Extraction of vanadium from black shale using pressure acid leaching

The extraction of vanadium from black shale was attempted using pressure acid leaching. The effects of the several parameters which included reaction time, concentration of sulfuric acid, leaching temperature, liquid to solid ratio and concentration of additive (FeSO₄) upon leaching efficiency of vanadium were investigated and a two-step counter-current leaching approach was developed. The results showed that the leaching efficiency of vanadium in the two-step process could reach above 90%. Vanadium was effectively separated and enriched by solvent extraction after leachate pretreatments, including the reduction of Fe³⁺ and adjustment of pH value. The extraction and stripping yields of vanadium were both > 98%. Ammonia was added to a stripping liquor to precipitate vanadium and then the ammonium poly-vanadate produced was calcined at 550 °C for 3 h to produce the high purity V₂O₅ powder. The overall yield of vanadium through all process stages was about 85%.     

Preparation of Sm2O3 and Co3O4 from SmCo magnet swarf by hydrometallurgical processing in chloride media

The recycling of rare earth elements(REE) from end-of-life REE based products is an environment friendly proposition. Waste Sm-Co based permanent magnet generated during machining is a good source for both Sm and Co. In the present study a simpler process of acid leaching at 80 ℃ followed by solvent extraction, oxalate precipitation and calcination is described for producing pure Sm_2 O_3 and Co_3 O_4. With either 10 vol% H_2SO_4 or 15 vol% HCI at 80 ℃ more than 95% Sm and Co are leached in 1 h.Extraction of Sm from sulphate leach liquor with TBP or Aliquat 336 was poor. Although extraction with TOPS-99 is quantitative but Sm from sulphate leach liquor precipitated as Sm_2(SO_4)_3·8 H_2 O. The chloride leach liquor at an initial pH of 2.5 and with 1.2 mol/L TOPS-99 shows requirement of 4-stages at A:O = 3:2. Stripping with oxalic acid precipitates Sm-oxalate which is calcined at 800 ℃ to produce Sm_2 O_3. Co_3 O_4 is recovered from the raffinate through oxalate precipitation followed by calcination at450℃.     

Production of high purity cobalt oxalate from spent ammonia cracker catalyst

A process was developed to produce > 99.9% pure cobalt oxalate from spent ammonia cracker catalyst pellets containing ∼ 20% Co generated at heavy water plants. A pilot plant for producing kilograms of cobalt oxalate of required purity has been set up. The process consists of leaching, oxidation, solvent extraction, ion exchange and oxalate precipitation. The major impurities present in the leach liquor after nitric acid oxidation were Fe³⁺, Al³⁺ and Ni²⁺. Fe³⁺ and Al³⁺ were separated from the leach liquor by using a solvent extraction (SX) process employing a mixed extractant system consisting of D2EHPA and TBP. Ni²⁺ was separated by ion exchange (IX) employing Dowex M4195 resin.     

Extraction of Ni (II) from Spent Hydrodesulfurization HDS Catalyst Through Leaching and Electroless Precipitation of Ni(OH)2

The extraction of nickel (II) from a spent hydro-desulfurization catalyst containing 11.6 pct Ni was carried out through sulfuric acid leaching. Variations of parameters such as the concentration of acid, temperature, and time, were studied and optimized. Nickel hydroxide was precipitated from the leach liquor via neutralization with 1 M sodium hydroxide up to pH 12 in three different methods: normal neutralization precipitation, and then neutralization precipitation followed by aging at 353 K (80 °C) for 4 hours and neutralization of the leach liquor with 10 pct (v/v) of 0.1 N sodium lauryl sulfate. X-ray diffraction (XRD) and transmission electron microscopy (TEM) microanalysis shows a difference in crystallinity with the method of precipitation. The nickel hydroxide contains Cu(II), Co(II), Zn(II), and Mn(II) as trace impurities. The discharge capacities of the precipitated nickel hydroxides were 120 mAhg^?1, 140.72 mAhg^?1, and 145.2 mAhg^–1 for aged sample, sample without surfactant, and with surfactant respectively.     

Removal of iron from quartz sands: A study by column leaching using a complete factorial design

The aim of the work was to decrease the iron content of ferrous quartz sands by fixed-bed column leaching with recycling of the leaching solutions in order to attain a product suitable for industrial use. Dissolution of iron was achieved by treating the sands in an acid medium with a reducing agent (oxalic acid) to convert Fe^III into Fe^II.The factors assumed to affect dissolution of iron, such as temperature, oxalic acid concentration, pH and flow-rate, were studied with a 2⁴ full factorial design in order to assess the main effects and the interactions among the factors.Removal of 46.1% iron gives a product containing 0.0163% Fe₂O₃ which is fit for industrial applications.     

