Solvent extraction of Ti(IV), Fe(III) and Fe(II) from acidic sulfate medium with di-o-tolyl phosphoric acid-benzene — hexan-1-ol system: A separation and mechanism study

The extraction of Ti(IV), Fe(III) and Fe(II) with di-o-tolyl phosphoric acid (HDTP, HA)-benzene-20% hexan-1-ol system was studied as function of contact time, concentrations of extractant in the organic phase and of H₂SO₄ in the aqueous phase, and temperature. The order of extractability under identical conditions is: Ti(IV) > Fe(III) ? Fe(II). The relative separations of these metal ions are also dependent on the above four factors. The maximum values of the separation factors, $\text{β}{}_1\text{=}\text{E}{}_{\mathrm{<mtext>Ti(IV)</mtext>}}{}^0\text{E}{}_{\mathrm{<mtext>Fe(IlI)</mtext>}}{}^0\text{= 175}$ at 0.10 M HDTP and 3.50 M H₂SO₄ concentrations, and $\text{β}{}_2\text{=}\text{E}{}_{\mathrm{<mtext>Ti(IV)</mtext>}}\text{E}{}_{\mathrm{<mtext>Fe(II)</mtext>}}{}^0\text{= 7800}$ at 0.10 M HDTP and 0.50 M H₂SO₄ concentrations, indicate that the separation of Ti(IV) from iron seems to be promising if iron is present in the divalent state. The mechanisms of extraction are discussed.     

Equilibria associated with cupric chloride leaching of chalcopyrite concentrate

The leaching of chalcopyrite in aqueous cupric chloride solutions is treated with a thermodynamic model which incorporates the equilibria associated with the formation of various cupric, cuprous and ferrous chloride complexes. The precipitate from a refluxing slurry of elemental sulfur and aqueous cuprous and ferrous chlorides contained predominantly covellite (CuS) and a small amount of pyrite (FeS₂). Mössbauer spectra indicate that there was no chalcopyrite formation. Cupric chloride leaching experiments on both chalcopyrite concentrate and cupric sulfide indicated that the extent of cuprous ion formation in these aqueous cupric chloride solutions is controlled by the thermodynamics of the equilibrium:

$\text{CuS+Cu}{}^{\mathrm{2+}}\text{? 2 Cu}{}^{\mathrm{+}}\text{+S}{}^0$

Conditions which favor the desired complete reduction of cupric ion are discussed.     

The effect of pH on cadmium extraction by Lix 34

The pH dependence of the extraction of cadmium from nitrate solution using LIX 34 diluted with Kermac 470B has been studied. The optimum pH for cadmium extraction in the system studied is in the neighborhood of 8.3. Below pH 7, the reagent forms a two-to-one ligand to cadmium ion chelate, whereas, at pH greater than 7, the extraction mechanism is extremely complicated due to the presence of NH₃, NH₄⁺ and some other unknown buffers. Moreover, the difference between the initial and equilibrated aqueous pH values, ΔpH, can be estimated by the following equation

$\text{Delta;}\text{pH}\text{=}\text{log}\text{1}\text{1 +(i?2)}\text{[}\text{cd}\text{]}{}_0\text{[}\text{NH}{}_3\text{]}{}^{\mathrm{<mtext>I</mtext>}}$

where i is the average co-ordination number of cadmium by ammonia in the aqueous phase. [Cd]₀ represents the equilibrium cadmium concentration in the organic phase and [NH₃]^I denotes the initial aqueous ammonia concentration.     

The extraction of nickel from ammoniacal media and its separation from copper,cobalt and zinc using hydroxyoxime extractants I. SME-529

Nickel extraction from ammoniacal media, containing, initially, 0.03 mol dm^?3 of nickel, by 10% v/v Shell metal extractant 529 in MSB 210 or Shellsol A, has been studied as a function of pH, counter anions (nitrate or sulphate), ammonium salt concentration, temperature (10 to 50°C), and the presence of copper, cobalt and zinc. The extraction of copper as a function of pH was studied in the presence of nickel. The pH for maximum nickel extraction decreased from 8.2 to 7.4 as the ammonium sulphate concentration increased from 0.1 to 3.4 mol dm^?3. Nickel extraction was lower from sulphate than from nitrate media and with Shellsol A, and at higher temperature. The enthalpy and entropy of extraction were negative. The presence of copper, cobalt and zinc decreased nickel extraction. Separation factors between copper and nickel are reported as a function of pH and indicate that initial removal of copper is desirable.From a consideration of the equilibrium, the extraction of nickel (with MSB 210 diluent) at 25°C could be expressed by the equation:

