Tungsten recovery from alkaline leach solutions as synthetic scheelite

The recovery of tungsten from alkaline leach solutions has been studied examining the effect of temperature, pH, Ca/WO₃ molar ratio and nature of the precipitated calcium tungstate. The precipitation kinetics of calcium tungstate, upon the addition of aqueous sodium tungstate to calcium solutions, was followed by potentiometric measurements using a calcium ion-selective electrode. Two models, a crystal growth model and a second-order reaction opposed by zero-order reaction, have been used to test the experimental data. Both models show that the apparent activation energy of CaWO₄ precipitation falls in the range 58 to 67 kJ mol^− 1.The kinetic data shows that the maximum recovery of precipitated calcium tungstate occurs at pH ≥ 8.5 with a 10% excess of CaCl₂ at 50 °C over a period of 20 min using sodium tungstate solutions of 100 g L^− 1 and 150 g L^− 1 WO₃.     

The precipitation of ammonium uranyl carbonate (AUC): Thermodynamic and kinetic investigations

The aim of this study is to determine the predominant chemical reaction during precipitation of ammonium uranyl carbonate (AUC) based on thermodynamic analysis and to investigate its kinetics. Four chemical reactions were considered. The Gibbs free energies, Δ_rG°(T) derived from the Ulich calculations as a function of temperature have been determined between 293.15 K and 353.15 K. The predominant chemical reaction of AUC precipitation was UO₂(NO₃)₂·6H₂O_(aq) + 6NH_3(g) + 3CO_2(g) → (NH₄)₄UO₂(CO₃)_3(s) + 2NH₄NO_3(aq) + 3H₂O_(l). According to the AUC precipitation kinetics results, the reaction best fits a second order rate equation. The rate constants k₂ were calculated at 313.15 K and 330.15 K and the activation energy E_a determined using the Arrehenius equation was found as 17.4 kJ/mol.     

The oxidation of thiosulfates with copper sulfate. Application to an industrial fixing bath

This paper presents the transformation of thiosulfate using Cu(II) salts, such as copper sulfate, at pH between 4 and 5. The nature and kinetics of this process were determined. In the experimental conditions employed, the reaction between thiosulfates and Cu(II) ions produces a precipitate of CuS and the remaining sulfur is oxidized to sulfate, according to the following stoichiometry: 1 mol thiosulfate reacts with 1 mol Cu²⁺ and 3 mol H₂O, generating 1 mol copper(II), 1 mol sulfate and 2 mol H₃O⁺. In the kinetic study, the apparent reaction order was ≈ 0 with respect to H₃O⁺ concentration, in the interval 1.0 · 10^? 4–1.0 · 10^? 5M H₃O⁺; of order 0.4 with respect to Cu²⁺ in the interval 0.21–0.85 g L^? 1 Cu²⁺; and of order 0 with respect to S₂O₃^2? in the interval 0.88–2 g L^?1 S₂O₃^2?. The apparent activation energy was 98 kJ mol^? 1 in the interval 15–40 °C. On the basis of this behavior an empirical mathematical model was established, that fits well with the experimental results. The thiosulfate transformation process using copper(II) sulfate was applied to an industrial fixing bath that proceeded from the photographic industry; after this, the resulting effluent contained less than 10 mg L^? 1 of thiosulfates.     

Kinetics and mechanism of stripping of Mn(II)–D2EHPA complex by sulphuric acid solution

