Solubility prediction of malachite in aqueous ammoniacal ammonium chloride solutions at 25 °C

The solubility of malachite in the presence of ammonia, ammonium chloride and their mixed solution is calculated by a geochemical modeling code and is measured in a series of dissolution experiments using synthetic malachite at 25 °C. The simulated results show a good agreement with the experimental data gained at 25 °C. The predicted and experimental results indicate that the precipitate CuO limits the copper solubility in aqueous ammonia and Cu(OH)_1.5Cl_0.5 in aqueous ammonium chloride. For a mixed solution containing ammonia and ammonium chloride, highest copper solubility can be achieved by adjusting the [NH₃]/[NH₄Cl] ratio to about 2:1. The thermodynamic model presented rationalises the interactions between the different components and predicts the influence of changes in the concentration of ammonia and ammonium chloride on the copper solubility of malachite.     

Leaching of vanadium from carbonaceous shale

The leaching of vanadium from carbonaceous shale using dilute H₂SO₄ was investigated, and the mechanism of leaching determined. The results showed that higher leaching efficiency of vanadium was obtained by increasing initial concentration of H₂SO₄, raising leaching temperature and prolonging leaching time. Addition of ammonium fluoride also enhanced recovery. A recovery of 92% was obtained using a liquid to solid ratio 4:1, initial H₂SO₄ concentration 18%, NH₄F addition 4.8 wt.% of carbonaceous shale, leaching temperature 95 °C and contact time 8 h; the recovery was only 56% without NH₄F. The presence of NH₄F enhanced the leaching of vanadiferous mica.     

The leaching of copper from sulphur activated chalcopyrite with cupric sulphate in nitrile-water mixtures

If chalcopyrite is roasted with sulphur at 400–450°C pyrite and idaite or bornite are produced. Bornite plus pyrite are also prepared by roasting a 1:1 mixture of chalcopyrite and covellite. These copper-iron sulphides were leached with acidified aqueous cupric sulphate solutions containing acetonitrile or hydracrylonitrile and the results are compared with leaching with acidified cupric chloride in brine. The nitrile route has the advantage of a less corrosive sulphate medium for subsequent copper recovery processes.Bornite appears to be the most attractive product from the roasting of sulphur and chalcopyrite because much of its copper can be readily leached. Iron reports to the solution only in the latter stages of extraction. Up to 80% of the copper in this bornite is leached with CuSO₄/RCN/H₂O at 60°C. Copper is recovered from the resulting cuprous sulphate solution by electrowinning with inert anode. The products are copper cathodes and cupric sulphate, which is recycled. The leach residue may be used to reactivate further chalcopyrite or is leached of its copper by established routes.     

Arsenic removal from copper ores and concentrates through alkaline leaching in NaHS media

Removal of arsenic impurity in ores and concentrates containing copper (Cu) through alkaline leaching in NaHS media was investigated in this work. Samples containing Cu from 10 to 40 wt.% and arsenic from 0.8 to 14 wt.% with enargite (Cu₃AsS₄) as main arsenic bearing mineral were used as starting materials and all leaching tests were conducted at 80 °C under normal atmospheric pressure. Solution and/or slurry potential and pH were maintained consistently below − 500 mV (SHE) and above 12.5 respectively with the addition of NaHS and NaOH, creating a reducing environment for arsenic dissolution and conversion of Cu₃AsS₄ to Cu₂S. Pulp density ranged from 100 to 1000 g/L, NaHS and NaOH reagents were added at 50–200 g/L each and leaching time varied from 10 min to 10 h.Characterization of solid samples (original and leach residue) by XRD and XRF analyses and chemical analysis of both solid and solution samples by ICP analysis showed that Cu₃AsS₄ in the starting material was completely decomposed or transformed to Cu₂S and arsenic released into solution as As (III)/As³⁺ ions (Na₃AsS₃). Over 90% of arsenic in the starting materials was removed within 1–3 h for materials with arsenic content from 1 to 4 wt.% and within 3–6 h for materials with arsenic content over 4–10 wt.%. Dissolution and analysis of leach residues obtained after leaching by ICP indicated that arsenic in the starting materials has been reduced in all cases to below 0.5 wt.%. In all test conditions dissolution of Cu and Fe into solution was not detected, indicating selective leaching of arsenic. NaHS application for removal of arsenic in Cu-ores and/or concentrates was demonstrated in this work and further research is in progress to develop a process to include treatment of arsenic leached into solution.     

