Arsenic stabilization of calcium arsenate waste by hydrothermal precipitation of arsenical natroalunite

The stabilisation of calcium arsenate waste from a copper smelter by precipitation of arsenical natroalunite has been investigated. This procedure could solve the problem of arsenical gypsum production because it is transformed into arsenic-free anhydrite. Natroalunite precipitation was studied at 180-200 °C from the slurry obtained after H₂SO₄-leaching and ozonation of the original waste - using sodium and aluminum sulfates as reagents. Calcium arsenate waste and final precipitates were characterized by chemical analysis (ICP), SEM-EDS and XRD. Solubility tests were also performed on original waste and selected precipitates.The effects studied in the hydrothermal treatment were: initial pH, Al/As molar ratio, As concentration, reaction time and prior gypsum removal. For (Al/As)_aq = 4.5, a complex natroalunite (~(Na,Ca)(Al,Fe)₃((S,As,P)O₄)₂(OH)₆)) was extensively formed. Under these conditions, As-substitution in TO₄ was 7-8% molar. Decreasing (Al/As)_aq increased As-substitution in natroalunite (up to ~ 14% molar) but mansfieldite ((Al,Fe)(As,P)O₄.2H₂O) co-precipitated. Other effects such as the arsenic concentration in the range 3.5-7.0 g/L and prior gypsum removal did not significantly alter the arsenic phase ratio and the composition of the arsenic phases. However, treatment at 180 °C increased mansfieldite precipitation. Calcium incorporation in natroalunite was small (~ 0.04 in formula) and for initial pH = 1, precipitation of Cu, Ni and Zn was insignificant.Arsenical natroalunite can be effective for long-term storage. At its natural pH (4-5), arsenic solubility remained stabilized at ~ 0.1 mg/L in 6 months. Mansfieldite was not stable under the same conditions with arsenic solubility between 5-10 mg/L. The major disadvantage of natroalunite is the limited arsenic partition between solid and aqueous phases (~ 0.45 molar). However, as the precipitation rate is fast (< 15 min), it seems technically feasible to improve the arsenic precipitation yield by semi-continuous addition of reagents in a single hydrothermal operation. Cadmium incorporated into the anhydrite structure causes cadmium release in leachates. It seems that this problem can only be overcome through prior removal of cadmium from the plant effluent.     

Differential precipitation of copper and nickel from acidic polymetallic aqueous solutions

During the pyrometallurgical treatment of copper concentrates several wastes (gaseous, solid and liquid) are generated. Wastewaters produced in a primary copper smelter are acidic polymetallic solutions characterized by intermediate to high sulphuric acid concentration. Additionally, they usually contain large quantities of valuable metals, such as Cu and Ni, as well as impurities, such as Pb, Zn, Fe, As, Sb, Bi, etc. Therefore, recovery of valuable metals from acidic polymetallic aqueous solutions is of great importance for every plant. This paper is dealing with the recovery of copper and nickel from the acidic polymetallic solutions generated in the copper smelter at Bor, Serbia. The process of differential precipitation of metals through neutralization with caustic soda was selected as the treatment alternative that combines simplicity, efficiency and reliability with low capital and maintenance costs. The concept was based on the development of a simple and efficient process that could be more attractive for industries located in countries suffering from economic recession. Theoretical analysis and experimental results showed that copper and nickel can be differentially removed from the acidic polymetallic solutions at pH = 7 and pH = 10, respectively, mainly contaminated by arsenic, which occurs at high concentrations in acidic polymetallic solutions. Both metals were precipitated as hydroxides, although a small portion of copper was removed also as copper sulfate salt. The resulted precipitates are rich in copper and nickel and can be recycled in copper, as well as in nickel smelters.     

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.     

Comparison of Metals Leaching from CCA- and ACQ-Treated Wood in Simulated Construction and Demolition Debris Landfills

Pressure-treated wood is often disposed of in landfills in the United States, very frequently in construction and demolition (C&D) debris landfills. C&D debris landfills in many states are not equipped with liner systems to protect groundwater. With the voluntary withdrawal of chromated copper arsenate (CCA) treated wood for most residential applications in January 2004, copper-based wood preservatives, including alkaline copper quaternary (ACQ), are more widely used. To evaluate the impact of metal losses from ACQ-treated wood disposed in C&D debris landfills and compare to those of CCA-treated wood under similar conditions, leachates from three simulated C&D debris landfills (lysimeters) were collected and analyzed for over a period of one year. The wood component in one lysimeter (the control lysimeter) contained pallet wood; the second lysimeter contained CCA-treated wood, and the third contained ACQ-treated wood. Each lysimeter was buried in an active landfill for temperature control. Several batch leaching tests [including the standardized toxicity characteristic leaching procedure (TCLP) and the synthetic precipitation leaching procedure (SPLP)] were also conducted for comparison purposes. Although the two lysimeters containing treated wood had elevated copper concentrations within the waste matrix, the concentration in the leachate samples from these lysimeters was below detection for Cu (<4?μg/L) throughout the duration of the experiment, likely a result of precipitation as copper sulfide mineral in the reducing conditions of the simulated C&D landfills. As expected, the lysimeter containing CCA-treated wood showed elevated concentrations of arsenic and chromium, with maximum concentrations of 1.16 mg/L and 0.2 mg/L respectively. Greater amounts of boron (B) leached from ACQ-treated wood than CCA-treated wood or pallet wood debris. The results suggest that copper leaching will not be a major concern upon the disposal of ACQ-treated wood in C&D debris landfills. Arsenic leaching from CCA-treated wood remains a concern for unlined C&D debris landfills.     

