Preparation of layered nickel aluminium double hydroxide from waste solution of nickel

The separation of nickel has been carried out from a waste solution containing 3.18 g/L Ni with other impurities such as Fe, Zn, Cu and As. Iron was removed by precipitation and Cu and Zn were removed by solvent extraction using LIX 622N and NaTOPS-99, respectively. After removal of all these impurities nickel was extracted by 1.5 M NaTOPS-99 in two counter-current stages at A:O ratio of 3:1 and the loaded organic was stripped with 30 g/L H₂SO₄ at phase ratio of unity. The strip solution of nickel was treated with Al₂(NO)₃ · 9H₂O for co-precipitation by increasing the pH of solution with 1 M NaOH up to 10. The Ni–Al layered double hydroxide was confirmed through XRD characterization.     

Thorium and uranium extraction from rare earth elements in monazite sulfuric acid liquor through solvent extraction

This paper describes the process of extraction of thorium and uranium from the sulfuric liquor generated in the chemical monazite treatment through a solvent extraction technique. The influence of the extractant type and concentration, contact time between phases, type and concentration of the stripping solution and aqueous/organic volumetric ratio were investigated. The results indicated the possibility of extracting, simultaneously, thorium and uranium from a solvent containing a mixture of Primene JM-T and Alamine 336. The stripping was carried out with a hydrochloric acid solution. After selecting the best conditions for the process, a continuous experiment was carried out in a mixer-settler circuit using four stages in the extraction step, five stages of stripping and one stage of the solvent regeneration. A loaded stripping solution containing 29.3 g/L of ThO₂ and 1.27 g/L of U₃O₈ was obtained. The metals content in the raffinate was below 0.001 g/L, indicating a thorium extraction of over 99.9% and a uranium extraction of 99.4%. The rare earths content in the raffinate was 38 g/L of RE₂O₃.     

Stripping conditions to prevent the accumulation of rare earth elements and iron on the organic phase in the solvent extraction circuit at Skorpion Zinc

The concentrations of the rare earth elements (particularly Y, Yb, Er and Sc) in the zinc-stripped organic phase streams in the solvent extraction process at Skorpion Zinc mine have increased gradually over the past four years. Iron is the only other impurity present in notable quantities in the organic phase after washing and scrubbing prior to zinc stripping. This project aimed to evaluate the effects that rare earth elements and iron in the organic phase have on the zinc solvent extraction process and to subsequently find appropriate stripping conditions for the removal of these elements from the zinc-stripped organic phase.Results obtained by performing laboratory scale batch tests indicated that the viscosity of the organic phase doubled and the phase disengagement time increased from 100 s to 700 s when the total rare earth elements and iron concentration in the organic phase was increased from 3100 mg/L to 6350 mg/L. The zinc loading capacity of the organic phase after two extractions furthermore decreased by a value between 1 g/L and 3 g/L, depending on the composition of the pregnant leach solution. The stripping of low concentrations of rare earth elements and iron from 40% di-2-ethylhexyl phosphoric acid (D2EHPA) diluted in kerosene was evaluated using two different stripping agents (H₂SO₄ and HCl) with concentrations between 1 M and 7 M, organic-to-aqueous ratios between 0.5 and 4, and temperatures between 30 °C and 50 °C. The highest stripping percentages were achieved at acid concentrations greater than 5 M, organic-to-aqueous ratios of less than 0.5 and high temperatures.     

Extraction of titanium (IV) from acidic media by tri-n-butyl phosphate in kerosene

The extraction of titanium (IV) from sulfate, and nitrate solutions has been studied using tri-n-butyl phosphate (TBP) in kerosene. Extraction of titanium was affected by acid concentration over the range of 0.5–4 mol L^?1. The titanium distribution coefficient reached a minimum between 1 and 2 mol L^?1 acid for both sulfate and nitrate solutions. Third phase formation was observed in the extraction of titanium from acidic media at all condition tested. At the next stage, the stripping of titanium was studied using H₂SO₄, H₂SO₄ + H₂O₂ and Na₂CO₃. The kinetics of the stripping were very slow for H₂SO₄. The use of complex forming stripping agents (H₂SO₄ + H₂O₂) and Na₂CO₃ significantly improved the kinetics of stripping. About 98% recovery was achieved by extracting titanium from an aqueous nitrate solution using TBP and stripping with sodium carbonate.     