Dissolution behaviour of a beryl ore for optimal industrial beryllium compound production

This study involves the leaching of the beryl ore with sulphuric acid (H₂SO₄) solution for predicting optimal beryllium extraction conditions with the aim of assessing the importance of leachant concentration, reaction temperature and particle size on the extent of dissolution. A kinetic model to represent the effects of these variables on the leaching rate was developed. It was observed that the dissolution of beryl ore increases with increasing H₂SO₄ concentration, temperature, decreasing particle size and solid to liquid ratio. At optimal leaching conditions, 89.3% of the ore was reacted by 1.25?mol/L at 75°C temperature and 120 minutes with moderate stirring, where 1612.0?mg/L Be²⁺, 786.7?mg/L Al³⁺, 98.1?mg/L Fe³⁺ and 63.4?mg/L Ag⁺ were found as major species in the leach liquor. The unleached products constituting about 10.7% were examined by X-ray diffraction (XRD) and found to contain primarily, siliceous compounds such as Xonotlite, Antigorite, Chrysolite and Kaolinite.     

Direct versus indirect bioleaching

The dissolution of metal sulfides is controlled by their solubility product and thus, the [H⁺] concentration of the solution, and further enhanced by several chemical mechanisms which lead to a disruption of sulfide chemical bonds. They include extraction of electrons and bond breaking by [Fe³⁺], extraction of sulfur by polysulfide and iron complexes forming reactants [Y⁺] and electrochemical dissolution by polarization of the sulfide [high Fe³⁺ concentration]. All these mechanisms have been exploited by sulfide and iron-oxidizing bacteria. Basically, the bacterial action is a catalytic one during which [H⁺], [Fe³⁺] and [Y⁺] are breaking chemical bonds and are recycled by the bacterial metabolism. While the cyclic bacterial oxidative action via [H⁺] and [Fe³⁺] can be called indirect, bacteria had difficulties harvesting chemical energy from an abundant sulfide such as FeS₂, the electron exchange properties of which are governed by coordination chemical mechanisms (extraction of electrons does not lead to a disruption of chemical bonds but to an increase of the oxidation state of interfacial iron). Here, bacteria have evolved alternative strategies which require an extracellular polymeric layer for appropriately conditioned contact with the sulfide. Thiobacillus ferrooxidans cycles [Y⁺] across such a layer to disrupt FeS₂ and Leptospirillum ferrooxidans accumulates [Fe³⁺] in it to depolarize FeS₂ to a potential where electrochemical oxidation to sulfate occurs. Corrosion pits and high resolution electron microscopy leave no doubt that these mechanisms are strictly localized and depend on specific conditions which bacteria create. Nevertheless, they cannot be called ‘direct’ because the definition would require an enzymatic interaction between the bacterial membrane and the cell. Therefore, the term ‘contact’ leaching is proposed for this situation. In practice, multiple patterns of bacterial leaching coexist, including indirect leaching, contact leaching and a recently discovered cooperative (symbiotic) leaching where ‘contact’ leaching bacteria are feeding so wastefully that soluble and particulate sulfide species are supplied to bacteria in the surrounding electrolyte.     

Leaching of Chalcopyrite Concentrate with Hydrogen Peroxide and Sulfuric Acid in an Autoclave System

In this study, leaching of chalcopyrite concentrate was investigated in an autoclave system using hydrogen peroxide and sulfuric acid. By decomposition of hydrogen peroxide, the active oxygen formed can provide both high oxidation potential and high pressure in a closed vessel for leaching. Preliminary studies showed that hydrogen peroxide can be used as an oxidant instead of oxygen gas in the autoclave. Central composite design (CCD) was used to examine the effects of the experimental parameters on the copper and iron extraction as a response. The proposed model equation using CCD showed good agreement with experimental data, the correlation coefficients R ² for copper and iron being 0.84 and 0.86, respectively. The optimum conditions to obtain the main goal of maximum copper and minimum iron extraction from chalcopyrite were determined as to be sulfuric acid concentration of 2.5 M, hydrogen peroxide concentration of 2.3 M, leaching time of 24 minutes, chalcopyrite amount of 3.17 g (in 50-mL solution), stirring speed of 630 rpm, and leaching temperature of 351 K (78 °C). Under the optimum condition, 76 pct of copper and 9 pct of iron were extracted from chalcopyrite concentrate. Extraction yield results of metals indicate that selective leaching of chalcopyrite can be achieved using hydrogen peroxide and sulfuric acid in an autoclave system.     

