用于从氰化矿浆中提取金的353E大孔双官能团树脂的制备

研究了用二甲胺和三甲胺的混合物作胺化剂,控制混合胺中各组成比例,使所制备的353E大孔双官能团阴离子交换树脂中2种官能团的比例适当.该树脂孔径适宜, 分布均匀,总交换容量高,机械强度好,动力学性能好,适用于从氰化矿浆中提取金.     

强碱性阴离子交换树脂吸附锆、铪分配系数研究

通过静态试验法,测试了201×7、D290、D296 3种离子交换树脂吸附硫酸锆(铪)溶液中的锆、铪的两相分配系数K_d(Zr)/K_d(Hf)的比值。试验证明D296大孔阴离子离子交换树脂吸附锆、铪的K_d(Zr)/K_d(Hf)的值最好,好于孔径较小的D290大孔阴离子交换树脂,即离子交换树脂的孔结构对两相分配系数比值有较为显著的影响。进一步研究了硫酸锆(铪)溶液中加入硫酸的量和溶液的温度对D296大孔阴离子交换树脂吸附锆、铪的K_d(Zr)、K_d(Hf)和二者比值的影响。研究表明,控制硫酸锆(铪)溶液的吸附温度在1~3℃和硫酸锆(铪)溶液中c(H_2SO_4)为1.6mol/L条件下,D296树脂对铪的吸附量最小,K_d(Zr)/K_d(Hf)的值达到最佳。研究结果表明控制D296树脂吸附硫酸锆(铪)溶液中的K_d(Hf)到较小的值,即K_d(Zr)/K_d(Hf)的值最大,有利于后续通过进一步淋洗试验分离得到核纯级的锆和铪。     

Selective flotation separation of spodumene from feldspar using new mixed anionic/cationic collectors

The selectivity in flotation separation of spodumene from other pegmatic aluminosilicates such as feldspar and mica, using a single fatty acid anionic collector, is found to be very difficult. It is attributed to the same surface Al site for interaction with fatty acid. In this study, a collector mixture of sodium oleate (NaOL) and dodecyl trimethyl ammonium chloride (DTAC) demonstrates a high selectivity for the flotation of spodumene from feldspar. The influences of important factors such as molar ratio of the mixed collectors, dosages of calcium chloride, sodium hydroxide and sodium carbonate, and pulp temperature on spodumene and feldspar flotation using the mixed collectors have been investigated through micro-flotation tests. The optimum molar ratio of NaOL and DTAC is found to be 9:1. Sodium hydroxide has to be used as a pH regulator. Without the use of depressants, no selectivity is observed as the flotation of spodumene and feldspar are activated by Ca²⁺. Sodium carbonate must be used as depressant of feldspar. A comparison of the flotation for a lithium pegmatite ore using mixed fatty acid soaps and mixed anionic/cationic collectors (NaOL/DTAC) was carried out by the batch flotation tests. The results indicated that NaOL/DTAC decrease collector consumption by two-thirds. The recovery and grade of Li₂O concentrates increase by 4.93% and 0.31%, respectively.     

Diagnostic pre-treatment procedure for simultaneous cyanide leaching of gold and silver from a refractory gold/silver ore

This study investigates the optimization of simultaneous dissolution of gold and silver from a refractory gold ore through determination of pre-treatment stages. Based on the mineralogical studies (thin layer and polished section) and chemical analysis on the ore sample, a “diagnostic leaching” procedure was designed. Results from diagnostic leaching suggest that the most effective pre-treatment agents for gold and silver are ferric chloride and sulfuric acid media, respectively. Optimum conditions for the simultaneous dissolution of gold and silver were determined using a two factorial design technique. Pre-treatments with sulfuric acid and ferric chloride reagents increased the efficiency of the dissolution of gold from 54.7% to 82% and silver from 37.4% to 81.6%.     