Solvent extraction and separation of palladium(II) and platinum(IV) from hydrochloric acid medium with dibutyl sulfoxide

The solvent extraction and separation performances of Pd(II) and Pt(IV) from hydrochloric acid solutions were investigated using dibutyl sulfoxide (DBSO) diluted in kerosene. Pd(II) was strongly extracted by a lower concentration DBSO in a lower concentration hydrochloric acid solution while the reverse was obtained for Pt(IV) extraction. Based on independent extraction and separation experiments of Pd(II) and Pt(IV), the separation parameters of Pd(II) and Pt(IV), including dibutyl sulfoxide concentration, contact time of aqueous and organic phases, organic/aqueous (O/A) phase ratio and H⁺ concentration of aqueous phase, were studied in detail, and the optimal separation parameters were obtained and summarized as the following: dibutyl sulfoxide concentration 0.6–1.2 mol dm^?3, organic/aqueous (O/A) phase ratio 0.6–1.0, H⁺ concentration of aqueous phase 1.0–1.5 mol dm^?3 and contact time of two phases 5 min. The as-prepared separation parameters were corroborated by the extraction and separation from a synthetic stock solution containing Pd(II), Pt(IV) as well as several common impurities like Fe(II), Cu(II) and Ni(II). The results revealed that Pd(II) could be separated efficiently from Pt(IV) with a high separation coefficient of Pd(II) an Pt(IV) (2.7 × 10⁴) by predominantly controlling dibutyl sulfoxide and hydrochloric acid concentrations. The extraction saturation capacity of Pd(II) was determined from 1.0 mol dm^?3 HCl solution with 3 mol dm^?3 dibutyl sulfoxide and its experimental value exceeded 14 g dm^?3 under the experimental conditions.Stripping of Pd(II) from loaded organic phase was performed using a mixed aqueous solution containing NH₄Cl and ammonia solutes. Pd(II) (99.2%) was stripped using the stripping solution containing 3% (m/v) NH₄Cl and 5 mol dm^?3 ammonia, respectively.     

Recovery of copper,nickel and cobalt from acidic pressure leaching solutions of low-grade sulfide flotation concentrates

Recovery of copper, nickel and cobalt from the acidic pressure leaching solutions of Jinbaoshan (YN Province, PRC) low-grade sulfide flotation concentrates was investigated. The proposed technique includes four major steps: (1) the acidity adjustment of the acidic pressure leaching solutions; (2) solvent extraction (SX) separation of copper by organic reagent XD5640, and then stripped from the loaded organic phase by H₂SO₄ solution for copper recovery; then (3) iron in raffinates after copper extracting is selectively removed by high-temperature hydrolysis precipitation in an autoclave; and lastly (4) nickel and cobalt are selectively precipitated by Na₂S from the final solutions after removing iron. The experimental results for treating 1 L acidic leaching solutions per batch by this new technique were reported, and some evaluation and further comparisons with previous investigations were also carried out. It was reported that the total percent recovery of Cu could reach 95% or more, and that of Ni and Co were all more than 99%. In the processing, the percent removal of impurities, such as Fe, Mg and Ca, were all also near to 99%.     

Solvent extraction of tetravalent hafnium from acidic chloride solutions using 2-ethyl hexyl phosphonic acid mono-2-ethyl hexyl ester (PC-88A)

Solvent extraction of Hf(IV) from acidic chloride solutions has been carried out with PC-88A as an extractant. Increase of acid concentration decreases the percentage extraction of metal indicating the ion exchange type mechanism. The plot of logD vs log[extractant], M is linear with slope 1.8 indicating the association of two moles of extractant with the extracted metal species. Plot of logD vs log[H⁺] gave a straight line with a negative slope of ∼2 indicating the exchange of two moles of hydrogen ions for every mole of Hf(IV). The effect of Cl⁻ ion concentration at constant concentration of [H⁺] did not show any change in D values. Addition of sodium salts enhanced the percentage extraction of metal and follows the order NaSCN > NaCl > NaNO₃ ∼ Na₂SO₄. Stripping of metal from the loaded organic (LO) with different acids indicated sulphuric acid as the best stripping agent. Regeneration and recycling capacity of PC-88A, temperature, extraction behavior of associated elements was studied.     

Leaching behaviour of natural and heat-treated brannerite-containing uranium ores in sulphate solutions with iron(III)

Uranium leaching tests were conducted on two naturally occurring, highly metamict brannerite ores from the Crockers Well and Roxby Downs deposits, South Australia. The ores were leached over a range of temperatures and Fe^(III) and H₂SO₄ concentrations. As well, samples of the ores were calcined at 1200 °C in air to investigate the effect of thermally induced recrystallisation on uranium dissolution. For the unheated samples, a maximum of ∼80% U dissolution was obtained using an Fe^(III) concentration of 12 g/L, an acid concentration of 150 g/L H₂SO₄ and a temperature of 95 °C. The heat treated samples performed poorly under identical conditions, with maximum uranium dissolution of <10% recorded. High uranium dissolution from natural brannerite can be achieved providing; (i) acid strength, oxidant strength and temperatures are maintained at elevated levels (compared to those traditionally used for uraninite leaching), and, (ii) the brannerite has not undergone any significant recrystallisation (e.g. through metamorphism).     

