Removal of cyanide in aqueous solution by oxidation with hydrogen peroxide in presence of copper-impregnated activated carbon

This work is devoted to the removal of free cyanide from aqueous solution by oxidation with hydrogen peroxide H₂O₂ catalyzed by copper-impregnated activated carbon. Effects of initial molar ratio [H₂O₂]₀/[CN⁻]₀, copper-impregnated activated carbon amount, pH and the temperature on cyanide removal have been investigated.The presence of copper-impregnated activated carbon has increased the reaction rate showing thus a catalytic activity. The rate of cyanides removal increases with the raise of the initial molar ratio [H₂O₂]₀/[CN⁻]0 and decreases with the increase in the pH from 8 to 12. The increase in the copper-impregnated activated carbon amount from 1.5 to 10 g/L in reaction solution has a beneficial effect. Beyond this value, the impact of activated carbon amount is not anymore significant. The temperature does not have a significant effect between 20 and 35 °C. The four successive times re-use of catalyst shows a good stability. The kinetics of cyanide removal has been found to be of pseudo-second-order with respect to cyanide and the rate constants have been determined. This process seems very interesting because the rate of cyanides removal is very fast, the reaction does not use soluble metal catalyst and it consumes only hydrogen peroxide as chemical product.     

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.     

Cyanide oxidation by ozone in a steady-state flow bubble column

The treatment of the effluent produced in the cyanidation of gold and silver ores is one of the main problems that the precious metals industry currently faces. These effluents contain variable amounts of free and complex cyanide (stable and unstable). Different methods for the cyanide elimination have been used; nevertheless, most of the times the consumption of reagents increases operating costs to prohibiting levels. As well, the process may give rise to by-products that are also toxic.Ozone oxidation is a promising alternative for cyanide control that offers several advantages. In this work, the oxidation of cyanide in synthetic alkaline solutions by ozone was studied in a countercurrent bubble column. Dissolved ozone, cyanide and cyanate profiles were obtained along the column, demonstrating that the cyanide/ozone reaction may take place all along the reactor. This is found to occur when the mole of O₃/mole of CN rate (specific ozone dose) is higher than one.The oxidation efficiency depends mainly on the specific ozone dose (moles of ozone fed per mole of cyanide), leading to 90% efficiency when the specific ozone dose is 1.2 mol O₃/mol CN or higher. The ozone consumption obtained in these tests (steady-state flow pilot-scale column) was similar to that obtained at the laboratory scale in a semi-batch mechanically stirred reactor (1.2–1.4 mol O₃/mol CN), suggesting that the size of the reactor has no effect on ozone consumption.     

Degradation of thiocyanate in aqueous solution by persulfate activated ferric ion

Thiocyanate formation from cyanidation of gold bearing ores is becoming a more common problem during gold processing. In this work, the application of an advanced oxidation process based on the use of persulfate (S₂O₈²⁻) as an environmentally friendly oxidant in the presence of ferric ion for destruction of a persistent and non-volatile inorganic contaminant, such as thiocyanate, in aqueous solutions is reported for the first time. The influence of various reaction parameters like ferric ion and persulfate dosage, initial thiocyanate concentration and the influence of radical scavenger are examined. An accelerated reaction using S₂O₈²⁻ to destroy thiocyanate can be achieved via chemical activation with Fe³⁺ to generate highly reactive sulfate anion-radicals (SO₄⁻). The results showed that degradation efficiency was negligible when persulfate was used alone, ferric ions significantly improved the degradation efficiency of thiocyanate at ambient temperature. Under the optimum molar ratios ([S₂O₈²⁻]:[SCN⁻] = 5:1 and [S₂O₈²⁻]:[Fe³⁺] = 1:0.2), 99% of thiocyanate present in aqueous solution at the initial concentration range of 1.72–17.2 mM was degraded within 60 min of reaction time. To evaluate the contribution of reactive free radicals generated through Fe(III)-mediated activation of persulfate to thiocyanate degradation, quenching experiments using methanol as the radical quenching agent were carried out. The obvious decrease in thiocyanate oxidation efficiency in the presence of methanol confirmed that the radical-based pathway was the dominant mechanism in Fe³⁺/S₂O₈²⁻ system. The degradation of thiocyanate was accompanied by the formation of cyanide as the main final product of the reaction. Thus the catalytic oxidation of thiocyanate makes it possible to return NaCN into the production process for leaching of precious metals. The work presents an efficient and environmentally acceptable wastewater treatment process applicable in mining facilities utilizing cyanidation of sulfide ores and/or concentrates.     

