Treatment of Mine Water for Sulphate and Metal Removal Using Barium Sulphide

Abstract.   The integrated barium sulphide process consists of: preliminary treatment with lime, sulphate precipitation as barium sulphate, H₂S-stripping, crystallization of CaCO₃, and recovery of barium sulphide. Our tests showed that during lime pre-treatment, sulphate was lowered from 2 800 mg/L to 1 250 mg/L by gypsum crystallization; metals were precipitated as hydroxides. The BaS treatment then lowered sulphate to less than 200 mg/L. Sulphide was lowered from 333 to less than 10 mg/L (as S) in the stripping stage, using CO₂ gas for stripping. The stripped H₂S-gas was contacted with Fe (III)-solution and converted quantitatively to elemental sulphur. The alkalinity of the calcium bicarbonate-rich water was reduced from 1 000 to 110 mg/L (as CaCO₃) after CO₂-stripping with air due to CaCO₃ precipitation. Fe (II), after sulphur production, was re-oxidized to Fe (III) using an electrolytic step. The running cost of the BaS process is R2.12/m³ (US$1 = SAR6.5) for the removal of 2 g/L of sulphate.     

A novel approach for recovery of rare earths and niobium from Bayan Obo tailings

It is difficult to economically recover rare earths (RE) and niobium (Nb) from Bayan Obo tailings by the existing metallurgical processes. In this study, a novel hydrometallurgy process was employed for separating and recovering RE and Nb from Bayan Obo tailings. Firstly, by sulfating roasting at 250 °C and subsequent leaching at 60 °C, the RE and Nb present in the polymetallic minerals can be efficiently extracted into the leach solution. Secondly, after the reduction of Ti⁴⁺ and Fe³⁺ ions (to Ti³⁺ and Fe²⁺ ions) with iron powders followed by hydrolysis at pH 2.01, the Nb can be efficiently precipitated from the leach solution. The impurities present in the precipitated product can then be removed by treating with NH₃⋅H₂O–H₂C₂O₄ system at pH 4.50. Thirdly, the RE can be efficiently precipitated at pH 7.15 from the filtrate of above hydrolysis reaction mixture. Finally, the impurities present in the crude RE can be removed by oxalate co-precipitation method. The yield of RE and Nb in this novel process reaches up to 90% and 78%, respectively. Both the Nb (60.67 wt% Nb₂O₅) and RE products (>88.65 wt% RE_xO_y) have high application value.     

Recovery of manganese from electric arc furnace dust of ferromanganese production units by reductive leaching

Recovery of manganese from electric arc furnace dust (EAFD) of a ferromanganese production unit was investigated using reductive leaching in sulfuric acid (H₂SO₄). Three different reducing reagents, oxalic acid, hydrogen peroxide and glucose, were tested. Effect of different leaching parameters on the leaching and separation selectivity of manganese from iron was investigated. Iron concentration in solution was also determined, since iron was the major contaminant in the leach liquor. From the results complete leaching of manganese was achieved at 0.31 mol/L oxalic acid concentration, 2 mol/L H₂SO₄, a liquid/solid ratio of 30/1 at 70 °C and with 90 min leaching time.     

Preliminary Studies on Precipitating Manganese Using Caro’s Acid and Hydrogen Peroxide

We report preliminary studies on the precipitation of manganese compounds by oxidation with Caro’s Acid (peroxomonosulphuric acid, H₂SO₅) or hydrogen peroxide (H₂O₂), from solutions with [Mn] = 1.2 g/L to achieve residual [Mn] <1 mg/L at a pH range of 5–9. It was found that with the addition of sodium carbonate and either Caro’s acid or hydrogen peroxide, it was possible to reduce manganese from 1.2 g/L to less than 1 mg/L in 60 min (batch reaction) at 25°C (at pH ≥ 5 using H₂SO₅, and pH = 9 using H₂O₂). By comparison, simple hydroxide precipitation under the same conditions at pH = 9 would only lower [Mn] to 165 mg/L.     

Fate of sulphate removed during the treatment of circumneutral mine water and acid mine drainage with coal fly ash: Modelling and experimental approach