Recovery of boron from borax sludge of boron industry

A two-step process for boron recovery from borax sludge is proposed in the present work. The borax sludge was leached with sulphuric acid solution. Then, for the removal of alkaline species from the leachate, calcium and magnesium were precipitated by adjusting the pH of leachate using 1.5 M NaOH and 1.5 M HCl solutions. The effects of pH, temperature, concentration and time on the precipitation process were investigated. It was determined that the calcium and magnesium concentrations in the leachate were reduced from 121 mg/L to 2.6 mg/L and from 145 mg/L to 4.0 mg/L, respectively, at a pH value of 12, a temperature of 70 °C, an initial boron concentration of 500 mg/L and a precipitation time of 3 h. Under these optimum conditions, it was observed that the boron concentration in the solution decreased very slightly. In this process, the alkaline species were successfully separated from the boron.Finally, borax pentahydrate (Na₂B₄O₇·5H₂O) was produced by the evaporation of the final solution obtained after precipitation process.     

Recycling of waste pickle acid by precipitation of metal fluoride hydrates

Stainless steel is pickled in mixed acid solutions (1–3 M HNO₃ and 0.5–4 M HF). The spent solution is usually neutralized with lime, and in Sweden about 18,000 tons/yr of metal hydroxide sludge is disposed as landfill waste. We are developing a cost-saving and environmentally friendly process, involving crystallization of β-FeF₃ · 3H₂O, where the metal content is recovered and the acid is recycled.Iron has been successfully separated from spent pickle bath solutions by precipitation of β-FeF₃ · 3H₂O in a continuous crystallizer (10 L scale) where the solution is concentrated by nanofiltration. The crystal growth rate of β-FeF₃ · 3H₂O has been determined in industrial pickle bath solutions at 50 °C and the results have been compared to previous measurements in pure HF/HNO₃ solutions prepared in the laboratory. The growth rate of β-FeF₃ · 3H₂O crystals at 50 °C is in the order of 10⁻¹¹ m/s in both industrial and pure acid mixtures.     

Oxidation of Fe(II) in a synthetic nickel laterite leach liquor with SO2/air

Under specific controlled conditions, the addition of SO₂ to oxygen or air produces the peroxy-monosulphate free radical in solution, which is a stronger oxidant than oxygen alone. In this study, the practical strategies required to optimise the oxidation of Fe(II) with SO₂/air was investigated at 75 °C as part of a process to remove iron as Fe(III) oxides from a synthetic nickel laterite high pressure acid leach solution containing 5 g/L Fe(II), 1 g/L Fe(III), 8 g/L Ni, 30 g/L Mg in sea water at pH about 2. The rate of Fe(II) oxidation was optimised in the pH range of 1.2–2.0 with respect to SO₂/air ratio and gas flow rates for minimum production of H₂SO₄ and maximum utilisation of SO₂. In order to minimise the air flow rates into the reactor vessel, the maximum rate of SO₂ addition that could be employed with air was established whilst maintaining oxidising conditions. The results provide strategies for commercial applications of the SO₂/air oxidising system and indicate important factors for reactor design.     

An environmentally-friendly technology of vanadium extraction from stone coal

A novel technology characterized by higher recovery of vanadium and which was environmentally-friendly was developed to recover vanadium from stone coal. Vanadium in stone coal could be leached by NaOH solution after roasting stone coal at 850 °C for 3 h. H₂SO₄, Mg(NO₃)₂ and ammonia were employed, respectively, in two steps to remove the impurities of Si and Al from the leach liquor. After extracting vanadium from the leach liquor with 10 vol% N₂₃₅, 20 vol% secondary octyl alcohol and 70 vol% sulfonated kerosene, 1.5 mol/L NaOH was used as a stripping agent to strip vanadium from extracting solution. Adding 80 g/L NH₄NO₃ to the stripping solution at 30–40 °C and pH 7.5, vanadium could be crystallized as ammonium metavanadate. Roasting ammonium metavanadate at 540 °C for 1 h, the purity of V₂O₅ met the standard specification. The total recovery of vanadium reached 67.39%, which was higher than the classical technology.     