The kinetics of stripping of Mn(II)–D2EHPA chelate (MnA₂.HA.H₂A₂) existing in kerosene phase by aqueous sulphuric acid solutions at constant sulphate ion concentration of 0.50 M have been investigated by the constant interfacial area stirred (Lewis) and non-stirred (Hahn) cells. Both pseudo-rate constant (q) or (rate / area) and flux (F) methods of the rate data treatment have been applied. The empirical rate equations have been derived. Results have been compared among themselves and other published works on Mn(II)–D2EHPA chelate stripping kinetics. Rate constants obtained from (q) or (rate / area) and (F) methods differ in magnitude and units and an explanation of this has been given.An analysis of rate equations obtained from the Lewis cell experimentation (F_b = 10^− 4.1[MnA₂.HA.H₂A₂]_(o)(1 + 0.005[H⁺]_(i)^− 1)^− 1 (1 + 7.94[H₂A₂]_(o)^0.5)^− 1) suggests that the process is chemically controlled at lower aqueous acidity and higher free D2EHPA concentration regions; whereas, it is diffusion-controlled at higher aqueous acidity and lower free D2EHPA concentration regions. In the investigated D2EHPA concentration and aqueous acidity regions, majority of data fall in the intermediate controlled region. The suggested mechanisms are supported by the values of activation energy (E_a). In the kinetic regime, the reaction step: MnA⁺ → Mn²⁺ + A⁻ occurring in the bulk aqueous phase is the slowest step which proceeds through S_N2 mechanism as indicated by the comparable higher negative value of ΔS^± in the stated condition.In Hahn cell technique, the empirical rate equation is: F_b = 10^− 5.9[MnA₂.HA.H₂A₂]_(o)(1 + 0.0025[H⁺]_(i)^− 1)^− 1(1 + 2.8[H₂A₂]_(o)^0.5)^− 1. Analysis of this equation together with the E_a value of 12–42 kJ/mol indicate that the stripping process is completely diffusion controlled in the low pH and extractant concentration region and in other parametric condition, that is intermediate controlled. For both cells, the ratio of (k_f) to (k_b) equals almost to K_ex of 10^− 2.44 obtained from the distribution study.     

Chemistry of the Ca–Se(IV)–H2O and Ca–Se(VI)–H2O systems at 25 °C

The Ca–Se(IV)–H₂O and Ca–Se(VI)–H₂O systems were studied by contacting either selenious acid or selenic acid solution with calcium oxide to attain equilibrium at 25 °C for one month. Analysis of the final solid phases and the associated solution, together with X-ray diffraction analysis and a study into the graphed relationships, showed the existence of three calcium selenites in the Ca–Se(IV)–H₂O system — Ca₂SeO₃(OH)₂·2H₂O (Se(IV) = 4.8 × 10^− 5–2.8 × 10^− 4 M); CaSeO₃·H₂O (Se(IV) = 2.8 × 10^− 4–0.86 M) and Ca(HSeO₃)₂·H₂O (Se(IV) > 0.86 M). It also showed four calcium selenates in the Ca–Se(VI)–H₂O system — Ca₂SeO₄(OH)₂ (Se(VI) = 0.21–0.39 M); CaSeO₄·2H₂O (Se(VI) = 0.40–9.1 M); CaSeO₄ (Se(VI) = 10.2 M) and CaSe₂O₇ (Se(VI) > 10.8 M). The X-ray diffraction analyses reported and SEM analyses indicate a high degree of crystallinity of all seven compounds. The stability and solubility regions for these compounds were defined versus pH, and the conventional solubility constants and conditional free energies of formation for the less soluble CaSeO₃·H₂O, Ca₂SeO₃(OH)₂·2H₂O, CaSeO₄·2H₂O and Ca₂SeO₄(OH)₂ were calculated from solubility data obtained.     

Product inhibition by sulphide species on biological sulphate reduction for the treatment of acid mine drainage

It is well recognised that the product of sulphate reduction, i.e. the sulphide species formed, may inhibit the biological process. In this paper, we further the kinetic study of biological sulphate reduction using the mixed population of complete oxidisers growing on acetate for which kinetic data has been reported previously as a function of sulphate concentration, temperature, dilution rate and volumetric sulphate loading using chemostat culture by Moosa et al. to provide kinetic insight into this inhibition.The effect of a feed sulphide concentration in the range 0.50 to 1.25 kg m^− 3 on the biological sulphate reduction process is established using chemostat culture at pH 7.0 ± 0.2. Further, the chemical speciation of sulphide as undissociated H₂S or dissociated HS⁻ on process inhibition is reported through the variation of operating pH across the range pH 6.0 to pH 7.5 at a sulphate feed concentration of 2.5 kg m^− 3. It is clearly shown that inhibition is chiefly mediated by the undissociated H₂S sulphide species, rather than the total sulphide concentration. This inhibition was shown to affect the maximum specific growth rate constant and the death rate constant in the Contois rate equation presented previously while having negligible effect on KS describing substrate affinity.     