Stability of 3CaO · Al2O3 · 6H2O in KOH + K2CO3 + H2O system for chromate production

In the novel clean chromate production process, the alkali liquor recycled to decompose the chromite ore mainly consists of KOH and K₂CO₃ which accumulates aluminium impurity and affects the quality of the product greatly. Aluminium impurity can be removed by adding CaO to precipitate as 3CaO · Al₂O₃ · 6H₂O (C₃AH₆). A study of the effects of various parameters, such as KOH concentration, K₂CO₃ concentration and temperature on the behaviour of C₃AH₆ show that C₃AH₆ is decomposed to 3CaO · Al₂O₃ · CaCO₃ · 11H₂O and CaCO₃ below 150 g L^− 1 KOH, and decomposed to Ca(OH)₂ above 150 g L^− 1 KOH. With K₂CO₃ concentration increasing, C₃AH₆ decomposes significantly, which results in more CaCO₃ or 3CaO · Al₂O₃ · CaCO₃ · 11H₂O produced. Temperature has a large positive effect on the decomposition of C₃AH₆ at 45 g L^− 1 KOH but has no significant effect at 150 g L^− 1 KOH. The optimal condition for removing aluminium impurity in the KOH + K₂CO₃ + H₂O system is 150 g L^− 1 KOH, 50 g L^− 1 K₂CO₃ and 80 °C.     

Extraction of vanadium from black shale using pressure acid leaching

The extraction of vanadium from black shale was attempted using pressure acid leaching. The effects of the several parameters which included reaction time, concentration of sulfuric acid, leaching temperature, liquid to solid ratio and concentration of additive (FeSO₄) upon leaching efficiency of vanadium were investigated and a two-step counter-current leaching approach was developed. The results showed that the leaching efficiency of vanadium in the two-step process could reach above 90%. Vanadium was effectively separated and enriched by solvent extraction after leachate pretreatments, including the reduction of Fe³⁺ and adjustment of pH value. The extraction and stripping yields of vanadium were both > 98%. Ammonia was added to a stripping liquor to precipitate vanadium and then the ammonium poly-vanadate produced was calcined at 550 °C for 3 h to produce the high purity V₂O₅ powder. The overall yield of vanadium through all process stages was about 85%.     

Pressure oxidation leaching of chalcopyrite. Part I. Comparison of high and low temperature reaction kinetics and products

The kinetics and products from the pressure oxidation of a chalcopyrite concentrate are compared under a range of reaction conditions promoted by various companies. The reaction conditions compared in this article, Part I, are referred to as the Phelps Dodge-Placer Dome and Activox® processes. The medium temperature processing of the concentrate will be reported in Part II.Experiments were conducted with the same concentrate over a temperature range of 108–220 °C with different salt and acid additions to compare the kinetics and recovery of copper, the speciation of sulfur and the deportment of iron-containing and other phases in the leach residues. The aim was to improve understanding of the mechanism and practical issues for the competing processes and to provide background knowledge often not available in the public domain.The chalcopyrite concentrate was found by Quantitative X-ray Diffraction (QXRD) analysis to contain about 80% chalcopyrite, 10% quartz, 6% pyrite and 2.5% talc and 1.5% clinochlore. It was demonstrated that greater than 94% of the copper could be extracted from the concentrate using either the Phelps Dodge-Placer Dome or Activox® process within 30 min. The extraction of the residual copper was strongly influenced by the presence of elemental sulfur.About 80–90% oxidation of sulfide to elemental sulfur occurred at 108 °C and was enhanced by the presence of the chloride ion. Above 180 °C there was complete oxidation to sulfate. However, in the presence of added chloride ion the rate of sulfate formation decreased.QXRD was employed to examine the leach residues. Iron was leached and re-precipitated forming a number of different phases depending upon the process temperature, acidity and salinity. At low temperature, in the presence of chloride, akaganéite was formed together with an uncharacterised amorphous hydrated iron oxide. Hematite formation was favoured at temperatures ≥ 150 °C, low acidity and low salinity; basic ferric sulfate formed at high temperature (220 °C), high acidity and low salinity. Goethite formation was favoured at ≤ 150 °C by low acidity and low salinity. Jarosite was formed at all temperatures under conditions of moderate to high acidity and its formation was enhanced in the presence of sodium ions.Several basic copper salts including atacamite (Cu₂(OH)₃Cl) and antlerite (CuSO₄·2Cu(OH)₂) were precipitated at 108 °C at low acidity, typically at pH > 2.8. Atacamite formed initially when the sulfate concentration was low but dissolved and the copper was re-precipitated to form antlerite as the sulfate (and copper) concentrations increased.     

