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.     

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.     

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.     

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)     

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.     

Extraction of rare earths and yttrium with high molecular weight carboxylic acids

The extraction behaviour of trivalent rare earths namely La, Ce, Pr, Nd, Sm, Gd, Dy and Ho including Y (M(III), where M represents rare earths and yttrium ) from chloride medium has been studied with the solutions of high molecular weight carboxylic acids such as cekanoic, naphthenic, neo-heptanoic and Versatic 10 in dodecane. The effects of equilibrium pH, extractant concentration, metal ion concentration etc have been investigated. Using slope analysis technique it has been inferred that the metal ions form monomeric complex of the type [M(HA₂)₃] with carboxylic acids (H₂A₂, the dimer form). The stoichiometry of the species has also been confirmed using non-linear least square regression method. The carboxylic acids show different behaviour for Y extraction, it resembles to that of heavy rare earth (Ho), for sterically hindered acids (neo-heptanoic and Versatic 10) and to that with lighter rare earths (Ce, Pr) for the less sterically hindered acids (cekanoic and napthenic). The extraction order for the rare earths has been found to be the same with the four acids, i.e., La < Ce < Pr < Nd < Sm < Gd < Dy < Ho. The extraction constant (K_ex) of the systems and the separation factor amongst the rare earth pairs have been evaluated.     

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.     

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.     

Leaching kinetics of synthetic CaWO4 in HCl solutions containing H3PO4 as chelating agent

In this study, the dissolution kinetics of synthetically prepared CaWO₄ in HCl solutions containing H₃PO₄ was studied. The effects of process parameters such as stirring speed, temperature and acid concentrations on the dissolution rate of CaWO₄ were investigated. The reaction rate was found to be of 1 / 3 and 2 / 3 order with respect to HCl and H₃PO₄ concentrations, respectively, and the activation energy for the dissolution reaction to be 60 kJ mol^− 1. The rate equation for the dissolution reaction was derived using the Avrami equation and the rate determining step was the chemical reaction on the surface of solid particles.     

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.     

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.     

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.     

Studies on the recovery of uranium from phosphoric acid medium using synergistic mixture of (2-Ethyl hexyl) Phosphonic acid,mono (2-ethyl hexyl) ester (PC88A) and Tri-n-butyl phosphate (TBP)

This paper describes the extraction of uranium from aqueous phosphoric acid medium using (2-Ethyl hexyl) Phosphonic acid, mono (2-ethyl hexyl) ester (PC88A) and tri-n-butyl phosphate (TBP) individually as well as their synergistic mixture in different diluents. The various experimental parameters are investigated to optimize optimise the suitable extraction conditions. Results indicate that a synergistic mixture of 0.90 M PC88A + 0.15 M TBP in xylene, can be used for the extraction of uranium from low phosphoric acid medium. Back extraction studies reveals that among all the common strippants used, 0.50 M solution of (NH₄)₂CO₃ was most suitable. The synergistic mixture of 0.90 M PC88A + 0.15 M TBP as extractant system and 0.5 M (NH₄)₂CO₃ as strippant is used to recover uranium from a conditioned wet process phosphoric acid and from actual radioanalytical waste generated during uranium analysis by modified Davies–Gray method. The recovery is found to be around 80% from conditioned WPA whereas better than 90% from modified Davies–Gray waste.     

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.     

Density model for sodium hydroxide–sodium aluminate solutions

The purpose of this paper is to develop a model of the densities of multi-component aqueous electrolyte solutions containing NaOH and NaAl(OH)₄. Coefficients for the Laliberte–Cooper model of multi-component electrolyte solutions were developed from published density data for the NaOH–NaAl(OH)₄–H₂O system. The density data were split into two groups, data for parameterization and data for validation. The model was able to predict the validation data well, with an R² greater than 0.99 for five of seven datasets and greater than 0.95 for all datasets. The model was shown to extrapolate to temperature and composition ranges outside those used for model parameterization. Similarly, using coefficients derived here for NaAl(OH)₄ and published coefficients for NaOH and NaNO₂, the model was shown to accurately predict the density of solutions that contain NaNO₂ in addition to NaOH–NaAl(OH)₄ (R² = 0.993). This indicates that the model coefficients developed here can be incorporated into models of diverse multi-component electrolyte solutions.     

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).     

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.     

Removal/recovery of hydrochloric acid using Alamine 336, Aliquat 336, TBP and Cyanex 923

Solvent extraction of hydrochloric acid has been studied from solutions containing 185.42 g/L acid using Alamine 336, Aliquat 336, TBP and Cyanex 923 as extractants. Kerosene was used as the diluent. Extraction of hydrochloric acid increased with the increasing concentration of the extractants. For all the extractants studied, the species extracted into the organic phase appears to be associated with one mole of extractant. For Alamine 336, the McCabe–Thiele construction indicated the possibility of > 99.5% HCl extraction in two counter-current stages at A:O = 1:4. For Cyanex 923, the McCabe–Thiele construction indicated the quantitative extraction of HCl in four counter-current stages at the A:O = 1:5 phase ratio. The acid loaded onto Aliquat 336, TBP and Cyanex 923 could be easily stripped with water. But acid loaded onto Alamine 336 could not be stripped with water. The stripping efficiency was ∼ 12% with dilute acid and quantitative with 1 N NaOH for acid loaded onto Alamine 336.     

Stripping of copper from CYANEX® 301 extract with thiourea–hydrazine–sodium hydroxide solution

The stripping of copper from the organic extract of bis(2,4,4-trimethylpentyl) phosphinodithioic acid, CYANEX® 301, using an aqueous mixture of thiourea, hydrazine and sodium hydroxide has been investigated. The optimal concentrations of the aqueous solution were found to be 1.0 M, 5 × 10^− 2 M and 5.0 M, respectively which led to 95% stripping of copper from CYANEX 301 with concomitant regeneration of the extractant. The characterization of the stripped copper product was done using a combination of microanalyses, cyclic voltammetry and X-ray diffractometry (XRD). The product was made up of two non-stoichiometric copper sulfides, Cu_xS with x = 1.60 and 1.77. Cu_1.60S was the major product. Hydrazine appears to play the role of reducing the disulfide species of CYANEX 301, R₂P(S)S–S(S)PR₂, formed during the extraction step, back into CYANEX 301, whereas thiourea provides a source of sulfur in the formation of the stripping products.     

Separation and preconcentration of trace indium(III) from environmental samples with nanometer-size titanium dioxide

This work assesses the potential of an adsorptive material, nanometer TiO₂, for the separation and preconcentration of trace indium ions from various aqueous media. The adsorption behavior of nanometer TiO₂ for indium ions was investigated. It was found that the adsorption percentage of the indium ions was more than 96% in pH 3.5–4.0, and the desorption percentage of In(III) ions was more than 99% in pH ≤ 1.5. Good relative standard deviate (1.5%) and lower analytical detection limit (0.45 µg?mL^? 1) were obtained. The adsorption equilibrium was well described by the Langmuir isotherm model with monolayer adsorption capacity of 4566 µg g^? 1 (25 °C). The accuracy of the method is confirmed by analyzing the certified reference material (GBW-07405, GBW07406). The results demonstrated good agreement with the certified values.