$\text{log}\text{D}{}_{\mathrm{<mtext>Ni</mtext>}}\text{= ?9.25 +}\text{log}\text{α}{}_0\text{+ 2}\text{log}\text{(C}{}_{\mathrm{HR}}\text{?2[}\text{Ni}\text{]}{}_{\mathrm{<mtext>org</mtext>}}\text{) + 2}\text{pH}$

where D_Ni represents nickel distribution coefficient, C_HR the total extractant concentration, [Ni]_org the organic nickel concentration (both in mol dm^?3) and α₀, the fraction of uncomplexed nickel in the aqueous phase calculated from the overall stability constants of Ni(NH₃)_i²⁺ complexes and the formation constant of NiSO₄ ion pairs.The limiting slope of ?1 for long D_Ni versus log [NH₃]_free plots indicated an average co-ordination number of 3 for nickel by ammonia in the aqueous phase.The uptake of SO₄^2?, NO₃^? and water by the organic phase was negligible. The uptake of ammonia was dependent upon pH and the ammonia content of the aqueous phase, and was negligible in the absence of metal.     

Extraction equilibrium and extraction kinetics of nickel from aqueous ammonium nitrate solution with versatic 10 in n-hexane

The mechanism of extraction of nickel from an aqueous ammonia-ammonium nitrate mixture by an n-hexane solution of Versatic 10 was investigated from the viewpoint of extraction equilibrium and extraction kinetics. From the equilibrium study, it was found that nickel is extracted according to the following two extraction reactions which are in accordance with the loading ratio of nickel to Versatic 10:

$\text{Ni}{}^{\mathrm{2+}}\text{+ 3 H}{}_2\text{R}{}_{\mathrm{2org}}\text{?NiR}{}_2\text{·4 HR}{}_{\mathrm{org}}\text{+2 H}{}^{\mathrm{+}}{}_{\mathrm{aq}}$

in the range of lower loading ratio, and

$\text{2 Ni}{}^{\mathrm{2+}}{}_{\mathrm{aq}}\text{+ 4 H}{}_2\text{R}{}_{\mathrm{2org}}\text{?(NiR}{}_2\text{·2 HR)}{}_{\mathrm{2org}}\text{+4 H}{}^{\mathrm{+}}{}_{\mathrm{aq}}$

in the range of higher loading ratio. The equilibrium constants for each reaction were also determined. From the study of extraction kinetics, the extraction rate was found to be proportional to the total concentrations of nickel and Versatic 10 and inversely proportional to that of hydrogen ion. A reaction mechanism is proposed in order to give a reasonable explanation for the observed concentration dependence of each of the reactant species.     

A study of chalcopyrite dissolution in acidic ferric nitrate by potentiometric titration

A pure chalcopyrite sample was studied using a potentiometric titration technique, which measures the oxidant consumption in the dissolution of sulphide mineral slurries in acidic ferric ion solutions. Rate measurements were made over a period of days at constant solution pH and oxidation potential at 25 and 40°C. Long term leaching followed the parabolic kinetics and dissolution stoichiometry previously described in the literature, but there was an initial reaction in which more iron than copper dissolved from the lattice and a metal-deficient “chalcopyrite-like” surface phase was rapidly formed. The subsequent dissolution reaction (observed specific reaction rate, $\text{s}{}_0\text{= 1.4 × 10}{}^{\mathrm{?11}}\text{mol cm}{}^{\mathrm{?2}}\text{s}{}^{\mathrm{<mtext>?</mtext><mtext>1</mtext><mtext>2</mtext>}}$ at 25°C; activation energy, E_a = 14 kcal mol^?1) was far more sensitive to temperature change and much slower than the rate of pore diffusion of oxidant ($\text{s}{}_0\text{≈ 10}{}^{\mathrm{?7}}\text{mol cm}{}^{\mathrm{?2}}\text{s}{}^{\mathrm{<mtext>?</mtext><mtext>1</mtext><mtext>2</mtext>}}$, E_a ≈ 2 kcal mol^?1) and was interpreted in terms of a solid state diffusion step within the crystal lattice.     