Kinetics of high-sulphur and high-arsenic refractory gold concentrate oxidation by dilute nitric acid under mild conditions

On one hand, high-sulphur and high-arsenic refractory gold concentrate (HGC) leads to regional ecological damage. On the other hand, it contains a lot of valuable elements. So utilization of it can bring social and environmental benefits. In this paper, the kinetics of HGC oxidation by dilute nitric acid under mild conditions was investigated. The effects of particle size (50–335 μm), reaction temperature (25–85 °C), initial acid concentration (10–30 wt.%) and stirring speed (400–800 rpm) on the iron extraction rate (C_r) were determined. It is obvious that C_r increases with the rise of initial nitric acid concentration, reaction time and stirring speed, but decreases with the increase of particle size. Oxidation kinetics indicates that the rate of reaction is diffusion controlled. The activation energies were determined to be 10.70 kJ/mol in the 10% HNO₃ and 12.25 kJ/mol in the 25% HNO₃.     

Leaching of chalcopyrite (CuFeS2) with an imidazolium-based ionic liquid in the presence of chloride

The effects of independent variables such as, temperature, concentration of ionic liquid (1-butyl-3-methyl-imidazolium hydrogen sulphate, [bmim][HSO₄]), chloride and sulphuric acid on copper extraction from chalcopyrite (CuFeS₂) ore were studied by surface optimization methodology. The Central Composite Face approach and a quadratic model were applied to the experimental design. The optimal copper extraction conditions given by the above methodology were 20% (v/v) of [bmim][HSO₄] in water, 100 g L⁻¹ chloride, and 90 °C. The concentration of chloride and the temperature together exert a synergistic effect in enhancing chalcopyrite dissolution. Experimental data were fitted by multiple regression analysis to a quadratic equation and analyzed statistically. A model was developed for predicting copper extraction from CuFeS₂ ore with variables such as Cl⁻, [bmim][HSO₄], H₂SO₄ concentrations and temperature in the range studied. The activation energy was calculated to be 60.4 kJ/mol (temperature range 30–90 °C), indicative of chemical control of the reaction and [bmim][HSO₄] acts as an acid in the reaction.     

Mechanisms of sulfide ion oxidation during cyanidation. Part II: Surface catalysis by pyrite

The mechanisms and the reaction products for the oxidation of sulfide ions in the presence of pyrite have been established. When the leach solution contains free sulfide ions, oxidation occurs via electron transfer from the sulfide ion to dissolved oxygen on the pyrite mineral surface, with polysulfides being formed as an intermediate oxidation product. In the absence of cyanide, the polysulfides are further oxidised to thiosulfate, whilst with cyanide present, thiocyanate and sulfite are also formed from the reaction of polysulfides with cyanide and dissolved oxygen. Polysulfide chain length has been shown to affect the final reaction products of polysulfide oxidation by dissolved oxygen.The rate of pyrite catalysed sulfide ion oxidation was found to be slower in cyanide solutions compared to cyanide free solutions. Mixed potential measurements indicated that the reduction of oxygen at the pyrite surface is hindered in the presence of cyanide. The presence of sulfide ions was also found to activate the pyrite surface, increasing its rate of oxidation by oxygen. This effect was particularly evident in the presence of cyanide; in the presence of sulfide the increase in total sulfur from pyrite oxidation was 2.3 mM in 7 h, compared to an increase of <1 mM in the absence of sulfide over 24 h.     

