Sulfuric acid leaching of polymetallic Abu Zeneima gibbsite-shale

The gibbsite shales belonging to the Paleozoic sequence of sedimentary rocks occurring at the Abu Zeneima area in SW Sinai Egypt has been found to host several economic metal values. These include Al, Cu, Zn, Co, Ni, REE, and U beside several Mn minerals. In this work, the ore was subjected to sulfuric acid leaching to maximize extraction of these metal values. The optimum leaching conditions were 800 g/L acid with a Solid/Liquid ratio of 1/2 at 100 °C for 4 h and using an ore size of − 60 mesh. Under these conditions, it was possible to achieve high leaching efficiencies with complete leaching of Al together with 93.0% Cu, 58.6% Zn, 69.0% Co, 92.6% Ni and 84.4% U. Extension of the leaching time to 8 and 12 h significantly improved the leaching efficiencies of Co and Zn to 91.0% and 84.0% respectively.     

Pressure leaching of high silica Pb-Zn oxide ore in sulfuric acid medium

The pressure leaching of high silica Pb-Zn oxide ore using air as pressurized gas in sulfuric acid medium is studied systematically in the present paper. Study parameters include the concentration of sulfuric acid, partial air pressure, leaching temperature, leaching time, and liquid-to-solid ratio. The experimental results obtained show that under optimum conditions, the extraction percentage of Zn is up to 96% and the extent of dissolution of Si, Pb, and Fe are as low as 1%, 2%, and 6%, respectively. Leaching pulp had good filter performance, and the silicic acid formed during sulfuric acid leaching was transformed into easily filterable dehydrated SiO₂ particles that can enter the slag. The filtration rate can be as high as 893 L/m² h.     

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%.     

Kinetics of reductive leaching of manganese oxide ore with molasses in nitric acid solution

The kinetics of reductive leaching of manganese from a low grade manganese ore in dilute nitric acid in the presence of molasses is examined. The rate is controlled by diffusion through the “product” layer composed of the associated minerals. The leaching process follows the kinetic model 1 − 2/3X − (1 − X)^2/3 = kt with an apparent activation energy of 25.7 kJ/mole. It is concluded that the concentration of HNO₃ and molasses as well as temperature are the main factors influencing the leaching rate. The results indicate a reaction order of 1.2 for HNO₃ concentration and 1.9 for molasses concentration.     

甲醛-硫酸亚铁协同还原浸出低品位氧化锰矿

在硫酸介质中以甲醛-硫酸亚铁为还原剂协同还原软锰矿,考察了甲醛-硫酸亚铁摩尔比、温度、反应时间、转速、硫酸浓度等因素对锰、铝的浸出率及溶液中铁和有机残留甲酸的影响.采用单因素实验获得较佳的还原工艺条件,采用HPLC测定溶液中的甲酸.结果表明,在固定转速为200 r/min、液固比为8 ml/g时,最佳反应条件为:甲醛-硫酸亚铁摩尔比1∶3(甲醛1.5 mol)、浸出时间3 h、硫酸浓度3 mol/L、温度90℃.在该条件下重复实验,锰的平均浸出率为93.51%,铝的平均浸出率为33.08%,铁的浓度为23.07 mol/L,甲酸浓度为0.001 g/L.     

Rare earth double sulfates from pre-concentrated bastnasite

A mixture of rare earth double sulfates was produced from a Turkish bastnasite-containing pre-concentrate (low grade concentrate) by sulfuric acid baking, subsequent water leaching and precipitation with sodium sulfate. The results of acid baking and leaching indicated that recoveries of rare earth elements up to 90% were readily obtained and the recovery of hydrofluoric acid as a by-product was also possible. Reasonable decontamination of the rare earth double sulfate salt from impurities such as Th, Fe, Al and Mg was possible by rapid precipitation at 50 °C using 1.25 times the stoichiometric amount of Na₂SO_4. The total rare earth double sulfate content (TREDS) was > 90% and analysed 17.3% La, 15.6% Ce, 3.2% Nd, 1.1% Pr, 0.3% Sm, 0.03% Eu, 0.01% Yb and 0.02% Y together with about 0.7% Ca, Fe, Al and other impurities.     

