Hydrometallurgical extraction of zinc from Jordanian electric arc furnace dust

Zinc was extracted in a Jordanian Steel Plant using an electric arc furnace dust. Sulfuric, nitric and hydrochloric acids were used at different concentrations to recover zinc from dust particles. The highest zinc extraction was obtained at low acid concentration of less than 1 mol/L in the order of H₂SO₄ followed by HNO₃ and then HCl. The kinetics of zinc extraction using H₂SO₄ showed a maximum zinc recovery of 72% obtained by using 0.1 mol/L acid concentration, 900 rpm agitation speed and 50 °C.     

The cyanidation of silver metal: Review of kinetics and reaction mechanism

Published rate data are analysed for the chemical and electrochemical dissolution of silver metal from rotating discs in aerated/oxygenated cyanide solutions at ≈25 °C, pH 11 and different partial pressures of oxygen. The current status of the reaction mechanism is also reviewed. Speciation analysis of 0.01 mM silver(I) in 1–100 mM cyanide solutions shows that Ag(CN)₂⁻ is the predominant complex (50%) at cyanide concentrations < 20 mM. However, at higher cyanide concentrations, Ag(CN)₃²⁻ (up to 60%) and Ag(CN)₄³⁻ (up to 10%) can be formed. Thus, it is important to consider a silver(I) : cyanide ion ratio of 2 or 3 in the Levich equation to calculate the diffusion coefficient of cyanide ion. Likewise, it is important to consider a silver(I) : oxygen ratio of 1 : 0.5 to calculate the diffusion coefficient of oxygen. This indicates the reduction of oxygen to hydrogen peroxide in the surface reaction. Analysis of exchange current density data for silver oxidation as a function of cyanide concentration shows the involvement of between 1 and 2 cyanide ions in the surface reaction. The limiting rate of silver dissolution at high cyanide concentrations (2.5 × 10^− 5 mol m^− 2 s^− 1 at ≈21 kPa oxygen pressure) represents the maximum surface coverage by cyanide. This value is in close agreement with the rate constant of the surface reaction 4 × 10^− 5 mol m^− 2 s^− 1 based on the pure kinetic current of the mixed “charge transfer plus diffusion” model proposed by Li and Wadsworth [Li, J., Wadsworth, M.E., 1993. Electrochemical study of silver dissolution in cyanide solutions. J. Electrochem. Soc. 140, 1921–1927].     

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.     

Studies on the effect of addition of silver ions on the direct oxidation of pyrite

The nature of the reaction between Ag⁺ and pyrite in 0.25 M H₂SO₄ solutions has been investigated in order to determine whether Ag⁺ can enhance the ferric sulfate leaching of this mineral. Analysis of reacted pyrite particles using scanning electron microscopy, X-ray photoelectron spectroscopy (XPS), and low-angle X-ray diffraction (XRD) indicates that elemental silver and elemental sulfur are the primary surface species formed by this interaction. Rest potential measurements of a pyrite electrode immersed in a solution containing 10⁻² M Ag⁺ are also consistent with what is expected for the deposition of metallic silver. Furthermore, the XRD data reveal that, at the most, only minor amounts of Ag₂S are being produced. The presence of Ag₂O has also been detected, but this is due to oxidation of silver after the experiment is complete and while the particles are being transferred for surface analysis. When 1 M ferric sulfate is contacted with pyrite which has been pretreated in a AgNO₃ solution, most of the silver immediately redissolves and does not redeposit while ferric ions are present. This indicates that the kinetics of the transfer reaction between Ag⁺ and pyrite is slower than the reaction between Fe³⁺ and pyrite and suggests that Ag⁺ does not likely enhance the ferric sulfate leaching.     

Solvent extraction of cadmium from sulfate solution with di-(2-ethylhexyl) phosphoric acid diluted in kerosene

In the present paper, solvent extraction process has been used for extraction of cadmium from sulfate solution using di-(2-ethylhexyl) phosphoric acid (D2EHPA) with 1% isodecanol in kerosene diluent expected from industrial effluents or leaching of ores/secondary materials. Different process parameters such as pH, contact time, extractant concentration, O/A ratio etc. were investigated. Results demonstrated that quantitative extraction of cadmium was feasible from 4.45 mM cadmium feed solution in single stage at equilibrium pH 4.5, time 2 min and O/A ratio 1:1 with 0.15 mM D2EHPA. The extraction mechanism of cadmium from sulfate solution by D2EHPA in kerosene could be represented at equilibrium by Cd²⁺ + 3/2 (H₂A₂)_org ⇔ CdA₂(HA)_org + 2H⁺. The loading capacity of 0.15 mM D2EHPA in sulfate solution was determined to be ∼ 8.9 mM cadmium. The loaded cadmium was effectively stripped using 180 g/L sulfuric acid. The metal or salt could be produced by electrolysis or crystallization from the stripped solution.     