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.     

The leaching and adsorption of gold using low concentration amino acids and hydrogen peroxide: Effect of catalytic ions,sulphide minerals and amino acid type

The leaching of gold using alkaline amino acids–hydrogen peroxide solutions at low concentrations has been studied. The application of alkaline amino acid–hydrogen peroxide system may offer an alternative and environmentally benign process for gold leaching, particularly in the context of leaching low grade gold ores in an in-situ or in heap leach processes. In the presence of an oxidant or oxidants, it was found that amino acids can dissolve gold at alkaline condition at low and moderate temperature. Heating the leach solution between 40 and 60 °C was found to enhance the gold dissolution significantly in alkaline amino acid–peroxide solutions. It was also found that gold dissolution increases by increasing amino acid concentration, peroxide and pH. Amino acids acts synergistically to dissolve gold. Although glycine showed the highest gold dissolution as a single amino acid compared to histidine and alanine, histidine was found to enhance gold dissolution when used in equimolar amounts with glycine. The presence of Cu²⁺ ion enhances gold dissolution in the glycine–peroxide solutions. The process will propose an environmentally benign process for gold treatment in order to replace the use of cyanide in heap or in-situ leaching. In the presence of pyrite, the amount of gold leached was lower due to the peroxide consumption in sulphide oxidation.     

TGA kinetic study on the hydrogen reduction of an iron nickel oxide

Reduction of a mixed iron nickel oxide by hydrogen to produce ferronickel alloy was investigated by thermo-gravimetric analysis (TGA). The iron nickel oxide, which contains 50 wt% Fe and 10 wt% Ni mainly in the form of hematite (Fe₂O₃) and nickel ferrite (NiFe₂O₄), is a residue produced from the water leaching of a selectively sulfation roasted nickel calcine. Continuous heating tests with a fixed heating rate as well as isothermal tests were performed. Results indicate that the reduction of the mixed oxide began above 350 °C. The reduction reactions occurred in a non-topochemical mode at low temperatures between 350 °C and 600 °C with its rate controlled by the solid–gas chemical reactions. The reduction rate increased with the increase in temperature from 350 °C up to 1100 °C. Between 600 °C and 1200 °C, the rate controlling step was the diffusion of reducing gas through the pores of the sample bed with an apparent activation energy of 34.1 kJ/mol. Due to the melting of the silicate material at 1200 °C which substantially reduced the sample porosity, the reduction rate decreased. Above 1200 °C, the reduction rate increased again due to the increased diffusion of the reducing gas through the molten sample at higher temperatures.     

Dissolution of gold during pyrite oxidation reaction

Numerous papers discuss the mechanism of alkaline oxidation of pyrite but there is limited information available describing the actual kinetics of the pyrite sulphide to thiosulphate reaction. A previous investigation in this series determined the rate of sulphide sulphur oxidation and thiosulphate yield in the reaction of pyrite with sodium hydroxide under various testing conditions. The goal of the current study is to validate these rates using two different gold-containing pyrite concentrates. A further objective of the current work is to investigate the simultaneous dissolution of gold with in situ formed thiosulphate during pyrite oxidation.It was found that at 20 psi oxygen overpressure and a temperature of 80 °C, the initial rate of sulphide oxidation and thiosulphate yield were close to 0.08 mol/h and 0.0155 mol/h, respectively. These rates are in agreement with previously published data. However, a shift from linearity occurred when the pH decreased below 12. A rapid decay of thiosulphate was evidenced at pH 8.3–9.2 while E_H was in the range of 22–141 mV. Based on relevant thermodynamic analysis of metastable thiosalts system, such rapid decomposition is not expected at these pH and E_H values. It is believed that the presence of unreacted pyrite acting as a catalyst caused this behaviour. It appears that under mildly alkaline conditions, the rate of oxidation of sulphide to thiosulphate becomes slower than the rate of thiosulphate degradation, which causes a net loss of thiosulphate in the system. The maximum extraction of gold and silver (96% and 75% respectively) was achieved under conditions of pH < 12.     