The solvent extraction separation of bismuth and molybdenum from a low grade bismuth glance flotation concentrate

The main purpose of this study was to characterize and to extract bismuth and molybdenum from a low grade bismuth glance concentrate. Selective leaching of bismuth could be obtained at a temperature range 60 to 85 °C for a leaching duration of 2 h with hydrochloric acid concentration of 150 gpl, lignin calcium concentration of 0.02 gpl and using a solid–liquid ratio 1/4 g/cc. Treatment of leach liquor for the solvent extraction of bismuth with N235 showed that 8.0 × 10^− 2 M N235 in kerosene, a 3 min period of equilibration and a pH 0.2 were sufficient for the extraction of Bi(III). This bismuth-loaded organic phase was almost completely stripped using 0.5 M EDTA solution. Treatment of leached residue was dealt with by roasting in the presence of slaked lime, and followed by hydrometallurgical treatment of the roasted products. In the lime roasting process, molybdenum recoveries of around 99% were achieved when an excess of 50% lime over stoichiometric requirement was roasted at 700 °C for 2 h and the calcine was leached with 4 M HCl, at 70–80 °C for 2 h. Molybdenum then was effectively extracted from the leached residual solution with N235. An optimum pH of 0.5 was determined for molybdenum extraction. From loaded solvent, this metal was easily stripped with ammonia solutions to give a pregnant solution suitable for final recovery of metal by salt precipitation. Under the optimized conditions, the ultimate recovery rate of bismuth and molybdenum was more than 99% and 98% respectively.     

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.     

Recovery of germanium and other valuable metals from zinc plant residues

The main purpose of this study was to characterize and to extract germanium from the copper cake of Çinkur Zinc Plant. The physical, chemical and mineralogical characterization of the ground copper cake sample obtained from Çinkur showed that it was 84% below 147 μm containing 700 ppm germanium. The copper cake also contained 15.33% Cu, 15.63% Zn, 1.66% Cd, 1.33% Ni, 0.64% Co, 0.35% Fe, 2.62% Pb, 12.6% As, 0.18% Sb and 3.42% SiO₂. The mineralogical analysis indicated the complex nature of the copper cake which was mainly composed of metallic and oxidized phases containing copper, arsenic, zinc, cadmium, etc. The sulfuric acid leaching experiments were performed under the laboratory conditions. The optimum collective extraction of germanium and other valuable metals was obtained at a temperature range 60 to 85 °C for a leaching duration of 1 h with sulfuric acid concentration of 150 gpl and using a solid–liquid ratio 1/8 g/cc. Under these conditions, the recovery of germanium was 92.7% while the other metals were leached almost completely. The optimum selective leaching conditions of germanium was determined as half an hour leach duration, 1/8 g/cc solid–liquid ratio, 100 gpl sulfuric acid concentration and a temperature range 40 to 60 °C. Under these conditions the leach recovery of germanium was 78%. The dissolution's of other metals like cobalt, nickel, iron, copper, cadmium and arsenic were almost low. So, germanium would be separated more selectively at the following precipitation by tannin stage.     

Chloride leaching of chalcopyrite

Two new process flowsheets have been developed which combine chloride leaching of copper from chalcopyrite with solvent extraction, to selectively transfer copper to a conventional sulfate electrowinning circuit. Chloride leaching with copper(II) as oxidant offers significant advantages for copper including increased solubility and increased rates of leaching. Both process flowsheets were similarly designed with a two stage counter-current leach but differ with respect to iron deportment. The goethite model flowsheet includes sparging of air or oxygen to the second leach stage to aid precipitation of iron as goethite (FeOOH). The hematite model flowsheet precipitates iron as hematite (Fe₂O₃) downstream from the leach in a dedicated autoclave. A mass balance has been completed for both process flowsheets and this determined the concentrations of copper and iron species in feed liquor returning to the leach following copper solvent extraction.The optimum leach extraction conditions were determined by varying grind size, temperature and residence time for both leach model scenarios. Leach tests were conducted using a chalcopyrite concentrate from Antamina in northern Peru, which contains a low to moderate amount of gangue material. The hematite model was also examined using a Rosario concentrate from Chile which contained chalcocite in addition to chalcopyrite and significant pyrite. Leach tests based on the hematite model were successful in achieving copper extractions > 95% in 4–6 h at 95 °C after fine grinding the concentrate (P₉₀ = 41 μm). However, copper extraction exceeded 99% from the finely ground Rosario concentrate (P₉₀ = 37 μm). In the goethite model leach tests, 89% copper extraction was achieved under optimum conditions in the atmospheric conditions tested.     

