Chloride assisted leaching of chalcocite by oxygenated sulphuric acid via Cu(II)–OH–Cl

The beneficial effect of the addition of sodium chloride upon the leaching kinetics of complex iron–nickel–copper sulphides at elevated temperatures and oxygen pressures has been widely reported since the late 1970s, but the role of chloride is still being investigated or debated. Previous researchers have considered chloride as: (i) a complexing agent for cuprous ions; (ii) a surfactant that disperses the molten sulphur and thus removes passivation of the mineral surface by elemental sulphur during pressure leaching; and (iii) a reagent which increases the surface area and the porosity of the insoluble product layer on the surface. A proper understanding of the role of chloride based on the leaching of individual sulphides of known composition in the absence of host minerals at low pulp densities would be useful for the development of chloride assisted sulphate leaching processes for complex sulphide ores, concentrates, and mattes. In the present study evidence for the formation of basic salts of Cu(II) and Fe(III) during leaching are presented. The published rate data are analysed for the leaching of copper from mono-sized chalcocite particles in oxygenated sulphuric acid solutions maintained at 85 °C, a temperature lower than the melting point of sulphur. The initial leaching follows a shrinking particle (sphere) model, and the apparent rate constants are first order with respect to the concentration of dissolved oxygen and chloride. The intrinsic rate constant for the surface reaction (0.2 m s⁻¹) is two orders of magnitude larger than the calculated mass transfer coefficient of oxygen (3 × 10⁻³ m s⁻¹). The proposed reaction mechanism considers the formation of an interim Cu(II)(OH)Cl⁰ species which facilitates the leaching process.     

Effect of water quality on the leaching of a low-grade copper sulfide ore

Copper extractions from a low-grade, ground copper sulfide ore (0.7% Cu) leached in three media were freshwater < seawater > double-strength seawater and pH 1.5 ⩾ pH 2; 84% extraction was achieved in pH 1.5 seawater in 28 days at 23 °C. Cu-oxide and carbonate dissolved completely and chalcocite was altered to secondary covellite, some of which persisted in all media for the duration of the 28-day experiment. Chalcopyrite and bornite were both oxidised more readily in saline water. Iron, sodium, potassium and sulfur (sulphate) concentrations in leach solutions diminished and the amounts of insoluble iron(III) reaction products increased with increased salinity and increased solution pH. While, overall, silicate dissolution was small, the amounts of poorly crystalline phases (both iron(III) and silica-rich phases) increased with increased salinity and were greater in pH 1.5 media. In the context of heap leaching, the increased amounts of secondary precipitates formed if saline water was used could result in lower extraction efficiency and the increased total dissolved solids, density and viscosity could result in increased energy costs for solution management at operations.The software package Geochemist’s Workbench was evaluated by modelling the synthetic seawater – pH 2 test. It was possible to predict the evolution of the solution composition, the main species and phase boundaries at the start and end of leaching, and the formation of three reaction products in accord with experimental data by applying the React sliding function.The tests were conducted using a pulverised ore sample to increase dissolution reaction kinetics, particularly for chalcopyrite. Future tests should be conducted using ore particle sizes appropriate to heap leaching. The copper distribution within particles indicated that the test ore may not be suited to heap leaching because the surface exposure of copper sulfide grains is limited. Therefore reactor designs better suited to smaller sized particles with/without pre-treatment should be considered.     

Leaching of low grade limonite and nontronite ores by fungi metabolic acids

This study was designed to investigate the nature of nickel and cobalt dissolution from limonite and nontronite ores. Leaching was achieved using 0.2–3 M of citric, lactic and malic acids. The effect of acid type, acid activity, oxygen reduction potential and pulp density on metal dissolution was studied systematically. The effect of leaching on the mineralogy of the ore was investigated by optical microscopy and by synchrotron based X-ray diffraction. From the experiments, it was shown that nickel and cobalt dissolution were dependent on acid activity, oxygen reduction potential, pulp density and mineralogy. The extent of metal dissolution was less affected by the type of acids. The trend of selectivity in limonite and nontronite are Co > Mn > Mg. Metal selectivity was intimately associated with the host minerals containing nickel and cobalt.     