Oxidation of cyanide in effluents by Caro’s Acid

Caro’s Acid (peroxymonosulphuric acid: H₂SO₅) is a powerful liquid oxidant made from hydrogen peroxide that has been adopted for the detoxification of effluents containing cyanides in gold extraction plants in recent years.The present work reports the findings of a study on the kinetics of aqueous cyanide oxidation with Caro’s Acid. Experiments were conducted in batch mode using synthetic solutions of free cyanide. The experimental methodology employed involved a sequence of two 2³ factorial designs using three factors: initial [CN⁻]: 100–400 mg/L; H₂SO₅:CN⁻ molar ratio: 1–1.5–3–4.5; pH: 9–11; each one conducted at one level of Caro’s Acid strength which is obtained with the H₂SO₄:H₂O₂ molar ratio used in Caro’s Acid preparation of 3:1 and 1:1. The objective was the evaluation of the effect of those factors on the reaction kinetics at room temperature. Statistical analysis showed that the three investigated variables were found to be significant, with the variables which affected the most being the initial [CN⁻] and the H₂SO₅:CN⁻ molar ratio. The highest reaction rates were obtained for the following conditions: H₂SO₅:CN⁻ molar ratio = 4.5:1; pH = 9; and Caro’s Acid strength produced from the mixture of 3 mol of H₂SO₄ with 1 mol of H₂O₂. These conditions led to a reduction of [CN⁻] from an initial value of 400 mg/L to [CN⁻] = 1.0 mg/L after 10 min of batch reaction time at room temperature. An empirical kinetic model incorporating the weight of the contributions and the interrelation of the relevant process variables has been derived as: −d[CN⁻]/dt = k [CN⁻]^1.8 [H₂SO₅]^1.1 [H⁺]^0.06, with k = 3.8 (±2.7) × 10⁻⁶ L/mg min, at 25 °C.     

The effect of [Fe]TOT on the dissolution of synthetic Pb-doped UO2 and Th-doped UO2

The dissolution of synthetic Pb-doped UO₂ and Th-doped UO₂ was systematically studied to determine the influence of leach parameters [Fe]_TOT and ORP under standard leach conditions of: T = 50 °C, [H₂SO₄] = 15 g/L (0.15 M), and UO₂ = 100 mg/L. Results demonstrated reduced uranium dissolution in both systems compared to pure UO₂. This effect was greatest for Th-doped UO₂. The decrease in uranium dissolution between the doped systems and pure UO₂ was attributed to the formation of precipitate layers at the surface of the solid, slowing down or blocking uranium release. In the case of Pb-doped UO₂, the formation of a Pb sulphate phase was directly detected but in the case of Th-doped UO₂, no layer was found. For the latter system it was postulated that passivation of the Th-doped UO₂ surface occurs due to the formation of oxidised Th-rich phases Th(OH)₄, ThO₂ and ThO₂·nH₂O at the surface of grains preventing uranium release. In tests varying the ORP, there was an approximately linear dependence of the dissolution rate on [Fe]_TOT for both systems however the rate orders indicated a step change between an ORP of 420 and 460 mV. The specific influence of Fe^II showed that both Pb–UO₂ and Th–UO₂ exhibited two distinct regions of dissolution rate dependency similar to that previously noted for pure UO₂.     

Removal of uranium(VI) under aerobic and anaerobic conditions using an indigenous mine consortium

Biological uranium removal was investigated using bacteria sourced from an uranium mine in Limpopo, South Africa. Background uranium concentration in the soil from the mine was determined to be 168 mg/kg using the ICP-OES calibrated against the uranium atomic absorption standard solution. Thus the bacteria isolated from the site were expected to be resistant to uranium-6 [U(VI)] toxicity. Preliminary studies using mixed cultures suggest that uranium reduction occurs under anaerobic conditions in most cases. U(VI) reduction by obligate aerobes isolated from the soil consortium was poor. The pure cultures namely; Pseudomonas sp., Pantoea sp. and Enterobacter sp. showed a high reduction rate at pH 5–6. The initial U(VI) reduction rate determined at 50% of added U(VI) was highest in the Pseudomonas sp. at 30 mg/L. Enterobacter sp. outperformed the other two species at 200 mg/L and 400 mg/L with a rate of 63 and 198 mg/L/h, respectively. Rapid reduction was observed in all cultures during the first 4–6 h of incubation with equilibrium conditions obtained only after incubation for 24 h. The results demonstrate the potential of microbial U(VI) reduction as an alternative technology to currently used physical/chemical processes for treatment and recovery of uranium in the nuclear industry.     