Arsenic removal from copper ores and concentrates through alkaline leaching in NaHS media

Removal of arsenic impurity in ores and concentrates containing copper (Cu) through alkaline leaching in NaHS media was investigated in this work. Samples containing Cu from 10 to 40 wt.% and arsenic from 0.8 to 14 wt.% with enargite (Cu₃AsS₄) as main arsenic bearing mineral were used as starting materials and all leaching tests were conducted at 80 °C under normal atmospheric pressure. Solution and/or slurry potential and pH were maintained consistently below − 500 mV (SHE) and above 12.5 respectively with the addition of NaHS and NaOH, creating a reducing environment for arsenic dissolution and conversion of Cu₃AsS₄ to Cu₂S. Pulp density ranged from 100 to 1000 g/L, NaHS and NaOH reagents were added at 50–200 g/L each and leaching time varied from 10 min to 10 h.Characterization of solid samples (original and leach residue) by XRD and XRF analyses and chemical analysis of both solid and solution samples by ICP analysis showed that Cu₃AsS₄ in the starting material was completely decomposed or transformed to Cu₂S and arsenic released into solution as As (III)/As³⁺ ions (Na₃AsS₃). Over 90% of arsenic in the starting materials was removed within 1–3 h for materials with arsenic content from 1 to 4 wt.% and within 3–6 h for materials with arsenic content over 4–10 wt.%. Dissolution and analysis of leach residues obtained after leaching by ICP indicated that arsenic in the starting materials has been reduced in all cases to below 0.5 wt.%. In all test conditions dissolution of Cu and Fe into solution was not detected, indicating selective leaching of arsenic. NaHS application for removal of arsenic in Cu-ores and/or concentrates was demonstrated in this work and further research is in progress to develop a process to include treatment of arsenic leached into solution.     

Leaching behavior of metals from a limonitic nickel laterite using a sulfation–roasting–leaching process

The leaching behavior of metals from a limonitic laterite was investigated using a sulfation–roasting–leaching process for the recovery of nickel and cobalt. The ore was mixed with water and concentrated sulfuric acid followed by roasting and finally leaching with water. Various parameters were studied including the amount of acid added, roasting temperature and time, sample particle size, addition of Na₂SO₄ and solid/liquid ratio in leaching process. More than 88% Ni, 93% Co and < 4% Fe are extracted under the determined conditions. Simultaneously, about 90% Mn and Cu, 70% Mg, 45% Al, 25% Zn, 4% Cr and Ca are extracted respectively. The pH of the leach solution is about 2. The leaching efficiency is independent of sample particle size due to decomposition of ferric sulfate formed during roasting. The roasted mass was characterized by various physico-chemical techniques such as DSC/TGA, XRD and SEM. This process provides a simple and effective way for the extraction of nickel and cobalt from laterite ore.     

Mechanochemical leaching of refractory zinc silicate (hemimorphite) in alkaline solution

A process was developed for extracting zinc from refractory zinc silicate (hemimorphite) by mechano-chemical leaching in sodium hydroxide solution with calcium oxide added to inhibit the leaching of silica. The effects of mechanical activation, calcium oxide dosage, temperature, leaching time, liquid–solid ratio and sodium hydroxide concentration were investigated. The dissolution of silica was considerably inhibited by adding CaO, but zinc extraction was almost unchanged. Under optimum conditions in 4.5 mol/L NaOH and twice the stoichiometric CaO at 383 K, the dissolution of Zn and SiO₂ were 93.6% and 7.3%, respectively, after 1 h using a liquid–solid ratio (L/S) of 10:1.     