A review of effects of silver,lead, sulfide and carbonaceous matter on gold cyanidation and mechanistic interpretation

Despite the wealth of published data on the beneficial or detrimental effects of silver, lead, sulfide, and carbonaceous matter on the rate of gold cyanidation at an anode or by dissolved oxygen, the lack of comparative studies on relative effects has hampered rationalisation of the role of these activators or passivators of gold. In the present study, the published rate data per unit surface area of gold, silver, and gold–silver alloys based on electrochemical or chemical dissolution of rotating discs or foils of constant surface area in aerated cyanide solutions at ambient temperatures are analysed on the basis of the Levich equation. The current status of the reaction mechanism is also reviewed and updated on the basis of species distribution and potential–pH diagrams, stoichiometric factors, and interim chemical species of gold(I), silver(I), and lead(II). The anodic peak potentials of reported voltammograms closely follow the potential–pH lines of Au(I)/Au(0) and Pb(II)/Pb(0) couples. Despite the formation of stable complexes between lead(II), nitrate, and hydroxide ions, the total calculated soluble lead(II) in alkaline solutions of pH range 10–11 saturated with lead hydroxide is shown to be < 0.1 g/m³. A comparison of the reported diffusion coefficients of cyanide ions and dissolved oxygen with the values based on the Levich plots of reported rates reveals the rate-controlling stoichiometric M/CN or M/O₂ molar ratios. The difference between some of these ratios and the generally accepted ratios of M/CN = 1/2 and M/O₂ = 1/0.5 or 1/0.25 based on the formation of M(CN)₂⁻, H₂O₂ or OH⁻ in the overall cyanidation reaction is attributed to the oxidation of cyanide to cyanate and passivation due to the formation of gold hydroxides/oxides. The alloyed or dissolved silver and lead eliminate passivation due to the involvement of mixed hydroxo–cyano complexes of silver and lead ions in the surface reaction. Gold dissolution by oxygen in cyanide-rich solutions is limited by oxygen diffusion, but enhanced by the presence of a low concentration of sodium sulfide due to the involvement of hydrosulfide ion in the surface reaction. However, excess lead or sulfide retards gold cyanidation due to surface blockage by metallic lead, lead hydroxide, or due to passivation by Au₂S/S. Even low concentrations of hydrosulfide passivate gold–silver alloys due to the formation of Ag₂S. This can be eliminated by adding stoichiometric quantities of lead(II) to precipitate sulfide as PbS. Large stoichiometric ratios of O₂/M for the cyanidation of graphite coated gold appears to be a result of the enhanced oxidation of cyanide by oxygen or hydrogen peroxide, leading to a cyanide deficiency at the surface and passivation of gold by hydroxide/oxide. The presence of excess cyanide or lead(II) does not override this effect.     

Arsenic stabilization of calcium arsenate waste by hydrothermal precipitation of arsenical natroalunite

The stabilisation of calcium arsenate waste from a copper smelter by precipitation of arsenical natroalunite has been investigated. This procedure could solve the problem of arsenical gypsum production because it is transformed into arsenic-free anhydrite. Natroalunite precipitation was studied at 180-200 °C from the slurry obtained after H₂SO₄-leaching and ozonation of the original waste - using sodium and aluminum sulfates as reagents. Calcium arsenate waste and final precipitates were characterized by chemical analysis (ICP), SEM-EDS and XRD. Solubility tests were also performed on original waste and selected precipitates.The effects studied in the hydrothermal treatment were: initial pH, Al/As molar ratio, As concentration, reaction time and prior gypsum removal. For (Al/As)_aq = 4.5, a complex natroalunite (~(Na,Ca)(Al,Fe)₃((S,As,P)O₄)₂(OH)₆)) was extensively formed. Under these conditions, As-substitution in TO₄ was 7-8% molar. Decreasing (Al/As)_aq increased As-substitution in natroalunite (up to ~ 14% molar) but mansfieldite ((Al,Fe)(As,P)O₄.2H₂O) co-precipitated. Other effects such as the arsenic concentration in the range 3.5-7.0 g/L and prior gypsum removal did not significantly alter the arsenic phase ratio and the composition of the arsenic phases. However, treatment at 180 °C increased mansfieldite precipitation. Calcium incorporation in natroalunite was small (~ 0.04 in formula) and for initial pH = 1, precipitation of Cu, Ni and Zn was insignificant.Arsenical natroalunite can be effective for long-term storage. At its natural pH (4-5), arsenic solubility remained stabilized at ~ 0.1 mg/L in 6 months. Mansfieldite was not stable under the same conditions with arsenic solubility between 5-10 mg/L. The major disadvantage of natroalunite is the limited arsenic partition between solid and aqueous phases (~ 0.45 molar). However, as the precipitation rate is fast (< 15 min), it seems technically feasible to improve the arsenic precipitation yield by semi-continuous addition of reagents in a single hydrothermal operation. Cadmium incorporated into the anhydrite structure causes cadmium release in leachates. It seems that this problem can only be overcome through prior removal of cadmium from the plant effluent.     