Use of quaternary ammonium salts to remove copper–cyanide complexes by solvent extraction

Cyanidation is one of the most common methods for the extraction of precious metals. In this process, effluents frequently contain relatively high concentrations of copper, which may react with cyanide to form cuprocyanide complexes adversely affecting the process.In this preliminary work, the use of solvent extraction to remove the copper–cyanide species from a synthetic solution similar to that of gold mill effluents was studied in order to permit the recycling of the solution into the process. For the extraction of these anions, the quaternary ammonium salts Quartamin TPR, Adogen 464 and Aliquat 336 were studied as extractants. The experimental results showed that for a synthetic solution of 710 mg/L copper and 1100 mg/L cyanide, it is possible to obtain a copper extraction of 99% when using 0.033 mol/L of the extractant Adogen 464 (organic/aqueous volume ratio (O/A) = 1) in the range of pH of 9–11. Up to 99% of the copper can be stripped from the organic solution after three contact times (5 min each) with 50 mL of sodium hydroxide 0.5 M (O/A = 1).     

Bioleaching of phosphorus from rock phosphate containing pyrites by Acidithiobacillus ferrooxidans

Pyrites can be oxidized by the bacterium Acidithiobacillus ferrooxidans (At. f.), producing H₂SO₄ and FeSO₄. Rock phosphate is dissolved by H₂SO₄, forming soluble phosphorus. Fe²⁺ in FeSO₄ is oxidized to Fe³⁺, producing energy to sustain the growth of At. f. The effects of four factors (rock phosphate dosage, pyrite dosage, culture temperature and time) on the fraction of phosphorous leached were investigated. It is suggested that the optimal conditions are as follows: rock phosphate dosage 1 g/L, pyrite dosage 30 g/L, culture temperature 30 °C, culture time 84 h. The fraction of phosphorous leached is up to 11.8%.     

Research on the recycling of valuable metals in spent Al2O3-based catalyst

A new technology was developed to recover multiple valuable elements in the spent Al₂O₃-based catalyst by X-ray phase analysis and exploratory experiments. The experiment results showed: In the condition of roasting temperature of 750 °C and roasting time of 30 min, mol ratio of Na₂O: Al₂O₃ 1.2, the leaching rate of alumina, vanadium and molybdenum in the spent catalyst is 97.2%, 95.8% and 98.9%, respectively. Vanadium and molybdenum in sodium aluminate solution can be recovered by barium hydroxide and barium aluminate, the precipitation rate of vanadium and molybdenum is 94.8% and 92.6%. Al(OH)₃ is prepared from sodium aluminate solution with carbonation decomposition process, and the purity of Al₂O₃ is 99.9% after calcinations, the recovery of alumina can reach 90.6% in the whole process. The Ni–Co concentrate was leached by sulfuric acid, a nickel recovery of 98.2% and over 98.5% cobalt recovery was obtained respectively under the experimental condition of 30% (w/w) H₂SO₄, 80 °C, reaction time 4 h, liquid:solid ratio (8:1) by weight, stirring rate of 800 rpm.     

Study of silver leaching with the thiosulfate–nitrite–copper alternative system: Effect of thiosulfate concentration and leaching temperature

Thiosulfate system is considered an interesting alternative leaching process for precious metals. Nevertheless, most of the literature published on these conventional thiosulfate leaching solutions has been focused on the use of ammonia and copper to generate the cupric tetraamine complex, which acts as a catalytic oxidant for silver. However, ammonia toxicity is also a detrimental issue in terms of the process sustainability. For that reason, thiosulfate–nitrite–copper solutions were studied as an alternative less toxic system for silver leaching.In this work, the effect of the thiosulfate concentration (0.07 M, 0.1 M and 0.15 M) and temperature (room temperature, 30, 35, 40 and 45 °C) on the metallic silver leaching kinetics is presented for the S₂O₃–NO₂–Cu system. The results show that the thiosulfate concentration plays an important role in the S₂O₃–NO₂–Cu–Ag system since it controls the silver leaching kinetics. On the other hand, an increase in temperature favors the silver recovery.Finally, the SEM–EDS analysis, the X-ray mapping and the X-ray diffractograms show that the solid silver particles are coated by a Cu, S and O layer for the 0.07 M and 0.1 M thiosulfate experiments, which is consistent with the formation of antlerite (Cu₃(SO₄)(OH)₄); while the 0.15 M thiosulfate scenario produced a layer composed only of Cu and S, revealing the formation of stromeyerite (CuAgS). The UV–Visible technique confirmed the in-situ generation of copper–ammonia complexes for the 0.07 M leaching condition; however, these complexes are not formed at the 0.15 M condition.