湿法炼锌渣中银的回收

在湿法炼锌工艺中锌精矿中的银主要富集在酸性浸出渣中,此矿样的浸出渣中Ag的品位约为234 g/t,还含有Zn、Pb等可重复利用金属,研究Ag、Zn、Pb等的回收再利用具有十分重要的意义。本文以酸性浸出渣为原料进行了物理分选、还原焙烧、直接熔炼法以及氧化焙烧-氰化提银的试验,重点研究了物理分选过程Ag、Zn、Pb的富集走向及氧化焙烧-氰化提银工艺中氯化钠用量、焙烧时间及温度对Ag浸出率的影响。研究得出:高温高酸浸出后浮选可使Zn和Ag得到富集;浸出渣酸浸后熔炼使粗铅中的Ag和Pb富集,Ag品位可提高6倍;并通过试验得到了较优的氧化焙烧和氰化浸出提银工艺参数。      

Investigations into complex processing of zinc cakes

Investigations are carried out into the processing of zinc cakes with the purpose of transferring zinc, copper, and iron from ferrites into a solution and of concentrating noble metals in the lead-containing silicate product. Zinc cakes are subjected to sulfatization by oleum followed by leaching with a concentrated H₂SO₄ solution. With the use of a planned multifactor experiment according to the Box plan, the effect of the H₂SO₄ content in the leaching solution on the cake leaching process, its duration, and temperature has been studied. Due to this, the models of the dependence of the listed factors on the extraction of Zn, Fe, and Cu into the solution have been obtained. The possibility of an almost complete extraction of these metals into the solution is demonstrated. The following leaching conditions are suggested as optimal: H₂SO₄ concentration of 8.33 g/l in the leaching and washing solutions, leaching time of 2 h, and leaching temperature of 75°C.     

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.     

Precious Metals Recovery from Waste Printed Circuit Boards by Gravity Separation and Leaching

ABSTRACT

This study focuses on the recovery of valuable metals, such as gold, silver, and copper, from the printed circuit boards of waste computers, using physical separation followed by leaching methods. Characterization studies revealed that resins and glass fibers were attached as grain together with base and precious metals. A hammer mill was employed as a second stage crusher to disintegrate the different components of the printed circuit board, thus improving the selectivity and recovery of metals. Separation studies using a laboratory-scale shaking table showed that 33% of the feed was removed as a light product and 96.8% of Au, 96.7% of Ag, and 97.7% of Cu were recovered in heavy fraction. Leaching the light fraction using 3 M H₂SO₄ and 0.33 M HNO₃ at 80°C for two h resulted in greater than 90% extraction of Cu. Au and Ag were dissolved at room temperature with a leaching solution of 0.2 M S₂O₃ ^2-, 0.02 M CuSO₄, and 0.2 M NH_3.H₂O, which provided a recovery of more than 59% of Au and 98% of Ag within eight h. Direct leaching tests using optimized conditions were also implemented on the crushed sample, and 45% of Au, 87.6% of Ag, and 70.8% of Cu were extracted.     

Metal Recovery from Bottom Ash of an Incineration Plant: Laboratory Reactor Tests

The bottom ashes, produced by the industrial incinerators, are an essential secondary raw material resource that has been drawing attention to recover economically important metals. In the present study, a hydrometallurgical process (chemical leaching) has been discussed to recover some economically important metals using lab scale reactors (600 mL and 2L capacity). Prior to the leaching tests, the material was characterized for chemical composition and mineralogical phase analysis through XRD. In addition, the conventional process of magnetic separation was applied before leaching tests to remove some easily separable parts. Two acidic reagents (HCl and H₂SO₄) and an alkali reagent (NaOH) have been used to compare the recovery of metal values from the pre-treated bottom ash samples at varying concentrations. Process parameters such as acid/alkali concentration, working volume and temperature have been optimized and the recoveries of metal values under optimum conditions were recorded. The studies showed that 3M H₂SO_4, 1.5 L working volume and 80°C were sufficient to leach 88% Mo, 82% V, 37% Ni, 37% Fe, and 28% Cu. Acid leaching tests using H₂SO₄ were found to be an economical and appropriate solvent for metal recovery from bottom ash.     

