This paper describes the leaching of gold with Geobrom® 3400 from a very rich (6.2 kg/tonne) black sand concentrate and its subsequent recovery by ion exchange resins and solvent extraction. The dissolution of gold from black sand ranged from 94 to 96%. Gold loadings and recoveries of nearly 100% were achieved by both ion exchange and solvent extraction methods. The kinetics and equilibrium of loading gold onto ion exchange resins and its elution from the resin were studied. The commercial ion exchange resins PAZ®- 4, DOVEX®- 21K, and Amberlite® XAD- 7, and Great Lakes Chemical experimental GLR-30 were evaluated in this work. Diethylene glycol dibutyl ether was used in all the solvent extraction work for gold recovery from Geobrom leach solution. The consumption of Geobrom was about 130 kg per tonne of the concentrate processed. Therefore, the cost of lixiviant to extract 3.11 x l0-2 kg of gold from the black sand was only about $1.00. Based on the electrochemical regeneration of bromine, the economics of bromine process vs. cyanidation were investigated.

Jan 1, 1990

By J. P. Allen

Cobalt, nickel, and manganese can be recovered quickly, cleanly, and selectively from cobalt-rich manganese crust by leaching in dilute acid solutions in the presence of hydrogen peroxide. Common selective precipitation methods can then be used to recover these metals from the leach liquor. The leaching reaction produces only oxygen as a byproduct, and the waste residue from leaching was tested and found to be non-hazardous under existing EPA specifications. The leaching process is effective in either fresh water or seawater. This research is a part of the Bureau of Mines program to identify the mineral resources of the Exclusive Economic Zone and devise methods for recovering their valuable components.

Jan 1, 1990

By J. P. Allen, D. A. Rice, H. L. Abercrombie

Cobalt, nickel and manganese can be recovered quickly, cleanly and selectively from cobalt-rich manganese ocean crust by leaching in dilute acid solutions in the presence of hydrogen peroxide. Common selective precipitation methods can then be used to recover these metals from the leach liquor. The leaching reaction produces only oxygen as a byproduct, and the waste residue from leaching was tested and found to be nonhazardous under existing EPA specifications. The leaching process is effective in either fresh water or seawater. This research is a part of the Bureau of Mines program to identify the mineral resources of the Exclusive Economic Zone and devise methods, for recovering their valuable components.

Jan 1, 1992

Existing processes for the recovery of platinum group metals (PGM) from UG-2 ores place stringent requirements with regard to chromite content upon feed material to the plant. An alternative hydrometallurgical process for which this is not an issue is proposed that involves reductive roasting of the ore followed by leaching with HCl/chlorine, ion exchange, and recovery of PGM from the ion exchange eluate by reduction. Problems associated with leaching of PGM are discussed and results of laboratory testwork for each of the process steps are given.

Jan 1, 2004

By Youze Xu

The leaching and reduction of Cr(VI) in soils contaminated by chromium-containing slag were investigated with column experiments. The results showed that Cr(VI) concentration in leachate decreased from 988 mg L-1 to 407 mg L-1 after 7 days when water was used as leaching agent. However, Cr(VI) in leachate in culture medium treatment was not detected at above periods. In comparison, Cr(VI) concentration in leachate from autoclaved soil remained 398 mg L-1 at the end of experiment. The results indicated that the indigenous microorganism was capable of reducing Cr(VI). The amount of Cr(VI) reduction was 335 mg at the whole experiment period and the reduction rate was 67.09 mg d-1. Carbon supply showed minor effect on Cr(VI) reduction. However the reduction capability of Cr(VI) could be improved by combination of the carbon and the nitrogen. The optimal quantity was 5 g L-1 of yeast extract and 4 g L-1 of glucose.

Jan 1, 2009

Mary Kathleen Uranium Ltd. recommenced operations in early 1976 following a twelve year period of care and maintenance. Several sections of the plant were modified or completely changed for the second operation. The most important change was the replacement of ion exchange with solvent extraction as the means of purifying and upgrading uranium rich Isolutions Most of the problems experienced in the solvent extraction system originate from the leach liquor which has a strong tendency to form stable emulsions. This has been countered by some careful control of leaching conditions and by closer observation of operations in the solvent extraction area. Most problems have now been resolved and plant recoveries are quite Satisfactory.

Jan 1, 1978

By E. E. Malouf

Leaching is rapidly becoming an important adjunct to mining operations. A summary of various leaching operations is presented showing the role of these activities to the successful operation of the mining venture. The summary will include recent developments in dump, heap, and in situ leaching of copper and uranium ores. It will include a discussion on the expanding impact of industrial and academic research on leaching as a mining tool, showing the interplay of chemical kinetics and biogenic activity.

