Metal separation from mixed types of batteries using selective precipitation and liquid-liquid extraction techniques

The purpose of this paper is to study metal separation from a sample composed of a mixture of the main types of spent household batteries, using a hydrometallurgical route, comparing selective precipitation and liquid-liquid extraction separation techniques. The preparation of the solution consisted of: grinding the waste of mixed batteries, reduction and volatile metals elimination using electric furnace and acid leaching. From this solution two different routes were studied: selective precipitation with sodium hydroxide and liquid-liquid extraction using Cyanex 272 [bis(2,4,4-trimethylpentyl) phosphoric acid] as extracting agent. The best results were obtained from liquid-liquid extraction in which Zn had a 99% extraction rate at pH 2.5. More than 95% Fe was extracted at pH 7.0, the same pH at which more than 90% Ce was extracted. About 88% Mn, Cr and Co was extracted at this pH. At pH 3.0, more than 85% Ni was extracted, and at pH 3.5 more than 80% of Cd and La was extracted.     

废弃SCR催化剂中钒和钨的浸出及回收

采用NaOH溶液一次性浸出废弃SCR催化剂中的钒和钨,并用硫酸对浸出液进行除杂,再利用NH4Cl和硫酸分步对浸出液中的钒和钨进行沉淀回收。在NaOH质量分数40%、液固比8、浸出时间4 h、浸出温度90℃的最佳碱浸条件下,钒和钨的浸出率分别达到90.44%和84.49%。除杂过程的铝去除率达到100%,硅去除率达到77.56%。在沉钒pH为8.0、n(NH_4~+)∶n(V)为4的最佳沉钒条件下,钒回收率达到82.79%。在n(SO_4~(2-))∶n(W)为2的最佳沉钨条件下,钨回收率达到76.41%。     

Reducing acid leaching of manganiferous ore: effect of the iron removal operation on solid waste disposal

The process of reducing acid leaching of manganiferous ore is aimed at the extraction of manganese from low grade manganese ores. This work is focused on the iron removal operation. The following items have been considered in order to investigate the effect of the main operating conditions on solid waste disposal and on the process costs: (i) type and quantity of the base agent used for iron precipitation, (ii) effective need of leaching waste separation prior to the iron removal operation, (iii) presence of a second leaching stage with the roasted ore, which might also act as a preliminary iron removal step, and (iv) effect of tailings washing on the solid waste classification. Different base compounds have been tested, including CaO, CaCO3, NaOH, and Na2CO3. The latter gave the best results concerning both the precipitation process kinetics and the reagent consumption. The filtration of the liquor leach prior to iron removal was not necessary, implying significant savings in capital costs. A reduction of chemical consumption and an increase of manganese concentration in the solution were obtained by introducing secondary leaching tests with the previously roasted ore; this additional step was introduced without a significant decrease of global manganese extraction yield. Finally, toxicity characteristic leaching procedure (TCLP) tests carried out on the leaching solid waste showed: (i) a reduction of arsenic mobility in the presence of iron precipitates, and (ii) the need for a washing step in order to produce a waste that is classifiable as not dangerous, taking into consideration the existing Environmental National Laws.     

工业含钒废水处理工艺的研究进展

针对大量企业采用传统提钒法制备钒产品产生的工业含钒废水的处理问题,综述了工业含钒废水处理工艺的研究进展,分别详述了沉淀法、溶剂萃取法、离子交换法、电解法、吸附法和生物法的处理工艺,并对比分析了它们的优势与不足。指出:工业含钒废水的处理既是环保的要求,又可实现有价资源的回收利用;在实际应用时,可依据工业含钒废水的p H选择相应的处理工艺及处理试剂。     

从废FCC催化剂和废汽车尾气净化催化剂中回收稀土的研究进展

以稀土元素(简称稀土)用量较大的废FCC催化剂和废汽车尾气净化催化剂为研究对象,分析总结了两类废催化剂中稀土的成分和回收方法。从废催化剂中回收稀土普遍采用先浸出、后分离提纯的方法。浸出普遍采用无机酸(多为盐酸),分离方法包括溶剂萃取法、化学沉淀法等,最后经焙烧得到稀土氧化物。为从废催化剂中高效回收稀土,可着重考察新浸出机制的引入、分离方式的选择、分离试剂的应用,以及浸出和分离条件的优化,从而为实现工业化回收提供技术支持。     

