Hydrogen reduction of copper bleed solution from an Indian copper smelter for producing high purity copper powders

A bench scale investigation on the hydrogen reduction of a highly acidic copper bleed solution to produce high purity copper powder is discussed in this paper. A titanium lined autoclave of 1 L capacity was used for this study. The parameters optimized on the bench scale were validated by performing hydrogen reduction of copper in a larger autoclave. Effect of various parameters viz., time of reduction, temperature of reduction, pressure variation, iron dose, volume of copper solution etc. were studied. Experiments were performed with synthetic and actual solution obtained from a copper plant. A 99% copper powder recovery is achieved by hydrogen reduction at a pressure of about 2400 kPa, reaction temperature of 453 K, stirring speed of 400 rpm for a reaction time of 2 h. The fine copper powder thus obtained had a good metallic lustre. Kinetics of reduction of copper was examined by drawing samples at different times and analyzing the percent copper reduction. The copper depleted solution was further purified with respect to the residual copper and can be processed for the recovery of nickel powder by hydrogen reduction.Properties of the copper powder obtained from the large-scale experiments from actual plant and synthetic solutions have been evaluated for powder metallurgical applications. The raw and annealed copper powders obtained from the synthetic copper solution were found to have an apparent density of 3.50 g/cm³, flow rate 35.6 g/min, hydrogen loss 0.2%, purity 99.8% and green density of 8.6 g/ cm³ while the powder from actual plant solution was found to have an apparent density of 3.59 g/ cm³, flow rate 46.0 g/min, hydrogen loss 0.6%, purity 99.4% and green density 8.6 g/ cm³. Thus, the copper powder produced by hydrogen reduction was found suitable for the application.     

Silver-catalyzed bioleaching of low-grade copper ores.: Part I: Shake flasks tests

A study of the effect of different variables (inoculation, pulp density, [Ag], nutrient medium, pH and [Fe³⁺]) on the silver-catalyzed bioleaching of a low-grade copper sulfide ore has been carried out in shake flasks. Chalcopyrite was the dominant copper mineral in the ore. Preliminary tests showed that addition of other ions (Sb, Bi, Co, Mn, Ni and Sn) did not enhance the copper dissolution rate. Conversely, an inoculation with mesophilic microorganisms and the addition of silver had a markedly catalytic effect on the extraction of copper. The kinetics of the silver-catalyzed chalcopyritic ore bioleaching was greatly affected by pulp density and silver concentration. Small amounts of silver (14.7 g Ag/kg Cu) dramatically accelerated the copper dissolution process while large amounts (294.12 g Ag/kg Cu) had an inhibitory effect. The copper dissolution rate was slightly affected in the range of pH between 1.2 and 2.5 but was significantly slower at pH 3.0. The effect of [Fe³⁺] in the presence of silver was studied both in abiotic and biotic conditions. High ferric iron concentrations in abiotic tests recovered similar copper amounts (∼ 95%) to those obtained without or with low [Fe³⁺] in the presence of bacteria. The leaching of copper from the low-grade copper ore can be very effectively enhanced with silver and mesophilic microorganisms. For that system, the onset of oxidizing conditions starts at an Eh value slightly higher than 650 mV. Above that critical value of potential the copper dissolution rate slows down. This also corresponds with the completion of the leaching process. As the potential rises past 650 mV, the copper extraction reaches a plateau.     

Organic acid leaching of base metals from copper granulated slag and evaluation of mechanism

A hydrometallurgical method is discussed to selectively extract base metals such as copper, cobalt, nickel and iron from the copper granulated slag (0.53% Cu) at atmospheric pressure. It involves first-stage leaching of slag with organic (citric acid) to selectively recover cobalt, nickel and iron. The residue containing high copper was subjected to second-stage leaching with inorganic (sulphuric) acid. Leaching parameters such as acid concentration, pulp density, temperature and time were optimised to extract metals from the granulated slag. A maximum recovery of 4.47% Cu, 88.3% Co, 95% Ni and 93.8% Fe were obtained in first-stage leaching with 2?N citric acid at room temperature using 10% pulp density (w/v) in 8–9?h. On subjecting the leach residue to the second-stage leaching with 2?M sulphuric acid, 66–72% Cu was recovered in 4?h. The kinetics of the metal leaching from the slag was established by the XRD and SEM–EDAX studies of the residues.     