Sorption behaviour and mechanism of ytterbium(III) on imino-diacetic acid resin

The sorption behaviour and mechanism of a novel chelate resin, imino-diacetic acid resin (IDAAR), for Yb(III) has been investigated in HAc–NaAc medium. The sorption of Yb(III) obeys the Freundlich isotherm. Optimum sorption for Yb(III) on IDAAR is at pH 5.13 and the statically saturated sorption capacity is 187 mg/g resin at 298 K. Yb(III) can be eluted using 1~2 mol L^− 1 HCl and the resin can be regenerated and reused without apparent decrease of sorption capacity. The apparent sorption rate constant is k₂₉₈ = 1.57 × 10^− 5 s^− 1; the apparent activation energy is 13.8 kJ mol^− 1 and the enthalpy change ΔH of IDAAR for Yb(III) is 29.8 kJ mol^− 1. The sorption mechanism of IDAAR for Yb(III) was examined by using chemical methods and IR spectrometry. The molar coordination ratio of the functional group of IDAAR to Yb(III) is 3:1 with the coordination compound formed between oxygen atoms in the functional group of IDAAR and Yb(III).     

Extraction of gold(III) in hydrochloric acid solution using monoamide compounds

The extraction and separation properties of Au(III) using two monoamide compounds, N,N-di-n-octylacetamide (DOAA) and N,N-di-n-octyllauramide (DOLA), which have different side chain lengths attached to the carbonyl carbon (CH₃ for DOAA and n-C₁₁H₂₃ for DOLA), were investigated. The solvent extraction of some precious and base metals (Au(III), Pd(II), Pt(IV), Rh(III), Fe(III), Cu(II), Ni(II) and Zn(II)) in HCl solutions was carried out using DOAA and DOLA diluted with n-dodecane and 2-ethylhexanol. A good selectivity for Au(III) extraction with 0.5 M extractant is obtained at lower HCl concentrations (< 3.0 M) in both systems. The extractability of Au(III) with DOAA is greater than that with DOLA. In the 0.5 M DOAA–3.0 M HCl system, a third phase is formed when the Au(III) concentration in the initial aqueous phase is over 39 g/L. In contrast, third phase formation is not found in the 0.5 M DOLA–3.0 M HCl system, and its loading capacity of Au(III) is about 79 g/L. The Au(III) extracted in the organic phase is effectively back-extracted by 1.0 M thiourea in 1.0 M HCl solution in both systems, while some thiourea is precipitated using the organic phase containing 20 g/L of Au(III). The back extraction of Au(III) using water is poor in the DOAA system, but possible in the DOLA system.     

Carrier-mediated transport of uranium from phosphoric acid medium across TOPO/n-dodecane-supported liquid membrane

Present studies deals with the application of supported liquid membrane (SLM) technique for the separation of uranium (VI) from phosphoric acid medium. Tri-n-octyl phosphine oxide (TOPO)/n-dodecane is used as a carrier and ammonium carbonate as a receiving phase for the separation of uranium (VI) from the phosphoric acid medium. Throughout the study PTFE membranes are used as a support. The studies involve the investigation of process controlling parameters like feed acidity of phosphoric acid, carrier concentration and stripping agents. The effect of nitric acid and sodium nitrate in feed is also studied. It is found that there is negligible transport of uranium (VI) from pure phosphoric acid medium but it increases to very significant amount if 2 M nitric acid is added to feed phase. More than 90% uranium (VI) is recovered in 360 min using 0.5 M TOPO/n-dodecane as carrier and 1.89 M ammonium carbonate as stripping phase from the mixture of 0.001 M H₃PO₄ and 2 M of HNO₃ as a feed. The flux and permeability coefficient are found to be 9.21 × 10^− 6 mol/m² s and 18.26 × 10^− 5 m/s, respectively. Lower concentration of phosphoric acid with 2 M HNO₃ and higher concentration of carrier is found to be the most suitable condition for maximum transport of uranium (VI) from its low-level sources like commercial phosphoric acid.     

