湿法炼锌净化钴渣新处理工艺

提出了氨－硫酸铵体系处理湿法炼锌净化钴渣的新工艺。该工艺首先用铵－氨水溶液浸出烘烤后的钴渣，再用锌粉对浸出液净化除杂并进行锌与镉、钴、铜及铜与钴的分离，在最佳技术条件下，金属浸出率（质量分数）分别为：Zn91.18%，Cu96.98%，Cd99.385,Co89.35%；净化所得的富钴渣含Co3.79%，富集比达8.4，从这种钴渣中可直接提取钴或钴盐；而净化液可直接制取活性锌粉，氨－硫酸铵法具有原料适应性强，设备防腐要求低，能常温操作，能耗低，除杂容易等优点，是处理湿法炼锌废渣、综合利用有价元素的较好方法。     

云南某低品位高结合率氧化铜矿选冶联合试验研究

云南某难选铜矿石属低品位高结合率氧化铜矿,采用选冶联合工艺实现了对该铜矿的高效回收.浮选试验研究表明,磨矿细度为-0.074 mm占80%,调整剂水玻璃用量800 g/t,组合捕收剂丁黄药+异戊黄(1∶1)用量240 g/t,铜回收指标最佳,经二粗二精闭路试验获得的铜精矿中铜品位9.17%,回收率17.76%.浮选尾矿经磁选除铁,磁选尾矿采用硫酸浸出,当磨矿细度为-0.045 mm占54%,硫酸用量60 kg/t,液固比3∶1,转速250 r/min,浸出温度45℃,浸出时间2h时,铜浸出率为81.88%,铜的综合回收率达86.55%.     

镍精矿氯气浸出液净化除铁工艺

硫化镍精矿氯气浸出－净化－电积工艺作为镍提取冶金的新方向，其净化过程大多采用溶剂萃取方法。该方法对于含铁较高的浸出液在一段时间后会出现萃取有机相性质变坏的问题，影响萃取剂的正常使用。因此，净化过程中必须先进行单独除铁。本试验研究了针铁矿法从镍精矿氯气浸出液中除铁的工艺。实验考察了在氯盐体系中氧化剂用量、反应温度、反应ｐＨ值、反应时间等因素对除铁效果的影响。确定了最佳工艺条件：氧化剂用量５ｇ／Ｌ，反     

采用硫酸铵焙烧方法从低品位碳酸锰矿中富集回收锰

研究了采用硫酸铵焙烧法从低品位碳酸锰矿富集回收锰的工艺．将碳酸锰矿与（NH4）2SO4通过研磨混合均匀，在马弗炉中300～500％焙烧0．5～3h．将焙烧过程铵盐分解产生的氨气发二氧化碳气体通过真空通入上一次的浸出液，将硫酸锰沉淀下来．焙砂采用60—90％热水，在液固为（3～10）：1的条件下浸取10—20min，得到硫酸锰浸出溶液，作为下一次焙烧过程的吸收液，对吸收液巾过滤得到的滤饼进行干燥后得到锰精矿．吸收液过滤后得到的滤液蒸发浓缩，结晶后又得到了（NH4）2SO4固体，可以复用．采用该工艺的锰回收率达80％以上，是一种从低品位碳酸锰矿富集回收锰的新工艺。     

压水堆主回路模拟腐蚀产物在柠檬酸,过氧化氢水溶液中的溶解特性研究

研究了压水堆主回路钴的模拟腐蚀产物在柠檬酸水溶液、过氧化氢水溶液、柠榨酸－过化氢水溶液中的溶解 特性，结果表明：Ｑ三模拟腐蚀产物在柠檬酸水中的溶解量随柠檬酸 增加而明显增大并且比铁的相应值高１０倍以上，在ＰＨ２．３３－ＰＨ４．５０范围内，ＰＨ３．００溶解量最大，在６０、７０．８０℃温度下，７０℃时的溶解量最高；钴模拟 物在Ｈ２Ｏ２水溶液中的溶解量比柠檬酸中的相应值低，钴模拟腐蚀产物在过氧化氢－柠檬     

低品位铜锌混合矿加压浸出研究

研究了低品位多金属铜锌混合矿的浸出条件．探讨了氧分压、酸度、温度、反应时间、添加剂等因素对铜锌浸出率的影响．结果表明，铜和锌的浸出率可达98％和99％．提高了矿产资源的利用率，利用直接加压浸出的方法取代传统的焙烧——浸出工艺。从生产源头上消涂了烟气污染．     

电镀污泥中铬的无害化处理及动力学分析

以某表面处理工业园电镀废水处理污泥为研究对象,以铬浸出率为指标,通过对重金属的浸出,分步回收达到无害化、资源化的目的.将污泥干燥、研磨,在不同浓度硫酸溶液中浸出,控制浸出时间、浸出温度和搅拌速率;浸出完成后抽滤使浸出液与残渣分离.采用正交试验法,确定对铬浸出效果影响因素的顺序为:硫酸浓度>搅拌速度>浸出时间>固液比.通过单因素优化试验,结果显示:当浸出温度为25 ℃、固液比为1∶15、浸出时间为20 min、搅拌速率为800 r/min、硫酸体积分数为30%时,铬的浸出率最高.最后用黄钠铁矾法除铁,用焦亚硫酸钠还原六价铬,用氢氧化钠分步沉淀铬、镍重金属,锌则继续留在溶液中.电镀污泥的浸铬实验的浸出动力学研究结果表明硫酸作为浸出剂的反应级数为1,反应的速率常数为:k=0.053 2e-4.52/RT.     