Agglomeration and column leaching behaviour of goethitic and saprolitic nickel laterite ores

Processability of complex, low-grade nickel (Ni) laterite ores via heap leaching is very limited due to some intractable geotechnical and hydrological challenges such as poor heap porosity/permeability and structural stability. This work presents some investigations on laboratory batch drum agglomeration and continuous column leaching behaviour of saprolitic (SAP) and goethitic (G) Ni laterite ores as part of the quest for an effective ore pre-treatment process for enhanced heap leaching. As a focus, the effect of ore mineralogy/chemistry on the agglomeration and column leaching behaviour of −2 mm (crushed from −15 mm run-of-mine) G and SAP Ni laterite ores was examined. To produce ∼5–40 mm agglomerates in <15 min, the SAP ore required a higher H₂SO₄ (30 wt.%) binder dosage compared with the G ore, although both ores displayed substantially similar, coalescence-controlled agglomeration mechanism. The resulting G agglomerates were more robust than the SAP ones based upon their compressive strength and acidic solution soak test measurements. However, over 100 days of continuous column leaching, the structural stability of the SAP agglomerate bed was slightly greater than that of G agglomerates, reflecting a lesser slump of the former. The pregnant leach solution analysis revealed greater Ni/Co extraction rates from the SAP than the G agglomerates. Whilst the total mass of acid consumed per ton dry ore processed was greater for the SAP ore, the total kg acid per kg Ni extracted was markedly lower. Incongruent leaching of gangue minerals’ constituent elements (e.g., Fe, Mn, Mg, Al and Si) occurred and contributed significantly to the overall acid consumption. The findings show the relevance of agglomeration and column leaching tests for providing useful information for plant designing and optimization of Ni laterite heap leaching operations.     

Adsorption of copper sulphate on PGM-bearing ores and its influence on froth stability and flotation kinetics

Copper sulphate is used as an activator in the flotation of base metal sulphides as it promotes the interaction of collector molecules with mineral surfaces. It has been used as an activator in certain platinum group mineral (PGM) flotation operations in South Africa although the mechanisms by which improvements in flotation performance are achieved are not well understood. Some investigations have suggested these changes in flotation performance are due to changes in the froth phase rather than activation of minerals by true flotation in the pulp zone. In the present study, the effect of copper sulphate on froth stability was investigated on two PGM containing ores, namely Merensky and UG2 (Upper Group 2) ores from the Bushveld Complex of South Africa. Froth stability tests were conducted using a non-overflowing froth stability column. Zeta potential tests and ethylenediaminetetraacetic acid (EDTA) tests were used to confirm the adsorption of reagents onto pure minerals commonly found in the two ores. The results of full-scale UG2 concentrator on/off copper sulphate tests are also presented. The UG2 ore showed a substantial decrease in froth stability in the order of reagent addition: no reagents > copper > xanthate > copper + xanthate, while Merensky ore showed a slight decrease. It was shown through zeta potential measurements that copper species were to be found on plagioclase, chromite, talc and pyrrhotite surfaces and through EDTA extraction that this copper was in the form of almost equal amounts of Cu(OH)₂ and chemically reacted copper ions on the Merensky and UG2 ore surfaces. In certain cases, the presence of copper sulphate and xanthate substantially increased the recovery, and therefore the implied hydrophobicity, of pure minerals in a frothless microflotation device. It was, therefore, proposed that increases in hydrophobicity beyond an optimum contact angle for froth stability, were the cause of instabilities in the froth phase and these were found to impact grade and recovery in a full-scale concentrator. Differences in the extent of froth phase effects between the different ores can be attributed to differences in mineralogy.     