Identification of arsenato antimonates in copper anode slimes

It was found that, in copper electrolyte, the combination of As(V) and Sb(V) can form arsenato antimonic acid (AAAc) and, the reactions of AAAc with As(III), Sb(III), and Bi(III) can produce the precipitates of arsenato antimonates. During copper electrorefining, the As, Sb, and Bi deposited into the anode slime from the electrolyte are dominant in the forms of arsenato antimonates. It is extremely difficult to separate pure arsenato antimonates from copper anode slimes, while it is easy to synthesize arsenato antimonates using H₂O₂ to oxidize As(III) and Sb(III) in copper refining electrolyte. The composition and structure of the arsenato antimonates were determined with chemical analysis, IR and XRD techniques. The characteristic bands in the IR spectra of arsenato antimonates are δ of As–OH and Sb–OH at 1126.8 cm^− 1, ν_as of As–OH at 1029.7 cm^− 1, ν_as of As–OX(X = As, Sb) at 819.5 cm^− 1, ν_as of Sb–OH at 618.4 cm^− 1, ν_as of Sb–OY(Y = As, Sb) at 507.2 cm^− 1, and ν_as of Sb–OBi at 470 cm^− 1.The arsenato antimonates form irregular masses of amorphous structure because there are many OH groups in AAAc, the OH groups bond with As(III), Sb(III), and Bi(III) at random, which makes the arsenato antimonates formed in copper refining electrolyte have no fixed ratios for As/Sb/Bi. The formation of the arsenato antimonates can be expressed as follows:aH₃AsO₄ + bH[Sb(OH)₆] + cMeO ⁺ →Me_cAs_aSb_bO_{(3a+5b+c/2+1)}H_{(a+5b−2c+2)}·xH₂O + cH⁺ + (a + b + c / 2 − 1 − x)H₂O, where Me = As(III), Sb(III) and Bi(III); a ≥ 1, b ≥ 1, c ≤ (3a + b)     

Kinetics of galena dissolution in nitric acid solutions with hydrogen peroxide

The extraction of lead from a galena concentrate in nitric acid solutions with additional hydrogen peroxide was studied taking stirring speed, temperature, hydrogen peroxide and nitric acid concentrations, and particle size as dissolution parameters. The dissolution curves followed the surface chemical reaction controlled shrinking core model over the whole range of parameters, except at high nitric acid concentrations where the reaction was diffusion-controlled. The activation energy of 42 kJ mol^− 1 and a linear relationship between rate and inverse particle size support the reaction controlled dissolution mechanism. Hydrogen peroxide addition accelerated the reaction compared with nitric acid alone. It was concluded that the dissolution process is favourable, since the acid consumed for oxidation of galena can easily be regenerated in the same reactor by means of hydrogen peroxide.     

Extraction of cadmium (II) from phosphoric acid media using the di(2-ethylhexyl)dithiophosphoric acid (D2EHDTPA): Feasibility of a continuous extraction-stripping process

The extraction of cadmium from phosphoric media has been studied. The D₂EHDTPA was used as extractant and dodecane as diluent. No third phase was observed in the investigated conditions.A continuous micro-pilot scale mixer-settler was successfully tested for both extraction and stripping. More than 99% extraction rate was obtained in steady-state conditions with a flow rate ratio Aqueous/Organic equal to 1.1. Continuous stripping was performed using HCl 4 M. More than 96% of the cadmium was stripped in one continuous mixer-settler stage for flow rate ratio equal to 0.7. Results were in good agreement with the predicted values based on the McCabe–Thiele method. Experimental mixer-settler stages behave as ideal ones (Murphree efficiency > 98%).An optimal flow sheet is proposed to purify the Wet Phosphoric Acid (WPA) and to recover a relatively concentrated cadmium solution (1 g L^? 1). Two ideal stages operating at phase ratio A/S equal to 5/1 are required for the extraction step leading to a very depleted raffinate (< 0.2 µg L^? 1). For the stripping step, six stages are required (S/A = 5/1). The recovered organic phase contains less than 2 µg L^? 1 and could be recycled in the extraction step.     

Cadmium extraction from chloride solutions with model N-alkyl- and N,N-dialkyl-pyridine-carboxamides