Solvent extraction of gold(I) from alkaline cyanide solutions by the cetylpyridinium bromide/tributylphosphate system

This paper explores the solvent extraction of KAu(CN)₂ from alkaline cyanide solutions using quaternary ammonium cetylpyridinium bromide (CPB) as an extractant with the addition of tributylphosphate (TBP) as a modifier. It also investigates the influence of several variables on gold extraction, including the molar ratio (β) of CPB to Au(I), the volume percentage of TBP (φ_TBP), NaCl concentration, phase ratio (AQ/ORG), and gold concentration in the aqueous phase. The results indicate that nearly all of the Au(I) (>98%) was transferred from the aqueous phase into the organic phase when β = 1 and φ_TBP = 30 vol%. We also carried out experiments for treating 20 L synthetic aurocyanide solution containing 10 mg/L Au(I) with column-shaped extraction equipment. The results demonstrated the recovery of more than 94.5% of Au(I) after two successive stages of extraction, and the Au(I) concentration in the raffinate was less than 0.5 mg/L. KSCN solution was used to strip the gold-loaded organic phase, and about 90% of Au(I) could be reverse extracted into the aqueous phase when the KSCN concentration reached 3.0 mol/L. The results obtained in this paper establish that the CPB/TBP extraction system has potential for practical application in the extraction and separation of gold from alkaline aurocyanide solutions.     

Simulation study of the optimal distribution of cyanide in a gold leaching circuit

The mineral industry has been using cyanidation to recover gold from ores for more than a century; however, a systematic study of the best reactant addition strategy in a cascade of agitated leaching tanks is not available in the open literature. A phenomenological mathematical model of the gold cyanidation process, calibrated with a set of industrial data from an Australian plant, together with an economic performance index is used to analyze this problem. The simulated results show that the best compromise between the two antagonistic effects, cyanide consumption and gold recovery, which are both function of cyanide concentrations, leads to a reagent distribution that depends on the leaching and cyanide consumption kinetics, pulp feed characteristics, and economic factors such as the gold market value. For the specific studied plant, in the operating range of low cyanide consumption and fast gold dissolution, all the cyanide must be added in the first tank; however, in the operating conditions of high cyanide consumption, cyanide has to be distributed in the first, second and third tanks.     

Selective leaching of arsenic and antimony from a tetrahedrite rich complex sulphide concentrate using alkaline sulphide solution

Removal of impurity elements in copper metallurgy is one of the major problems encountered today since pure copper ore reserves are becoming exhausted and the resources of unexploited ores often contain relatively high amounts of impurity elements like antimony, arsenic, mercury and bismuth, which need to be eliminated. The present work is aimed at pre-treating a tetrahedrite rich complex sulphide concentrate by selective dissolution of the impurities, therefore, upgrading it for pyrometallurgical processing. To accomplish this, dissolution of antimony and arsenic by an alkaline sulphide lixiviant from the concentrate were investigated. The lixiviant proved selective and effective to dissolve these impurity elements from the concentrate with good recoveries. Further investigations on the factors influencing the leaching efficiency of the lixiviant were studied. The parameters considered were sulphide ion and hydroxide ion concentrations, mineral particle size, reaction temperature and leaching time. Analysis of the leach residue indicates that copper content of tetrahedrite has transformed into copper sulphides with the average chemical formula Cu_1.64S. The grade and economic value of the concentrate were improved greatly after sulphide treatment, and therefore, suitable as a feedstock for smelting. The impurities have been reduced to low levels which are tolerable in the smelting furnace and consequently reduce both the treatment and environmental problem encountered when such concentrate is processed pyrometallurgically.     

对某铀矿碱浸液离子交换工艺中树脂转型有关问题的探讨

摘要:某铀矿碱(碳酸盐)浸液采用离子交换回收铀的工艺,贫树脂采用NaHCO3转型,由于转型液中ρ(NaHCO3)、ρ(Cl)高,ρ(U)较高,故难以有效利用.在对该离子交换回收铀工艺中存在的树脂转型有关问题进行探讨的基础上,提出了分段储存转型液和改变淋洗终点控制的措施.措施实施后,取得显著效果:转型液的利用率由原来的不到10％增加到20％左右,NaHCO3消耗量降低约30％.     

A review of copper cyanide recovery technologies for the cyanidation of copper containing gold ores

Many gold producers are today processing gold ores containing significant amount of cyanide soluble copper. Typically, cyanide destruction is used to prevent the discharge of copper cyanide into tailings storage facilities. This imposes a significant financial cost to producers from the additional cyanide used to solubilize the copper and the cost of cyanide destruction reagents. Therefore, the recovery of copper as a valuable by-product and the recycle of cyanide to the leach circuit have the potential for significant economic and environmental benefits. This includes enabling the treatment of gold ores with even higher soluble copper. Over the years, a variety of processes have been developed or proposed to recover the copper and/or cyanide including acidification based technologies such as AVR and SART, direct electrowinning, activated carbon, ion exchange resins, solvent extraction, polychelating polymers, and membrane technologies. In this paper, these processes are critically reviewed and compared, with particular focus on the advantages and limitations, and the separation of copper from cyanide. Ultimately, there is no universal process solution and the choice is highly dependent on the nature of the stream to be treated and integration with the whole processing plant.     