Pressure leaching of copper converter slag using dilute sulphuric acid for the extraction of cobalt, nickel and copper values

In our earlier studies [1–4], conditions were optimized for leaching converter slag with ferric chloride/dilute sulphuric acid for the recovery of cobalt, nickel and copper. By using both leachants most of the copper, nickel and cobalt values could be solubilized. Subsequent treatment of the leach liquors for separation and recovery of metals was difficult due to the presence of large quantities of iron in relation to other metal concentrations. In the present work, an attempt has been made to develop a process based on pressure leaching of the slag with dilute sulphuric acid in which iron contamination could be minimized by oxidation and hydrolysis. Various parameters including leaching time, pulp density, particle size, concentration of acid and oxygen partial pressure were studied to optimize the solubility of metal values. Under optimum conditions about 90% copper and more than 95% each of nickel and cobalt could be extracted with only 0.8% extraction of iron.     

Selective recovery of metals in leachate from chromated copper arsenate treated wastes using electrochemical technology and chemical precipitation

This research explores different ways of recovering metals (As, Cr and Cu) and treating leachates from an acid leaching process used for the decontamination of chromated copper arsenate (CCA)-treated wood. Total precipitation at pH = 7 with ferric chloride and an anionic polymer (Magnafloc 10), leads to 99% removal of the three metals from the leachate. Copper electrodeposition is very efficient with more than 99% of copper recovery, while arsenic was observed as a disruption in the process of copper deposition. Arsenic and copper are able to deposit as Cu₃As while applying an electrical current. In order to remove arsenic in the leachate prior to electrochemical treatment, coagulation at pH = 4 was tested, followed by pH readjustment and electrodeposition. This process produces a good quality of elemental copper by electrodeposition (+ 99% pure copper), whereas 99% of arsenic and 88% of chromium are concentrated in the metallic sludge resulting from the coagulation–precipitation treatment. These results show that these two techniques are promising for the management of CCA leachate, one being very simple as it consists of only one precipitation step, the other one allowing recycling of valuable copper via electrochemical treatment. Operation costs have been estimated at 92 $/ton of treated wood (t.t.w.) undergoing leaching treatment followed by precipitation with sodium hydroxide at pH = 7. Acid leaching followed by precipitation at pH = 4 and electrolytic deposition costs 110 $/t.t.w. with copper sale revenue of 15 $/t.t.w.     

Copper recovery from chalcopyrite concentrates by the BRISA process

The technical viability of the BRISA process (Biolixiviación Rápida Indirecta con Separación de Acciones: Fast Indirect Bioleaching with Actions Separation) for the copper recovery from chalcopyrite concentrates has been proved. Two copper concentrates (with a copper content of 8.9 and 9.9 wt.%) with chalcopyrite as the dominant copper mineral have been leached with ferric sulphate at 12 g/L of ferric iron and pH 1.25 in agitated reactors using silver as a catalyst. Effects of temperature, amount of catalyst and catalyst addition time have been investigated. Small amounts of catalyst (from 0.5 to 2 mg Ag/g concentrate) were required to achieve high copper extractions (>95%) from concentrates at 70 °C and 8–10 h leaching. Liquors generated in the chemical leaching were biooxidized for ferrous iron oxidation and ferric regeneration with a mixed culture of ferrooxidant bacteria. No inhibition effect inherent in the liquor composition was detected. The silver added as a catalyst remained in the solid residue, and it was never detected in solution. The recovery of silver may be achieved by leaching the leach residue in an acid-brine medium with 200 g/L of NaCl and either hydrochloric or sulphuric acid, provided that elemental sulphur has been previously removed by steam hot filtration. The effect of variables such as temperature, NaCl concentration, type of acid and acidity–pulp density relationship on the silver extraction from an elemental sulphur-free residue has been examined. It is possible to obtain total recovery of the silver added as a catalyst plus 75% of the silver originally present in concentrate B (44 mg/kg) by leaching a leach residue with a 200 g/L NaCl–0.5 M H₂SO₄ medium at 90 °C and 10 wt.% of pulp density in two stages of 2 h each. The incorporation of silver catalysis to the BRISA process allows a technology based on bioleaching capable of processing chalcopyrite concentrates with rapid kinetics.     

