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.     

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.     

Density functional theory study of α-Bromolauric acid adsorption on the α-quartz (1 0 1) surface

Adsorption mechanism of collector α-Bromolauric acid (CH₃(CH₂)₉CHBrCOOH, α-BLA) on α-quartz (1 0 1) surface has been investigated by first-principles calculations based on density functional theory (DFT). The interaction energies of H₂O molecule, calcium ions (Ca²⁺), hydroxyl ions (OH⁻), calcium hydroxyl ions (Ca(OH)⁺), and α-BLA⁻ ions with α-quartz (1 0 1) surface were in the order of Ca(OH)⁺ < Ca²⁺ < OH⁻ < H₂O < α-BLA⁻. The results revealed that the collector α-BLA cannot adsorb on α-quartz (1 0 1) surface due to the hindrance of hydration shell of quartz surface, while Ca(OH)⁺ could repulse the hydration shell and consequently adsorb on quartz surface, which further leads to the adsorption of the collector α-BLA⁻ anions on Ca(OH)⁺-activated quartz surface. Mulliken populations analysis of the external oxygen atom (O2) of quartz surface, calcium atom (Ca) of Ca(OH)⁺, and oxygen atom (O1) of collector α-BLA⁻ (–OH group) shows that the electron transfer between the Ca–O1 and Ca–O2 atoms. The overlap area of electron density between Ca–O1 and Ca–O2 atoms indicates strong interactions among the three atoms of Ca, O1, and O2, suggesting that Ca(OH)⁺ ions act as a bridge between the α-quartz (1 0 1) surface and the α-BLA collector.     

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.     

Single and binary component sorption of the fission products Sr2+, Cs+ and Co2+ from aqueous solutions onto sulphate reducing bacteria

This study investigates the removal of the fission products Sr²⁺, Cs⁺ and Co²⁺ in single and binary metal solutions by a sulphate reducing bacteria (SRB) biomass. The effect of initial concentration and pH on the sorption kinetics of each metal was evaluated in single metal solutions. Binary component equilibrium sorption studies were performed to investigate the competitive binding behaviour of each metal in the presence of a secondary metal ion. Results obtained from single metal equilibrium sorption studies indicated that SRB have a higher binding capacity for Sr²⁺ (q_max = 416.7 mg g^?1), followed by Cs⁺ (q_max = 238.1 mg g^?1), and lastly Co²⁺ (q_max = 204.1 mg g^?1). Among the binary systems investigated, Co²⁺ uptake was the most sensitive, resulting in a 76% reduction of the sorption capacity (q_max) in the presence of Cs⁺. These findings are significant for future development of effective biological processes for radioactive waste management under realistic conditions.     

An innovative process for extracting boron and simultaneous recovering metallic iron from ludwigite ore

Ludwigite ore has not yet been utilized on an industrial scale due to its complex mineralogy and fine mineral dissemination in China. Boron–iron separation and dissolution activity of boron-bearing minerals in alkaline liquor are the two key issues in the utilization of ludwigite ore, governing the boron recovery as well as operating cost. This paper proposes an innovative process for extraction of boron and iron from ludwigite ore based on coal-based direct reduction process with sodium carbonate (Na₂CO₃). The novel process involves reduction roasting, combined leaching and grinding of reduced ludwigite ore, followed by magnetic separation of leach residue, and experimental validation for each of the processing steps is demonstrated. Alkali-activation of boron and metallization of iron were synchronously achieved during carbothermic reduction of ludwigite ore in the presence of Na₂CO₃. Consequently, boron was readily extracted in the form of sodium metaborate (NaBO₂) with water at room temperature during ball mill grinding, and metallic iron powder was recovered from the leaching-filtering residue by magnetic separation. Boron extraction of 72.1% and iron recovery of 95.7% with corresponding iron grade of 95.7% in the magnetic concentrate were achieved when ludwigite ore was reduced with 20% sodium carbonate at 1050 °C for 60 min.     