Innovative technology for processing saprolitic laterite ores by hydrochloric acid atmospheric pressure leaching

An innovative technology for processing saprolitic laterite ores from the Philippines by hydrochloric acid atmospheric leaching and spray hydrolysis is proposed. The factors that affect the hydrochloric acid atmospheric leaching of the laterite ores and spray hydrolysis of the atmospheric acid leach solution are investigated. Experimental results show that the leaching of Ni, Fe, and Mg is 98.9 wt%, 97.8 wt%, and 80.9 wt%, respectively, under optimal acid leaching conditions. The hydrolysis of Ni and Fe by the atmospheric acid leach solution approaches 100 wt% at the temperature range of 450–500 °C. Characterization results show that a serpentine mineral, nominally Mg₃Si₂O₅(OH)₄, is the major component and goethite, FeO(OH), is the minor one in the laterite ores. Treatment by hydrochloric acid atmospheric leaching breaks the mineral lattices of the laterite ores and makes amorphous silica the primary product in the atmospheric acid leach residue. The grade of Ni in the hydrolyzate increases to 4.55%. The hydrolyzate with high Ni content can be utilized for ferro-nickel production.     

Statistical analysis of bioleaching copper,cobalt and nickel from polymetalic concentrate originating from Kamoya deposit in the Democratic Republic of Congo

The effects of five parameters, temperature, pH, leaching duration, stirring speed and pulp density on the bioleaching of copper, cobalt and nickel from a polymetallic flotation concentrate were investigated. The leaching was carried out according to the L₂₅ (5⁵) orthogonal design. The optimal values of the parameters were determined using a Taguchi method through signal-to-noise analysis. ANOVA was applied to verify the individual contribution of each parameter and their degree of significance. It was found out that pulp density was the most influential factor on the bioleaching yield of the three metals altogether, followed by pH and temperature. For the copper bioleach, the following optimal parameters were determined: temperature – 37.5 °C, pH 1.6, leaching duration – 20 days, stirring speed – 350 rpm and pulp density – 7.5%. Verification experiments conducted according to these optimal parameters brought copper yield of 72.6%. For the cobalt bioleach, SEM observations of pure carrolite indicated a progressive bacterial colonization of mineral surface with time.     

Study of silver leaching with the thiosulfate–nitrite–copper alternative system: Effect of thiosulfate concentration and leaching temperature

Thiosulfate system is considered an interesting alternative leaching process for precious metals. Nevertheless, most of the literature published on these conventional thiosulfate leaching solutions has been focused on the use of ammonia and copper to generate the cupric tetraamine complex, which acts as a catalytic oxidant for silver. However, ammonia toxicity is also a detrimental issue in terms of the process sustainability. For that reason, thiosulfate–nitrite–copper solutions were studied as an alternative less toxic system for silver leaching.In this work, the effect of the thiosulfate concentration (0.07 M, 0.1 M and 0.15 M) and temperature (room temperature, 30, 35, 40 and 45 °C) on the metallic silver leaching kinetics is presented for the S₂O₃–NO₂–Cu system. The results show that the thiosulfate concentration plays an important role in the S₂O₃–NO₂–Cu–Ag system since it controls the silver leaching kinetics. On the other hand, an increase in temperature favors the silver recovery.Finally, the SEM–EDS analysis, the X-ray mapping and the X-ray diffractograms show that the solid silver particles are coated by a Cu, S and O layer for the 0.07 M and 0.1 M thiosulfate experiments, which is consistent with the formation of antlerite (Cu₃(SO₄)(OH)₄); while the 0.15 M thiosulfate scenario produced a layer composed only of Cu and S, revealing the formation of stromeyerite (CuAgS). The UV–Visible technique confirmed the in-situ generation of copper–ammonia complexes for the 0.07 M leaching condition; however, these complexes are not formed at the 0.15 M condition.     

Process development for the recovery of europium and yttrium from computer monitor screens

This work describes the development of a process for the recovery of Eu and Y from cathode ray tubes (CRTs) of discarded computer monitors with the proposition of a flow sheet for the metals dissolution. Amongst other elements, europium and yttrium are presented in the CRTs in quantities – 0.73 w/w% of Eu and 13.4 w/w% of Y – that make their recovery worthwhile. The process developed is comprised of the sample acid digestion with concentrated sulphuric acid followed by water dynamic leaching at room temperature. In the CRTs, yttrium is present as oxysulphide (Y₂O₂S) and europium is an associated element – Y₂O₂S:Eu³⁺ (red phosphor compound). During the sulphuric acid digestion, oxysulphide is converted into a trivalent Eu and Y sulphate, in solid form, with the liberation of H₂S. In the second step, metals are leached from the solid produced in the acid digestion step by dynamic leaching with water. This study indicates that a proportion of 1250 g of acid per kg of the sample is enough to convert Eu and Y oxysulphide into sulphate. After 15 min of acid digestion and 1.0 h of water leaching, a pregnant sulphuric liquor containing 17 g L⁻¹ Y and 0.71 g L⁻¹ Eu was obtained indicating yield recovery of Eu and Y of 96% and 98%, respectively. Both steps (acid digestion and water leaching) may be performed at room temperature.     