Jan 1, 1973

By P. Zhu

The leaching behavior of copper powders recovered from waste printed circuit boards (WPCBs) was investigated using electrogenerated chlorine and aeration in a sulfuric acid solution with added NaCl and CuSO4. Various parameters were studied to understand the mechanism of the copper leaching. When the current density increased from 0 to 0.1 A/cm2, the temperature from 30 to 60° C, agitation speed from 100 to 400 rpm, and the size ranged from below 75 to 150 µm, the leaching ratios of copper powders increased. The leaching kinetics were analyzed using a shrinking core model: 1-(1-X)1/3 = kt, and the leaching process was chemically controlled with an activation energy of 14.7 kJ/mol due to Cl- ion catalysis. The ratios of [Cu+]/[Cu2+] at various temperatures (30, 40, 50, 60 and 70° C) showed a periodic variation with the increasing of the leaching time. The redox potentials of the leaching solution and copper leaching ratios in different conditions were measured. The result indicates that dissolved oxygen achieved the dissolution of coppers and electrogenerated chlorine can promote the dissolution of copper powders and improve the reaction rate of kinetics in the presence of oxygen.

By P. Latostenmaa, T. Hailemariam, M. Lundström, P. Halli

Doré process is unique way to treat precious metals in the copper electrorefining plant, since it uses anode slime as a raw material. In this pyrometallurgical process, also slag is produced together with Doré metal. However, the process is never ideal, and a part of metal is always lost into the slag. By using Doré slag as raw material for simply mineral acid leaching, some of the lost elements can be extracted and recovered, thus, increasing the sustainability and keeping metals in the circular economy loop. The Doré process itself is developed to convert almost all precious metals and PGMs in the Doré metal, leaving other elements into the slag fraction. The Doré slag consist of slag formers, such as borax and silica based minerals, but contain also traces of sealing materials. In addition, the Doré slag is rich in metals present in the anode slime, such as Cu, Bi and Sb, which are not converted into the Doré metal. There are only few published applications at the precious metal plants, where conservative grinding and flotation circuits have been utilized for the enrichment of valuable metals present in the Doré slag fraction (Godbehere, P.W., 1983; Segawa, C. and Kusakabe, T., 1996). However, the present study shows an alternative approach for the recovery of metals present in the Doré slag by hydometallurgical approach. The study focus on direct mineral acid (aqua regia) leaching of metals present in the Doré slag. The initial dissolved concentrations of Cu, Bi and Sb from Doré slag in the temperature range of 30-50 °C under oxygen purging with addition of 1 vol-% of H2O2 are presented. Chemical and mineralogical analyses were performed for the raw material, leaching residue, and leaching solutions in order to ascertain the leaching behavior of the metals of Bi, Cu and Sb.

Jan 1, 2018

By Kenneth N. Han, Peter Mam-soo Kim

The extraction behavior of gold in ammoniacal solutions in the presence of bromine as an oxidant has been studied. The effect of concentration of ammonia, bromine, and pH of the solution on the rate of dissolution of gold was investigated. The dissolution behavior has been analyzed using electrochemical method, and its results were compared with direct leaching tests using gold powder. The overall dissolution of gold in ammoniacal solution containing bromine was found be affected significantly by pH of the solution and the concentration of bromine. The rate expression found in this study has the following form.

Jan 1, 2005

By Qi P. H

Gold dissolution can be achieved with a number of complexing ligands, all with varying degrees of stability. Processes based on the use of halogens have been known for many years. Beside providing the necessary complex- ing role, the halogens under the appropriate conditions are very effective oxidants for gold. Of the halogens, the gold iodide complexes are the most stable in aqueous solution. Iodine can establish the desired oxidation poten- tial for oxidizing gold, and unlike chlorine and bromine, the oxidizing potential of iodine is mild enough to avoid the oxidation of impurity constituents like metal sulfides. A series of experiments to investigate the fundamental electrochemistry of gold in different halide solutions was performed. Cyclic voltammograms of gold in the presence of 10-2 M chloride, bromide and iodide, respectively show that for a given potential the rate of oxidation is much faster in iodide solution than in either chloride or bromide. An overlay of linear sweep voltammograms for the anodic oxidation of gold in iodide and the cathodic reduction of iodine species on gold provides an effective method for examining the dissolution of gold in iodideftodine solution. The dissolution rates obtained from both electrochemical and actual leaching results are compared. Optimum leaching conditions are analyzed in terms of the iodide/iodine concentrations. Conceptual processes for the recovery of gold and iodide and the regeneration of iodine are discussed.