Effect of phosphate on lead removal during a copper recycling process from wastes using ammoniacal chloride solution

Techniques for the removal of lead have been studied in order to develop a hydrometallurgical copper recycling process consisting of copper leaching from wastes using an ammoniacal chloride solution and subsequent copper electrowinning. The solubility of Pb(II) in the ammoniacal chloride solution increased with ammonia concentration; this was attributable to the formation of a lead ammine complex. The lead dissolution was depressed from the order of 10^− 3 M to the order or 10^− 5 M by the addition of phosphate into the leaching solution because of the precipitation of chloropyromorphite (Pb₅(PO₄)₃Cl), while no significant effect was observed by the addition of carbonate. Linear sweep voltammetry and potentiostatic electrolysis in the solution containing Pb(II) revealed that lead was deposited during the copper electrowinning, even in the potential region more positive than the equilibrium redox potential for the Pb/Pb(II) couple on the lead electrode, because of the alloy formation with copper. In a galvanostatic electrolysis, however, the lead content at the electrodeposited copper cathode was found to be lower than 5 ppm at the current density range of 125–400 A/m², when the Pb(II) concentration in the electrolyte was 5 × 10^− 5 M. Since this Pb(II) concentration was achieved by the phosphate addition, these results indicated the effectiveness of phosphate for lead removal in the copper recycling process using the ammoniacal chloride solution.     

Precipitation of sodium silicofluoride (Na2SiF6) and cryolite (Na3AlF6) from HF/HCl leach liquors of alumino-silicates

The HF + HCl leach liquor generated from the dissolution of silica, alumina and silicate gangue minerals in a low-grade molybdenite concentrate contains H₂SiF₆ and H₃AlF₆. Studies were conducted to recover the two valuable fluorides as Na₂SiF₆ and Na₃AlF₆ (synthetic cryolite) by precipitation with Na₂CO₃ from the leach liquor. An initial investigation was carried out to determine the precipitation conditions for Na₂SiF₆ and Na₃AlF₆ from their individual acid solutions. Subsequently, the conditions were determined for the selective precipitation of the two fluorides from a synthetic mixed acid solution similar to the leach liquor. When the acid solution was neutralized with 3 mol/L Na₂CO₃, Na₂SiF₆ precipitated first at pH 1.35 whilst Na₃AlF₆ required an increase in pH above 2.2 before it precipitated. Maximum recovery of the two fluorides was best achieved at about 50 °C. A similar trend was observed for the precipitation of Na₂SiF₆ and Na₃AlF₆ from the leach liquor of molybdenite upgrading. Phases of precipitated fluorides were identified by XRD and surface morphology by SEM. The purity of the Na₂SiF₆ precipitate was 99.5% whereas Na₃AlF₆ was contaminated with Na₃FeF₆.     