N-alkyl- and N,N-dialkyl-pyridine-carboxamides with the amide group at the 2nd, 3rd or 4th position were used to recover cadmium(II) from acid chloride solutions. It was found that N,N-dialkyl-pyridine-carboxamides in toluene diluent are effective extractants for the recovery of cadmium from chloride solutions at pH < 2. Cadmium extraction ability rises as the distance of the amide group from the pyridine nitrogen increases. In strongly acid chloride media, extractants (L) with N,N-dialkyl-amide group at position 3 or 4 in the pyridine ring form cadmium ion pairs: (LH⁺)₂(CdCl₄²⁻) but the dialkyl derivative of picolinamide probably forms the complex (LH⁺)₂(CdCl₄²⁻)(LHCl). Monoalkyl-pyridine-carboxamides are not suitable for cadmium extraction from chloride solutions because in weak-acid systems (pH > 3, [Cl⁻] = 0.02–4 M) the N-alkyl-pyridine-3-carboxamides form very slightly soluble cadmium complexes and the N-alkyl-pyridine-2-carboxamides do not extract cadmium(II) from chloride solutions under these conditions. In strongly acid systems ([HCl] = 0.01–4 M) N-alkyl-pyridine-2-carboxamide with a branched carbon chain is ineffective, but the hydrochlorides and cadmium complexes of the rest of the N-alkyl-amides are slightly soluble in hydrocarbon diluents. In ethanol solutions hydrophobic pyridine-3-carboxamides form (CdCl₂)_nL₂ complexes (n = 2 or 3) with cadmium chloride.     

Characterization of the iron arsenate–sulphate compounds precipitated at elevated temperatures

To help clarify the nature of the iron arsenate–sulphate compounds produced during the autoclave treatment of refractory gold ores and concentrates, systematic synthesis studies were undertaken; in addition to scorodite and Fe(SO₄)(OH), two other compounds, designated as Phase 3 and Phase 4, were identified. Whereas Fe(SO₄)(OH) is predominantly an orthorhombic compound, Phase 3 can have the same composition but is predominantly the monoclinic polytype, the formation of which is promoted by the solid-solution uptake of As; substitution of As results in a corresponding decrease in the OH required to maintain the charge balance; e.g., Fe[(SO₄)_0.60(AsO₄)_0.40]_∑1.00[(OH)_0.6(H₂O)_0.4]_∑1.00. Phase 4 corresponds to Fe(AsO₄)·¾H₂O. In 0.4 M Fe(SO₄)_1.5 (22.3 g/L Fe), 0.41 M (40 g/L) H₂SO₄, 0.09 M (7 g/L) As(V) solutions, sulphate-containing scorodite was formed at 150–175 °C. Phase 3 precipitated at 175–210 °C, but mixtures of Phase 3 and Fe(SO₄)(OH) formed above 210–220 °C. The Fe content of Phase 3 is about 30 mass %, whereas the AsO₄ and SO₄ contents vary widely and in an inversely proportional manner, reflecting the extensive mutual structural substitution of these anions. At 205 or 215 °C, Fe(SO₄)(OH) was precipitated from 0.4 M Fe(SO₄)_1.5 (22.3 g/L Fe), 0.41 M (40 g/L) H₂SO₄ solutions containing < 0.03 M (2 g/L) As(V). Increasing As(V) concentrations enhance the precipitation of Phase 3, but only Phase 4 was precipitated from solutions containing > 0.33 M (25 g/L) As(V). The composition of Phase 4 is nearly constant and it contains < 1 mass % SO₄. Acid concentrations > 0.2 M H₂SO₄ had little effect on the composition of the precipitates. At 205 °C in 0.41 M (40 g/L) H₂SO₄, 0.09 M (7 g/L) As(V) media, mixtures of scorodite and Phase 4 precipitated from 0.0–0.1 M Fe(SO₄)_1.5 (0.0–5.6 g/L Fe) solutions; for Fe(SO₄)_1.5 concentrations > 0.1 M, only Phase 3 formed. To provide a preliminary indication of the solubility of Phase 3 and Phase 4 in tailings impoundments, the various precipitates were leached at room temperature for 40 h in water. The As concentrations dissolved from Phase 3 were consistently < 0.1 mg/L, which suggests that Phase 3 might be an acceptable medium for arsenic disposal. In contrast, the soluble As concentrations from Phase 4 were 1–3 mg/L.     