酸浸-萃取-氨沉淀法从石煤钒矿中提取钒

采用氧化焙烧脱炭-硫酸氧化浸出-P204+TBP溶剂萃取-氨水沉钒的工艺方法,从江西某石煤钒矿中提取V2O5,考察了硫酸用量、萃取及反萃次数、反萃液pH值对工艺过程的影响。试验结果表明:钒矿破碎后在硫酸溶液中用氯酸钠进行氧化浸出,钒的浸出率可达到96%以上;用P204+TBP溶剂萃取和稀硫酸溶液反萃,再用氨水沉淀钒,最终得到纯度98.0%以上的V2O5产品;从石煤钒矿到V2O5的总收率可达86.14%～93.09%。该工艺对钒的回收效果明显,操作简单,生产成本低,对环境污染较小。     

Application of P507 and isooctanol extraction system in recovery of scandium from simulated red mud leach solution

The current recovery technique of Sc was complicated and the chemical consumption was high. This was due to the low content of Sc in resources and the difficulty of stripping. In this research, the isooctanol was added into the 2-ethylhexyl phosphonic acid mono-2-ethylhexyl ester (P507) extraction system to reduce the extraction and improve the stripping of Sc. The maximum stripping ratio of Sc from loaded organic phase by sulfuric acid can increase from 10% (without isooctanol) to 99% (with 15 vol% isooctanol). In the extraction test of the simulated red mud leaching liquor, the separation factors between Sc and Zr, Sc and Ti are 36 and 350, separately. At the same time, other metals are almost not extracted. The high selectivity and stripping of Sc suggest that the P507 with isooctanol extraction system can be applied in the practical Sc recovery process.     

用LIX84从富钴结壳硫酸浸出液中选择性萃取铜

采用LIX84作萃取剂、硫酸作反萃剂 ,从大洋富钴结壳常温常压活化硫酸浸出除铁后液中萃取铜。试验考察了相比、平衡水相pH值、时间等因素对LIX84萃铜的影响。结果表明 ,相比、平衡水相 pH值、混合时间都对铜的萃取率有一定影响。最后优化出的萃取工艺条件为 (体积百分数 )有机相 12 %LIX84+ 88%煤油 ,室温 ,相比 (O/A)=1/ 2 0 ,出口水相pH2 60± 0 0 5 ,萃取级数为 2级 ,每级混合时间 5min。经过 2级萃取、1级洗涤、3级反萃后 ,可以得到完全符合电解沉积要求的硫酸铜溶液 ,从而使浸出液中的铜与其它金属彻底分离     

Microbial Dissolution of Zn-Pb Sulfide Minerals Using Mesophilic Iron and Sulfur-Oxidizing Acidophiles

The purpose of this study is to test the feasibility of using mixed culture of iron and sulfur-oxidizing bacteria for the dissolution of metals from high-grade zinc and lead sulfide ore. Considering that the roll crusher could reduce the ore size to less than 2 mm, this size fraction was selected in order to study the possibility of removing mill circuit. Effects of parameters such as pulp density, initial pH, Fe²⁺, oxidation–reduction potential (ORP), and pH fluctuations were investigated, as well. The maximum Zn dissolution was achieved under the conditions of initial pH 2, initial 75 g/L FeSO₄ · 7H₂O, and pulp density of 50 g/L. The results indicated that under the optimum conditions, about 68.8% of zinc was leached during 24 days of bacterial leaching treatment. The lead recoveries were low (about 1%), because of precipitation of Pb as lead arsenate chloride. Furthermore, the surface studies by using SEM images showed that during chemical leaching the ore dissolution starts from surface discontinuities, but in bacterial leaching all surface becomes involved. In addition, in another process the ore was leached separately with sulfuric acid and sodium hydroxide, and then final results were compared to the bacterial leaching tests in order to find the optimum hydrometallurgical method to extract zinc and lead from these ores.     