Speciation of vanadium (V) extracted from acidic sulfate media by trioctylamine in n-dodecane modified with 1-tridecanol

The speciation of vanadium (V) extracted from acidic sulfate media by protonated trioctylamine in n-dodecane modified with 5% (wt) 1-tridecanol has been investigated by Fourier transformed infrared spectroscopy (FTIR) and ⁵¹V nuclear magnetic resonance spectroscopy (⁵¹V NMR). In aqueous sulfate solutions, vanadium (V) exists both as VO₂⁺ and VO₂SO₄⁻ ions. The FTIR spectra of 0.2 mol kg^− 1 protonated trioctylamine in n-dodecane modified with 5% (wt) 1-tridecanol, loaded with various concentrations of vanadium (V) by extraction from 1 mol kg^− 1 H₂SO₄, indicate that vanadium (V) exists in organic phases as polyvanadates, likely as decavanadates. The condensed nature of the extracted form of vanadium (V) was neither confirmed nor precluded by ⁵¹V NMR as the micellar structure of these organic phases imposes local conditions which allow the transformation of VO₂⁺ and VO₂SO₄⁻ into polyvanadates, but also modify the chemical shifts compared to the ones observed in bulk aqueous solutions for mononuclear and polynuclear vanadium (V) species.     

Chloride leaching of chalcopyrite

Two new process flowsheets have been developed which combine chloride leaching of copper from chalcopyrite with solvent extraction, to selectively transfer copper to a conventional sulfate electrowinning circuit. Chloride leaching with copper(II) as oxidant offers significant advantages for copper including increased solubility and increased rates of leaching. Both process flowsheets were similarly designed with a two stage counter-current leach but differ with respect to iron deportment. The goethite model flowsheet includes sparging of air or oxygen to the second leach stage to aid precipitation of iron as goethite (FeOOH). The hematite model flowsheet precipitates iron as hematite (Fe₂O₃) downstream from the leach in a dedicated autoclave. A mass balance has been completed for both process flowsheets and this determined the concentrations of copper and iron species in feed liquor returning to the leach following copper solvent extraction.The optimum leach extraction conditions were determined by varying grind size, temperature and residence time for both leach model scenarios. Leach tests were conducted using a chalcopyrite concentrate from Antamina in northern Peru, which contains a low to moderate amount of gangue material. The hematite model was also examined using a Rosario concentrate from Chile which contained chalcocite in addition to chalcopyrite and significant pyrite. Leach tests based on the hematite model were successful in achieving copper extractions > 95% in 4–6 h at 95 °C after fine grinding the concentrate (P₉₀ = 41 μm). However, copper extraction exceeded 99% from the finely ground Rosario concentrate (P₉₀ = 37 μm). In the goethite model leach tests, 89% copper extraction was achieved under optimum conditions in the atmospheric conditions tested.     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

Luminescence modification of Eu3+-activated molybdate phosphor prepared via co-precipitation

Eu-activated CaMoO4 phosphors were co-precipitated in an aqueous solution, and NH3·H2O, NH4HCO3 and (NH2)2CO as pre-cipitating aid agents were compared based on the morphology and particle size distribution of the phosphor samples. Sm3+ as sensitizer ion was investigated on the luminescence of CaMoO4:Eu, and it could strengthen the 406 nm absorption of this phosphor. At last, the scheelite CaMoO4:Eu and wolframite ZnMoO4:Eu were selected to compare their host absorption. The result showed that the scheelite molybdate host exhibited stronger UV absorption than wolframite one.     

Copper recovery from chalcopyrite concentrates by the BRISA process

The technical viability of the BRISA process (Biolixiviación Rápida Indirecta con Separación de Acciones: Fast Indirect Bioleaching with Actions Separation) for the copper recovery from chalcopyrite concentrates has been proved. Two copper concentrates (with a copper content of 8.9 and 9.9 wt.%) with chalcopyrite as the dominant copper mineral have been leached with ferric sulphate at 12 g/L of ferric iron and pH 1.25 in agitated reactors using silver as a catalyst. Effects of temperature, amount of catalyst and catalyst addition time have been investigated. Small amounts of catalyst (from 0.5 to 2 mg Ag/g concentrate) were required to achieve high copper extractions (>95%) from concentrates at 70 °C and 8–10 h leaching. Liquors generated in the chemical leaching were biooxidized for ferrous iron oxidation and ferric regeneration with a mixed culture of ferrooxidant bacteria. No inhibition effect inherent in the liquor composition was detected. The silver added as a catalyst remained in the solid residue, and it was never detected in solution. The recovery of silver may be achieved by leaching the leach residue in an acid-brine medium with 200 g/L of NaCl and either hydrochloric or sulphuric acid, provided that elemental sulphur has been previously removed by steam hot filtration. The effect of variables such as temperature, NaCl concentration, type of acid and acidity–pulp density relationship on the silver extraction from an elemental sulphur-free residue has been examined. It is possible to obtain total recovery of the silver added as a catalyst plus 75% of the silver originally present in concentrate B (44 mg/kg) by leaching a leach residue with a 200 g/L NaCl–0.5 M H₂SO₄ medium at 90 °C and 10 wt.% of pulp density in two stages of 2 h each. The incorporation of silver catalysis to the BRISA process allows a technology based on bioleaching capable of processing chalcopyrite concentrates with rapid kinetics.     