山东某焙烧氰化尾渣综合回收工艺研究

针对山东某焙烧氰化尾渣的性质,采用高温氯化焙烧工艺对其进行了综合回收试验研究。结果表明:在氯化钙添加量7%、焙烧温度1 100℃、焙烧时间2 h的工艺条件下,有价金属Au、Ag、Cu、Pb、Zn的挥发率分别达到97.52%、71.78%、85.92%、98.59%、93.81%,指标较好。同时,根据试验结果提出了尾渣综合回收有价金属的工艺流程,为后续研究奠定了基础。     

Fluidized Bed Selective Oxidation-Sulfation Roasting of Nickel Sulfide Concentrate: Part II. Sulfation Roasting

The fluidized bed sulfation roasting process followed by water leaching was investigated as an alternative process to treat nickel sulfide concentrate for nickel production. The effects of several roasting parameters, such as the sulfation gas flow rate, roasting temperature, the addition of Na₂SO₄, and the roasting time, were studied. 79 pct Ni, 91 pct Cu, and 95 pct Co could be recovered with minimal dissolution of Fe of 4 pct by water leaching after two-stage oxidation-sulfation roasting under optimized conditions. The sulfation roasting mechanism was investigated, showing that the outermost layer of sulfate melt and the porous iron oxide layer create a favorable sulfation environment with high partial pressure of SO₃. Sulfation of the sulfide core was accompanied by the conversion of the sulfide from Ni_1?x S to Ni₇S₆ as well as inward diffusion of the sulfation gas.     

污泥中重金属的回收工艺研究

研究了重金属污泥中的4种重金属离子(Cu2+、Zn2+、Pb2+、Cd2+)的回收方法,考察了酸的种类、酸浓度、温度、时间、固液比对浸出效果的影响。结果表明,4mol/L的HNO3与750℃焙烧3h后的污泥在固液比1∶20,转速600r/min,80℃下反应3h,Cu2+、Pb2+、Zn2+、Cd2+的浸出率分别为92.0%、99.5%、99.5%、99.3%。     

从多金属复杂金精矿焙砂中分离提取有价金属

对某多金属复杂金精矿的焙砂进行了"酸浸提取铜锌—盐浸提取铅银—氰化提取金银"工艺试验,获得了各工序的最佳工艺条件。在最佳工艺条件下的综合试验表明,金、银、铜、锌、铁的回收率分别为94.63%、65.12%、90.45%、82.87%、98.92%,有效实现了各有价金属的综合回收。     

氰渣低温微熔锍化资源化技术

黄金行业产出的氰渣已列为固体危废,其脱毒处置技术是所有黄金企业不得不面对的严峻挑战和必须尽快解决的重大难题。探究了利用低温微熔锍化法处置低品位氰渣并高效回收贵金属及有价金属的技术可行性。考察了焙烧温度、焙烧时间及添加剂用量等条件对贵金属回收率的影响,阐明了在促进剂、低温和还原气氛作用下形成低熔点金银及铜铅锌等有价金属锍化物晶核长大再次矿化、骤冷后浮选回收的机理。在优化条件下处理含金1.71 g/t、银54 g/t、铁32.71%的氰化尾渣,渣中金品位可降至0.14 g/t,浮选精矿中金、银的品位分别达到35.46、737.14 g/t,金、银的平均回收率分别达到92.44%、80.67%,磁选所得铁精矿中铁的品位为63.57%,回收率83%,产率68%。有效回收贵金属的同时,实现了低品位氰渣的无害化处置和资源化利用。     

铜碲渣中碲的回收工艺研究

采用硫酸化焙烧—水浸脱铜—碱浸提碲—电积工艺回收铜碲渣中的碲,考察了NaOH浓度、温度等对碲浸出的影响,并对水解脱杂和碲电积工艺进行探索。结果表明,经过硫酸化焙烧、水浸后,铜脱除率为75%;电积产品碲纯度为99.90%,综合碲回收率约93%;金、银、铂和钯等贵金属进一步富集在碱浸渣中。     

The co-precipitation of copper and zinc with lead jarosite

The formation of lead jarosite, Pb_0.5Fe₃(SO₄)₂(OH)₆, in the presence of dissolved copper and/or zinc results in a significant substitution of these metals in the jarosite phase; the co-precipitation is most pronounced in sulphate media but also occurs, to a lesser degree, in chloride solutions. The copper and/or zinc substitute for iron, and under extreme conditions the product approaches beaverite, Pb(Cu,Zn)Fe₂(SO₄)₂(OH)₆, in structure and composition. The extent of co-precipitation increases sharply with increasing concentrations of dissolved CuSO₄ or ZnSO₄ and slightly with either an increasing stoichiometric ratio of PbSO₄/Fe³⁺ or increasing ionic strength. The co-precipitation of copper or zinc is not significantly affected by acid concentration although the yield of product declines with increasing concentration of H₂SO₄. The extent of reaction is relatively insensitive to reaction temperatures in the range 130–180°C and to reaction times in excess of 2 h. Copper is strongly co-precipitated in preference to zinc from solutions containing both metals. Other divalent base metals such as Co, Ni and Mn are also co-precipitated with lead jarosite although not to the same degree as copper or zinc.     