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.     

Electrochemical modeling and study of copper deposition from concentrated ammoniacal sulfate solutions

A fundamental investigation of the electrolytic deposition of copper from concentrated aqueous ammoniacal solutions has been carried out based on the thermodynamic analysis of the system Cu–NH₃–H₂O. The speciation of copper vs. pH and redox potential was modeled in high ionic strength solutions, in which the activity coefficients of the system species were estimated according to the Modified Bromley's Methodology. The electrochemical behavior of the redox system Cu(0)/Cu(I)/Cu(II) in concentrated aqueous ammoniacal solutions was studied at pH = 9.5 and the cathodic reactions in these solutions were determined. It was found that metallic copper was formed under strongly reductive redox conditions, while under mildly reductive to mildly oxidative conditions the cuprous di-ammine complex species dominate. Under highly oxidative conditions the cupric tetra-ammine complex species predominated. According to the theory and results, the cathodic deposition of copper from concentrated aqueous ammoniacal solutions proceeds in a two-step reduction mechanism. The cupric ammine species are first reduced to cuprous di-ammine, which in turn is reduced to metallic copper. The electrochemical experiments revealed that copper deposition over time follows a sigmoid-type curve, verifying the two-step mechanism. The main feature of these sigmoid curves was the presence of an induction period with negligible copper deposition, followed by an acceleration period where the copper deposition rate gradually increased. By increasing the applied cell voltage, the induction period was significantly reduced or disappeared.     

含砷铜物料中砷、铜分离试验研究

采用焙烧-水浸联合工艺路线处理复杂含砷铜物料,试验结果表明,在加碱焙烧温度550℃,碱系数1.5倍,时间1.0h;焙烧渣水浸液固比5∶1,温度60~65℃,时间1.0h的条件下,99%砷可与铜分离,从而富集得到含铜76%的高含铜原料,可直接并入铜冶炼系统。     

Effect of phosphate on lead removal during a copper recycling process from wastes using ammoniacal chloride solution

Techniques for the removal of lead have been studied in order to develop a hydrometallurgical copper recycling process consisting of copper leaching from wastes using an ammoniacal chloride solution and subsequent copper electrowinning. The solubility of Pb(II) in the ammoniacal chloride solution increased with ammonia concentration; this was attributable to the formation of a lead ammine complex. The lead dissolution was depressed from the order of 10^− 3 M to the order or 10^− 5 M by the addition of phosphate into the leaching solution because of the precipitation of chloropyromorphite (Pb₅(PO₄)₃Cl), while no significant effect was observed by the addition of carbonate. Linear sweep voltammetry and potentiostatic electrolysis in the solution containing Pb(II) revealed that lead was deposited during the copper electrowinning, even in the potential region more positive than the equilibrium redox potential for the Pb/Pb(II) couple on the lead electrode, because of the alloy formation with copper. In a galvanostatic electrolysis, however, the lead content at the electrodeposited copper cathode was found to be lower than 5 ppm at the current density range of 125–400 A/m², when the Pb(II) concentration in the electrolyte was 5 × 10^− 5 M. Since this Pb(II) concentration was achieved by the phosphate addition, these results indicated the effectiveness of phosphate for lead removal in the copper recycling process using the ammoniacal chloride solution.     

Processing of Cobalt-Copper-Arsenic Complex Ore Concentrate

Extraction of cobalt from complex ore flotation concentrates obtained from the Blackbird Mine. Idaho. USA is reviewed. After flotation of a primary copper concentrate, a bulk concentrate is recovered containing major amounts of cobalt, arsenic, and iron, with minor amounts of copper and nickel. This concentrate can be upgraded during flotation by removing iron, but with considerable cobalt loss. Chemical extraction difficulties are caused by the high arsenic and iron content of the concentrate. The historical approach, including a short lived plant, has been pressure oxidative leaching followed by ferric arsenate precipitation, solution purification, and cobalt electrowinning. Smelting and sulfation roasting followed by leaching are unsatisfactory but also discussed. The most recent studies, showing some promise, have been on oxygen-calcium chloride leaching and on biooxidative leaching at moderate pH with simultaneous iron hydrolysis and ferric arsenate precipitation.     