Leaching characteristics of rare metal elements and chlorine in fly ash from ash melting plants for metal recovery

In terms of resource recovery and environmental impact, melting furnace fly ash (MFA) is attracting much attention in Japan due to its high metal content. The study aims to obtain fundamental information on using a water extraction method not only to concentrate valuable rare metals but also to remove undesirable substances such as chlorine for their recovery from MFA. The composition and leaching characteristics of MFA was investigated. The results revealed that the metal content in MFA is nearly equal to raw ore quality. The content of Ag, In, Pd, Pb, and Zn is, in fact, higher than the content of raw ore. As for leaching behavior, Ag, Bi, In, Ga, Ge, Sb, Sn, and Te showed the lowest release at a neutral pH range. Pd was leached constantly regardless of pH, but its concentration was quite low. On the other hand, most of the Tl was easily leached, revealing that water extraction is not appropriate for Tl recovery from MFA. Major elements Cl, Ca, Na, and K, occupying about 70% of MFA, were mostly leached regardless of pH. Base metal elements Cu, Pb, and Zn showed minimum solubility at a neutral pH. The leaching ratio of target rare metal elements and base metal elements suggests that the optimal pH for water extraction is 8-10, at which the leaching concentration is minimized. The water extraction process removed most of the Cl, Ca, Na, and K, and the concentration of rare metals and base metals increased by four or five times.     

Hydrometallurgical recycling of cobalt from zinc plants residue

A large amount of hot filter cake (HFC) is annually generated in Iranian zinc plants. It contains 1% zinc, 16–30% manganese, 5–25% calcium and 1–4.5% cobalt. Usually, zinc is selectively leached by an alkaline medium and its residue is known as alkaline leached HFC (ALHFC). In the present study, the possibility of cobalt extraction from ALHFC was investigated using a creative hydrometallurgical process. At the first stage, zinc and cadmium were selectively removed with sulfuric acid. At the second stage, it was deeply focused on the possibility of selective reductive leaching of cobalt by H₂O₂ as a reductant in the presence of manganese. As results, several differences were found between the mechanism of cobalt and manganese leaching. Accordingly, cobalt leaching was more affected by acid concentration and manganese leaching was more affected by reductant concentration. Consequently, with manipulating these important parameters, it was made possible to selectively separate cobalt from manganese. Based on the obtained results, 90.9% of cobalt and only 10.04% of manganese were leached with 1% of H₂O₂. At the third stage, pregnant cobalt solution was successfully purified through a solvent extraction process with D2EHPA. Finally, cobalt hydroxide as our final product with a purity of more than 99% was precipitated from the pure pregnant solution at 70 °C.     

生态修复植物蜈蚣草中砷的回收

采用管式炉高温热解-NaOH-Na₂CO₃混合液碱浸-CuSO₄·5H₂O沉淀的方法回收生态修复植物蜈蚣草中的砷，最终得到产品砷酸铜。该方法的最佳工艺条件为：热解温度600 ℃，热解时间30 min，CaO加入量（CaO与蜈蚣草的质量比）8%； m（NaOH）∶m（Na₂CO₃）=1∶3，碱浸温度70 ℃， 碱浸时间2 h， 固液比1∶10； 沉淀反应pH 5， 沉淀反应温度70 ℃。采用该方法处理生态修复植物蜈蚣草，得到产品砷酸铜的纯度为93%，砷回收率达88%。     

有机溶剂浸取石油污染土壤中的多环芳烃

采用索氏提取法和气相色谱分析了某炼油厂周边土壤中的多环芳烃的种类及含量,考察了6种有机溶剂对土壤中多环芳烃的浸取效果,探讨了溶剂与溶质溶解度参数差异对浸取效果的影响。结果表明：土壤试样中含有蒽、荧蒽、芘、 、苯并[a]蒽、苯并[b]荧蒽、苯并[k]荧蒽和苯并[a]芘8种多环芳烃,其含量分别为4.20,22.05,10.62,4.26,5.54,0.80,0.94,4.18 mg/kg;筛选出二氯甲烷作为土壤中多环芳烃的浸取溶剂;最佳浸取条件为浸取时间10 min、浸取温度30 ℃、溶剂与土壤的液固比5:1（mL/g）、土壤含水量8%,在此条件下,总多环芳烃浸出率为83.0%,各种多环芳烃的浸出率分别为蒽97.8%、荧蒽78.2%、芘99.9%、 98.5%、苯并[a]蒽81.1%、苯并[b]荧蒽47.6%、苯并[k]荧蒽14.8%、苯并[a]芘58.7%。     