Processing copper–vanadium precipitate formed from crude TiCl4 in titania and titanium sponge production

Copper wire is used to remove vanadium from crude TiCl₄ in titania and titanium sponge production which produces a copper–vanadium precipitate. The recovery of copper and vanadium from this precipitate was studied. Experiments found that the precipitate can be naturally oxidized by stacking for one month in air, converting > 90% metallic copper contained in the original precipitate into CuCl₂·2H₂O, Cu₂Cl(OH)₃ and Cu₂(OH)₃Cl. The copper oxy-chlorides were easily converted to Cu(OH)₂ by stirring in dilute NaOH at pH 11 and 80 °C under a liquid-to-solid ratio of 4:1. When the pH was lowered to about pH 2.5 by sulfuric acid, iron, titanium and vanadium oxides remained in the first acid leach residue and copper was selectively leached into solution. By evaporating and cooling the leach solution, a product of CuSO₄·5H₂O with 99.7% purity was obtained.To recover the vanadium, the filter cake was roasted with Na₂CO₃ at 700 °C for 3 h under the stoichiometric proportion of 2.5 for V. The calcine was then leached with water at 70 °C and NH₄VO₃ was precipitated by the addition of NH₄Cl. Calcination of NH₄VO₃ at 550 °C for 2 h produced V₂O₅ with a purity of 98.6%. After vanadium recovery, the residue was leached once again with sulfuric acid and the total recoveries of copper and vanadium were 98.6% and 95.7% respectively.     

Microbial extraction of nickel from Sukinda chromite overburden by Acidithiobacillus ferrooxidans and Aspergillus strains

In this study, the recovery of nickel from a low-grade chromite overburden was attempted by employing two fungal strains, Aspergillus niger and Aspergillus fumigatus, and a mixed culture of mesophilic acidophiles (predominantly Acidithiobacillus ferrooxidans). Various factors were studied for bioleaching of chromite overburden such as, temperature, pH and pulp density. It was found that the At. ferrooxidans culture solubilized nickel effectively at temperatures ranging from 30 °C to 37 °C, whereas the organism was not able to solubilize nickel at higher temperatures, such as 45 °C. The use of higher pulp density resulted in a decrease of the percent nickel recovery whereas lower pulp density resulted in higher recovery values. Besides, increased supplemental ferrous iron increased the leaching efficiency of the At. ferrooxidans culture. The maximum nickel solubilization was 40%, at 2% pulp density, and 24%, at 10% pulp density, at 30 °C after 28 days leaching at 150 rpm.In the case of fungal strains, a comparison of leach ability of chromite overburden and roasted overburden was made. The factors studied were pulp density and reaction time. The adapted fungal strain showed better leaching results as compared to the unadapted strains. The in situ nickel leaching efficiency of a laboratory stock culture of A. niger showed maximum recovery of 34% nickel with roasted chromite overburden, at 2% pulp density, while 32% nickel was solubilized by A. fumigatus, under the same conditions at 30 °C and 150 rpm, after 28 days incubation.     

Hydrothermal purification and enrichment of Chilean copper concentrates. Part 2: The behavior of the bulk concentrates

The hydrothermal treatment of Chilean Codelco-type copper concentrates with copper sulfate solutions was investigated as a mean of removal of impurities and subsequent increase of the copper assay. The behavior of the mineral phases (digenite, chalcopyrite, covellite, bornite, pyrite and sphalerite) was similar to those obtained in previous works from pure mineral samples. An almost complete transformation of bornite, chalcopirite, covellite and sphalerite into Cu_2 ? xS phases was obtained at 225 °C–240 °C. The highest degree of elimination (around 80%) of impurities was in Zn, Cd, Tl and Bi. An intermediate elimination (40–70%) was achieved for Pb and Te, with only moderate elimination (20–40%) of Mo, Hg, Sb and As. Temperature was the variable having the greatest influence on the elimination of the impurities. A concentrate containing 33% Cu, 33% S, 22% Fe and 2% Zn was converted to a highly enriched concentrate containing 70% Cu, 19% S and 3% Fe. The advantages of a concentrate of this type would include: (1) raising by more than twice the smelting capacity due to the high copper content, (2) generation of a minimum amount of slag, (3) reduction by almost 50% in sulfur emissions, (4) substantial reduction of wastes containing hazardous metals and, finally (5), retention of the option to hydrometallurgical copper recovery since the neo-formed Cu_2 ? xS phases are more reactive than chalcopyrite to the chemical or biochemical leaching.     