Jan 1, 1991

By Ning Wang, Tengfei Guo, Wanyan Li, Hanjie Wen, Hannian Gu, Zehai Li

Red mud has been reported as a potential resource for lithium. To investigate the leaching behavior of lithium, a lithium-rich red mud was neutralized with hydrochloric or oxalic acid and then leached with acetic acid. When red mud was neutralized using dilute hydrochloric acid, about 58 to 60 percent of lithium was leached with 25 percent acetic acid at 95 °C for 60 min, while more than 95 percent of sodium and 85 percent of calcium were dissolved in the leachate solution. When oxalic acid was used to neutralize red mud prior to acetic acid leaching, lower lithium leaching efficiencies of about 42 to 47 percent were observed. Nevertheless, these leaching efficiencies of lithium were close to that of calcium and much higher than those of sodium under the same conditions. In this study, 25 percent (v/v) of acetic acid at 85 °C for 60 min was demonstrated to be the optimum conditions for lithium extraction from oxalic acid-neutralized red mud using acetic acid. The results are useful and of interest for lithium recovery and purification from red mud.

By Ning Wang, Tengfei Guo, Wanyan Li, Hanjie Wen, Hannian Gu, Zehai Li

Bauxite residue (red mud) is an alkaline waste derived from the industrial process for extraction of alumina from bauxite ore. Some bauxite residues contain considerable concentrations of critical metals, for example, rare earth elements. Lithium in red mud was also reported as a potential resource. In the current study, a Li-rich red mud was leached using acetic acid to investigate the leaching behavior of lithium followed by neutralizing with hydrochloric or oxalic acid. By neutralizing red mud using diluted hydrochloric acid (0.01 mol/L), 58.04–60.27% of lithium was leached using 25% acetic acid at 95 °C for 60 min. In the meantime, more than 95% of sodium and 85% of calcium were also dissolved in the leachate solution. While oxalic acid (0.5 mol/L) was used to neutralize red mud prior to acetic acid leaching, lithium performed lower leaching efficiencies (42.41–46.88%) in comparison of using of hydrochloric acid–neutralized red mud. Nevertheless, these leaching efficiencies of lithium were close to that of calcium and much higher than those of sodium under the condition. In this study, 25% (v/v) of acetic acid, at 85 °C for 60 min, was demonstrated as optimum conditions for lithium extraction from oxalic acid–neutralized red mud using acetic acid. The results would be useful and of interest for lithium recovery and purification from red mud.

By Ning Wang, Tengfei Guo, Wanyan Li, Hanjie Wen, Hannian Gu, Zehai Li

Bauxite residue (red mud) is an alkaline waste derived from the industrial process for extraction of alumina from bauxite ore. Some bauxite residues contain considerable concentrations of critical metals, for example, rare earth elements. Lithium in red mud was also reported as a potential resource. In the current study, a Li-rich red mud was leached using acetic acid to investigate the leaching behavior of lithium followed by neutralizing with hydrochloric or oxalic acid. By neutralizing red mud using diluted hydrochloric acid (0.01 mol/L), 58.04–60.27% of lithium was leached using 25% acetic acid at 95 °C for 60 min. In the meantime, more than 95% of sodium and 85% of calcium were also dissolved in the leachate solution. While oxalic acid (0.5 mol/L) was used to neutralize red mud prior to acetic acid leaching, lithium performed lower leaching efficiencies (42.41–46.88%) in comparison of using of hydrochloric acid–neutralized red mud. Nevertheless, these leaching efficiencies of lithium were close to that of calcium and much higher than those of sodium under the condition. In this study, 25% (v/v) of acetic acid, at 85 °C for 60 min, was demonstrated as optimum conditions for lithium extraction from oxalic acid–neutralized red mud using acetic acid. The results would be useful and of interest for lithium recovery and purification from red mud.

By K. N. Han

The leaching behavior of palladium in ammoniacal solutions was investigated by using the rotating-disk technique in an autoclave. The study included the effects of rotation speed, temperature, oxygen pressure, composition of ammoniacal solutions and catalysts on the overall leaching rate. The dissolution of palladium in ammoniacal solutions was discussed in terms of the thermodynamics and a kinetic model. Even though the Eh-pH diagram suggests that palladium metal can dissolve in ammoniacal solutions at room temperature, it was found that the dissolution was very slow at temperatures below 100 °C but was significantly improved at higher temperatures. It was also found that the rotating speed of the palladium disk had no effect on the leaching rate. The apparent activation energy was found to he 69.7 kJ/mole. Based on the results obtained, it was concluded that the dissolution of palladium is controlled by chemical reaction.

Jan 1, 1999

By K. N. Han, X. Xie

The dissolution behavior of palladium powder in thiourea solutions was investigated. The variables studied were temperature, stirring speed and the concentrations of thiourea, acid and ferric ion. Palladium was found to be soluble in thiourea solutions in the presence of an oxidant such as ferric ion. The dissolution of palladium in thiourea solutions was believed to be chemical-reaction controlled, with an apparent activation energy of about 25.3 kJ/mol.