钛白废酸与赤泥联合提取氧化钪的工艺研究

以钛白废酸、赤泥为原料,经过浸出、一次萃取反萃、酸溶水解、二次萃取、酸洗、二次反萃、草酸沉淀、精制等提纯工序,可得到纯度99.99%的Sc2O3。赤泥和钛白废酸中钪的回收率分别达到57.8%和93.3%。     

Separation of aluminum from rare earth by solvent extraction with4-octyloxybenzoic acid

Leaching method is usually used to extract rare earth(RE) elements from ion adsorbed RE ores.In the leaching process,some impurities such as aluminum(Al) enter the leaching solution.The separation of Al from RE by carboxylic acid extractant 4-octyloxybenzoic acid(POOA) was studied in this article.By changing the pH value,temperature,solvent,saponification degree and other parameters,the extraction and separation performance of POOA in chloride system was systematically studied.Through specific e...     

Leaching Kinetics of Bauxite in Hydrochloric Acid

The Bayer process is currently used to produce cell-grade alumina from bauxite. However, if the reactive silica content in the bauxite exceeds 7%, losses of caustic and aluminium as sodalite (2 Na₂0.2 Al₂O₃·3 SiO₂·2 H₂O) become economically unacceptable. As Australia possesses large tonnages of bauxite containing more than 7% reactive silica, a process for the production of cell-grade alumina horn these bauxites using hydrochloric acid leaching was developed by the authors.

The process consists of the following steps: calcination of bauxite, leaching of calcined bauxite in hydrochloric acid, filtration of residue, crystallization of A1C1₃.6 H₂O, decomposition of AlCl₃.6 H₂O crystals to produce A1₂O₃, regeneration of hydrochloric acid for recycle into the leaching step.

The kinetic studies of the dissolution showed that the extraction of aluminium is independent of the solid to leachant ratios studied. The acid concentration has a marked effect on leaching kinetics of bauxite and the leaching time decreases substantially as the concentration of acid increases. The order of the reaction is greater than one. The rate equation can be described by the following least squares line of best fit.

r = 4.1 × 10^?5 HCl^0.5 + 1.27 × 10^?5 HCl² where

r = the initial reaction rate, g Al extracted/sec.

Aluminium extraction greater than 90% can be achieved from bauxite in 4 h using 25% HCl while 30% acid is considered the maximum acid strength due to the crystallization of A1C1₃.6 H₂O from leach solutions.

The rate of leaching increases as the temperature of leaching is increased. An Arrhenius plot of the initial reaction rate (g Al extracted/sec) against the inverse absolute leaching temperature produced a straight line. The activation energy for the reaction was determined to be 83.3 kJ/mole with a standard deviation of 2.9 kJ/mole. The correlation between the initial reaction rate and the inverse absolute temperature can be described by the following regression equation:

r = 4.71×l0⁸e^83820/RT

The large activation energy value suggests that the reaction is chemically controlled rather than diffusion controlled. The additives such as NaCl and FeCI₃ do not effect the leaching rate.     

Treatment of Chloride Waste Pickle Liquor by Solvent Extraction for the Recovery of Iron

A systematic study of the extraction of Fe(III) from chloride waste pickle liquor has been investigated using Cyanex 923 diluted with kerosene to recover iron values from the pickle liquor. Various parameters were studied to optimize the conditions for maximum recovery of iron. Extraction increases with increasing concentration of both hydrochloric acid and extractant. The species extracted into the organic phase appears to be HFeCl₂ with 1 M of the solvent. Effect of various salts as additives on Fe extraction was also studied and it was found that addition of NaCl enhanced the extraction about 2.5 times as compared to that without its addition.

Saturated loading capacity was found to be 60.9 g/L Fe in four contacts at O/A of 1. The stripping of Fe(III) with different concentration of hydrochloric acid and nitric acid from the loaded Cyanex 923 was found to increase up to 1 M of both the acids and then decrease with further increase in acid concentration up to 10 M. However, 100% stripping efficiency of Fe(III) was achieved with 0.8 M oxalic acid in two countercurrent stages at an aqueous:organic phase ratio of 3:1. Extraction parameters for maximum extraction of Fe(III) were optimized.     

Hydrometallurgical Process for Copper Recovery from Waste Printed Circuit Boards (PCBs)