Effect of solution composition on the optimum redox potential for chalcopyrite leaching in sulfuric acid solutions

The leaching rate of chalcopyrite (CuFeS₂) by Fe³⁺ in H₂SO₄ solutions depends on the redox potential determined by the Fe³⁺/Fe²⁺ concentration ratio, and there is a maximum leaching rate at an optimum redox potential. The present study investigated the effects of solution composition on the optimum redox potential by electrochemical measurements using a CuFeS₂ electrode and electrolyte solutions containing 0.01–1 kmol m^− 3 of H₂SO₄, Fe²⁺, and Cu²⁺ at 298 K in nitrogen.Anodic-polarization curves of the CuFeS₂ electrode showed that there was a current peak on the curves in the presence of Cu²⁺ and Fe²⁺, corresponding to the maximum leaching rate. The redox potential of the peak increased markedly with increasing Cu²⁺ concentration, while it was little affected by the H₂SO₄ and Fe²⁺ concentrations. These results agree with the results of leaching experiments reported previously, and indicate that the optimum redox potential for chalcopyrite leaching is a function of the Cu²⁺ concentration. An empirical equation for the optimum redox potential for CuFeS₂ leaching is proposed.     

The performance,kinetics and microbiology of sulfidogenic fluidized-bed treatment of acidic metal- and sulfate-containing wastewater

A sulfidogenic fluidized-bed reactor (FBR) process was developed for treating acidic metal- and sulfate-containing wastewater. The process operating parameters were determined and the bacterial diversity of the FBR was described. The process was based on sulfate reduction by sulfate-reducing bacteria (SRB), precipitation of metals as sulfides with the biogenic H₂S and neutralization of the water with biologically produced bicarbonate alkalinity. The lactate- and ethanol-utilizing FBRs precipitated over 600 mg Zn l^− 1 day^− 1 and 300 mg Fe l^− 1 day^− 1 at a hydraulic retention time of 6.5 h. Metal precipitation was over 99.8% and effluent soluble Zn and Fe concentrations were below 0.1 mg l^− 1. Zinc and iron precipitated predominantly as ZnS, FeS₂ and FeS. The wastewater pH was increased from 2.5–3 to 7.5–8.5 during the treatment. Acetate oxidation was the rate-limiting step in ethanol oxidation. Ethanol oxidation was more affected by sulfide toxicity than was acetate oxidation. The FBR microbial communities contained SRB related to members of the genera Desulfovibrio, Desulfobulbus, Desulfotomaculum, Desulfobacca and Desulforhabdus, and also many species that do not reduce sulfate. The FBR communities contained many previously undescribed bacteria. This study demonstrated the feasibility of the sulfidogenic FBR for the concomitant removal of acidity, metals and sulfate from wastewaters.     

Leaching of waste battery paste components. Part 1: Lead citrate synthesis from PbO and PbO2

In this study, as part of developing a new process which can avoid smelting and electro-winning, citric acid based reagents in aqueous media were reacted with PbO and PbO₂. These two oxides are important components in the spent lead-acid battery paste and together account for up to 50% of the paste by weight. PbSO₄, the main component in a spent battery paste accounting for the remaining 50%, is dealt with in Part 2 in a separate paper. Reaction between PbO and C₆H₈O₇·H₂O or PbO₂ with a mixture of C₆H₈O₇·H₂O and H₂O₂ yielded lead citrate, Pb(C₆H₆O₇)·H₂O, which was characterised by XRD, SEM and FT-IR analysis. Optimal synthesis conditions were determined by investigating the effect of time, temperature, concentration, and the starting Pb oxide/water ratio. The optimal condition for leaching a mol of PbO at room temperature (20 °C) was found to be: 1 mol of (C₆H₈O₇)·H₂O solution; 1/3 as the starting PbO/water ratio and 15 min of reaction time. Pure citrate product, Pb(C₆H₆O₇)·H₂O was rapidly crystallized from the solution, in the leaching process. Leaching of PbO₂ required the use of a mild reducing agent. For each mole of PbO₂, the optimum condition at 20 °C was found to be: a solution containing 4 mol of C₆H₈O₇·H₂O and 2 mol of H₂O₂; 1/5 as the starting solid PbO₂/water ratio; and 60 min of reaction time. The product, as with PbO, was pure Pb(C₆H₆O₇)·H₂O compound. The remaining lead content of the filtrate solution was 0.017% and 1% corresponding to recoveries of 99.98% and 99% of lead as citrate, after the leaching/crystallization/filtration process with PbO and PbO₂, respectively. Asymmetric stretching vibrations between 1599 and 1662 cm^? 1, whereas symmetric vibrations between 1520 and 1327 cm^? 1 for lead citrate synthesised from PbO and asymmetric stretching vibrations between 1600 and 1642 cm^? 1 as well as symmetric vibrations between 1517 and 1326 cm^? 1 for the product obtained from PbO₂ revealed the strong IR adsorptions associated with a carboxylate structure. XRD data was identical to the well documented crystalline Pb(C₆H₆O₇)·H₂O compound from both the oxides. SEM revealed the formation of plate/sheet like morphologies. The difference in the column size of the Pb(C₆H₆O₇)·H₂O formed from the two lead oxides can be related to difference in the rate of the respective reactions.     