Dissolution of uranium from silicate-apatite ore by Acidithiobacillus ferrooxidans

Bioleaching of a low-grade Indian silicate-apatite uranium ore containing 0.024% U₃O₈ and 10.6% iron with minor amounts of base metals has been reported. The studies involved extraction of uranium using enriched culture containing Acidithiobacillus ferrooxidans (A. ferrooxidans) derived from the source mine water employing bio-chemically generated ferric ion as an oxidant. Parameters such as particle size of the ore, pulp density, and pH of lixiviant media were optimised. Maximum uranium bio-dissolution of 98% was achieved using ore of mixed particles of < 76 μm size. Uranium bio-recovery was found to be 96% at the pulp density (PD) of 10% (w/v) and 20% (w/v) with the particles of < 76 μm size in 40 days at 2.0 pH and 35 °C temperature. At 1.7 pH and 20% (w/v) PD, 98% uranium bio-recovery was achieved with a rise in redox potential from 595 mV to 715 mV in 40 days. Uranium bio-dissolution may be correlated with the generation of ferric ions through the bio-chemical action on the ore. The work illustrated the efficacy of leaching of uranium by the involvement of bacteria by indirect mechanism.     

AN OVERVIEW OF PROCESS OPTIONS AND BEHAVIORAL ASPECTS OF THE COPPER VALUES RECOVERED FROM THE COPPER BLEED STREAM OF A COPPER SMELTER DEVELOPED AT THE NATIONAL METALLURGICAL LABORATORY

In the present research, an effort has been made to prepare copper salt/powder from the copper bleed stream generated during the electrowinning of pure copper from the copper anode in a copper smelter. Various approaches have been opted for the complete recovery of copper values such as: evaporation–crystallization, electrolytic process, and direct hydrogen reduction. Physical and chemical properties of copper powder/salt produced from the large-scale experiments from actual plant and model solutions have been evaluated for P/M applications and compared with the standard properties. Thus, mixed crystal suitable for recycling back to the system as a makeup salt containing nickel in a tolerable range could be recovered by evaporation and crystallization of the bleed stream up, to 50%. Copper powder recovery by the electrolysis process at a current density of 700 A/m² was about 95%. Scanning electron microscope examination showed that the powder was dendritic in nature. On annealing, the purity of the copper powder was found to be 99.95%. The annealed powder had apparent density of 3.04 g/cc, hydrogen loss 0.72%, and acid insoluble as 0.27%. On compaction of <104-µm sized powder, the green density was found to be 8.7 g/cc. Similarly, the recovery of the copper powder obtained from the model copper solution by the hydrogen reduction process was found to be >99% and the annealed powder had an apparent density of 3.50 g/cc, flow rate 35.6 g/min, hydrogen loss 0.195%, purity 99.8%, and green density of 8.57 g/cc while the powder from the actual plant solution was found to have an apparent density of 3.49 g/cc, flow rate 46.0 g/min, hydrogen loss 0.598%, purity 99.4%, and green density 8.57 g/cc for the powder < 100 µm in size. Thus, the properties of copper powder produced by hydrogen reduction and electrolytic route were compared and were found to be suitable for the P/M applications.     

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.     

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.     

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.     