Kinetic and mechanistic study of FeS dissolution in aqueous sulfur dioxide solution

FeS leaching was carried out in aqueous sulfur dioxide solution at low temperature and pressure (T = 60°CC;P_SO2 = 51.7 kPa) using both a constant-pressure gas-uptake apparatus and conventional glass-leaching equipment. In the initial stages of the leaching reaction the ratio [SO₂]/[Fe²⁺] was found equal to 1.3. This suggests a stoichiometry given by the following sequence of equations

$\text{FeS + SO}{}_2\text{·H}{}_2\text{O ? FeS + H}{}^{\mathrm{+}}$

$\text{FeS · HSO}{}_3{}^{\mathrm{?}}\text{+ H}{}^{\mathrm{+}}\text{→ Fe}{}^{\mathrm{2+}}\text{+ H}{}_2\text{S + SO}{}_3{}^{\mathrm{2?}}$

As the reactions progressed a build up of various oxy-sulfur species was observed.The leaching process appears to proceed by two concurrent dissolution paths, namely acid dissolution and aquated sulfur dioxide dissolution. In both cases the rate determining step (r.d.s.) involves a surface desorption reaction. In the constant leaching region, the kinetics can be represented by the following rate expression:

At low [H+] relative to [SO₂-H₂O], path I predominates. This can be represented by:

$\text{|}{}_{\mathrm{s}}\text{FeS +SO}{}_2\text{· H}{}_2\text{O ? |}{}_2\text{FeS SO}{}_2\text{· H}{}_2\text{O}$

$\text{|}{}_{\mathrm{s}}\text{FeS SO}{}_2\text{· H}{}_2\text{O ? |}{}_{\mathrm{s}}\text{FeS HSO}{}_3{}^{\mathrm{?}}\text{+ H}{}^{\mathrm{+}}$

$\text{|}{}_{\mathrm{s}}\text{FeS HSO}{}_3{}^{\mathrm{?}}\text{→ Fe}{}^{\mathrm{2+}}{}_{\mathrm{(aq)}}\text{r.d.s.}$

At high [H⁺] relative to [SO₂·H₂O], path II predominates. This can be represented by:

$\text{|}{}_{\mathrm{s}}\text{FeS + H}{}^{\mathrm{+}}\text{? |}{}_{\mathrm{s}}\text{FeS H}{}^{\mathrm{+}}$

$\text{|}{}_{\mathrm{s}}\text{FeS H}{}^{\mathrm{+}}\text{→ Fe}{}^{\mathrm{2+}}{}_{\mathrm{(aq)}}\text{r.d.s.}$

The apparent activation energy was found to be 41.6 kJ per mole.     

Reaction kinetics for the leaching of MnO2 by sulfur dioxide

The reaction kinetics for the leaching of MnO₂ by sulfur dioxide have been studied using studied using monosize particles at dilute solid phase concentrations in a stirred reactor to determine the important chemical factors which govern the kinetic response of the system. The conclusion that the reaction rate is limited by a chemical reaction at the MnO₂ surface is supported by: (1) an apparent activation energy of 35.9 kJ/mol (8.6 kcal/mol), (2) the inverse first-order relationship between the rate constant k and the initial particle diameter, (3) the independence of the reaction rate on stirring speed and, more importantly, the magnitude of the calculated reaction velocity constant (～10^?3 cm/s compared to predicted mass-transfer coefficients of 10^?2 cm/s), and (4) the one-half order dependence of the reaction rate with respect to the SO₂ concentration.The rate-limiting step is considered to be an electrochemical surface reaction, a conclusion which is substantiated by electrode half cell potential measurements. Using the Butler-Volmer equation, a theoretical analysis of the electrochemical reaction resulted in the following rate equation:

which was consistent with the experimental results.     

The extraction of mercury(II) from concentrated HBr solutions by benzene. A comparison of various representations of single-ion activity coefficients in concentrated aqueous solutions

The extraction of tracer Hg(II) from aqueous HBr solutions can be described by the reaction

$\text{HgBr}{}_2\text{(aq)}\text{?}\text{HgBr}{}_2\text{(org)}$

This reaction has been used to test various conventions for the single-ion activity of bromide ion as well as chloride ion using data already published. It is found that {X^?} = y_x?C (y_x? given by a simple semiempirical equation cf. below) and {X^?} = C give a good fit, but the popular function {X^?} = y_±C does not work at all. It thus seems that it is better to use the stoichiometric concentration than its product with the mean ionic activity coefficient, and this simple convention for ligand activity might be a useful starting point when evaluating complex formation in concentrated electrolytes.The extraction of HgBr₂ by TLAHBr (TLA = trilaurylamine) dissolved in benzene can be fairly well described by the reaction

$\text{TLAHBr(org) + HgBr}{}_2\text{(aq)}\text{?}\text{TLAHBr·HgBr}{}_2\text{(org)}$

and the stoichiometric distribution coefficient D can be given by D = KD° (α_1,0 + α_1,0) where K is a constant and α_1,0 + α_1,1 the fraction of TLAHBr not containing acid in excess of the 1:1 composition.     