Cathodic reactions of Cu2+ in cupric chloride solution

In this study we demonstrate the kinetics of Cu²⁺ reduction in concentrated cupric chloride solutions. Experiments were carried out near the boiling point of the solution ([NaCl] = 280 g/l and [Cu²⁺] = 1–40 g/l) at T = 90 °C, atmospheric pressure, pH = 2. Electrochemical methods such as cathodic polarization curves and cyclic voltammetry were used to investigate the cathodic reactions of copper complexes. To identify the nature and the rate-controlling steps of the reactions, rotating disk electrode (RDE) experiments were conducted. The chemical environment studied was similar to that of the Outokumpu HydroCopper^TM process, which uses a cupric chloride solution to leach copper from the mineral chalcopyrite.The results suggest that the cathodic reactions are the reduction of [CuCl]⁺ to the complex [CuCl₃]²⁻, the reduction of [CuCl₃]²⁻ to solid copper and hydrogen evolution. The diffusion coefficient and the unit rate constants for the solution species were calculated. The exchange current density and rate constant for electron transfer were also estimated. A simulation was made of the cathodic polarization curve and it was in good agreement with the experimental data.     

Hydrothermal purification and enrichment of Chilean copper concentrates: Part 1: The behavior of bornite,covellite and pyrite

The nature and basic kinetics of the hydrothermal reactions of bornite, covellite and pyrite with copper sulfate solutions were investigated, as a previous study on the behavior of the bulk Chilean copper concentrates. The reaction of bornite produced digenite at < 160 °C and djurleite above this temperature. The reaction products formed a continuous, non porous layer. The rates were of zero order with respect [Cu²⁺]_(aq), and the activation energy was ～ 100 kJ/mol. Covellite was transformed to digenite at < 200 °C and to chalcocite (Q and M) at > 200 °C. The reaction was characterized for an irregular nucleation and growth of the products, which were non-protective. The order with respect the [Cu²⁺]_(aq) was 0.6 and the activation energy ～ 110 kJ/mol. The pyrite reaction was significant at > 200 °C and extensive at 240 °C. It produced a mixture of digenite/chalcocite-Q, with small amounts of djurleite. The transformation has a high Philling–Betworth ratio (～ 1.6), causing the break of the product layers. The reaction order and the activation energy were also 0.6 and ～ 110 kJ/mol, respectively. The rate of the bornite reaction should be controlled by solid-state ionic diffusion, whereas for covellite and pyrite the rates should be limited by electrochemical processes. Towards the hydrothermal transformation, the reactivity of the main phases present in Chuquicamata-type concentrates was: bornite > chalcopyrite > covellite > sphalerite > pyrite.     

The incongruent dissolution of scorodite — Solubility,kinetics and mechanism

This work reports the results of a laboratory investigation on the long-term stability of crystalline scorodite conducted at fixed pH (5–9) and temperature (22 °C, 50 °C and 75 °C). The scorodite used in this work was prepared via a hydrothermal synthesis procedure. The dissolution of scorodite at 22 °C was found to be extremely slow. At neutral pH, the arsenic concentration stabilized after 24 weeks at 5.9 mg/L. Analysis of the solubility data as a function of temperature yielded a scorodite solubility model equation. The solubility product of scorodite was recalculated as 10^− 25.4 using the solubility data generated by the study and the geochemical code PHREEQC for solution modelling. As scorodite dissolved, the iron re-precipitated as 2-line ferrihydrite. The growth and re-crystallization of ferrihydrite was apparently retarded by arsenate adsorption. The dissolution rate of scorodite was modeled with a decreasing exponential equation. The initial rate approached first order dependency on OH⁻ concentration while the apparent activation energy suggested that scorodite dissolution is chemically controlled.