Prediction of cyanide recovery from silver leaching tailings with AVR using multivariable regression analysis

In this study, the tailings of the silver leaching plant, Eti Gumus AS, Turkey, containing high amounts of cyanide were tested to determine optimum recovery of cyanide, using packed columns for volatilization stage of AVR (acidification, volatilization or stripping, reneutralization). All recovery tests were performed in a pilot plant constructed using random packing with column internals and practised for mass transfer calculations. The main parametrical results of each test were measured in the laboratory and statistical analysis was performed using multivariable regression analysis. The parameters investigated were the viscosity of slurry, specific density, flow time of slurry in the funnel, air velocity, pulp density (wt%), pH, gas/liquid flow ratio (G/L), operating time, recovery rates and ambient temperature. The complex effects of the parameters were evaluated by multivariable linear regression analysis, favourable models then being corrected by using variance analysis. Verification of the predictive models to prove their practical validity and residual analysis of data were performed for preliminary treatment using the AVR process of leached tailings.     

Development and application of a first order rate equation for modelling the dissolution of gold in cyanide solution

A first order rate equation has been shown to accurately model the rate of gold extraction by cyanidation. The equation has the following form: E = EI(1 − e^−kt)where E is the gold extraction (%) at time t (hours), EI is the ultimate gold extraction (%) and k is the rate constant (h⁻). Cyanidation of four distinct ore samples indicated that the equation predicted the gold extraction as a function of time with a mean error of less than 3%.An algorithm has been developed so that the rate constant and ultimate fold extraction can be estimated from only four data points collected within the initial 28 hours of the leach. The model equation can also be used to statiscally quantify the effect of process variables on the rate of gold extraction and ultimate recovery. An example of the effect of dissolved oxygen concentration on the rate of gold extraction using Casa Berardi East zone ore is included in the report. The precision with which the extraction can be predicted could be increased by an improved sampling technique to minimize the error at leach times from 0 – 8 hours.     

Selective flotation of scheelite from calcite and fluorite using a collector mixture

Collector 733, a sodium soap (C_12–16COONa) is widely used industrially for scheelite flotation. Low selectivity of 733 collector is always observed. In this study, a collector mixture of 733 and MES (sodium fatty acid methyl ester sulfonate) demonstrated a high selectivity for the flotation of scheelite from calcite and fluorite. An optimal mass ratio 4:1 of 733:MES was found, producing a 65.76% WO₃ concentrate grade with a recovery of 66.04% from a feed material containing only 0.57% WO₃. In addition, the effect of water hardness and water glass addition were studied. The results indicated that the presence of Ca²⁺ or Mg²⁺ had little effect on the adsorption of the collector mixture at the scheelite surface. Addition of water glass for depressing calcite and fluorite had no significant effect on the adsorption of the collector mixture on the scheelite surface. The advantages of this new collector mixture (733+MES) include lower cost, low dosage, high tolerance against water hardness and high selectivity, and this collector mixture has great potential for industrial application.     

离子交换树脂在设备间转移过程中的磨损试验研究

设计加工了1套试验装置来研究树脂在输送过程中的磨损问题。考察树脂在转移过程中管道直径及水流量与树脂磨损的关系,树脂转移100次后测量其磨损量及磨损率,分析影响树脂磨损的因素,认为:树脂在输送过程中,当管径很大时(如Ф100mm),其内含空气,对树脂转移产生干扰,增大了树脂磨损量;当管径小时含空气量较小,在相同水流量下,管径越小,树脂在连接管中就越密集,树脂相互之间摩擦更剧烈,树脂与管壁的接触更频繁,磨损就越严重。在相同的连接管径下,流体流量越大,流速就越大,树脂与连接管管壁碰撞越剧烈,磨损也就越严重。     