The treatment of acid mine drainage (AMD) and circumneutral mine water (CMW) with South African coal fly ash (FA) provides a low cost and alternative technique for treating mine wastes waters. The sulphate concentration in AMD can be reduced significantly when AMD was treated with the FA to pH 9. On the other hand an insignificant amount of sulphate was removed when CMW (containing a very low concentration of Fe and Al) was treated using FA to pH 9. The levels of Fe and Al, and the final solution pH in the AMD–fly ash mixture played a significant role on the level of sulphate removal in contrast to CMW–fly ash mixtures. In this study, a modelling approach using PHREEQC geochemical modelling software was combined with AMD–fly ash and/or CMW–fly ash neutralization experiments in order to predict the mineral phases involved in sulphate removal. The effects of solution pH and Fe and Al concentration in mine water on sulphate were also investigated. The results obtained showed that sulphate, Fe, Al, Mg and Mn removal from AMD and/or CMW with fly ash is a function of solution pH. The presence of Fe and Al in AMD exhibited buffering characteristic leading to more lime leaching from FA into mine water, hence increasing the concentration of Ca²⁺. This resulted in increased removal of sulphate as CaSO₄·2H₂O. In addition the sulphate removal was enhanced through the precipitation as Fe and Al oxyhydroxysulphates (as shown by geochemical modelling) in AMD–fly ash system. The low concentration of Fe and Al in CMW resulted in sulphate removal depending mainly on CaSO₄·2H₂O. The results of this study would have implications on the design of treatment methods relevant for different mine waters.     

尿素量对CeO2粒子结构、形貌和光催化活性的影响

以尿素为沉淀剂,采用微波回流法制备了CeO₂粒子。利用X射线衍射(XRD),扫描电子显微镜(SEM),孔隙比表面积(BET)以及荧光光谱(PL)等对样品的物理化学性能进行了表征。研究了不同尿素添加量对样品结构、形貌的影响,并通过降解苯酚溶液考察了所制备的CeO₂粒子的光催化活性。结果表明,所制备的二氧化铈粒子形貌为规则的菱形,随尿素添加量的增加,CeO₂粒子的尺寸先增大后减小且逐渐变得均匀。当硝酸铈与尿素物质的量比为1:7时,所得CeO₂-4样品的PL光谱强度最低,比表面积最大。在紫外光照射下,CeO₂-4样品的光催化性能最好,这可以归结为其具有较高的比表面积以及有效的电子-空穴对的分离效率。因此,合适的尿素添加量有利于光催化活性的提高。     

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.     

A novel separation process for detoxifying cadmium-containing residues from zinc purification plants

A novel separation process for detoxifying cadmium-containing residues arising from zinc hydrometallurgical processes has been developed. The solution of sulfuric acid and ammonium citrate ((NH₄)₃C₆H₅O₇) is used as the lixiviant to partially extract copper and other metals from the residue at room temperature. Copper in the leachate is recovered by solvent extraction (SX) with LIX 973. Cobalt in the raffinate from the copper SX process is recovered by precipitation with α-nitroso-β-naphthol. Zinc remaining in the solution, at high concentrations, is further separated from cadmium by solvent extraction with P204 (di(2-ethylhexly)phosphoric acid, D2EHPA). The results show that around 50% of the copper and virtually all the zinc, cadmium and cobalt in the residue can be leached under the experimental conditions used here. Virtually all dissolved copper and cobalt can be recovered in the subsequent solvent extraction and precipitation processes. The presence of (NH₄)₃C₆H₅O₇ could significantly facilitate the extraction of zinc by D2EHPA. More than 97% of the dissolved zinc can be recovered through a two-stage counter current solvent extraction process. After separating copper, cobalt and zinc, cadmium can easily be recovered from the solution either by cementation with zinc powder or by electrowinning, while the purified solution can be recycled back to the leaching process.     

Using Calcium Carbonate/Hydroxide and Barium Carbonate to Remove Sulphate from Mine Water

This study evaluated the effectiveness of using barium bicarbonate to remove sulphate from neutralized AMD. The Ba(HCO₃)₂ was produced by dosing a BaCO₃ solution with CO₂ to form Ba(HCO₃)₂. This greatly increased the barium ion concentration, which rapidly removed sulphate linked to either calcium or magnesium. Following sulphate removal, the Ca(HCO₃)₂ or Mg(HCO₃)₂ containing water can be stabilised by CO₂ stripping with air, which results in CaCO₃ precipitation. The MgCO₃ remains in solution.     

Treatment of Acid Leachate from Coal Discard using Calcium Carbonate and Biological Sulphate Removal

Abstract.   An integrated approach is proposed for treating acidic coal discard leachate, consisting of CaCO₃ handling and dosing, CaCO₃-neutralization, and biological sulphate removal. It was found that: powdered CaCO₃ can be slurried to a constant density and used to neutralize the acid water, remove Fe (II), Fe (III), and Al, and partially remove the sulphate (to saturation level); biological sulphate removal can be used to lower the sulphate to less than 200 mg/L using ethanol as the carbon and energy source; CO₂ produced during calcium carbonate treatment can be used for H₂S-stripping and; H₂S gas recovered in the sulphate removal stage can be used for iron removal.     