The anodic dissolution of ZnS electrodes in sulfuric acid solutions

The anodic dissolution of reagent grade ZnS and ZnS concentrate has been studied. The majority of the research was conducted in H₂SO₄ although a limited number of tests were made in HCl and KOH. Polarization studies showed that both Zn²⁺ and S were products in the acidic solutions. ZnS passivated in KOH solutions. The electrodes were fabricated by hand pressing mixtures of ZnS + pitch (5–15%) and sintering at 800°C in a N₂ atmosphere. The open circuit potentials were considered to be mixed potentials resulting from the anodic dissolution of ZnS and cathodic reduction of S or O₂. Current efficiencies and $\text{Zn}{}^{\mathrm{2+}}\text{S}$ ratios were determined at 0.5 and 0.85 V vs. SHE. The results indicated the occurrence of both electrochemical and chemical dissolution steps as well as the further oxidation of H₂S, namely,

$\text{ZnS(s)=Zn}{}^{\mathrm{2+}}\text{(aq)+S(s)+2e}$

$\text{ZnS(s)=2H}{}^{\mathrm{+}}\text{(aq)=Zn}{}^{\mathrm{2+}}\text{(aq)+H}{}_2\text{S(aq)}$

$\text{H}{}_2\text{S(aq)+4 H}{}_2\text{O(I)=HSO}{}_4\text{(aq)+9H}{}^{\mathrm{+}}\text{(aq)+8e}$

The overall dissolution appears to be mass transfer-limited, probably either by the diffusion of Zn²⁺ from the reaction interface, through the reacted layer to the bulk solution, or the dissolution of precipitated Zn(OH)₂ in the reacted layer by the diffusion of H⁺ into the layer. The concentrate anodes polarized more drastically than the reagent grade ZnS anodes, possibly due to the presence of PbS impurity which would form PbSO₄ in the anode pores.     

The oxidation of Fe(II) in aqueous sulphuric acid under oxygen pressure

An experimental investigation is presented on the oxidation of Fe(II) in aqueous sulphuric acid under oxygen pressure. The effects of partial oxygen pressure, p_O2, temperature and initial concentrations of H₂SO₄ and Cu(II) on the oxidation process rate have been determined within the ranges applied in a new hydrometallurgical process for copper recovery from its sulphide concentrates [13–16]. The Fe(II) oxidation reaction rate was found to be described by a second-order kinetic equation at a ferrous iron concentration exceeding 3–8 g/l:

     

Fluidised-bed anodic dissolution of covellite

The anodic dissolution of synthetic covellite (CuS) has been investigated using a fluidisedbed anode technique. In sulphuric acid electrolyte the dissolution of CuS is accompanied by excessive oxygen evolution at the Pt feeder electrode unless the applied current is maintained at unrealistically low values. In mixed electrolytes, such as sulphuric acid-sodium chloride or sulphuric acid-potassium bromide, the anodic dissolution of CuS proceeds at acceptable values of applied current via a charge transfer mechanism provided by the $\text{Cl}{}^{\mathrm{?}}\text{-}\text{1}\text{2}\text{Cl}{}_2\text{or Br}{}^{\mathrm{?}}\text{-}\text{1}\text{2}\text{Br}{}_2$ redox couple. The results of the present study are compared with previous work on the fluidised-bed anodic dissolution of chalcocite (Cu₂S).     

Galvanic interaction between chalcopyrite and pyrite during bacterial leaching of low-grade waste