含金银复杂多金属高硫铜矿粗矿再磨浮选回收铜金银硫

某铜金银多金属复杂高硫铜矿铜品位2.52%,含S 27.59%,伴生Au、Ag分别达7.8g/t、585.8g/t,金属回收价值高,硫化矿含量接近60%,铜硫分离困难。采用混合浮选技术,在粗磨基础上,对粗精矿进行再磨处理,药剂制度上采用新型抑制剂STY和CaO组合使用,并且加入少量硫化钠与活性炭进行脱药处理。采用单因素方法探索了药剂用量和药剂制度对浮选效果的影响。结果表明,在脱药剂(活性炭+Na_2S)用量为(1 500+100)g/t、石灰4 500g/t、STY 1 200g/t最佳药剂制度条件下可得到铜品位21.27%、回收率高达92.43%,含银高达4 115.8g/t、含金达34.9g/t的铜精矿,硫品位45%、含金7.5g/t、含银153.2g/t的硫精矿;铜硫分离效果很好,有价金属铜、硫、金、银均可得到高效回收。     

Bioleaching of gold and copper from waste mobile phone PCBs by using a cyanogenic bacterium

Chromobacterium violaceum (C. violaceum), a cyanide generating bacterium has been used to leach out gold and copper from the waste mobile phone printed circuit boards (PCBs) containing ∼34.5% Cu and 0.025% Au in YP (yeast extract and polypeptone with glycine) medium. The bioleaching was carried out in an incubator shaker (150 rpm) at 30 °C and 15 g/L pulp density in the pH range 8-11. Dissolution of gold and copper increased from 7.78% (0.225 ppm) to 10.8% (0.46 ppm) and 4.9% (419 ppm) to 11.4% (879 ppm) in 8 days with increase in pH from 8 to 11 and 8 to 10 respectively. Supplementing oxygen with 0.004% (v/v) H₂O₂ increased the copper leaching to 24.6% (1743 ppm) at pH 10 in 8 days whereas improvement in gold leaching was insignificant with the recovery of 11.31% Au at pH 11.0. The waste PCBs can thus be recycled in environmental friendly manner.     

Factors influencing the rate of gold cyanide leaching and adsorption on activated carbon, and their impact on the design of CIL and CIP circuits

The carbon in pulp (CIP) and carbon in leach (CIL) processes became firmly established in the gold mining industry in the 1980s, initially in South Africa and Australia, from where they spread rapidly to all the gold producing regions of the world. The percentage of annual global gold production by activated carbon-based processes grew from zero in the 1970s to almost 70% by the turn of the century, which represented a phenomenal rate of acceptance of a new technology by a traditionally conservative industry. The main reason for this rapid acceptance of the new technology was the fact that the first few large industrial plants in South Africa convincingly demonstrated better gold recoveries than the traditional filtration/Merrill Crowe process, with lower capital and operating costs. And as the plants developed an operating track record over their first few years of life, they proved to be remarkably robust mechanically, and highly tolerant of plant upsets, changes in feed composition and solution phase contaminants that had caused great problems in Merrill Crowe plants.These stellar attributes of the carbon-based gold plants have led to complacency and laziness in the industry, both at the new plant design stage, and with on-going optimization of existing plants. In many cases, basic “rules of thumb” that were developed as design criteria for the early CIP plants are still used today, with no appreciation of the factors that may cause one plant to perform quite differently from another. There seems to be little incentive to improve performance when it is well known that most CIP and CIL plants operate quite well with minimal optimization and, in many cases, minimal understanding of the factors that influence performance. Consequently, almost all CIP and CIL plants are overdesigned at the construction stage and are then operated sub-optimally. This can lead to higher gold losses and/or higher capital and operating costs than necessary.This paper examines the factors that influence CIP and CIL plant design and performance, and demonstrates a very simple methodology that can be used to arrive at something close to an optimum plant design. It can also be used as an on-going tool by plant metallurgists to transform a fairly well run plant into an exceptionally well run plant.     

