某含银高硫铜矿的综合回收试验

某含银高硫铜矿含铜0.76%、硫24.35%及银34.92 g/t,有价矿物种类多、矿石性质复杂,采用抑硫优先浮选铜-活化浮选硫的原则工艺流程进行试验,配合石灰作为硫化铁矿物抑制剂以及筛选出丁基黄药+酯-105作为硫化铜矿物的组合捕收剂,强化了银在铜精矿中的富集。在选定工艺条件下,可获得铜品位21.60%、银品位602.84 g/t的铜精矿(铜和银回收率分别为89.30%和54.39%),硫品位45.60%、银21.55g/t的硫精矿(硫和银回收率分别为89.79%和29.59%),实现了铜、硫和银的综合回收利用。     

富银铜锌多金属矿选矿试验研究

某富银铜锌多金属硫化矿,其银品位为125 g/t,铜品位为0.26%,锌品位为1.13%,硫品位为3.04%;90%的银赋存于银黝铜矿、黝铜矿等铜矿物中。采用“优先浮银铜﹣锌硫混浮﹣锌硫分离”工艺流程进行处理,选择高效银捕收剂SAC强化银的回收,全流程实验获得的银(铜)精矿含银10094 g/t,含铜16.67%,银回收率88.78%,铜回收率80.23%,锌精矿含锌45.46%,锌回收率81.81%。实现了矿石中银铜锌的综合回收。     

焙烧-氰化工艺处理含金硫精矿提高银回收率的研究

采用焙烧-氰化工艺处理复杂含金硫精矿时,银的回收率偏低,为了提高该过程银回收率,详细的研究了焙烧制度、焙烧料加添加剂、盐浸条件和氰化条件等对银回收率的影响。试验结果表明,焙烧过程中加添加剂和焙烧温度是影响银回收率的2个关键因素;提高银回收率适宜的条件为:焙烧时加入5%添加剂A,温度控制在450~500℃,所得焙砂经酸浸铜、盐浸铅后,在液固比2~3,氰化物质量分数0.4%,搅拌氰化24 h,弃渣含银40 g/t,银的总回收率81%。     

高砷难处理硫精矿氰化浸出提银实验研究

某硫精矿中的银矿物嵌布粒度非常细小,包裹于硫化物中的银约占50.46%,属于难处理高砷含银硫精矿。采用细磨后化学预处理氰化浸出,银浸出率仍然低于80%;硫精矿经氧化焙烧后,As、S的脱除率都达到90%以上,但银浸出率却较低;对该含银硫精矿添加钠盐焙烧预处理,再采用常规氰化法浸出,银浸出率显著提高,达到85.15%,同时氰化钠耗量降低至2.0 kg/t。     

高银铅铜混合精矿选矿试验研究

某高银铅铜混合精矿为铜铅混浮产品,表面受到浮选药剂严重污染,导致铜矿物与铅矿物可浮性差异微小,给铜铅分离带来非常不利的影响,其银品位达到4208.0 g/t,铅品位43.28%,铜品位3.39%,98.80%的银赋存于方铅矿。采用“活性炭脱药-抑铅浮铜”工艺流程处理,环保高效的方铅矿抑制剂GYC,高效铜捕收剂SAC,全流程试验获得银品位4839.2 g/t,银回收率97.36%,铅品位50.05%,铅回收率97.90%的银铅精矿;铜品位20.16%,铜回收率91.23%,银品位724.6 g/t,银回收率2.64%的铜精矿,实现了混合精矿中银铅与铜的高效回收。     

云南某含银铜铅混合精矿分离试验研究

云南某铜铅混合精矿含Cu 8.14%、Pb 38.57%、Ag 251.62 g/t，对其进行浮选试验研究铜铅的分离。通过条件试验，确定在磨矿细度为-200目含量为93.85%的情况下，抑制剂CMC+亚硫酸钠用量选择1000 g/t，捕收剂Z-200用量选择10 g/t。采用“抑铅浮铜”一粗三精一扫的闭路试验流程，获得铜品位24.73%、回收率87.24%、含铅品位6.23%的铜精矿；铅品位62.71%、回收率84.48%，含铜0.86%的铅精矿。银在铜铅精矿中进一步富集的总回收率为73.04%，实现了该铜铅混合精矿的分离及银的进一步富集。     

某微细粒难选金矿氰化尾矿的浮选提金工艺试验研究

新疆某金矿历年来生产的氰化浸出尾矿,组成较为复杂,金、银、硫集中分布在-0.045 mm粒级中,通过对该金矿氰渣的实验室试验和扩大连续浮选试验研究,取得了金精矿品位19.50 g/t,回收率66.50%的工艺指标,有效地回收了该矿石中的微细粒金和包裹金。     