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.     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).     

Effect of fluidized-bed carrier material on biological ferric sulphate generation

High biomass hold-up and high iron oxidation rates of a biological ferric sulphate generating fluidized-bed reactor (FBR) requires a carrier material with high specific surface area, high porosity and inertness. In this work, the effect of activated carbon (AC), diatomaceous earth (Celite) and Al₂O₃ (Compalox) carrier materials on the ferric sulphate generation in FBRs were studied. Compalox dissolved during the experiments and formed an unfluidizable aggregate, and was therefore rejected. The slow dissolution of Celite resulted in a light, fine-grained, layer on top of the fluidized bed that had to be changed into fresh Celite. AC resisted well the friction caused by fluidization. The iron oxidation in the continuous-flow FBRs became limited by oxygen supply already at loading rates of 2.5 kg Fe²⁺ m⁻³ h⁻¹. Iron oxidation rates of 27.6 and 25.7 kg m⁻³ h⁻¹ were obtained in batch FBR experiments with AC and Celite, respectively.Biomass accumulation of 6.2 × 10¹⁰, 2.4 × 10¹⁰ and 8.0 × 10⁹ cells per g of carrier was detected on Celite, AC and Compalox, respectively. The bacterial community structures on the carrier materials were revealed by Polymerase Chain Reaction and Denaturating Gradient Gel Electrophoresis (PCR-DGGE) followed by partial sequencing of the 16S rRNA gene. Two bacterial strains, Leptospirillum ferriphilum and a strain similar to a strain unofficially named “Ferrimicrobium acidiphilum”, were detected. Examination of the carrier material surfaces with scanning electron microscopy (SEM) coupled with energy dispersive spectroscopy (EDS) revealed that all carrier materials were covered with jarosite precipitates and that the bacteria were mainly retained on the jarosite covered areas. In conclusion, AC was the most promising carrier material for a large-scale biological ferric sulphate generating FBR based on its availability, durability and the achieved high iron oxidation rates.     

Electrochemical enhanced oxidative decomposition of chromite ore in highly concentrated KOH solution

A novel method which introduces an electrochemical field to enhance the oxidative decomposition of chromite in a KOH sub-molten salt medium was proposed and proven to be feasible and efficient. Under optimal reaction conditions (slot current density 750 A/m², alkali concentration 60 wt.%, reaction temperature 150 °C, alkali-to-ore mass ratio 6:1, and particle size <200 mesh), the extraction rate of chromium reached 99%, after reacting for 480 min. In comparison with the current liquid-phase oxidation technologies, the reaction temperature in the new approach is 150–250 °C lower, and the alkali concentration of the reaction medium is lower by more than 20%, showing substantial advantages in terms of energy efficiency, equipment corrosion alleviation and prospects for industrial application. The reaction kinetics study shows that the extraction process under optimal reaction conditions is jointly governed by surface chemical reaction and solid product layer diffusion with the apparent activation energy calculated to be 17.56 kJ/mol.     

The influence of pyrite pre-oxidation on gold recovery by cyanidation

The influence of pyrite pre-oxidation in alkaline solutions on gold recovery by cyanidation from Twin Creek refractory gold ore in which pyrite was identified as the major sulfide mineral has been investigated with the aid of electrochemical measurements, leaching experiments, and direct analysis of reaction products for selected residues. It was found that gold recovery by cyanidation in bottle roll experiments mainly depended on the extent of pyrite pre-oxidation. The rate of pyrite oxidation in alkaline solutions measured by electrochemical measurements, including chronoamperometry and linear sweep voltammetry, increased with an increase in pH, potential, and temperature. All alkaline reagents used for the electrochemical measurements, NaOH, NH₄OH, Na₂CO₃ and Ca(OH)₂, showed a similar effect on pyrite oxidation kinetics. However, the results of alkaline pre-oxidation for pyrite of the Twin Creek refractory gold ore suggested that NaOH and Na₂CO₃/Ca(OH)₂ were superior to Ca(OH)₂. Without pre-oxidation, cyanide leachable gold was found to be only 20% which could be increased to 70% under appropriate pre-oxidation conditions. At the same time, cyanide consumption decreased from 2.5 kg/t ore to 1.5 kg/t ore.Selected residues after pre-oxidation and cyanidation were examined by X-ray diffraction. Backscattered electron images of pyrite particles in these residues were taken. The reaction products at the surface of pyrite particles were found to be iron-, silicon-, and calcium-bearing compounds with variable amounts of sulfur as determined by X-ray energy dispersion analysis. Additionally, some mineral fines, such as aluminum and/or potassium-bearing minerals, were found to be present at the partially oxidized pyrite surface.     