Selective adsorption of precious metals from hydrochloric acid solutions using porous carbon prepared from barley straw and rice husk

Porous carbon was prepared by carbonization from agro-waste such as rice husk and barley straw to evaluate the adsorption of precious and base metals from metal solutions. The effects of hydrochloric acid concentration, metal ion concentration, and contact time on adsorption were examined. Rice husk carbon was found to be highly selective for Au(III) and inert to Pt(IV), Pd(II) and other base metals. Barley straw carbon adsorbed these three precious metal ions, but was inert to base metal ions such as Cu(II), Fe(III) and Ni(II). The maximum adsorption capacity of rice husk carbon for Au(III) was 0.76 mol/kg and the maximum adsorption capacity of barley straw carbon for Au(III), Pt(IV) and Pd(II) was 1.47, 0.39 and 0.64 mol/kg, respectively. The effectiveness of recovery of precious metals from industrial solution was also tested and barley straw carbon was found to be highly efficient and selective for the targeted metal ions in the presence of excess of other metal ions. Rice husk and barley straw carbon are thus potential alternatives to commercially available activated carbon as they have high selectivity and are efficient with low production costs.     

The use of ion exchange resins for the treatment of cyanidation tailings part 1––process development of selective base metal elution

This work investigates the use of oxidative acid eluents for the elution of base metals from strong base ion exchange resins. Eluents composed of a mixture of H₂O₂ and H₂SO₄ were tested for eluting base metals from resins loaded with mixtures of base and precious metal cyanides. This process removed 100% of Cu and Zn loaded on the resin, without affecting the precious metal loading. It was found that copper could be removed separately from the other base metals. The elution technique was not effective for removing iron from the resin. Cyanide associated with base metals was recovered as NaCN. Some oxidation of cyanide was noticed, subject to the elution conditions.This oxidative acid elution process could be used in commercial operations for the selective elution of base metals from a strong base ion exchange resin bed operating in alternative adsorption/base metal elution cycles. Thus, virtually all metal cyanide species could be recovered from cyanide leached solutions or slurries to give relatively clean tailings without compromising precious metal recovery efficiency. The process also caters for cyanide recovery and recycling.     

The recovery of nickel from high-pressure acid leach solutions using mixed hydroxide product – LIX®84-INS technology

This route for the recovery of nickel from HPAL laterite solutions was successfully demonstrated by the Centaur Mining Cawse Nickel Operation and has been selected by BHPBilliton for their Ravensthorpe Nickel–Yabulu Refinery circuit. The mixed hydroxide route is also well suited to treatment of solutions by other leach processes such as Activox^®.The technology can use an acid strip coupled to nickel electrowinning or an ammonia strip and basic nickel carbonate production. The solvent extraction technology involved in these two circuits is reviewed and compared. The ancillary operations of pregnant leach solution clarification, reagent remediation, reductive stripping of cobalt, zinc and copper transfers are discussed. The treatment of ammoniacal leach solutions containing higher copper concentrations than those normally seen in laterite processing is also addressed.Possible routes for recovery of cobalt and nickel from the solvent extraction circuit are also discussed.     

Prevention of acid drainage from gold mining in the western United States and impacts on water quality

Water quality regulations recently adopted by states in the western USA have increasingly stringent limitations on allowable changes in the quality of surface water and groundwater. The “nondegradation of water” provisions in state regulations require accurate predictions and control of the quantity and quality of acid mine water and strictly limit the entry of acid water into natural water systems. Long-term water quality impacts from mine waste rock, spent ore from heap leaching, tailings and open pits must be considered in design, operation and reclamation of proposed or expanded mining operations. Acid-base testing, humidity cells, column testing and shake flask tests have been used with mixed success to predict the extent of acid water production. The types and forms of sulfide minerals present, bacterial catalysis of the sulfide oxidation reaction and configuration of the reclaimed facilities are all important elements in accurately predicting acid mine drainage. A critical factor in prediction of acid mine impacts is a pathway and fate analysis which includes geochemical reactions with aquifer materials and dilution and dispersion of parameters in the leachate plume. Of particular concern is the production and transport of arsenic, metals and residual cyanide from mined areas. Evaluation of three major operating gold mines in the northwestern United States shows the relationship between production of acidic water, movement of this water in aquifers and impacts on groundwater and surface water. Column testing showed reduction in concentrations of most metals by 50 to over 90 percent during travel through aquifers. Clays and silt zones were very effective in adsorbing metals. Operational control of water/rock reactions and reclamation design can significantly reduce or eliminate acid drainage. Soil cover, revegetation and slope are the major components that limit long-term acid drainage and metal contamination of surface water and groundwater. Compliance with water quality limits can be achieved only by design and operation of mining facilities to minimize the formation of acidic waters.