云南某含银铜矿伴生银矿物强化回收实验研究

云南某含银铜矿石银品位为21.77 g/t,主要赋存于黄铜矿、游离银和磁黄铁矿中,在选别过程中铜硫分离时兼顾银的回收具有一定的难度。采用抑硫浮铜流程,研究筛选出石灰及Z200作为合适的调整剂和高效捕收剂,在抑制黄铁矿、磁黄铁矿的同时强化银的回收。最终闭路试验获得了品位23.92%、回收率97.10%的铜精矿,且含银136.38 g/t,银的回收率达到77.05%,实现了铜硫分离的同时强化银的回收。     

水热浸出-浮选综合回收铅银渣中金银实验研究

为了提高资源利用效率,实现二次资源综合利用,对铅银渣进行了实验研究。铅银渣中银品位为245 g/t,金品位为1.52 g/t,金、银含量均较高。银的赋存状态研究表明,银主要以再造矿物铜蓝、硫化银混合相存在。结合银的赋存状态进行流程探索,确定采用水热浸出-浮选工艺流程。条件实验研究表明,液固比为2:1,浸出温度为70℃,浸出时间为2 h进行水热浸出,金、银在渣中富集率较高。当磨矿细度-0.044 mm占85%,粗选T19用量为4000 g/t,硫酸铜用量为600 g/t,酯-30用量为500 g/t时,经过一次粗选、两次精选、一次扫选的闭路实验流程,可获得银品位为3805 g/t,银回收率为86.82%的银精矿,银精矿中金的品位为25.8 g/t,金回收率为94.96%的较好指标,实现了铅银渣的综合回收。     

湿法炼锌酸浸出渣浮选回收银试验

传统锌精矿伴生银回收,均是从挥发窑渣中进行,由于矿物性质变化,导致银无法有效回收。进行了从锌酸浸出渣中浮选回收银的研究,解决了锌精矿中伴生银回收的问题。某湿法炼锌浸出渣含Ag 350 g/t,Au0.01 g/t,Pb 3.87%,Zn 17.45%,Cu 1.38%,银主要以自然银存在,占60.13%。采用高效捕收剂HT-1#、起泡剂HT-2#进行浮选,经过一粗四精三扫工艺,获得银精矿产率4.39%,含银6616 g/t,银回收率82.98%,取得良好的经济技术指标。     

铂钯精矿湿法分离Bi、Fe富集贵金属工艺

以铜阳极泥分铜液还原所得铂钯精矿为原料,根据其矿物特性选择HCl作为浸出剂湿法脱除Bi、Fe等主要贱金属元素,富集Au、Ag、Pt、Pd等贵金属;通过热力学计算绘制Bi(Ⅲ)、Fe(Ⅲ)在盐酸体系的组分分布图,实验考察了HCl浓度、Cl-浓度、反应温度和时间等因素对Bi、Fe浸出率的影响.结果表明:在HCl浓度为2 m...     

THE FLOTATION OF COPPER OXIDIZED ORES WITH MIXED REAGENT OF HYDROXAMATE AND XANTHATE

羟肟酸钠与丁黄药联用,可以提高矽卡岩氧化铜矿的浮选效率,降低药剂消耗.特别是羟肟酸-黄药“混合剂”的合理使用,促成了氧化铜矿浮选技术的新进展:处理含孔雀石和假孔雀石矿石,能获得精矿品位约26%Cu,回收率近80%,从高硅泥质铜矿中,可直接浮出精矿含铜≥25%.自由氧化铜回收率≥85%,选择合适的调整剂及相应的流程结构更有利于羟肟酸-黄药“混合剂”在Cu-Fe共生难选养化矿浮选中发挥最佳效用,不仅可以得到品位达36%的优质铜精矿.而且能预先浮出富含Ag的金精矿(每吨含Ag为340g,Au为108g),并有助于从铜扫选尾矿中回收铁.     

铜铅多金属硫化矿中伴生金的强化回收实验研究

采用铜铅混选富集-抑铜浮铅浮选分离的工艺,对某铜铅多金属硫化矿中的伴生金进行强化回收研究。在铜铅混选阶段,弱碱性条件(pH=9)下,用Z-200(30 g/t)做捕收剂,金在铜铅混合精矿中有效富集;在铜铅分离阶段,以硫化钠(4000 g/t)预先脱药,用乙硫氮(30 g/t)浮选铅矿物,金在铜精矿中进一步富集。工艺闭路实验获得含铜18.69%、含金42.70 g/t的含金铜精矿,铜和金的回收率分别达到92.58%和56.84%;还同时获得含铅61.45%的铅精矿。可实现铜铅多金属硫化矿中伴生金的强化回收。     