Management of thiosalts in mill effluents by chemical oxidation or buffering in the lime neutralization process

Laboratory studies were conducted to investigate the removal or management of thiosalts within the lime-neutralization process, to prevent or minimize the adverse effects of thiosalts that cause delayed acidity to downstream environment. The oxidizing reagent hydrogen peroxide (H₂O₂) and the pH stabilizing (buffering) reagents carbon dioxide (CO₂), sodium bicarbonate (NaHCO₃) and sodium carbonate (Na₂CO₃) were examined for removal and management of thiosalts, respectively. Chemical oxygen demand (COD) was determined to be a proxy for thiosalts and was employed for their rapid assessment. The Target Level of thiosalts harmless to aquatic life was found to be 30 mg/L or less. The optimized lime-neutralization process required a pH level of 9.5–10 and aeration. Over-liming to pH levels >11 did not provide excess alkalinity, hardness, or a decrease in thiosalt levels.Addition of H₂O₂ to either the acid or lime-neutralized water at a molar H₂O₂:S₂O₃ ratio of 1–1.5 removed thiosalts to safe levels. About 10–15 min. at room temperature was ample time low temperatures slowed down the process but the dosages were not affected. Removal of thiosalts from 170 to 30 mg/L caused a decrease in pH from 9.6 to 6.5. Among the buffering reagents studied, both NaHCO₃ and Na₂CO₃ provided adequate buffering and a stable pH of 7 to the lime-neutralized water; whereas CO₂ resulted in poor buffering and an unstable pH that remained below 6. In cold temperatures, NaHCO₃ and Na₂CO₃ also outperformed CO₂ with higher alkalinity and hardness. Na₂CO₃ addition to lime neutralized water at pH 9.5 was found to be the most cost-effective option. Other methods could have niche applications, depending on seasonal variations and temperature.     

Determination of specific heat capacity of sulphide materials at temperatures below 100 °C in presence of moisture

The specific heat capacity (C_p) of one copper and three nickel concentrates was determined using a self-heating apparatus and by drop calorimetry over the temperature range 50 to 80 °C in the presence of 6% moisture. The C_p values from both techniques were comparable and shown to be measuring the same property. The C_p values were similar for all four concentrates increasing from ca 0.4 to 1.4 J g⁻¹ K⁻¹ as temperature increased from 50 to 80 °C. Uses of C_p to identify self-heating risk and to modify the Rosenblum standard test are discussed.     

Contact angle and bubble attachment studies in the flotation of trona and other soluble carbonate salts

Trona, Na₂CO₃ · NaHCO₃ · 2H₂O, is mined as the primary source for sodium carbonate production in the United States. Recent studies have shown that the flotation method can be used for pre-processing of trona ore to remove insoluble mineral contaminants for the production of soda ash (sodium carbonate). Studies with carbonate salts suggest that certain important factors can affect their flotation response, including viscosity of the brine and interfacial water structure. Flotation studies showed that contrary to the strong flotation of NaHCO₃ with both anionic and cationic collectors, Na₂CO₃ does not float at all. Based on the analysis of interfacial water structure in saturated brines, Na₂CO₃ was found to act as a strong water structure maker, whereas NaHCO₃ acts as a weak water structure maker. Bubble attachment time measurements suggest that collector adsorption at the surface of NaHCO₃ induces flotation; this is not the case for Na₂CO₃. Contact angle measurements indicated that the surface of Na₂CO₃ is hydrated to a great extent, whereas the NaHCO₃ salt surface is less hydrated. These results reveal that there is a strong relationship between the interfacial water structure and the contact angle of these salts. The less stable NaHCO₃ surface is ascribed to the interfacial water structure which allows for NaHCO₃ flotation with both anionic and cationic collectors.     

Application of ocean manganese nodules for the adsorption of potassium ions from seawater

Extracting potassium from seawater has great economic potential, although conventional methods offer low separation capacity and selectivity. In this study, a series of novel potassium ionic sieves (PISs) were synthesized using ocean manganese nodules (OMN) as raw materials. The PISs were characterized by XRD, SEM, and nitrogen adsorption–desorption. The potassium adsorption capacities and separation factors of PISs and OMN in KCl solution and sea brine showed that KMnO₄ treatment will result in the highest adsorption and separation performance. The resulted sample OMN-C exhibit major composition of birnessite-type potassium manganese oxides and high micropore volumes. The adsorption capacities of OMN-C to K⁺ in KCl solution and sea brine were 35.2 mg g⁻¹ and 22.1 mg g⁻¹, respectively. The separation factor of OMN-C was α(K⁺/Na⁺) = 108.6 and the sieve did not adsorb Mg²⁺, indicating its relatively high separation selectivity to K⁺. Therefore, OMN-C can selectively extract potassium from sea brine effectively. This study not only utilized the abundant OMN resources, but also prepared effective PISs, which showed great potential in the utilization of seawater.     

Critical coalescence concentration of inorganic salt solutions

Critical coalescence concentration (CCC) was determined in a laboratory-scale mechanical flotation cell for a series of coalescence inhibiting inorganic salts (KCl, NaCl, Na₂SO₄, CaCl₂ and MgSO₄) compared to two commercial frothers (methyl isobutyl carbinol, Dowfroth 250C). The salt CCC values ranged from 0.07 M (MgSO₄) to 0.31 M (KCl and NaCl) and correlated with ionic strength. The CCC values are compared to transition concentrations in the literature. The effect of salts on gas dispersion in flotation systems is discussed.     