Mineral carbon storage in pre-treated ultramafic ores

Mineral carbon sequestration (MCS) is a type of carbon storage based on natural rock weathering processes where CO₂, dissolved in rainwater, reacts with alkaline minerals to form solid carbonates. Although MCS has advantages over other carbon storage techniques, an economic MCS process has not yet been developed. Two approaches were taken in this work to attempt to reduce the cost of MCS. The first approach was to use a waste material, serpentine waste from ultramafic nickel ore processing, as a feedstock. The second approach was to develop pre-treatments to increase the carbon storage capacity of the feedstock. Two pre-treatments were investigated in this work, including microwave pre-treatment and leaching with ligands at neutral to alkaline pH. The carbon uptake of ultramafic ores was found to increase with increasing microwave pre-treatment after a threshold heating time of 4 min was surpassed. A maximum carbon uptake of 18.3 g CO₂/100 g ore (corresponding to a carbonate conversion of 36.6%) was observed for microwave pre-treated ore. The increase in carbon uptake was attributed primarily to the conversion of serpentine to olivine in ultramafic ores that occurs as result of microwave pre-treatment. The effect of five different ligands (catechol, citrate, EDTA, oxalate and tiron) on the carbon uptake of ultramafic ores was investigated. Of the ligands tested, only catechol and tiron were found to both improve the leaching of magnesium from the ores and the quantity of CO₂ stored. A maximum carbon uptake of 9.7 g/100 g ore (corresponding to a carbonate conversion of 19.3%) was observed for ultramafic ore pre-leached and carbonated in tiron solution at pH 10. This is the first time ligands have been reported to improve the carbon uptake of mineral carbon sequestration feedstock. Although process optimization work was not conducted, both microwave pre-treatment and leaching with ligands at neutral to alkaline pH show promise as ways to lower the cost of MCS.     

Copper leaching from chalcopyrite concentrate in Cu(II)/Fe(III) chloride system

This study was conducted to develop a novel process for copper recovery from chalcopyrite by chloride leaching, simultaneous cuprous oxidation and cupric solvent extraction to transfer copper to a conventional sulfate electrowinning circuit, and hematite precipitation to reject iron. Copper leaching from chalcopyrite concentrate in ferric and cupric chloride system was investigated using a two-stage countercurrent leach circuit under a nitrogen atmosphere at 97 °C to minimize the concentrations of cupric and ferric ions in pregnant leach solution for subsequent copper solvent extraction while maintaining a maximum copper extraction. A high calcium chloride concentration (110–165 g/L) was used to maintain a high cuprous solubility and enhance copper leaching. With 3–4 h of leaching time for each stage, the copper extraction reached 99% or higher while that of iron was around 90%. With decreasing concentrate particle size from p80 of 26 to 15 μm, the copper extraction increased by about 0.2% while the iron extraction increased by about 2.0%. The concentration of Cu(II) + Fe(III) in the pregnant leach solution was able to be reduced to 0.04 M. When the cupric concentration fell below the above limiting value, the elemental sulfur present was reduced by cuprous ions to form copper sulfide, eventually stopping the leaching of copper. Under this condition, only iron was leached. A very small amount of sulfur (1.2–1.4%) was oxidized to sulfate, resulting in an increase from 3 to 9 g/L in HCl concentration. The extractions of trace metals (Cr, Pb, Ni, Ag and Zn) were 96–100%.     