Selective silver electroseparation from ammoniacal thiosulfate leaching solutions using a rotating cylinder electrode reactor (RCE)

A systematic electrochemical study at controlled potential, using a rotating cylinder electrode reactor, was carried out on a synthetic solution with composition 0.2mol/L Na₂S₂O₃, 0.05mol/L CuSO₄, 0.025mol/L EDTA, 0.6 mo/L NH₃, 1 × 10^{− 3}mol/L AgNO₃, which simulates actual leaching solution, to determine the conditions which permit the selective silver electroseparation with a minimum interference from the copper ions. The presence of 10 ppm of gold (5 × 10^{− 4}M Au(I)) was found to enhance the gold–silver alloy deposition, with negligible effect on the copper ions. Those results were compared to that obtained with solutions originating from a previously characterized sulfide concentrate leach, in order to determine the influence of other impurities present in the pregnant solution, on the selective silver electrodeposition, obtaining averaging current efficiencies of 80%, with 90% of silver in the formed deposit and less than 5% of both, copper and lead, with only minor quantities of iron and zinc. Additionally, it was shown that the resulting solution could be recycled back to the leach stage, obtaining similar performances to those produced by the fresh solution, implying that it was not significantly altered in its composition, by the electroseparation process.     

Leaching of sulfide copper ore in a NaCl–H2SO4–O2 media with acid pre-treatment

A study was made of the leaching of a sulfide copper ore in a NaCl–H₂SO₄–O₂ media after pre-treatment by agglomeration with H₂SO_4(conc) and NaCl. The leaching variables evaluated included the amount of NaCl to be employed, the percentage of solids in the leaching solution, particle size of the raw mineral to be leached, and the preferable method of agitation in the leaching system. Mineralogical characterization of the material to be leached included analysis of the raw ore and of the leached ore residue using reflected-light microscopy, X-ray diffraction, and scanning electron microscopy. The soluble species included djurleite and digenite. The most important parameters in the leaching process proved to be particle size and type of agitation. A total percentage of copper extraction of 70% was achieved using mechanical stirring, which increased to 78% when using compressed air agitation. The best extraction of the copper was achieved when leaching with 3 g/L of chloride, room temperature of 20 °C, and when all particles were < 1.65 mm in diameter.     

Leaching of zinc sulfide in alkaline solution via chemical conversion with lead carbonate

A novel hydrometallurgical process for recovering Zn from ZnS in alkaline solution via chemical conversion with PbCO₃ was developed in this work. The S originally present in ZnS can be converted into PbS, while the Zn can be converted into Na₂Zn(OH)₄ in the alkaline solution in the presence of PbCO₃. And then, the Pb in PbS deposited in the leach residues can be converted into PbCO₃ again in the Na₂CO₃ solution. It was found that over 90% of Zn can be extracted from ZnS when the leaching process is operated in 6 mol/L NaOH solution at 90 °C with PbCO₃ as additive, and over 95% of PbS in the leach residues can be converted into PbCO₃ by stirring the leach residues in Na₂CO₃ solution with air bubble at a temperature of 80 °C. The leaching solution can be used to produce metallic zinc powder by electrowinning after chemical separation of impurities.     