Study of extraction and purification of Ni, Co and Mn from spent battery material

In spent battery material, there are plenty of valuable metals, such as copper, nickel, cobalt, manganese. Recovery of valuable metals from spent battery material not only protects the environment but also improves the utilization of resources and decreases the cost of battery material. In this study, hydrochloric acid is used as lixivant with characteristics of faster leaching rate and being recycled easily. The optimal conditions are that hydrochloric acid concentration is 6 mol/L, reaction temperature is exactly 60 °C, liquid/solid ratio is 8:1, (H₂O₂)_mol/(MeS)_mol = 2, and the leaching time is 2 h, the results show that the dissolution yields of Ni, Co and Mn can be 95 wt.% at least. The basic purification concept of the leaching solution includes that copper is removed through replacement by iron powder followed by iron precipitation in goethite method. The results show that Cu and Fe can be removed 99 wt.% at the least. At the same time, the loss of Ni, Co and Mn is not beyond 2 wt.%, 3 wt.％ and 2 wt.%, respectively. This method makes the preparation of pure Ni_xCo_yMn_z ternary system precursor economical. The process seems to be able to claim base metals from waste in a reliable and feasible way.     

铝土矿盐酸浸出过程研究

以澳大利亚难处理三水铝土矿为对象、盐酸为浸出剂进行无焙烧浸出试验,考察浸出温度、矿物粒度以及浸出时间对氧化铝浸出率的影响。结果表明,优化工艺条件为:矿物粒度-55μm、浸出温度100~110℃、浸出时间120 min、盐酸浓度10%、浸出液固比100∶7。此条件下氧化铝的浸出率为95.49%,氧化铁的浸出率为96.72%。以该酸浸液为原料,使用TBP-苯体系进行铝铁分离萃取试验,在萃取温度25℃、相比O/A=1∶1、盐酸浓度1.5 mol/L,萃取时间10 min的条件下,经单级萃取,溶液中铁元素的萃取率可达95%,铝元素损失率为6%,铁铝萃取分离系数为408,经3级以上逆流萃取,铁铝分离系数可达800以上。以纯水作为反萃剂,在温度25℃、相比O/A=1∶1、反萃时间5 min、单级反萃,铁的反萃率达95%。     

杂质离子对萃取法从石煤酸浸液中分离纯化钒的影响

含钒石煤经焙烧、硫酸浸出后,酸浸液中含有高浓度的Fe3+、Fe2+、Al 3+和Mg2+等杂质离子。以D2EHPA和TBP为萃取剂,磺化煤油为稀释剂,采用萃取法对该酸浸液进行钒的提纯试验,考察杂质离子对钒萃取率的影响。结果表明,在配制的纯溶液中,V4+的萃取率明显高于V5+;Fe3+质量浓度大于5g/L时会显著降低V4+萃取率;Al 3+和Mg2+的质量浓度低于10g/L时,其共萃率明显降低。对实际酸浸液进行还原处理后,99%的V4+能够被萃取回收,而大部分Fe2+、Al 3+和Mg2+则存在于萃余液中。少数共萃的Fe2+在反萃作业后留在贫有机相中,Al 3+和Mg2+在沉钒后留在沉钒尾水中,不影响V2O5纯度。     