Complex treatment of mining and metallurgical wastes for recovery of base metals

A complex process for the recovery of copper and zinc from mining and metallurgical wastes has been investigated and proposed. It includes sulfuric acid leaching of old pyrite flotation tailings to produce ferric containing leach solution; followed by ferric leaching of copper converter slag flotation tailings with the leach solution. A sample of old pyrite flotation tailings from the concentrator containing 0.36% of copper and 0.23% of zinc was leached with 10% sulfuric acid in the column. Recovery of copper and zinc reached 47.1% and 47.2%, respectively. The pregnant leach solutions contained 15.9 g/L of ferric iron. The subsequent ferric leaching of copper converter slag flotation tailings containing 0.53% copper and 2.77% zinc with the pregnant leach solution was conducted. The effects of various process parameters on the leaching dynamics of metals under batch conditions were investigated. Under the best conditions (temperature 70 °C, pulp density 30%, ferric iron concentration 15.9 g/L, initial pH of the pulp 0) the recovery of copper and zinc reached 79.6% and 43.7%, respectively. It was concluded that acid leaching of base metals from old pyrite flotation tailings with pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective and promising technique for the complex treatment of mining and metallurgical wastes.     

Copper leaching from waste electric cables by biohydrometallurgy

This study examines the leaching of copper from waste electric cables by chemical leaching and leaching catalysed by Acidithiobacillus ferrooxidans in terms of leaching kinetics and reagents consumption. Operational parameters such as the nature of the oxidant (Fe³⁺, O₂), the initial ferric iron concentration (0–10 g/L) and the temperature (21–50 °C) were identified to have an important influence on the degree of copper solubilisation. At optimal process conditions, copper extraction above 90% was achieved in both leaching systems, with a leaching duration of 1 day. The bacterial leaching system slightly outperformed the chemical one but the positive effect of regeneration of Fe³⁺ was limited. It appears that the Fe²⁺ bio-oxidation is not sufficiently optimised. Best results in terms of copper solubilisation kinetics were obtained for the abiotic test at 50 °C and for the biotic test at 35 °C. Moreover, the study showed that in same operating conditions, a lower acid consumption was recorded for the biotic test than for the abiotic test.     

Selective leaching of penalty elements from copper concentrates: A review

Custom copper smelters impose substantial financial penalties for the presence of deleterious impurity elements in copper concentrates and can outright reject concentrates which contain impurity elements in concentrations that exceed specified values. Hence, there is strong motivation to remove penalised impurity elements from copper concentrates at the mine site before shipping to custom smelters. A number of leach systems have been developed for the selective extraction of penalty elements from copper concentrates, including: alkaline sulphide leaching (ASL); hypochlorite leaching; dilute sulphuric acid leaching with aluminium sulphate; and combined pressure oxidation (POX) leaching with copper precipitation leaching. This paper reviews these four systems with emphasis on the leaching behaviour of penalty elements. ASL has previously been employed in industry for the selective extraction of As and Sb from tetrahedrite-rich copper concentrates. Sodium sulphide solution leaches As, Sb, and Hg from a large range of minerals, however, does not leach arsenopyrite, a mineral which often contains a significant portion of the total As in copper concentrates. Hypochlorite leaching extracts As associated with enargite minerals. This leach system benefits from superior rates of As extraction when compared with ASL, and for this reason, has gained recent interest within the research community. Two major issues have been identified with hypochlorite leaching of copper concentrates. These are poor reagent selectivity towards As-bearing minerals and high levels of hypochlorite consumption. Unless these two issues are resolved it is unlikely that hypochlorite leaching will be employed in commercial processes. Dilute sulphuric acid leaching with aluminium sulphate is used to extract F associated with fluorite. This leach system also extracts F associated with apatite and chlorite. Laboratory-scale experiments and extensive operating experience have indicated that fluorite can be substantially leached from copper concentrates without addition of aluminium sulphate provided that the concentration of sulphuric acid in the leach solution is sufficiently high (at least 40 g L⁻¹). POX/copper precipitation leach systems have potential to extract a large number of penalty elements from copper sulphide concentrates while simultaneously upgrading the concentration of copper in the concentrate. Two patented POX/copper precipitation leach processes have been specifically developed for the deportment of penalty elements. These two processes are reviewed in detail.     