Inhibition of bacterial activity in acid mine drainage

Acid mine drainage water give rise to rapid growth and activity of an iron- and sulphur- oxidizing bacteriumThiobacillus ferrooxidians which greatly accelerate acid producing reactions by oxidation of pyrite material associated with coal and adjoining strata. The role of this bacterium in production of acid mine drainage is described. This study presents the data which demonstrate the inhibitory effect of certain organic acids, sodium benzoate, sodium lauryl sulphate, quarternary ammonium compounds on the growth of the acidophilic aerobic autotroph Thiobacillusferrooxidians. In each experiment, 10 milli-litres of laboratory developed culture ofThiobacillus ferrooxidians was added to 250 milli-litres Erlenmeyer flask containing 90 milli-litres of 9-k media supplemented with FeSO₄ 7H₂O and organic compounds at various concentrations. Control experiments were also carried out. The treated and untreated (control) samples analysed at various time intervals for Ferrous Iron and pH levels. Results from this investigation showed that some organic acids, sodium benzoate, sodium lauryl sulphate and quarternary ammonium compounds at low concentration (10-2 M, 10–50 ppm concentration levels) are effective bactericides and able to inhibit and reduce the Ferrous Iron oxidation and acidity formation by inhibiting the growth ofThiobacillus ferrooxidians is also discussed and presented     

Mechanochemical processing of chrysocolla with sodium sulphide

Autogenous solid state processing of chrysocolla with sodium sulphide has been carried out at different CuSiO₃–Na₂S-ratios, particle sizes and temperatures. Water soluble sodium silicate (Na₂SiO₃) and covellite (CuS) were determined by means of chemical analysis and X-ray examination as reaction products. More than 84% of chrysocolla and 50% of associated malachite were converted to covellite at 80 °C, atmospheric pressure, 10 min contact time, stochiometric CuSiO₃–Na₂S-ratio and particle sizes of <100 μm. The mechanochemical treatment using a vibrating mill increases the conversion rate of chrysocolla and malachite up to 90% and 70%, respectively. After washing of the products, the sodium silicate solution was converted into sodium sulphate (Na₂SO₄) by adding sulphuric acid, with precipitation of silica as by-product. In addition, the carbothermic reduction of crystallised sodium sulphate led to sodium sulphide necessary for solid state processing.     

Oxidative Precipitation of Manganese from Dilute Waters

Various metals may be removed from mine water by precipitation of their hydroxides. However, over the pH range of 7–9, it is more efficient to oxidize Mn II to Mn IV than to precipitate Mn II as its hydroxide. Chlorine (or hypochlorite) can be used to do this, but its use may not be appropriate for mine waters that will be recycled as process solutions or discharged into a receiving body. Consequently, there has been interest in reducing chlorine use or its total replacement by non-chlorinated oxidants, such as oxygen and peroxygens, for treatment of mine water. In this work, we report the results of a comparative investigation of the following oxidants: NaClO (as reference), O₂; H₂O₂; Caro’s acid (H₂SO₅), and the combination of H₂O₂ with NaClO, with initial [Mn] = 10 mg/L, at 25 °C, with 100 and 300 % excess oxidant above the stoichiometric requirement. It was found that the reaction pH has to be greater than 8 to obtain effective precipitation. It is possible to reach a final [Mn] below 1.0 mg/L in 60 min of batch reaction time, using either NaClO, Caro’s Acid, or a combination of NaClO + H₂O₂. Using only O₂ or H₂O₂ was ineffective.     

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.     

Cobalt recovery from mixed Co–Mn hydroxide precipitates by ammonia–ammonium carbonate leaching

The research work presented in this paper focused on the recovery of cobalt from mixed Co–Mn hydroxide precipitates (obtained from sulphate leach liquors of nickel oxide ore), using ammonia–ammonium carbonate leaching. The characterization of the initial mixed hydroxide precipitates, as well as the corresponding leached residue was carried out by X-ray Diffraction, TG–DTA and Scanning Electron Microscopy.Cobalt and manganese precipitation was based on the statistical design and analysis of experiments, in order to determine the main effects and interactions of the precipitation factors, which were the equilibrium pH and the temperature. Co and Mn were precipitated as hydroxides at pH = 10.5 and T = 25 °C, using 5 M NaOH as a neutralizing agent, by 99.9% and 99.5%, respectively. The main mineralogical phases were, Mn₃O₄ (Hausmannite), γ-Mn₃O₄ and CoMn₂O₄, while Co(OH)₂ and Mn(OH)₂ (Pyrochroite) were also present as minor constituents.Cobalt and manganese separation was based on selective cobalt recovery by ammonia–ammonium carbonate leaching of the produced mixed hydroxide precipitate. The factors studied were the ammonia–ammonium carbonate concentration and the solid to liquid ratio. The cobalt recovery efficiency reached 93%. Mn₃O₄ (Hausmannite) was the main mineralogical phase of the leached residue, while MnCO₃ (Rhodochrosite) and Mn₂O₃ were also present. Small quantities of cobalt were also observed in the residue as CoMn₂O₄.     