Processing of Egyptian boiler-ash for extraction of vanadium and nickel

Proposed technique in this investigation is given for vanadium and nickel enrichment in the Egyptian boiler ash. Among the various concepts for recovery of vanadium and nickel from boiler ash, the pyro-metallurgical approach is technically feasible, but is not cost-effective from an operational economy standpoint. Another technically viable process which, however, needs further development and presented in this investigation, is the hydrometallurgical processing that involves acid leaching under oxygen pressure of ground ash, followed by electrolytic separation of nickel from sulphate solution and vanadium is then neutralized and precipitated by adjustment the pH value and calcined to produce V2O5.     

Process development for recovery of copper and precious metals from waste printed circuit boards with emphasize on palladium and gold leaching and precipitation

A novel hydrometallurgical process was proposed for selective recovery of Cu, Ag, Au and Pd from waste printed circuit boards (PCBs). More than 99% of copper content was dissolved by using two consecutive sulfuric acid leaching steps in the presence of H₂O₂ as oxidizing agents. The solid residue of 2nd leaching step was treated by acidic thiourea in the presence of ferric iron as oxidizing agent and 85.76% Au and 71.36% Ag dissolution was achieved. The precipitation of Au and Ag from acidic thiourea leachate was investigated by using different amounts of sodium borohydride (SBH) as a reducing agent. The leaching of Pd and remained gold from the solid reside of 3rd leaching step was performed in NaClO-HCl-H₂O₂ leaching system and the effect of different parameters was investigated. The leaching of Pd and specially Au increased by increasing the NaClO concentration up to 10 V% and any further increasing the NaClO concentration has a negligible effect. The leaching of Pd and Au increased by increasing the HCl concentration from 2.5 to 5 M. The leaching of Pd and Au were endothermic and raising the temperature had a positive effect on leaching efficiency. The kinetics of Pd leaching was quite fast and after 30 min complete leaching of Pd was achieved, while the leaching of Au need a longer contact time. The best conditions for leaching of Pd and Au in NaClO-HCl-H₂O₂ leaching system were determined to be 5 M HCl, 1 V% H₂O₂, 10 V% NaClO at 336 K for 3 h with a solid/liquid ratio of 1/10. 100% of Pd and Au of what was in the chloride leachate were precipitated by using 2 g/L SBH. Finally, a process flow sheet for the recovery of Cu, Ag, Au and Pd from PCB was proposed.     

Cobalt separation from waste mobile phone batteries using selective precipitation and chelating resin

Separation of cobalt from mixed-waste mobile phone batteries containing LiCoO₂ cathodic active material was investigated using selective precipitation and chelating resin. Cobalt was recovered from the active powder materials containing 47 % Co oxide together with Mn, Cu, Li, Al, Fe, and Ni oxides. The metal ions were leached sufficiently using 4 M HCl. The metal ions detected spectrophotometrically were removed from the leaching solution by selective precipitation at pH 5.5, with cobalt loss of 27.5 %. Conditions for achieving a recovery of Co in the filtrate by chelating resin were determined experimentally by varying the pH and time of the reaction, as well as the initial resin-to-liquid ratio. The cobalt was efficiently determined by absorption spectrometry at λ _max 510 nm. Chelating polyamidoxime resin was synthesized by polymerization of acrylonitrile and followed by amidoximation reaction. Physically cross-linked gel of polyacrylonitrile was made by a cooling technique.     

Hydrometallurgical processing of Nd–Fe–B magnets for Nd purification

An evaluation of various metal purification processes subsequent to the leaching processing of the neodymium (Nd) product from neodymium–iron–boron (Nd–Fe–B) magnets has been conducted. These post-leaching purification processes included precipitation; replacement and electrolysis were studied in order to check the purity of the recovered neodymium. A hydrometallurgical investigation was adopted to digest the metal content of the scrap Nd–Fe–B magnets for the recovery of valuable Nd metal and other metals such as Fe, B, Co and Ni. The effect of leaching conditions such as solid-to-liquid ratio and temperature were optimized and 100 % Nd, 100 % Fe, 100 % B and 85.87 % Co leaching efficiencies were achieved under these conditions. The coating material of the magnet, Ni, achieved 50 % impregnation after increasing the reaction temperature to 70 °C. The metals present in the optimal leaching solution were recovered 99 % by pH adjustment. However, the replacement had the highest separation efficiency for the recovery of Nd metal. Further, the optimal leaching Nd–Fe–B solution was subjected to the electrolysis processes in order to verify the recovery efficiency for all metals.     