The dissolution of galena in ferric chloride media

The dissolution of galena (PbS) in ferric chloride-hydrochloric acid media has been investigated over the temperature range 28 to 95 °C and for alkali chloride concentrations from 0 to 4.0 M. Rapid parabolic kinetics were observed under all conditions, together with predominantly (>95 pet) elemental sulfur formation. The leaching rate decreased slightly with increasing FeCl₃ concentrations in the range 0.1 to 2.0 M, and was essentially independent of the concentration of the FeCl₂ reaction product. The rate was relatively insensitive to HCl concentrations <3.0 M, but increased systematically with increasing concentrations of alkali or alkaline earth chlorides. Most significantly, the leaching rate decreased sharply and linearly with increasing initial concentrations of PbCl₂ in the ferric chloride leaching media containing either 0.0 or 3.0 M NaCl. Although the apparent activation energy was in the range 40 to 45 kJ/mol (∼10 kcal/mol), this value was reduced to 16 kJ/mol (3.5 kcal/mol) when the influence of the solubility of lead chloride on the reaction rate was taken into consideration. The experimental results are consistent with rate control by the outward diffusion of the PbCl₂ reaction product through the solution trapped in pores in the constantly thickening elemental sulfur layer formed on the surface of the galena.     

硫-石灰溶液中氧压浸出金精矿

研究了石灰液体系中元素硫加压氧化浸金过程。通过四变量三水平的实验设计,发现元素硫与氢氧根的摩尔比及氧耗量是影响金浸出的最重要因素。单变量试验表明,适宜的S0/OH摩尔比为1.0～1.2,最优氧耗量为0.38～0.55 NL/g S0,浸出过程中终pH 应为3.5～6.0。最优条件下金的一段浸出率为82%～85%,二段浸出可达91%。     

高砷含锑难浸金精矿提金工艺的研究

研究了高砷含锑难浸金精矿的中温低压氧化酸浸- 石灰液或氨浸氧化的联合预处理技术的工艺和优化操作条件,讨论了催化氧化酸浸、石灰液氧化浸出及氧化氨浸中主要影响因素的作用规律。预处理后, 金的氰化浸出率由未预处理时的几乎为零提高到90%以上。同时提出了一种酸浸渣直接氧压浸金的新过程,论述了酸浸渣中的元素硫在石灰液中氧压浸出的过程原理及主要条件因子的影响规律。结果表明, 金的浸出率基本上与联合预处理后金的氰化率相同。     

Reaction mechanism for the acid ferric sulfate leaching of chalcopyrite

The acid ferric sulfate leaching of chalcopyrite, CuFeS₂ + 4Fe⁺³ = Cu⁺² + 5Fe⁺² + 2S⁰ was studied using monosize particles in a well stirred reactor at ambient pressure and dilute solid phase concentration in order to obtain fundamental details of the reaction kinetics. The principal rate limiting step for this electrochemical reaction appears to be a transport process through the elemental sulfur reaction product. This conclusion has been reached in other investigations and is supported by data from this investigation in which the reaction rate was found to have an inverse second order dependence on the initial particle diameter. Furthermore, the reaction kinetics were found to be independent of Fe⁺³, Fe⁺², Cu⁺² and H₂SO₄ in the range of additions studied. The unique aspect of this particular research effort is that data analysis, using the Wagner theory of oxidation, suggests that the rate limiting process may be the transport of electrons through the elemental sulfur layer. Predicted reaction rates calculated from first principles using the physicochemical properties of the system (conductivity of elemental sulfur and the free energy change for the reaction) agree satisfactorily with experimentally determined rates. Further evidence which supports this analysis includes an experimental activation energy of 20 kcal/mol (83.7 kJ/mol) which is approximately the same as the apparent activation energy for the transfer of electrons through elemental sulfur, 23 kcal/ mol (96.3 kJ/mol) calculated from both conductivity and electron mobility measurements reported in the literature. formerly Metallurgy Graduate Student, University of Utah.     