Ionic liquid-based modified cold-induced aggregation microextraction (M-CIAME) as a novel solvent extraction method for determination of gold in saline solutions

Modified-cold-induced aggregation microextraction (M-CIAME) was used for determination of gold in saline solutions. It is robust against the much higher concentration of salt (up to 40%). In this method sodium hexafluorophosphate (NaPF₆) was added to the sample solution containing Au-TMK complex and a very small amount of 1-hexyl-3-methylimidazolium tetrafluoroborate [Hmim][BF₄]. Afterward the solution was cooled in an ice bath and a cloudy solution was formed. After centrifuging, the extractant phase was analyzed using a spectrophotometric detection method. Under the optimum conditions, the limit of detection (LOD) was 0.7 ng mL^?1 and the relative standard deviation (RSD) was 1.65% for 50 ng mL^?1 gold. The method was applied for the determination of trace amount of Au in mineral and seawater with satisfactory results.     

Froth zone characterization of an industrial flotation column in rougher circuit

In this investigation the froth zone of an industrial column (4 m “diameter” × 12 m “height”) in rougher circuit was characterized. Experiments were carried out at Miduk copper concentrator, Iran. Miduk is a unique copper processing plant which utilizes columns in rougher circuit. Cleaning and selectivity actions in the rougher froth were illustrated using solids and grade profiles along with RTD data. The impact of froth depth (FD) on overall rate constant (k) and k–S_b relationship was evaluated. Dependency of overall flotation kinetics on froth depth and gas velocity (J_g) was modeled by k = 4.97(FD)^?0.87(J_g)^0.80. Froth recovery (R_f) was estimated and modeled in terms of froth residence time of slurry (FRT_Slurry) as R_f = R_f,maxexp(?k × FRT_Slurry). Finally, the correlation between k, S_b (indicative of the collection zone performance) and FRT_Slurry (indicative of the froth zone performance) was modeled by k = 0.02 (FRT_Slurry)^?0.62(S_b)^0.82.     

Self-heating activation energy and specific heat capacity of sulphide mixtures at low temperature

Energy of activation (E_a) and specific heat capacity (C_p) for mixtures of sulphide minerals that on their own do not self-heat (SH), sphalerite/pyrite, pyrite/galena, chalcopyrite/galena and sphalerite/galena, were determined using a self-heating apparatus at temperatures below 100 °C in the presence of moisture. The mixtures all gave E_a ranging from 22.0 to 27.8 kJ mol⁻¹, similar to the range reported for Ni- and Cu-concentrates. The E_a is close to that for partial oxidation of H₂S which adds to the contention that the partial oxidation of H₂S contributes to SH of sulphides at low temperature. The C_p values ranged from 0.152 to 1.071 JK⁻¹ g⁻¹ as temperature rose from 50 °C to 80 °C, similar to the reported findings on Ni- and Cu-concentrates. The role of galvanic interaction in promoting SH is tested by examining correlations with the rest potential difference of the sulphides in the mixture.     

Selective flotation of cassiterite with benzohydroxamic acid

The flotation of cassiterite mineral from gangue with a collector benzohydroxamic acid (BHA), and the interactions between the BHA and cassiterite have been investigated. It is shown through microflotation that the BHA is able to flot cassiterite very well, calcite quite limitedly, and quartz not at all, so the selective separation of cassiterite–quartz mixture was readily achieved; while for the efficient separation of cassiterite–calcite mixture containing 48.94% SnO₂, sodium hexametaphosphate (SHMP) was needed as a depressant for the gangue, and under the condition of the BHA 100 mg L⁻¹, SHMP 3.5 mg L⁻¹, a cassiterite concentrate with the grade of 85.50% SnO₂ was obtained with the recovery of SnO₂ 95.5%. Batch flotation further demonstrated that for an industrial tin slime, which contained 0.42% Sn, 13.65% SiO₂, 24.14% CaO, 16.60% MgO, 4.50% Al₂O₃ and 6.58% Fe, the tin recovery of 84.5% after one separation was reached with the concentrate grade of 1.84% Sn under the condition of the BHA 178 mg L⁻¹, SHMP 27 mg L⁻¹. In terms of zeta potential and infrared spectra studies the main interactions between the collector BHA and the mineral cassiterite in a flotation system are chemisorption with the formation of Sn–BHA compounds rather than electrostatic attractions between them.     