Simultaneous recovery of vanadium and nickel from power plant fly-ash: Optimization of parameters using response surface methodology

Simultaneous recovery of vanadium (V) and nickel (Ni), which are classified as two of the most hazardous metal species from power plant heavy fuel fly-ash, was studied using a hydrometallurgical process consisting of acid leaching using sulfuric acid. Leaching parameters were investigated and optimized in order to maximize the recovery of both vanadium and nickel. The independent leaching parameters investigated were liquid to solid ratio (S/L) (5–12.5 wt.%), temperature (45–80 °C), sulfuric acid concentration (5–25 v/v%) and leaching time (1–5 h). Response surface methodology (RSM) was used to optimize the process parameters. The most effective parameter on the recovery of both elements was found to be temperature and the least effective was time for V and acid concentration for Ni. Based on the results, optimum condition for metals recovery (actual recovery of ca.94% for V and 81% for Ni) was determined to be solid to liquid ratio of 9.15 wt.%, temperature of 80 °C, sulfuric acid concentration of 19.47 v/v% and leaching time of 2 h. The maximum V and Ni predicted recovery of 91.34% and 80.26% was achieved.     

Treatment of exhaust fluorescent lamps to recover yttrium: experimental and process analyses

The paper deals with recovery of yttrium from fluorescent powder coming from dismantling of spent fluorescent tubes. Metals are leached by using different acids (nitric, hydrochloric and sulphuric) and ammonia in different leaching tests. These tests show that ammonia is not suitable to recover yttrium, whereas HNO(3) produces toxic vapours. A full factorial design is carried out with HCl and H(2)SO(4) to evaluate the influence of operating factors. HCl and H(2)SO(4) leaching systems give similar results in terms of yttrium extraction yield, but the last one allows to reduce calcium extraction with subsequent advantage during recovery of yttrium compounds in the downstream. The greatest extraction of yttrium is obtained by 20% w/v S/L ratio, 4N H(2)SO(4) concentration and 90°C. Yttrium and calcium yields are nearly 85% and 5%, respectively. The analysis of variance shows that acid concentration alone and interaction between acid and pulp density have a significant positive effect on yttrium solubilization for both HCl and H(2)SO(4) medium. Two models are empirically developed to estimate yttrium and calcium concentration during leaching. Precipitation tests demonstrate that at least the stoichiometric amount of oxalic acid is necessary to recover yttrium efficiently and a pure yttrium oxalate n-hydrate can be produced (99% grade). The process is economically feasible if other components of the fluorescent lamps (glass, ferrous and non-ferrous scraps) are recovered after the equipment dismantling and valorized, besides the cost that is usually paid to recycling companies for collection, treatment or final disposal of such fluorescent powders.     

酸浸—萃取—沉淀法回收废锂离子电池中的钴

采用酸浸—萃取—沉淀法回收废锂离子电池中的钴。实验结果表明:废锂离子电池在600℃下煅烧5 h可将正极材料上的有机黏结剂与正极活性物质分离;正极活性物质在Na OH溶液浓度为2.0 mol/L、n(Na OH)∶n(铝)=2.5、碱浸温度为20℃的条件下碱浸反应1 h后,铝浸出率达99.7%;已除铝的正极活性物质在硫酸浓度为2.5 mol/L、H_2O_2质量浓度为7.25 g/L、液固比为10、酸浸温度为85℃的条件下酸浸反应120 min,钴浸出率高达98.0%;酸浸液在p H为3.5、萃取剂P507与Cyanex272体积比为1∶1的条件下,经2级萃取,钴萃取率为95.5%;采用H_2SO_4溶液反萃后在硫化钠质量浓度为8 g/L、反萃液p H为4的条件下沉淀反应10 min,钴沉淀率达99.9%。     