Uranium recovery from strong acidic solutions by solvent extraction with Cyanex 923 and a modifier

Uranium stripping with strong acid solution is always highly desired due to its simple operation and less pollution. However, intensive acid neutralisation for uranium precipitation in the subsequent step limited its application. A new solvent extraction process has been developed to transfer uranium from strong to weak sulphuric acid solutions suitable for uranium precipitation without intensive neutralisation. An organic system consisting of 10% Cyanex 923 and 10% isodecanol as the modifier in ShellSol D70 was optimised for the process. It was found that uranium was extracted efficiently from 4 to 6 M H₂SO₄ solutions with the organic system, and it could be efficiently stripped with 0.2–0.5 M H₂SO₄ solutions. Both extraction and stripping kinetics of uranium were very fast, reaching the equilibrium within 0.5 min. Temperature between 30 and 60 °C has slight effect on uranium extraction and stripping. Four theoretical stages could effectively extract more than 98% uranium from a solution containing 17.5 g/L U and 6.0 M H₂SO₄ at an A/O ratio of 1:1.5, and it could generate a loaded organic solution containing about 12 g/L U. More than 99% U could be stripped from the loaded organic solution containing 14.6 g/L U with 0.5 M H₂SO₄ using five stages at an A/O ratio of 1:3. As a result, the loaded strip liquor containing more than 40 g/L U would be obtained which is suitable for uranium recovery by precipitation using hydrogen peroxide. A conceptual process has been proposed for uranium transfer from strong to weak sulphuric acid solutions for its recovery.     

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.     

Adsorbing flocs in expanded/fluidised bed reactors: A new basis for pollutants removal

The present work describes studies concerning a new adsorption technique based on the use of adsorbent flocs in an expanded/fluidised bed reactor for the removal of pollutants from aqueous solutions. The technique, based on flocculation of aqueous suspensions of powdered adsorbent materials, when conducted in an expanded/fluidised bed reactor takes advantage of conducting adsorption and solid–liquid separation in one single stage. Studies were performed using flocculated powdered activated carbon and natural zeolites, alone and in mixtures, for phenol and ammonia adsorption. A reactor with cylindrical–conical geometry was used for the pollutants adsorption in flocs beds (pure and mixed), as well as the regeneration/recycle of the exhausted adsorbents. Results proved the high adsorption efficiency of powdered natural zeolites and activated carbon flocs for the uptake of ammonia (11 mg NH₃–N g⁻¹) and phenol (132 mg g⁻¹), at 38 and 19 m h⁻¹ loading rates, respectively. Regeneration/recycle of the pollutant-saturated beds was possible for the ammonia/natural zeolites adsorption case, using sodium sulphate as regenerator. Use of mixed flocs beds was efficient, showing advantages such as multiple-pollutants adsorption in one single stage, higher loading rates when using light materials (activated carbon) combined with heavier ones (natural zeolites) and use of small adsorbent concentrations (not possible otherwise). Economical and environmental issues regarding the technique are also discussed in the paper. The new technique shows great potential as an alternative physicochemical adsorption process for pollutant removal from aqueous solutions using low-cost and highly-available powdered adsorbent materials.