Jan 1, 1997

By D. Hong, K. N. Han

The leaching behavior of rhenium in ammonium iodide solutions with iodine as oxidant was studied. Experiments were carried out in an autoclave to investigate the effects of stirring speed, temperature, solution pH, and iodine and iodide concentrations on the dissolution of rhenium. It was found that a high temperature and a high solution pH were essential for the efficient leaching of rhenium. The dissolution rate was found to be dependent upon the concentration of iodine and iodide. It was concluded that the dissolution of rhenium under the conditions studied was controlled by the mass transfer of iodine with an apparent activation energy of 15.8 kJ/mole.

Jan 1, 1999

By D. Yu, T. A. Utigard, M. Barati

"Sulfation roasting of nickel concentrate to generate water-soluble metal sulfates was investigated as an alternative process to smelting for the recovery of valuable metals. The kinetics of dissolution of various species of the calcine in three leachants was studied; hot water, dilute HCl, and concentrated HCl. The results indicated that dissolution of sulfates by hot water at 90 ºC is achieved within 30 minutes. In order to increase the recovery of the valuable metals, acid leaching is necessary due to the existence of ferrite in the calcine.INTRODUCTIONInvestigations on the sulfation roasting of nickeliferous sulfide ores or concentrates were mainly carried out in 1950’s to 1990’s to generate water soluble non-ferrous metal sulfates followed by metal recovery by leaching (Stephens, 1953; Thornhill, 1954, 1961; Fletcher and Shelef, 1964a, 1964b; Fletcher and Hester, 1964; Frankiewicz, 1977; Norrgran, 1983). Commercialization of the sulfation roasting process was only performed by Falconbridge Ltd. (Thornhill, 1961; Boldt, 1967), which employed a sulfation roast process on nickeliferous pyrrhotite at 680 oC where Ni, Cu, and Co were leached and Fe2O3 was produced for the iron industry. In 1965, 118,342 tons of pyrrhotite was processed, however because of the huge SO2 emissions (no acid plant) relative to the nickel production and the poor economics, this plant was shut down.Nickel production from sulfide ores via pyrometallurgical means, especially the electric furnace smelting process, is a major emitter of SO2 to the environment and a big energy consumer (Warner, Diaz, Dalvi, Mackey, Tarasov and Jones, 2007). Nowadays, due to the more stringent environmental regulations as well as the need to reduce energy consumption, metal production by sulfation roasting is gaining more attention as an alternative process to smelting. The main focus of this paper is to investigate the extent and rate of leaching of valuable metals from the sulfation roasted nickel calcine, aiming to maximize the metal recovery and minimize the dissolution of iron species."

Jan 1, 2012

By E. Fried

Cemented paste backfill (CPB) is an innovative tailings management method that consists in returning considerable amounts of mine tailings into the underground mine openings. The filled stopes thus participate to the ground support that allows maximizing orebody extraction and a safer environment for the mine workers. There are also significant environmental advantages of using CPB. Indeed, up to 50% of the concentrator tailings can be backfilled, leading to a smaller volume of problematic waste materials to be managed at the surface. Once the CPB is placed, or when the mine is flooded at its end-life, the backfill is subjected to underground water flows that may leach and spread the pollutants (mainly metals, sulphates, and acidity) which can contaminate the underground waters. The aim of this study is to characterize the leaching properties of CPB material and to identify the cement phases that ensure the metal fixation. A laboratory procedure was implemented consisting in the preparation of various paste backfill mixtures using sulphide-rich and sulphide-free materials, different types of binder, and mixing waters having various sulphates, copper and zinc concentrations. After 90-days curing time, the samples were submitted to column leaching tests. The CPB materials were flushed each day with 100 ml of deionised water during 100 days. The collected leachates were then analyzed for pH, Eh, electric conductivity and chemical composition. For some samples, a differential acid neutralisation test was carried out. The objective of this experiment is to dissolve selectively the hydrated phases by adding incremental acid concentrations. By analyzing punctually the equilibrated solutions, the composition of the dissolved phases and their metal content can be estimated. The leaching test results showed that the pH of the leachates remains alkaline during the test period. When the pH decreases under 10 (after 30 days), different behaviours in terms of chemical element release (especially for zinc) were observed. During the test, the oxidation of pyrite seems to slow down, probably due to passivating phenomena. Moreover, the differential acid neutralisation test results allowed a better understanding of the CPB mineralogy and the Zn-bearing cement phases. Zinc and the other metals seem to be fixed within a calcium-magnesium silicate.

May 1, 2007