Leaching behaviour of natural and heat-treated brannerite-containing uranium ores in sulphate solutions with iron(III)

Uranium leaching tests were conducted on two naturally occurring, highly metamict brannerite ores from the Crockers Well and Roxby Downs deposits, South Australia. The ores were leached over a range of temperatures and Fe^(III) and H₂SO₄ concentrations. As well, samples of the ores were calcined at 1200 °C in air to investigate the effect of thermally induced recrystallisation on uranium dissolution. For the unheated samples, a maximum of ∼80% U dissolution was obtained using an Fe^(III) concentration of 12 g/L, an acid concentration of 150 g/L H₂SO₄ and a temperature of 95 °C. The heat treated samples performed poorly under identical conditions, with maximum uranium dissolution of <10% recorded. High uranium dissolution from natural brannerite can be achieved providing; (i) acid strength, oxidant strength and temperatures are maintained at elevated levels (compared to those traditionally used for uraninite leaching), and, (ii) the brannerite has not undergone any significant recrystallisation (e.g. through metamorphism).     

The effect of mechanical–physical pretreatment on hydrometallurgical extraction of copper and tin in residue from printed circuit boards from used consumer equipment

This work is focused on studying mechanical–physical pretreatment of printed circuit boards from used consumer equipment followed by extraction of copper and tin from residue fractions by leaching in hydrochloric acid solutions. Mechanical–physical pretreatment was realized in three different procedures. Key processes were electro-dynamic separation, cross-flow air sifter separation and air table separation, respectively. Leaching experiments were carried out in 1 M and 2 M HCl at 80 °C. The results show that cross-flow air sifting leads to the highest accumulation of non-ferrous metals in a residue fraction. From this fraction, the highest extraction of tin with minimal copper extraction was achieved.     

Recovery of zinc and cadmium from industrial waste by leaching/cementation

Metallic zinc production from sulfide zinc ore is comprised by the stages of ore concentration, roasting, leaching, liquor purification, electrolysis and melting. During the leaching stage with sulfuric acid, other metals present in the ore in addition to zinc are also leached. The sulfuric liquor obtained in the leaching step is purified through impurities cementation. This step produces a residue with a high content of zinc, cadmium and copper, in addition to lead, cobalt and nickel. This paper describes the study of selective dissolution of zinc and cadmium present in the residue, followed by the segregation of those metals by cementation. The actual sulfuric solution, depleted from the electrolysis stage of metallic zinc production, was used as leaching agent. Once the leaching process variables were optimized, a liquor containing 141 g/L Zn, 53 g/L Cd, 0.002 g/L Cu, 0.01 g/L Co and 0.003 g/L Ni was obtained from a residue containing 30 wt.% Zn, 26 wt.% Cd, 7 wt.% Cu, 0.35 wt.% Co and 0.32 wt.% Ni. The residue mass reduction exceeded 80 wt.%. Cementation studies investigated the influence of temperature, reaction time, zinc concentration in feeding solution, pH of feeding solution and metallic zinc excess. After that such variables were optimized, more than 99.9% of cadmium present in liquor was recovered in the form of metallic cadmium with 97 wt.% purity. A filtrate (ZnSO₄ solution) containing 150 g/L Zn and 0.005 g/L Cd capable of feeding the electrolysis zinc stage was also obtained.     

Atmospheric acid leaching of siliceous goethitic Ni laterite ore: Effect of solid loading and temperature