Recovery of yttrium from deep-sea mud

Deep-sea mud rich in rare earth yttrium has received lots of attention from the international community as a new resource for Y. A novel process, which mainly includes acid leaching, solvent extraction, and oxalic acid precipitation-roasting, is proposed for recovery of Y from deep-sea mud. A series of experiments were conducted to inspect the impacts of various factors during the process and the optimum conditions were determined. The results show that the Y of deep-sea mud totally exists in apatite minerals which can be decomposed by hydrochloric acid and sulfuric acid solution. The highest leaching efficiency of Y is 94.53% using hydrochloric acid and 84.38% using sulfuric acid under the conditions of H~+concentration 2.0 mol/L, leaching time 60 min, liquid-solid ratio 4:1 and room temperature 25 ℃(only in case of sulfuric acid, when using hydrochloric acid, the leaching temperature should be 60 ℃). Because of the much lower leaching temperature, sulfuric acid leaching is preferred. The counter current extraction and stripping tests were simulated by a cascade centrifugal extraction tank device. Using 10 vol% P204,15 vol% TBP and 75 vol% sulfonated kerosene as extractant, 98.79% Y~(3+) and 42.60% Fe~(3+) are extracted from sulfuric acid leaching liquor(adjusted to pH = 2.0) after seven-stage counter current extraction with O/A ratio of 1:1 at room temperature, while other metals ions such as Al~(3+), Ca~(2+), Mg~(2+)and Mn~(2+) are almost not extracted. The Y~(3+) in loaded organic can be selectively stripped using 50 g/L sulfuric acid solution and the stripping efficiency reaches 99.86% after seven-stage counter current stripping with O/A ratio of 10:1 at room temperature, while only 2.26% co-extracted Fe~(3+) is stripped. The Y~(3+) of loaded strip liquor can be precipitated by oxalic acid to further separate Y~(3+) and Fe~(3+). The precipitation efficiency of Y~(3+) in loaded strip liquor can be 98.56% while Fe~(3+) is not precipitated under the conditions of oxalic acid solution concentration 200 g/L, quality ratio of oxalic acid to Y of 2, and precipitation time 0.5 h. And the precipitate was roasted at 850 ℃ for 3 h to obtain the oxide of Y in which the purity of Y_2 O_3/REO is 79.02% and the contents of major non-rare earth impurities are less than 0.21%.Over the whole process included acid leaching, solvent extraction, and oxalic acid precipitation-roasting,the yttrium yield is 82.04%.     

核桃壳还原浸出软锰矿的动力学

探究以核桃壳为还原剂硫酸浸出氧化锰矿过程的动力学。考察了搅拌速度、反应温度、硫酸浓度、反应时间以及核桃壳用量对锰浸出率的影响。结果表明,锰的浸出率随着搅拌速度、硫酸浓度、核桃壳用量的增大和温度的升高而增大。浸出前60 min浸出率的增长速度较快。在反应温度为369 K、硫酸浓度3.5 mol/L、核桃壳加入量40 g/L、反应时间2.5 h、转速200 r/min时,锰浸出率达93.18%。浸出过程属于化学反应控制,对应的活化能为45.5 kJ/mol,硫酸浓度和核桃壳用量的反应级数分别为0.897、0.2。     

废锂电池酸浸液非萃取除杂研究

主要研究了废锂电池酸浸液杂质除杂原理,采取“两段酸浸—中和除Fe、Al—絮凝除F—深度除杂”非萃取除杂工艺。在除杂过程中,由于酸浸液含高浓度Ni、Co、Mn离子,中和剂的种类和浓度将影响Ni、Co、Mn的损失率。通过工艺控制及中和剂调试,选择10%CaCO3作为除Fe、Al、F中和剂,Fe、Al、Cu、F含量可分别从0.20、9.76、0.58、1.66 g/L降至0.01、0.02、0.01、29.86 mg/L,达到三元前驱体溶液杂质标准要求。此时,Ni、Co、Mn的损失率分别仅为0.96%、0.04%、0.01%,均在接受范围之内。     

Leaching of a sphalerite concentrate with H2SO4–HNO3 solutions in the presence of C2Cl4

The enhanced leaching of sphalerite concentrates in H₂SO₄–HNO₃ solutions and the extraction of sulfur with tetrachloroethylene were studied. Variables of the process were investigated including leaching temperature, reaction time, liquid / solid ratio, and tetrachloroethylene concentration. The number of cycles that tetrachloroethylene could be recycled did not have a significant effect on zinc extraction. The results indicated that 99.6% zinc extraction was obtained after three hours of leaching at 85 °C and 0.1 MPa O₂, when 20 g of sphalerite concentrate were leached in a 200 ml solution containing 2.0 mol/L H₂SO₄ and 0.2 mol/L HNO₃, in the presence of 10 ml C₂Cl₄. Leaching rates were significantly improved under these conditions.     