Factors Controlling the Migration of Tailings-Derived Arsenic: A Case Study at the Yara Siilinjärvi Site

The behaviour of arsenic (As) derived from tailings was investigated at the Yara Siilinjärvi apatite mine and industrial site in eastern Finland. The study assessed factors influencing the migration and fate of As and compared the anthropogenic As load to the natural geogenic background. Environmental risks related to As were assessed by examining the As concentrations in humus, glacial till, aquatic sediments, groundwater, and surface water. The occurrence and fractionation of As and the presence of secondary precipitates and geochemical transformations in the tailings and in the ambient soil and sediment were evaluated by selective extraction. The water-derived emissions were evaluated by field measurements, hydrogeochemical analysis, and modelling. Results indicate elevated environmental risks due to dust and seepage emissions from the tailings since the concentrations and mobility of As and other potentially harmful elements (PHEs) such as Co, Ni, and Zn were elevated relative to the geogenic background. These elements were mainly associated with Fe (oxy)hydroxides in the soil and their mobility was closely linked to Fe biogeochemistry. Additionally, although the concentrations of As and PHEs were high in the tailings pond and seepage water, they decreased in ambient groundwater and surface water, indicating Fe (oxy)hydroxide stability. This was supported by hydrogeochemical modelling, which indicated precipitation of Fe oxides and hydroxides. According to speciation modelling, As was present mainly as toxic trivalent arsenious acid (H₃AsO₃) in groundwater and as the less toxic pentavalent As acid (H₂AsO₄ ^? and HAsO₄ ^2?) in surface water.     

Mineral carbonation of PGM mine tailings for CO2 storage in South Africa: A case study

The ultra-fine milled tailings generated during the processing of PGM ores in South Africa have a theoretical potential to sequester significant amounts of CO₂ (∼14 Mt per annum) through mineral carbonation. Mg-bearing orthopyroxene is the major sequestrable mineral in these tailings, which also contains significant quantities of Ca-bearing plagioclase, as well as minor quantities of clinopyroxene, olivine, serpentine and hornblende. In this study, the feasibility of using PGM tailings to sequester CO₂ has been investigated empirically using the two-step, pH swing method. The rates and extents of cation (Ca, Mg and Fe) extraction and subsequent carbonation were determined and compared. Both organic (oxalic and EDTA) and HCl solutions were utilised in the cation extraction step, which was conducted at time periods up to 8 h and at a temperature of 70 °C. The extents of cation dissolution were relatively low under all experimental conditions investigated, particularly for the case of Mg (between 3.3% and 5.0% extraction). A comparison of the extents of leaching with the mineralogical composition of the tailings indicated that the extracted Mg originated primarily from clinopyroxene, with the orthopyroxene remaining relatively inert under the experimental conditions. Subsequent carbonation of the acid leach solution after pH adjustment with NaOH resulted in the rapid formation of a number of carbonate minerals, including gaylussite (Na₂Ca(CO₃)₂·5(H₂O)), magnesite (MgCO₃), hydromagnesite (Mg₅(CO₃)₄(OH)₂·4H₂O), dolomite (CaMg(CO₃)₂), ankerite (Ca(Fe,Mg)(CO₃)₂), and siderite (FeCO₃). On the basis of these findings, further studies will be focused on developing a better understanding of the factors affecting the dissolution of Mg-bearing orthopyroxene minerals, and on exploring alternative leach reagents and conditions, with a view to developing a more effective process for the accelerated carbonation of PGM tailings.     

Production of η-alumina from waste aluminium dross

Processing of aluminium dross is one of the most challenging tasks because of its toxic nature. The dross generated while melting at various facilities is generally remelted with salts to recover residual metal values. The remaining residue dross contains mostly aluminium oxide, alloying elements and salts such as NaCl or KCl. This residue dross while stock piling creates pollution of the adjoining area as salts leach out to water stream and also emits harmful gases. In the present study domestic aluminium dross was treated for developing a suitable process flow sheet to obtain η-alumina a high valued product. Initially H₂SO₄ leaching was carried out for both un-washed and washed dross. With un-washed dross the leaching efficiency achieved was ∼71% but washing of dross followed by leaching raised the recovery to ∼84%. Washing of dross is essential to have higher alumina recovery and also to recover salt for recycling. The liquor obtained after treatment of the dross with acid was further processed to obtain aluminium hydroxide of amorphous nature by hydrolyzing aluminium sulphate with aqueous ammonia. The aluminium hydroxide was then subjected to calcinations which resulted in the formation of η-alumina at 900 °C.     

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.