In this study, atmospheric acid leaching behaviour of siliceous goethitic nickel (Ni) laterite ore is investigated. Specifically, the effect of −200 μm feed solid loading (30 vs. 45 wt.%) and temperature (70 vs. 90 °C) on leach kinetics, acid consumption capacity and Ni and cobalt (Co) extraction was studied under isothermal, batch (4 h) leaching conditions at pH 1. Incongruent leaching was observed for constituent elements reflecting slow but steady release of value (Ni and Co) and some of gangue metals such as Fe, Mg and Al accompanied by faster and sharp release of Na and Si. Higher temperature and lower pulp solid loading, both led to a 40–50% increase in overall Ni and/or Co extraction and higher acid consumption. At 70 °C and 45 wt.% solid loading, Ni/Co extraction after 4 h was the lowest (∼14/16%) whilst the highest extraction (∼67/56%) was observed at 90 °C and 30 wt.% solid loading. Temperature appeared to have dramatic influence on Ni/Co and other impurity metals’ extractions revealing the chemical reaction controlled nature of the leaching. Higher solid loading and longer leaching time also both slowed down the leach kinetics. A two-stage chemical reactions-controlled leaching mechanism involving a faster initial leaching kinetics followed by a slower leaching at lower rate constants and higher activation energies was established for release of Ni, Co, Fe and Mg. The mechanism reflects the fast leaching of reactive host mineral phases (e.g., clays and Mg–silicates) during first 30 min followed by slow leaching of more refractory mineral phases (e.g., goethite and quartz) during the rest of leaching period. The findings provide a greater understanding for enhanced atmospheric acid leaching process of siliceous goethitic laterite ores.     

Textural,mineralogical and chemical characteristics of copper reverb furnace smelter slag of the Okiep Copper District,South Africa

The Okiep Copper District in South Africa has produced more than 110 million tons at a grade of 1.71% Cu from several small mafic ore bodies. The ore was smelted on site and generated ∼5 mt of slag. During the life of mine attempts to recover copper from the slag by flotation had limited success. After mine closure the challenge of environmental rehabilitation and the possible disposal of the slag, triggered a reinvestigation into the viability of slag as a copper resource. Characterisation of the slag as a contribution to the potential copper recovery is the objective of this study.The slags are hard, vitreous with a matrix of Si–Fe–Al–Mg–Ca glass and laths of Mg–Fe–olivine, Fe–Mg–orthopyroxene and minor Cr-spinel. Copper grade varies between 0.11% and 0.42% with minor nickel, cobalt, molybdenum, zinc and tungsten. All economic elements are hosted by disseminated spheroidal prills which consist mainly of the copper sulphides bornite, chalcocite, covellite and chalcopyrite with exsolved sulphide phases of the minor base metals as well as rhenium and silver. Prills consisting of metallic copper and alloys are minor constituents. Prill diameter is highly variable with most in the 40–60 μm range and the historically poor copper recovery is attributed to the small prill size. Crushing of slag to −45 μm as opposed to the previous −75 μm should significantly increase sulphide liberation and recovery of copper and minor base metal sulphides by conventional flotation.Provided the operation is economically viable, redistribution of the processed slag to environmentally acceptable sites will resolve the present pollution and rehabilitation challenge related to the dumps in the Okiep Copper District. The operation will also have a positive socio-economic impact on this poverty-stricken part of South Africa.     

Metals recovery from galvanic sludge by sulfate roasting and thiosulfate leaching

The purpose of this work is the selective recovery of Au, Ag, Cu, and Zn from two types of galvanic sludge using a mixed process of sulfate roasting and sodium thiosulfate leaching. In the experiments, the sludge was mixed with a sulfate promoter (sulfur, iron sulfate, or pyrite) and treated by pyrometallurgical processes at temperatures up to 750 °C. At this stage, this agent is thermally oxidized, turning the furnace atmosphere into a reducing one and the metallic oxides into water-soluble sulfates. Afterward, the sulfates can be treated by leaching with water for recovery of Ag, Cu, and Zn. The gold does not form sulfates in this reaction and was recovered through a second leaching stage using sodium thiosulfate, an effective reagent and less harmful to the environment than cyanide. Different parameters such as the sulfate promoter that achieves the highest recovery of metals, the proportion of galvanic sludge to sulfating agent, the temperature, the heating time in the oven, and the leaching time were evaluated. Additionally, a comparison of gold recovery using cyanide versus sodium thiosulfate was performed. The configuration that showed the best metal recovery included a 1:0.4 ratio of sludge to sulfur, an oven temperature of 550 °C, a roasting time of 90 min, and a water leaching time of 15 min. Using these parameters, recovery rates of 80% of the silver, 63% of the copper, and 73% of the Zn were obtained. The sodium thiosulfate leaching resulted in a recovery of 77% of the Au, close to the values obtained using cyanide.