重铬酸钾氧化-硫酸亚铁铵返滴定法测定铜闪速冶炼烟尘中砷

准确测定铜闪速冶炼烟尘中的砷对于炉前配料的计算和生产控制具有重要的作用。采用硝酸-氯酸钾饱和溶液、氟化铵溶液、高氯酸溶解样品,再用硫酸(1+1)驱除硝酸后,在盐酸介质中以硫酸铜为催化剂,用次亚磷酸钠把溶液中的砷离子还原为单质砷,过滤分离其他杂质。以过量的重铬酸钾标准滴定溶液溶解单质砷,以N-苯代邻氨基苯甲酸溶液为指示剂,用硫酸亚铁铵标准滴定溶液滴定过量的重铬酸钾标准滴定溶液,建立了硫酸亚铁铵返滴定法测定铜闪速冶炼烟尘中砷的方法。采用次亚磷酸钠将溶液中的砷还原为单质砷沉淀时,可能会有部分砷因未被还原而被过滤到溶液中,试验考察了滤液中残存的砷量对砷测定结果的影响。研究表明,滤液中砷的质量分数小于0.01%,相对于样品中的砷可以忽略不计。共存元素的干扰试验表明,样品中共存的铜、铅、铁、锌等元素对砷测定的影响可忽略不计。将实验方法应用于测定3个铜闪速冶炼烟尘样品中的砷,并进行加标回收试验,测定结果的相对标准偏差(RSD,n=11)在0.35%～2.6%之间,加标回收率在99%～101%之间。采用实验方法测定2个铜闪速冶炼烟尘样品中的砷,测定结果与微波消解-电感耦合等离子体原子发射光谱法(ICP-AES)相符。     

重铬酸钾氧化-硫酸亚铁铵返滴定法测定铜闪速冶炼烟尘中砷

准确测定铜闪速冶炼烟尘中的砷对于炉前配料的计算和生产控制具有重要的作用。采用硝酸-氯酸钾饱和溶液、氟化铵溶液、高氯酸溶解样品,再用硫酸(1+1)驱除硝酸后,在盐酸介质中以硫酸铜为催化剂,用次亚磷酸钠把溶液中的砷离子还原为单质砷,过滤分离其他杂质。以过量的重铬酸钾标准滴定溶液溶解单质砷,以N-苯代邻氨基苯甲酸溶液为指示剂,用硫酸亚铁铵标准滴定溶液滴定过量的重铬酸钾标准滴定溶液,建立了硫酸亚铁铵返滴定法测定铜闪速冶炼烟尘中砷的方法。采用次亚磷酸钠将溶液中的砷还原为单质砷沉淀时,可能会有部分砷因未被还原而被过滤到溶液中,试验考察了滤液中残存的砷量对砷测定结果的影响。研究表明,滤液中砷的质量分数小于0.01%,相对于样品中的砷可以忽略不计。共存元素的干扰试验表明,样品中共存的铜、铅、铁、锌等元素对砷测定的影响可忽略不计。将实验方法应用于测定3个铜闪速冶炼烟尘样品中的砷,并进行加标回收试验,测定结果的相对标准偏差(RSD,n=11)在0.35%～2.6%之间,加标回收率在99%～101%之间。采用实验方法测定2个铜闪速冶炼烟尘样品中的砷,测定结果与微波消解-电感耦合等离子体原子发射光谱法(ICP-AES)相符。     

某难处理金矿热压氧化渣预处理试验研究

高铜难处理金矿经酸性热压氧化后,铜基本被浸出进入溶液中,消除了铜对氰化过程的影响,而银在热压处理过程中易与生成的黄钾铁矾相结合,生成难处理的银铁矾[AgFe₃（SO₄）₂（OH）₆],在随后的常规氰化试验中,金回收率达99%以上,但银回收率不足10%。针对银回收率低的问题,系统考察了矿浆浓度、NaCN浓度、石灰用量、预处理温度和时间、氰化时间及炭密度等因素对金、银浸出率的影响,进而确定了最佳浸出条件。试验结果表明：在85~90 ℃、矿浆浓度为40%、石灰用量为40 kg/t的条件下,对氧化渣进行碱性预处理,随后在NaCN用量为0.10%的条件下浸出8 h,银回收率得到大幅提高（达到85%）,金浸出率也保持在99%以上。     