The bacterial leaching of a low-grade chalcopyrite waste rock in a lixiviant containing thermophilic, Sulfolobus-like microorganisms at 60°C and a lixiviant containing Thiobacillus ferrooxidans at 28°C has been compared with the leaching in sterile lixiviant in terms of copper solubilized in elapsed time and the conversion of $\text{Fe}{}^{\mathrm{3+}}\text{Fe}{}^{\mathrm{2+}}$. Bacterial action has been shown to drastically increase the ratio $\text{Fe}{}^{\mathrm{3+}}\text{Fe}{}^{\mathrm{2+}}$ with elapsed time of leaching. Direct observations of the associated pyrite and chalcopyrite surface corrosion, utilizing scanning electron microscopy, showed that during the leaching of these sulfides as separate, non-contacting phases, the pyrite corroded more rapidly than the chalcopyrite in both sterile and inoculated media. This effect was more pronounced at elevated temperature and in the presence of bacteria. When the pyrite and chalcopyrite were in contact, the resulting galvanic interaction caused the chalcopyrite to corrode more rapidly than the pyrite, which was effectively passivated. The leaching of chalcopyrite is thereby enhanced in contact with pyrite. This effect is accelerated in the presence of bacteria. The corrosion of chalcopyrite was also markedly enhanced as a result of the oxidation of elemental sulfur (formed during the reaction) to sulfuric acid. This reaction was also accelerated by bacterial catalysis. The important implications of the enhanced chalcopyrite corrosion by galvanic interaction in the leaching of low-grade chalcopyrite waste and other galvanic-contact regimes involving metal sulfides are identified and discussed.     

Single cell extraction of zinc and manganese dioxide from zinc sulphide concentrate and manganese ores

The hydrometallurgical processing of zinc sulphide concentrates with sulphuric acid in the presence of manganese dioxide (manganese ore has been employed) and subsequent electrolytic co-deposition of cathodic zinc metal and anodic manganese dioxide is described.The influence of various parameters on the reaction

$\text{ZnS + MnO}{}_2\text{+ 2H}{}_2\text{SO}{}_4\text{= ZnSO}{}_4\text{+ MnSO}{}_4\text{+ S}{}^0\text{+ 2H}{}_2\text{O}$

has been studied. Optimum conditions for rapid and efficient reaction have been determined.The simultaneous electrowinning of zinc at the cathode and γ-MnO₂ at the anode from the leach liquor was studied. The effects of variation of current density, temperature, electrolyte composition etc. have been described in detail. During leaching 99% extraction of zinc, 98% of manganese, and 96% liberation of elemental sulphur was achieved. 80–90% anodic and cathodic current efficiencies can be obtained under optimum conditions with impurity levels of only a trace of manganese in the zinc deposit and vice-versa.The anodically deposited manganese dioxide was the γ-battery active variety and was found to be satisfactory.The results indicate the potential for the development of a technique for zinc and manganese dioxide production in a single cell.     

The kinetics of the sulphuric acid leaching of nickel and magnesium from reduction roasted serpentine

Nickel may be extracted with partial selectivity over magnesium from laterites containing serpentine by reduction roasting followed by sulphuric acid leaching. This paper describes the results of a kinetic study of the sulphuric acid leaching of nickel and magnesium from the reduction roasted serpentine component of a typical laterite. The serpentine used in this work analyzed 1.65% nickel, 20.2% magnesium and 6.10% iron.Initially, leaches were carried out at temperatures of 30, 50 and 70° C to determine the acid requirement for complete nickel extraction using practical leaching conditions (25% solids) under which the acidity drops to a low level. A minimum acid addition of 60 g/l was needed, which gave 80% to 83% nickel extraction from material in which 85% of the nickel was reduced using hydrogen at 700° C. Under these conditions, about 17% of the magnesium was leached at each of the temperatures studied.To facilitate an understanding of leaching kinetics, leaches were then performed using constant acidities (0.1% solids) of 60, 30 and 15 g/l acid at temperatures of 30, 50 and 70° C. Closely sized particles (?65 + 100 mesh) were leached so that magnesium dissolution rates could be tested against established “shrinking core” models.The main conclusions are that, under the experimental conditions, nickel dissolution rates were chemically controlled by either $\text{2}\text{H}{}^{\mathrm{+}}\text{+}\text{1}\text{2}\text{O}{}_2\text{+ 2e →}\text{H}{}_2\text{O}$ or 2 H⁺ + 2e → H₂ occurring at the surfaces of the 40% nickel/iron alloy platelets formed during reduction. The rate controlling process had an activation energy of 11kcal/mole.During extraction of metallic nickel, the dissolution of magnesium follows a “shrinking core” kinetic model, which assumes the reaction is unimpeded by a surface layer of silica - a reaction product. Rates of magnesium dissolution during this stage of the reaction (up to 25% dissolved) were chemically controlled with an activation energy of 12 kcal/mole.At magnesium extractions above 25%, at which point all the metallic nickel had been extracted, the rate of silicate attack was limited by diffusion through a silica coating attached to particle surfaces.The results indicate that selectivity for nickel dissolution over that of magnesium does not depend strongly on acidity and temperature at the levels investigated.