Dissolution and passivation mechanisms of chalcopyrite during bioleaching: DFT calculation,XPS and electrochemistry analysis

In this work, density functional theory (DFT) calculation, X-ray photoelectron spectroscopy (XPS) and electrochemistry analysis were carried out to investigate the dissolution process and passivation mechanisms of chalcopyrite in the presence of sulfur and iron oxidizing microorganisms. Both DFT calculation and XPS analysis indicated that the formula of chalcopyrite should be Cu + Fe³ + (S²⁻)₂. Disulfide (S₂²⁻) and polysulfide (S_n²⁻) can be easily formed on the surface of chalcopyrite due to the surface reconstruction. The dissolution process of chalcopyrite in bioleaching was mainly dependent on redox potential. Chalcopyrite was predominantly directly oxidized to polysulfide when redox potential was lower than about 350 mV vs. Ag/AgCl and resulted in low dissolution rate. When redox potential was in the range of about 350–480 mV vs. Ag/AgCl, chalcopyrite was mainly transformed to intermediate species of Cu₂S rather than polysulfide, thus resulting in high dissolution rate. When redox potential was higher than about 480 mV vs. Ag/AgCl, chalcopyrite was principally directly oxidized to polysulfide which caused the passivation of chalcopyrite. Finally, a model of dissolution and passivation mechanisms of chalcopyrite in the presence of sulfur and iron oxidizing microorganisms was provided.     

The effect of seawater based media on copper dissolution from low-grade copper ore

Due to the scarcity of water in the north of Chile, there is interest in small-scale mining using seawater to leach the minerals. This situation has led to this research aimed to determine the effect of different process variables on the extraction of copper based on the ore type of these small-scale mining activities. The extraction of copper from finely ground (<150 μm) low-grade mixed ore (0.36% Cu) was studied in different acidic media (H₂SO₄ and HCl). The effects of water quality (tap water, seawater and synthetic process water) and lixiviants on copper leaching were investigated at three temperatures (25 °C, 35 °C and 45 °C). Synthetic process water was prepared by adding Na⁺, Cl⁻, Cu²⁺ and SO₄²⁻ to seawater. Copper extractions between 70% and 80% were achieved in 48 h using seawater, similar to the extractions obtained using tap water. Rapid copper dissolution occurred when synthetic process water was used, from 90% to 97%. This marked increase was related to the addition of Cu²⁺, which promoted the formation of CuCl⁺. Seawater was comparable to freshwater in terms of leaching kinetics and yield potential by raising the chloride concentration and increasing the formation of copper chloride ions. The findings of this study also expanded our understanding of the consequences of substituting seawater for freshwater at industrial leaching operations.     

Recovery of zinc from an industrial zinc leach residue by solvent extraction using D2EHPA

A hydrometallurgical treatment involving solvent extraction of zinc using di-2-ethylhexyl phosphoric acid (D2EHPA) has been investigated to recover zinc from an industrial leach residue. The residue was leached with sulfuric acid producing leach liquor which was subjected to solvent extraction for enrichment of zinc and removal of impurities. Operating variables, such as pH, D2EHPA concentration, temperature, aqueous/organic (A/O) phase ratio, tri-butyl phosphate (TBP) concentration and sodium sulfate (Na₂SO₄) concentration in aqueous phase were studied. Practically, all zinc was extracted from the aqueous solution at pH 2.5 with 20% w/w D2EHPA in kerosene. Increasing either TBP concentration up to 5%, or Na₂SO₄ concentration up to 0.2 M, increased the zinc extraction. Zinc could be extracted at one theoretical stage at A/O of 1/1, as calculated by McCabe–Thiele method.     

Alkaline digestion of dunite for Mg(OH)2 production: An investigation for indirect CO2 sequestration

The sequestration of CO₂ by carbonating natural minerals has a great potential for secure reduction of net CO₂ emissions. Feedstock Mg–silicate minerals are usually converted into Mg rich solutions or Mg(OH)₂ before the carbonation process, due to the slow reaction kinetic of direct carbonation. The present work studied the alkaline digestion of Mg–silicate minerals into Mg(OH)₂ for CO₂ sequestration. Powdered dunite containing ∼73 ± 2 wt% of forsterite (Mg₂SiO₄) was dissolved using highly concentrated NaOH aqueous systems at 90 and 180 °C with varied NaOH concentration and duration of reaction. Thermal analysis and Rietveld Refinement Quantitative Phase Analysis (QPA) confirmed that an effective digestion of dunite was possible at 180 °C achieving 80 wt% of Mg(OH)₂. It was found that NaOH concentration in solution, temperature and duration of reaction significantly influence the progress of digestion.