Kinetic model for the cyanidation of silver ammonium jarosite in NaOH medium

A synthesis of silver ammonium jarosite has been carried out obtaining a single-phase product with the formula: [(NH₄)_0.71(H₃O)_0.25Ag_0.040]Fe_2.85(SO₄)₂(OH)_5.50. The product consists on compact spherical aggregates of rhombohedral crystals. The nature and kinetics of alkaline decomposition and also of cyanidation have been determined. In both processes an induction period followed by a conversion period have been observed. During decomposition, the inverse of the induction period is proportional to [OH⁻]^0.75 and an apparent activation energy of 80 kJ mol^− 1 was obtained; during the conversion period, the process is of 0.6 order (OH⁻ concentration) and an activation energy of 60 kJ mol^− 1 was obtained. During cyanidation, the inverse of the induction period is proportional to [CN⁻]^0.5 and an apparent activation energy of 54 kJ mol^− 1 was obtained; during the conversion period the process is of 0 order (CN⁻ concentration) and an activation energy of 52 kJ mol^− 1 was obtained. Results obtained are consistent with the spherical particle model with decreasing core and chemical control, in the experimental conditions employed. For both processes and in the basis of the behaviour described, two mathematical models, including the induction and conversion periods, were established, that fits well with the experimental results obtained. Cyanidation rate of different jarosite materials in NaOH media have also been established: this reaction rate at 50 °C is very high for potassium jarosite, high and similar for argentojarosite and ammonium jarosite, lower for industrial ammonium jarosite and negligible for natural arsenical potassium jarosite and beudantite. These results confirm that the reaction rate of cyanidation decreases when the substitution level in the jarosite lattice increases.     

Pressure acid leaching of a Chinese laterite ore containing mainly maghemite and magnetite

The high pressure acid extraction of nickel and cobalt from a Chinese laterite containing mainly maghemite and magnetite was studied. X-ray diffraction (XRD) and scanning electron microscopy/X-ray energy dispersive spectroscopy (SEM/EDS) were employed to characterize the residues. The factors influencing the dissolution of maghemite and magnetite, nickel and cobalt extractions and iron precipitation were investigated. The results show that after 75 min at 270 °C with an acid/ore ratio of 0.55, maghemite and magnetite completely dissolved, liberating 98% Ni and 88% Co into the leach liquor. EDS analysis reveals that some nickel may be associated with the amorphous silica and/or basic ferric sulfate, resulting in a minor loss of nickel. The presence of a cobalt-containing phase in the residues, believed to be ringwoodite, is mainly responsible for the incomplete extraction of cobalt. Both maghemite and magnetite dissolved gradually with the increase in temperature from 200 to 270 °C. Maghemite dissolved more slowly than magnetite at 270 °C which also produced ferrous sulfate in the leach liquor and increased the total iron extraction. Increasing temperature and/or agitation accelerated the hydrolysis of ferric sulfate. The leaching of maghemite and magnetite corresponds to a dissolution-precipitation mechanism. In both high and low acidic environments, the precipitation of ferric sulfate proceeds through the initial formation of basic ferric sulfate and its conversion to hematite. The extent of conversion depends largely upon residual acidity and reaction time.     

Recovery of vanadium during ammonium molybdate production using ion exchange

The recovery of vanadium in ammonium molybdate production with raw sodium molybdate solution was studied. Experimental results showed that the vanadium and some of impurities P, As and Si can be adsorbed along with the molybdenum from the feed liquid by the weak base resin D314 in the pH range of 2.5-3.5 and then they can be eluted from the loaded resin with ammonia liquor. The vanadium can partially natural-precipitate from the eluted solution. The lower the pH value is and, the longer the standing time is, the less the vanadium remained in the solution will be. Standing for 24 h in pH value 6.9, the vanadium in the solution was reduced to 0.51 g/L V₂O₅. It was found that the precipitate is ammonium isopoly-vanadate and it is impure. By washing the precipitate with hydrochloric acid and ammonia solution sequentially and then roasting at 500 °C for 2 h, the product of V₂O₅ with the purity 99.12% was obtained. The impurities P, As and Si in the stood solution were removed by purifying with MgCl₂ under the pH value range of 8.0-9.0 at 60-80 °C for about 2 h, while the removal of the vanadium in the solution was performed by adsorbing with the strong base resin D296 in pH value about 7.0. The devanadiumized solution can be used to produce high-quality ammonium molybdate. The loaded vanadium resin was eluted with HCl 6 mol/L and, the eluted solution was returned to adjust the sodium molybdate solution pH value. The vanadium can be effectively separated and recovered in the process.     