Present paper focuses on the selective recovery of copper from the enriched ground printed circuit boards (PCBs) using leaching and solvent extraction. The metal-enriched ground sample obtained from the beneficiation of the sized PCBs in a laboratory scale column type air separator contained mainly 49.3% Cu, 3.83% Fe, 1.51% Ni, 5.45% Sn, 4.71% Pb, and 1.85% Zn. The leaching of the enriched sample with 3.5 mol/L nitric acid dissolved 99% copper along with other metals at 323 K temperature and 120 g/L pulp density in 1 h time. The composition of the leach liquor with wash solution was found to be 42.11 g/L Cu, 2.12 g/L Fe, 4.02 g/L Pb, 1.58 g/L Zn, and 0.4 g/L Ni. The McCabe–Thiele plot indicated the requirements of three counter-current stages for maximum extraction of copper from the leach liquor at pH 1.5 using 30, 40, and 50% (v/v) LIX 984 N at the phase ratios (A/O) of 1:3, 1:2, and 1:1.5, respectively. The counter-current simulation studies show the selective extraction of 99.7% copper from the leach liquor feed of 1.5 pH in three stages with 50% LIX 984 N at A/O phase ratio of 1:1.5. The stripping of copper from the loaded organic with sulfuric acid produced copper sulfate solution from which copper metal/powder could be recovered by electrolysis/ hydrogen reduction.     

Recovering REEs from NdFeB wastes with high purity and efficiency by leaching and selective precipitation process with modified agents

A leaching and selective precipitation approach is proposed in this work to recover rare earth elements (REEs) from NdFeB magnet wastes collected from industry. Hydrochloric acid and oxalic acid were employed as the leaching and precipitation agents, respectively. Hexamethylenetetramine (HMTA) or tartaric acid was used as the chelating agent during leaching. Both leaching and precipitation processes were optimized individually. For leaching process, the effects of two different chelating agents, the concentrations of leaching agent, chelating agent, and temperature on the extraction and recovery yields were investigated. The optimized process based on the factorial experiment was determined to be the hydrochloric acid concentration of 6 mol/L, the tartaric acid concentration of 50 g/L, and the temperature of 313 K, by which the extraction yields of Fe and REEs up to 67.99% and 99.27%, respectively, are obtained. For the precipitation process, the optimized oxalic acid dosage and pH value were also determined. The produced RE oxide products have the purity and recovery yield up to 95.83% and 90.18%, respectively. These results indicate that the present method with low acid consumption and high product purity has advantages over many other approaches for REE recovery.     

High-efficiency simultaneous extraction of rare earth elements and iron from NdFeB waste by oxalic acid leaching

Iron can not be recovered at high value because only rare earth elements are effectively recovered from NdFeB waste via oxidation roasting-hydrochloric acid leaching process.In this study,a new method for leaching NdFeB waste with oxalic acid was developed.The high-efficiency,simultaneous and high-value recovery of rare earth elements and iron was realized to simplify the process and improve the economic benefit.Results of the oxalic acid leaching experiments show that under the optimum leaching conditions at 90℃ for 6 h in the aqueous solution of oxalic acid(2 mol/L) with a liquid-solid ratio of60 mL/g,the iron leaching efficiency and precipitation rate of rare earth oxalate reach 93.89% and 93.17%,respectively.Rare earth oxalate and Fe(C₂O₄)3^3- were left in the residue and the leaching solution,respectively.The leaching mechanism was further analyzed by characterising the leach residues obtained through X-ray powder diffraction(XRD) and scanning electron microscopy-energy dispersive X-ray spectroscopy(SEM-EDS).Results of the leaching kinetics study indicate that the process of oxalic acid leaching follows the shrinking nucleus model,and the leaching kinetics model is controlled by the mixed factors of diffusion and chemical reaction.The leaching residue was calcined at 850℃ for 3 h and then decomposed into rare earth oxide,which can be directly used to prepare rare earth alloy via molten salt electrolysis.For the leaching solution,ferric oxalate solution was reduced using Fe powder to prepare the ferrous oxalate(FeC₂O₄-2H₂O).     

The ferric fluosilicate leaching of lead concentrates: Part I. Kinetic Studies

A new hydrometallurgical leaching process, which dissolves lead concentrates with acidified ferric fluosilicate solution, has been investigated for the selective extraction of lead and zinc from lead concentrates containing galena. The leaching of the Pine Point lead concentrate by ferric fluosilicate solutions was studied under various experimental conditions in the temperature range 20 °C to 95 °C. Temperature had a pronounced effect on the dissolution of the concentrates. The rates of lead leaching are very rapid over the temperature range 38 °C to 95 °C. The kinetics of zinc extraction are much lower than those of lead extraction. The reaction rates for the dissolution of galena were found to be controlled by surface chemical reaction. The apparent activation energy of the leaching reaction was calculated to be 62.1 kJ/mol. The initial concentrations of Pb²⁺, H⁺, and Fe³⁺ in the lixiviant do not have a significant effect on the rate or extent of lead extraction under the experimental conditions in this study.