Cathodic reactions of Cu2+ in cupric chloride solution

In this study we demonstrate the kinetics of Cu²⁺ reduction in concentrated cupric chloride solutions. Experiments were carried out near the boiling point of the solution ([NaCl] = 280 g/l and [Cu²⁺] = 1–40 g/l) at T = 90 °C, atmospheric pressure, pH = 2. Electrochemical methods such as cathodic polarization curves and cyclic voltammetry were used to investigate the cathodic reactions of copper complexes. To identify the nature and the rate-controlling steps of the reactions, rotating disk electrode (RDE) experiments were conducted. The chemical environment studied was similar to that of the Outokumpu HydroCopper^TM process, which uses a cupric chloride solution to leach copper from the mineral chalcopyrite.The results suggest that the cathodic reactions are the reduction of [CuCl]⁺ to the complex [CuCl₃]²⁻, the reduction of [CuCl₃]²⁻ to solid copper and hydrogen evolution. The diffusion coefficient and the unit rate constants for the solution species were calculated. The exchange current density and rate constant for electron transfer were also estimated. A simulation was made of the cathodic polarization curve and it was in good agreement with the experimental data.     

转底炉直接还原铜渣回收铁、锌技术

采用转底炉直接还原工艺,将铜渣含碳球团在高温条件下直接还原得到金属化球团和高品位氧化锌粉尘,再通过熔分或磨矿磁选方式将铁回收,得到的铁产品可作为冶炼含铜钢的原料.转底炉中试结果表明:采用"转底炉直接还原—燃气熔分"流程处理铜渣,可获得TFe品位94%以上、铁回收率93%以上的熔分铁水;采用"转底炉直接还原—磨矿磁选"流程处理铜渣,可获得TFe品位90%以上、铁回收率85%以上的金属铁粉;采用两种流程处理铜渣,均可获得锌品位60.02%的ZnO粉尘.结果表明,经过转底炉直接还原,铜渣中的铁橄榄石Fe_2SiO_4和磁铁矿Fe_3O_4相转变为含有金属铁Fe、二氧化硅SiO_2和少量辉石相Ca(Fe,Mg)Si_2O_6的金属化球团,具备通过磨选或熔分进行进一步富集的条件.     

The Effect of Bath Parameters on the Electrocrystallisation of Cox–Cu100−x Alloys on Stainless Steel Cathode

The electrocrystallisation of the alloys of Co_x–Cu_100?x onto stainless steel cathode was investigated by performing cyclic voltammetry (CV) to understand the mechanism of deposition. The deposit morphology and crystal structure of deposit were analysed using scanning electron microscopy (SEM) and X-ray diffraction (XRD), respectively. The kinetic parameters were obtained from the cathodic polarisation of the CV to predict the electron transfer mechanism in the process. The transfer coefficient value (α) of the kinetic parameter revealed that both cathodic and anodic processes were unsymmetrical. It was demonstrated that the current efficiency of the deposit increased from 96.8% at pH 4.0 to 99.2% at pH 7, and then it dropped to 89.7% at pH 8. Before the deposition of the Co–Cu alloy, the initial copper deposition occurred at ??0.24 V and peaked at ??0.66 V. This was followed by the deposition of the Co–Cu alloy at ??1.04 V, which occurred after the deposition potential of Cu²⁺ (??0.24 V) and Co²⁺ (??0.89 V). The current then increasesd with a small increment in applied potential due to subsequent diffusion-controlled copper reduction along with the co-deposition of Co. The variation in the kinetic parameters was also reflected in the current efficiencies, the deposit morphologies, the crystallographic orientations and the nucleation overpotential values. The percentage of cobalt content in the alloy was observed to decrease in at.% from 54.35% at pH 4 to 49.86% at pH 6 and further to 20.62% at pH 8. The structure of the deposited alloy confirmed the formation of a single solid solution phase having different planes such as (222), (311), (220), (200) and a sharp peak due to face-centred cubic structure with (111) plane. This strong peak along with other similar peaks were observed in all the XRD of the deposit obtained at pH 4, 6 and 8. The morphology of the deposit characterised by the SEM showed that the deposit changed from a bitter gourd to a regular cauliflower-like structure as the pH value changed from 4 to 8.