山东某含金磁黄铁矿浮选尾矿磁选回收试验研究

山东某含金磁黄铁矿原矿金品位1.60 g/t,硫品位1.86%,属含金硫铁矿。矿石性质研究结果表明,部分以磁黄铁矿为载体的金,矿物含量为0.96%,金品位8.25 g/t,原矿金分配率5.25%。生产流程对以磁黄铁矿为载体的金矿物的回收水平仍有提高空间。为了解决这一问题,开展了从生产原矿和生产尾矿中回收以磁黄铁矿为载体的金的对比试验,结果表明,磁选不宜用于原矿、重选不宜用于尾矿中载金磁黄铁矿的回收;尾矿磁选流程可以实现含金磁黄铁矿的有效富集,最终选择全粒级磁选工艺流程,获得了金品位1.52 g/t,硫品位2.87%的含金磁黄铁矿。尾矿金、硫回收率分别为52.09%、62.93%,对原矿回收率分别为12.27%、18.56%,实现了以磁黄铁矿为载体的金矿物的综合利用。     

Process for Complete Utilization of Zinc Leach Residues

Filter residues are mixed in high proportion with pyrite cinders and subjected to a modified chloridizing and sulfatizing roast in standard multiple-hearth furnaces. By additions of sodium chloride and of suitable sulfur-bearing materials, the metals Cu, Zn, and Cd, as well as Ag and Au, are rendered soluble to a large extent. By leaching with acid, Zn, Cu, Ag, Cd, and Au are eluted. The remaining iron oxide is conveyed into sintering or pelletizing plants.If the residues contain in addition undesirable elements such as As, Sb, and Sn, they are first subjected to a chloridizing and sulfatizing roast; Cu, Zn, and the noble metals are extracted by leaching. The fine-grained iron oxide, containing As, Sb, and Sn, is either discarded or purified from these elements by means of chloridizing volatilization in a reducing atmosphere. Given sufficiently high contents, these accompanying elements can be recovered economically.In recovering the nonferrous metals from the solutions, the classical precipitation and cementation processes, well tried under plant conditions, are compared with ion exchange processes with regard to cost and process efficiency. Pilot plant results of the application of ion exchangers to solutions of higher concentration are given.     

FeCl3浸取含Au复杂硫化矿的动力学研究

研究了FeCl_3溶液浸取复杂含金硫化精矿中有价金属Cu,Ag,Pb和Zh的过程,考察了主要过程参量对浸取速率的影响.实验数据用非线性回归方法处理,给出动力学模型,说明精矿中Cu的溶解过程由表面化学反应控制,其速率是Fe~(3+)的0.5级和Cl~(-)的零级反应.实验条件下Cu,Ag,Ph和Zn的最终浸取率分别达到96,95,90和91%,超过90%的Fe进入浸取液,总硫的70%以上转化为元素硫,而Au的浸取率则小于3%表明FeCl_3溶液浸取是含Au复杂硫化精矿预处理并回收有价金属的有潜力的方法     

Surface enoblement by dissolution of Cu,Ag and Zn from single phase gold alloys

Single phase alloys with 30 at.% Au and with different proportions of Ag, Cu and Zn as additional alloying elements have been investigated with regard to selective dissolution. Measurements of rest potentials and ESCA analyses show that Cu and Zn are preferentially dissolved in aqueous solutions of chloride and small concentrations of sulphides leaving behind a surface significantly enriched in Au with a thickness of approximately 1 nm. No significant selective dissolution of Ag was observed in the solutions investigated. Only alloys containing substantial amoungs of Ag tended to become tarnished in sulphur containing solutions. It is suggested that Ag₂S is more easily formed than sulphides of Cu and Zn because also other oxidative reactions to oxides/ hydroxides of these elements take place, which is not the case for Ag. These oxides/hydroxides of Cu and Zn dissolve into the solution creating an Au-surface enrichment also in the diluted sulphide-containing solutions.     

PURIFICATION OF COBALT ANOLYTE USING THE NOVEL SOLVENT EXTRACTION SYSTEM

In present research, a novel extractant system (D2EHPA + naphthenic acid + pyridine- ester) was used to purify cobalt anolyte and a simulated industrial production were carried out. This novel extraction system can extract Cu and/or Ni against Co from chloride medium solutions at pH range of 2.5-4.5. About 2g/l nickel and 0.2g/l copper were removed from the cobalt chloride anolyte containing about 100g/l cobalt and 200g/l chloride ions respectively, the raffinate contains nickel and copper less than 0.03g/l and 0.0003g/l respectively and can be used to electrolyze high-purity cobalt. About 5.5t cobalt anolyte was purified in the simulation industrial experiment and kilogram quantities of cobalt of 99.98% purity and about 95% recovery have been produced.