EFFECT OF CERTAIN SURFACTANTS ON THE LEACHING AND PRECIPITATION PROCESSES IN ZINC FERRITE RESIDUE TREATMENT

The effect of a surfactant mixture of nonylphenolpolyethylene glycol (D1), dinaphthylmethane-4,4′-disulphonic acid (D2), and polyethylene glycol with molecular weight 400 (D3) on the dissolution of zinc and metal impurities present in zinc ferrite residue in dilute sulfuric acid (160 g L^?1 H₂SO₄) as well as on both jarosite and goethite precipitation was studied at 90°C. The following influences of the surfactant mixture (D1 + D2 + D3), determined by comparing the results obtained in the presence and absence of surfactants, were found. Adsorption of surfactants on zinc ferrite residue surface decreases the dissolution of zinc and metal impurities (Fe, Cu, Cd, As, Sb, and Co). Their extraction efficiencies at the end of the super hot leaching process carried out with surfactants are 4.85–6.29% lower than without them. The formation of a sulfur “sponge” layer on the surface of liquor during the dissolution of ZnS present in zinc ferrite residue is hindered by the surfactants due to their effect as wetting agents and sulfur dispersants. The presence of surfactants reduces the amount of zinc and metal impurities (Fe, Cu, Cd, and As) remaining in the solution after jarosite or goethite precipitation by 5.33–5.86% or 8.03–9.93%, respectively. The volume of jarosite and goethite precipitates increases in the presence of surfactants due to their effect as wetting and flocculation agents. On the other hand, D1 and D3 act as complexing agents. The abovementioned effects of surfactants improve the sorption capacity of both jarosite and goethite, thus ensuring better purification of zinc sulphate solutions, but hindering zinc leaching.     

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.     

Effect of roasting activation of rare earth molten salt slag on extraction of rare earth,lithium and fluorine

A new process was proposed to extract rare earth elements(REEs),Li and F from electrolytic slag of rare earth molten salt by synergistic roasting and acid leaching.Firstly,the thermodynamic analysis of roasting reaction was carried out,then the effects of roasting factors on leaching REEs,Li and F in slag were investigated.In additions,the mineral phase and morphology of molten salt slag,roasting slag and acid leaching slag were characterized,and the migration mechanism of REES,Li and F minerals...     

Leaching reaction and kinetics of zinc from indium-bearing zinc ferrite under microwave heating

Leaching of zinc from indium-bearing zinc ferrite (IBZF) under microwave heating (MH) has been investigated. The result showed microwave intensified the leaching reaction of IBZF in the MH process. Microwave had a great nonthermal effect on the leaching reaction. The effective collision and the H₂SO₄ activation under the action of microwave belonged to the nonthermal microwave effect. Particle size of IBZF in the range from 45 to 150 μm almost had no effect on the zinc leaching in the MH process. Leaching temperature and leaching time had important effects on the zinc leaching. Zinc leaching in the MH process obeyed the unreacted shrinking core model very well, and the activation energy was 73.747 kJ/mol. The kinetic equation was \(1 - (1 - x)^{1/3} = 8.82 \times 10^8 e^{ - 73.747 \times 10^3 /RT} t\). The ratio of frequency factor of K_0(In)/K_0(Zn) was up to 4.69, indicating the effect of microwave intensification on the indium leaching was greater than that on the zinc leaching.     

The dissolution of sphalerite in ferric sulfate media

The dissolution of sphalerite, (Zn,Fe)S, in ferric sulfate media was investigated using closely sized fractions of crushed sphalerite crystals. Linear kinetics were observed, and the rate increased in proportion to the surface area, as the average particle size of the sphalerite decreased. The predominant reaction products are ZnSO₄, FeSO₄, and elemental sulfur. The leaching rate increases with increasing temperature, and the apparent activation energy is 44 kJ/mol. The relatively high apparent activation energy suggests that the rate is chemically controlled, a conclusion supported by the insensitivity of the rate of the rotation speed that was observed in complementary rotating disk experiments. The rate increases as the 0.3 to 0.4 power of the Fe(SO₄)_1.5 concentration, and is nearly independent of the pulp density, in the presence of a stoichiometric excess of ferric sulfate. In 0.3 M Fe(SO₄)_1.5 media, the rate increases with increasing acid concentrations >0.1 M H₂SO₄, but is insensitive to more dilute acid concentrations. In the absence of ferric ions, the rate increases rapidly with increasing H₂SO₄ concentrations, and relatively rapid rates are observed in solutions containing >0.5 M H₂SO₄. The rate decreases with increasing initial concentrations of ZnSO₄, MgSO₄, or FeSO₄ in the ferric sulfate leaching solution, and this emphasizes the importance of maintaining the dissolved iron in a fully oxidized state in a commercial leaching operation.     

杂铜阳极泥回转窑氧化焙烧酸浸工艺研究

采用回转窑氧化焙烧—酸浸工艺回收杂铜阳极泥金属铜,研究了不同试验条件对铜浸出率的影响。结果表明:在氧化焙烧温度700℃、焙烧时间20 min、原料粒度-5 mm、空气流量0.5 L/min的条件下,铜浸出率高达97.10%,镍浸出率>90%,大部分铅、锡、锑、铋及贵金属金、银、钯残留在浸出渣中,可以作为后续提取有价金属及贵金属的原料。