Recovery of nickel and cobalt from laterite leach solutions using direct solvent extraction: Part 2: Semi- and fully-continuous tests

In Part 1 of this paper, two synergistic solvent extraction systems consisting of Versatic 10/LIX63/TBP and Versatic 10/4PC were assessed in batch tests for the separation and purification of nickel and cobalt from synthetic laterite leach solution after iron removal. In Part 2, semi- and fully-continuous tests are reported for the Versatic 10/LIX63/TBP system, with conditions optimised for separating nickel and cobalt from manganese, magnesium and calcium.Semi-continuous extraction tests were conducted using the synergistic organic system consisting of 0.50 M Versatic 10, 0.45 M LIX63 and 1.0 M TBP in Shellsol D70. With a pH profile of 5.5/6.1/6.5 for the three stages EX1/EX2/EX3 at 40 °C, the nickel and cobalt extractions were 99.9% with only 5 mg/L nickel and < 1 mg/L cobalt left in the raffinate. With two stages of scrubbing and a pH profile of 5.4/5.0 at 40 °C, about 2 mg/L manganese and less than 1 mg/L magnesium and calcium were left in the scrubbed organic solution. With two stripping stages and an O/A ratio of 10 at 40 °C using 50 g/L H₂SO₄ as strip solution, the stripping efficiencies of nickel and cobalt were over 95%.A fully-continuous pilot plant was operated for 280 h. With an O/A ratio of about 2 and a pH profile of 5.5/5.8/6.0/6.3 for the four stages EX1/EX2/EX3/EX4 at 40 °C, both nickel and cobalt were almost completely extracted. The nickel and cobalt concentration in the raffinate was lower than detection limit of 0.2 mg/L. The manganese, magnesium and calcium concentrations in the loaded organic solution were 34, 8 and 1 mg/L, respectively. Using a pH profile of 5.4/5.0 for SC1/SC2 at an O/A ratio of 10 and 40 °C, the manganese scrubbing efficiency was over 96% and the concentrations of manganese and magnesium in the scrubbed organic solution were < 5 mg/L and that of calcium 1 mg/L. Using three strip stages and a strip solution containing 50 g/L H₂SO₄ and 55 g/L Ni at an O/A ratio of 10 and 40 °C, over 98% Ni and 99% Co were stripped with only 64 mg/L Ni in the stripped organic solution. The nickel concentration in the loaded strip liquor was 86 g/L, giving a ΔNi of 31 g/L. The loaded strip liquor contained less than 1 g/L acid.     

Intracellular recovery of gold by microbial reduction of AuCl4 ions using the anaerobic bacterium Shewanella algae

Microbial reduction and intracellular precipitation of gold was achieved at 25 °C and pH 7 by using the mesophilic anaerobic bacterium Shewanella algae with H₂ as the electron donor. The reductive precipitation of gold by S. algae was a fast process: 0.1–1 mol/m³ AuCl₄⁻ ions were completely reduced to insoluble gold within 30 min. The biogenic precipitates were crystalline gold nanoparticles of 10–20 nm present in the periplasmic space. The reducing power of S. algae at 3.2 × 10¹⁵ cells/m³ and 25 °C was comparable to that of aqueous citric acid solution (chemical reductant) at 20 mol/m³ and 50 °C. The intracellular recovery of gold is potentially attractive as an environmentally friendly alternative to conventional methods.     