Electrochemical modeling and study of copper deposition from concentrated ammoniacal sulfate solutions

A fundamental investigation of the electrolytic deposition of copper from concentrated aqueous ammoniacal solutions has been carried out based on the thermodynamic analysis of the system Cu–NH₃–H₂O. The speciation of copper vs. pH and redox potential was modeled in high ionic strength solutions, in which the activity coefficients of the system species were estimated according to the Modified Bromley's Methodology. The electrochemical behavior of the redox system Cu(0)/Cu(I)/Cu(II) in concentrated aqueous ammoniacal solutions was studied at pH = 9.5 and the cathodic reactions in these solutions were determined. It was found that metallic copper was formed under strongly reductive redox conditions, while under mildly reductive to mildly oxidative conditions the cuprous di-ammine complex species dominate. Under highly oxidative conditions the cupric tetra-ammine complex species predominated. According to the theory and results, the cathodic deposition of copper from concentrated aqueous ammoniacal solutions proceeds in a two-step reduction mechanism. The cupric ammine species are first reduced to cuprous di-ammine, which in turn is reduced to metallic copper. The electrochemical experiments revealed that copper deposition over time follows a sigmoid-type curve, verifying the two-step mechanism. The main feature of these sigmoid curves was the presence of an induction period with negligible copper deposition, followed by an acceleration period where the copper deposition rate gradually increased. By increasing the applied cell voltage, the induction period was significantly reduced or disappeared.     

Leaching of enargite in H2SO4–NaCl–O2 media

This paper discusses the leaching of enargite (Cu₃AsS₄) in sulfuric acid–sodium chloride solution using oxygen as oxidant at atmospheric pressure. The dissolution of arsenic from enargite in this medium proceeds with elemental sulfur as a reaction product according to:2Cu₃AsS₄ + 6H₂SO₄ + 5.5O₂ → 6CuSO₄ + 2H₃AsO₄ + 8S° + 3H₂OThe dissolution rate of arsenic was found to be very slow. About 6% of arsenic was dissolved in 7 h of leaching at 100 °C in a solution containing 0.25 M H₂SO₄ and 1.5 M NaCl under a flow of oxygen of 0.3 l/min.The kinetics of arsenic dissolution is well represented by a shrinking core model for spherical particles controlled by surface reaction. An apparent energy of activation of 65 kJ/mol was obtained for the temperature range 80 to 100 °C.     

Mineralogical characterization of a copper anode and the anode slimes from the la caridad copper refinery of mexicana de cobre

A detailed mineralogical study was carried out to characterize a copper anode, the anode-face slimes, the slimes on the bottom of the refining cell, and the autoclave-leached slimes from the La Caridad refinery of Mexicana de Cobre. The objective was to identify possible Pb-Sb-Bi and As-Sb-Bi interactions that could control the Sb and Bi concentrations of the electrolyte. Although some Pb, As, Sb, and Bi can be found in solid solution in the copper crystals of the anode, these elements are mostly present as Cu-Pb-As oxide and Cu-Pb-As-Sb-Bi oxide inclusions at the grain boundaries. During electrorefining, the Pb, As, Sb, and Bi in solid solution dissolve. Part of the Pb, As, Sb, and Bi in the oxide inclusions also dissolves, but part reacts in situ to form PbSO₄ and Pb₅(AsO₄)₃(OH,Cl). Some of the dissolved elements reprecipitate as PbSO₄, SbAsO₄, Sb-As oxide, Sb-As-Bi oxide, Pb₅(AsO₄)₃(OH,Cl), and an oxidate phase of mainly Cu-Ag-AsO₄-SO₄ composition. Thus, high As contents facilitate the precipitation of Sb and Bi from the electrolyte. Although Pb-Sb oxide and Pb-Bi oxide species were only rarely detected, a high Pb content in the anode may retard the dissolution of the Cu-Pb-As-Sb-Bi oxide inclusions, thereby retaining some Sb and Bi in the raw anode slimes. Autoclave leaching dissolves part of As, Sb, and Bi, but the SbAsO₄ and Sb-As-Bi oxide species remain in the leach residue. The Pb is converted almost entirely to PbSO₄, which is present as subhedral crystals in the autoclave leach residue.