The effect of aromatic and aliphatic amines on copper electrowinning from acidic sulphate electrolyte

Copper electrodeposition in the presence of various types of aromatic and aliphatic amines was studied. The effects of operating variables including organic additives concentrations and temperature on the limiting current were investigated by the potentio-dynamic polarization technique. The effects of amines on the surface tensions of the solutions were measured; the results showed that amines reduce solution surface tension. The adsorption of all inhibitors on copper cathode was found to obey Temkin, Flory-Huggin and kinetic adsorption isotherm. The calculated free energy of adsorption (ΔG_ads.) of inhibitor on copper surface indicated that the adsorption reactions were spontaneous (ΔG_ads. < 0). The thermodynamic activation parameters (E_a, ΔH*, ΔS* and ΔG*) were estimated and discussed. It was found that activation energy values for copper electrodeposition in the inhibited solutions were higher than that for the uninhibited solution. The high inhibitor efficiency was discussed in terms of the strong adsorption of inhibitor molecules on the copper surface.     

某高砷高铜金精矿综合回收金银铜试验研究

某高砷高铜金精矿含砷高达9.42%,采用加压氧化—氰化工艺处理,铜、金、银浸出率分别为96%～97%、99%、78%,加压氧化过程80%以上的砷固化在氧化渣中。同时开展了铜萃取、萃余液处理、毒性浸出等工艺单元试验,打通整体流程。毒性浸出试验表明,氰化渣、中和渣毒性浸出液中的重金属、砷浓度达标。采用加压氧化工艺处理高砷高铜金精矿是可行的。     

从电镀污泥中回收镍、铜和铬的工艺研究

采用硫酸浸出-硫化沉铜-两段中和除铬-碳酸镍富集工艺,从电镀污泥中综合回收铜、铬和镍.考察了各工序过程中的影响因素,获得了最佳工艺条件:酸浸过程中反应时间为0.5 h,反应温度为50℃,硫酸加入量为理论量的0.8倍;沉铜过程中,沉铜剂加入量为理论量的1.2倍,反应时间和反应温度分别为1 h和85℃;采用两段除铬工序有效降低了沉淀过程中的镍损失.整个工艺中,铜、铬和镍的回收率分别达到98%、99%和94%以上.     

高砷氯氧锑碱浸脱砷试验研究

采用常温碱浸脱砷方法处理高砷氯氧锑,就水洗、液固比、终点pH值等因素对脱砷效果的影响进行了研究。结果表明:浸出前调浆水洗可降酸除铜,浸出时液固比选择12∶1、终点pH值控制在10.0左右可脱除氯氧锑渣中90%以上的砷。含砷浸出液经石灰乳沉砷后可返回浸出工序重复使用。     

Studies on the leaching of an arsenic–uranium ore

Preliminary leaching studies were carried out to develop a suitable method for the recovery of uranium and the elimination of arsenic from a low grade carbonate/silicate ore containing 64 ppm U and 2446 ppm As, as well as some Cu, Pb, Ni and Zn. An examination of the mineralogy found mostly uranium(VI) minerals, such as uraninite, and various base metal sulfides and arsenates in veins and fissures. Roasting the ore at 500–800 °C to volatilize arsenic proved to be unsuitable. Therefore, the ground ore was subjected to direct leaching with sulfuric acid, sodium sulfide and ferric chloride at 80–90 °C with a liquid to solid ratio of 1:1. With sulfuric acid at a concentration of 180 kg/t ore, complete recovery of both uranium and arsenic was achieved giving undesirable arsenic in the leach liquor. The maximum recovery of uranium and arsenic by leaching with sodium sulfide was only 20% and 18%, respectively. However, 3 M ferric chloride leached approximately 92% U(VI) and precipitated arsenic as ferric arsenate. Therefore, maximum uranium can be extracted and arsenic eliminated as impurity by selective leaching with ferric chloride.     

Separation of zinc from spent zinc-carbon batteries by selective leaching with sodium hydroxide

Physical methods such as crushing and sieving followed by magnetic separation steps were applied to separate non-magnetic material (88.4 wt.%) from zinc-carbon spent batteries. The oversize material was processed by eddy current separation to recover zinc sheet, carbon rods, and plastics. The undersize fraction (− 2.36 mm) with a metal composition of 15.5% Zn, 17.5% Mn, and 1.4% Fe was used for the leaching experiments under different conditions such as concentration of sodium hydroxide, temperature, agitation speed, and pulp density. Selective leaching using 4 mol dm^− 3 NaOH at 100 g dm^− 3 solid/liquid ratio, 80 °C and 200 rpm gave a zinc extraction of 82% and a manganese extraction of less than 0.1%. An overall zinc recovery was about 88.5% by the processes described in this study.