A novel recovery process of metal values from the cathode active materials of the lithium-ion secondary batteries

A novel process was conducted with experiments which separated and recovered metal values such as Co, Mn, Ni and Li from the cathode active materials of the lithium-ion secondary batteries. A leaching efficiency of more than 99% of Co, Mn, Ni and Li could be achieved with a 4 M hydrochloric acid solution, 80 °C leaching temperature, 1 hour leaching time and 0.02 gml^− 1 solid-to-liquid ratio. For the recovery process of the mixture, firstly the Mn in the leaching liquor was selectively reacted and nearly completed with a KMnO₄ reagent, the Mn was recovered as MnO₂ and manganese hydroxide. Secondly, the Ni in the leaching liquor was selectively extracted and nearly completed with dimethylglyoxime. Thirdly, the aqueous solution in addition to the 1 M sodium hydroxide solution to reach pH = 11 allowed the selective precipitation of the cobalt hydroxide. The remaining Li in the aqueous solution was readily recovered as Li₂CO₃ precipitated by the addition of a saturated Na₂CO₃ solution. The purity of the recovery powder of lithium, manganese, cobalt and nickel was 96.97, 98.23, 96.94 and 97.43 wt.%, respectively.     

Intensification of sulphation and pressure acid leaching of nickel laterite by microwave radiation

The microwave is a clean and environmentally acceptable energy and various microwave-assisted metal extraction processes have been developed. The novel application of microwave irradiation for sulphation of the nickel laterite ore prior to pressure sulphuric acid leaching has been investigated. Under optimum conditions, about 92% of nickel can be extracted from the ore and above 90% of the iron precipitates as hematite with the residual acid content in the leachate lower than 31 g/L. The microwave sulphation proceeds rapidly and a relatively low leaching temperature and low mass ratio of sulphuric acid to ore can be used in the subsequent pressure leaching to achieve a satisfactory nickel extraction. This may potentially reduce the subsequent heavy capital and operation cost of the pressure acid leaching process.     

P204从高铝粉煤灰硫酸浸液中萃取除铁

主要针对高铝粉煤灰酸浸液中铁铝萃取分离性能进行研究,通过模拟高铝粉煤灰酸浸液(C(Al3+)=37.1 g/L,C(Fe3+)=5.1 g/L),采用P204+260#溶剂油萃取体系,对组成液中铁铝进行萃取分离,系统研究了P204体积分数、温度、反应时间、溶液初始pH值、相比V(O):V(A)等因素对萃取和反萃的影响,...     

Acid Refining of Synthesised Alloys Within the Quaternary System Fe–Si–Ca–AI: Effect of Ferric Ions

A systematic study of synthesis of different chemical composition of alloys within the quaternary Fe–Si–Ca–Al system, versus acid leaching refining using HCl with or without FeCl₃ has been undertaken. Such study has the aim of achieving an adequate structural composition of Fe–Si alloys, in order to obtain high purity and fine particle size silicon by acid leaching, at more reduced costs. The experimental results show that an adequate control of Fe/Al and Si/Ca concentration ratios can hinder the formation of tetragonal and orthorhombic FeSi₂ phases, and consequently promotes a successful refining of ferro-silicon alloys. The presence of ferric ion influences on a distinct way the rate of removal of Al–Fe–Si and Fe–Al–Si–Ca phases, increasing the first and decreasing the second one. The suitable starting composition which could give the expected results must be the following: 92–94% Si; 3–3.5% Fe; 3.5–4% Al and Ca ≤ 0.5%.