锰矿浸渣中水溶性锰的回收

采用水洗—生石灰沉淀法回收锰矿浸渣中的水溶性锰，考察了液固比、水洗时间对锰矿浸渣中水溶性锰洗出率（洗出液与锰矿浸渣中水溶性锰的质量比）的影响，并探讨了n（生石灰）∶n（水溶性锰）、聚丙烯酰胺絮凝剂加入量、反应温度和反应时间对洗出液中水溶性锰回收率（沉淀与洗出液中锰的质量比）的影响。实验结果表明：水洗工段，在去离子水体积（mL）与锰矿浸渣质量（g）比为5∶1、水洗时间5 min的条件下，水溶性锰洗出率达到92%;生石灰沉淀工段，n（生石灰）∶n（水溶性锰）对水溶性锰回收率的影响最大，其次为反应温度、絮凝剂加入量和反应时间，在n（生石灰）∶n（水溶性锰）=1.8、絮凝剂加入量0.2 mg/L、常温、反应时间10 min的条件下，水溶性锰回收率达90%以上;锰矿浸渣中水溶性锰的总回收率达83%以上。     

Ammoniacal leaching and recovery of copper from alloyed low-grade e-waste

The paper concerns a hydrometallurgical method for selective recovery of copper from low-grade electric and electronic wastes. The following consecutive stages were proposed: smelting of the scraps to produce Cu–Zn–Ag alloy, leaching of the alloy in ammoniacal solution, and selective copper electrowinning. Cu–Zn–Ag alloy was a polymetallic and five-phase system. It was leached in chloride, carbonate, sulfate and thiosulfate solutions. This resulted in the separation of the metals, wherein metallic tin and silver as well as lead salts remained in the slimes, while copper and zinc were transferred to the electrolyte. Copper was selectively recovered from the ammoniacal solutions by the electrolysis, leaving zinc ions in the electrolyte. The best conditions of the alloy treatment were obtained in the ammonia–carbonate system, where the final product was copper of high purity (99.9 %) at the current efficiency of 60 %. Thiosulfate solution was not applicable for the leaching of the copper alloy due to secondary reactions of the formation of copper(I) thiosulfate complexes and precipitation of copper(I) sulfide, both inhibiting dissolution of the metallic material.     

Designing a purification process for chromium-, copper- and arsenic-contaminated wood

The disposal of chromated copper arsenate (CCA)-treated wood is becoming a serious problem in many countries due to increasing levels of contamination by the hazardous elements, chromium, copper and arsenic. The present experiment was conducted as a preliminary step toward one-step solvent extraction of CCA-treated wood. Because chromium, copper and arsenic have different chemical characteristics, it is best to consider them separately prior to designing a one-step extraction process. As a basis, various two-step extraction processes were first designed and tested experimentally to determine feasibility. Among these combinations, the treatment combining oxalic acid as the 1st step and a sodium oxalate solution under acidic conditions (pH 3.2) as the 2nd step was found to be an effective way of extracting CCA elements from treated wood. Extraction efficiency reached 100% for arsenic and chromium and 95.8% for copper after a 3-h sodium oxalate treatment, following a 1-h pre-extraction process with oxalic acid. On the other hand, the same combination under alkaline conditions (pH 11.2) during the 2nd step was ineffective for copper removal, indicating that pH plays an important role in complexation with sodium oxalate solution. The present results suggest that the extraction of CCA elements using a combination of oxalic acid and acidic sodium oxalate solution is a promising basis for application to a one-step extraction method.     

Leaching assessment of road materials containing primary lead and zinc slags

Characterisation of the leaching behaviour of waste-containing materials is a crucial step in the environmental assessment for reuse scenarios. In our research we applied the multi-step European methodology ENV 12-920 to the leaching assessment of road materials containing metallurgical slag. A Zn slag from an imperial smelting furnace (ISF) and a Pb slag from a lead blast furnace (LBF) are investigated. The two slags contain up to 11.2 wt% of lead and 3.5 wt% of zinc and were introduced as a partial substitute for sand in two road materials, namely sand-cement and sand-bitumen. At the laboratory scale, a leaching assessment was performed first through batch equilibrium leaching tests. Second, the release rate of the contaminants was evaluated using saturated leaching tests on monolithic material. Third, laboratory tests were conducted on monolithic samples under intermittent wetting conditions. Pilot-scale tests were conducted for field testing of intermittent wetting conditions. The results show that the release of Pb and Zn from the materials in a saturated scenario was controlled by the pH of the leachates. For the intermittent wetting conditions, an additional factor, blocking of the pores by precipitation during the drying phase is proposed. Pilot-scale leaching behaviour only partially matched with the laboratory-scale test results: new mass transfer mechanisms and adapted laboratory leaching tests are discussed.