Copper removal by chelating adsorption in solution purification of hydrometallurgical zinc production

Use of the chelating adsorbent CuWRAM® in the copper removal step of hydrometallurgical zinc process has been studied. This adsorbent contains 2-(aminomethyl)pyridine groups anchored on a polyamine-silica composite and it binds copper and other transition metals by a chelating adsorption mechanism. Equilibrium binding capacity of metal sulfates and sulfuric acid from synthetic and authentic ZnSO₄ process solution was determined at 25-90 °C using batch adsorption measurements. The copper removal efficiency was tested using a laboratory-scale fixed-bed column.Results of the equilibrium measurements show that the selectivity of CuWRAM® is sufficient for feasible separation of copper in the presence of 250-fold zinc excess. Increasing the operation temperature from 25 °C to 60 °C affects only slightly the binding capacity of copper and at the same time decreases the capacity of zinc. In column separation, increasing temperature substantially improves copper removal efficiency from the ZnSO₄ process solution. The improvement is mainly due to enhanced intra-particle mass transport. The positive effect is further amplified by marked decrease in viscosity of the feed solution. The optimum temperature for copper removal appears to be around 60 °C. According to the results of this study, copper can be separated from the authentic ZnSO₄ solution by the chelating adsorbent, while nickel, cobalt and cadmium must be separated by means of conventional methods like cementation with zinc dust. A process scheme is proposed for the solution purification step in the zinc process.     

Henrian activity coefficient of Pb in Cu-Fe mattes and white metal

A transpiration method was used to evaluate the Henrian activity coefficient of Pb (γ°_Pb) in Cu-Fe mattes and white metal. Values for the activity coefficient of Pb (γ _Pb) have been evaluated as a function of the Cu/Fe molar ratio from 1 to ∞, as a function of the sulfur deficiency (defined as SD=X _S−1/2X _Cu−X _Fe, where X ₁ is the mole fraction of the ith species) from −0.02 to +0.02, and at temperatures between 1493 and 1573 K. Analysis of γ _Pb as a function of the trace element concentration reveals that the activity coefficient is independent of Pb content at weight percents less than 0.2. Dependence of γ _Pb on temperature was found to be slight, and as such, comparison of data obtained by other investigators at 1473 K was possible. Agreement in the data is excellent, and all the data have been used to generate the empirical equation

Hydrochloric acid-hydrogen peroxide leaching and metal recovery from a Greek zinc-lead bulk sulphide concentrate

A zinc-lead bulk sulphide concentrate from Kirki (Greece) was leached in aqueous solution with HClH₂O₂at atmospheric pressure and 95°C to extract up to 97% zinc, 40% lead, 80% silver and less than 12% iron after 6 h. Highly pure PbCl₂ crystallized from the leach filtrate on cooling. Sulphur was oxidized to the elemental form; its loss as sulphate ion in solution and residue was 7.5%. During leaching no emission of H₂S or SO₂ was detected. Conditions were determined for producing a pregnant solution (130 g/L Zn) low in iron and lead, to facilitate zinc extraction by electrolysis. Leaching experiments were conducted at 40% solids, 0.55 g HCl and 0.26 g H₂O₂ per gramme concentrate in a 1-L reactor. After a second leaching of the residue in aqueous solution with HCl (1 M) for 1 h at 90°C, the residual lead sulphate was extracted, so that total lead recovery was over 98%. Other values, such as silver and sulphur, could be recovered by additional treatment of either the leach solution or leach residue.     

Properties of electrolyte solutions relevant to high concentration chloride leaching. III. Solubility of pertinent solids and iron(III)/iron(II) redox potential measured in concentrated magnesium chloride solutions

Results of solubility measurements of nickel chloride, manganese chloride, iron(II) chloride, hematite and akaganeite in aqueous solutions of MgCl₂ (0.5–3.5 mol L^− 1) at temperatures of 60 and 90 °C are reported. Solubilities of metal(II) chlorides decrease almost linearly with MgCl₂ concentration due to the common ion effect. Nickel chloride and iron(II) chloride solubilities are very similar, while manganese chloride is about 30% more soluble.Hematite is more stable (i.e. less soluble) than akaganeite under all conditions investigated in this study, while ferrihydrite is considerably less stable. In other words, there is no change in the relative stabilities of these phases effected by the presence of high magnesium chloride concentrations. The solubility of all of these phases decreases with temperature and, for each temperature, the solubility constants increase linearly with the MgCl₂ concentration. The present results allow the prediction of the iron concentration as a function of the H⁺ and MgCl₂ molality at equilibrium with hematite or akaganeite.The Fe(III)/Fe(II) redox behaviour has been characterized in concentrated aqueous solutions of MgCl₂ (1.5–3.5 mol L^− 1) at a temperature of 25 °C. Standard redox potentials are ca. 100 mV lower than at infinite dilution and change linearly by only 13 mV in the range 2–4 mol L^− 1 MgCl₂.