FeSO_4-ZnSO_4-H_2O体系中水热赤铁矿法沉铁的研究

针对高铁闪锌矿的矿物学特点开发的还原浸出-置换沉铜-中和沉铟-水热赤铁矿法沉铁新工艺可实现该矿物中有价元素的高效浸出与回收,并获得环境友好型赤铁矿法沉铁渣.FeSO_4-ZnSO_4-H_2O体系中Fe(Ⅱ)氧化水热水解赤铁矿法沉铁研究结果表明:温度是影响亚稳态铁矾物相形成和转化的关键因素,升高温度亚稳态铁矾物相的热力学稳定性降低并转变为赤铁矿;增加Zn~(2+)浓度、提高氧分压、添加晶种、增大搅拌转速均会促进Fe(Ⅱ)氧化水热水解为赤铁矿.在反应温度为190℃、锌离子浓度为100 g/L、晶种添加量为20 g/L、反应时间为4 h、氧分压为0.6 MPa、搅拌速度为500 r/min的优化工艺条件下,Fe(Ⅱ)氧化水热水解赤铁矿法沉铁过程中除铁率为95%,获得含铁58.3%大颗粒的纯净沉铁渣.     

废高磁合金钢中钴,镍的分离和利用

用硫酸，盐酸和硝酸溶解废高磁合金钢，并将Ｆｅ＾２＋氧化为Ｆｅ＾３＋，先用黄铁矾法除去大部分铁，再用尿素除去少量的铁及铝，钛，铜，最后在ＮＨ３－ＮＨ４Ｃｌ体系中分离钴，镍，并制成相应的盐，钴，镍的回收率分别为８１．５％，８９．７％。     

AC法处理高锑低银类铅阳极泥——铜和铋的回收

用AC法处理高锑低银类铅阳极泥，其干馏渣水浸液经两段置换、硫酸浸铜、稀盐酸浸铋，综合回收了铜、铋、锑，并使其得到了较大程度的分离，Cu，Bi和Sb的置换率分别为99．75％，96．74％和99．45％；置换渣含铜53．73％，含铋20．79％．用硫酸浸铜法，实现了铜-锑、铜-铋的有效分离，最终铜以硫酸铜产出，品位为93．74％～96．21％，Fe含量为1．13％～1．47％，回收率为93．33％；用稀盐酸浸出铋-锑渣，铋以含Bi69．70％的铋精矿产出，直收率及总回收率分别为90．87％和94．73％，此外还产出Sb含量为36．21％的锑渣，返回氯化浸出过程，总回收率为94．06％．     

Technological conditions and kinetics of leaching copper from complex copper oxide ore

The kinetic behavior of leaching copper from low grade copper oxide ore was investigated. The effects of leaching temperature, H₂SO₄ concentration, particle size of crude ore and agitation rate on the leaching efficiency of copper were also evaluated. And the kinetic equations of the leaching process were obtained. The results show that the leaching process can be described with a reaction model of shrinking core. The reaction can be divided into three stages. The first stage is the dissolution of free copper oxide and copper oxide wrapped by hematite-limonite ore. At this stage, the leaching efficiency is very fast (leaching efficiency is larger than 60%). The second stage is the leaching of diffluent copper oxides, whose apparent activation energy is 43.26 kJ/mol. During this process, the chemical reaction is the control step, and the reaction order of H₂SO₄ is 0.433 84. The third stage is the leaching of copper oxide wrapped by hematite-limonite and silicate ore with apparent activation energy of 16.08 kJ/mol, which belongs to the mixed control.     

Extracting indium and preparing ferric oxide for soft magnetic ferrite materials from zinc calcine reduction lixivium

A hydrometallurgical process for indium extraction and ferric oxide powder preparation for soft magnetic ferrite material was developed.Using reduction lixivium from high-acid reductive leaching of zinc oxide calcine as raw solution,copper and indium were firstly recovered by iron powder cementation and neutralization.The recovery ratios of Cu and In are 99% and 95%,respectively.Some harmful impurities that have negative influences on magnetic properties of soft magnetic ferrite material are deeply removed ...     

Techniques of copper recovery from Mexican copper oxide ore

Mexican copper ore is a mixed ore containing mainly copper oxide and some copper sulfide that responds well to flotation. The joint techniques of flotation and leaching were studied. The results indicate that an ore containing 19.01% copper could be obtained at a recovery ratio of 35.02% by using sodium sulfide and butyl xanthate flotation. Over 83.33% of the copper oxide can be recovered from the tailings by leaching in suitable conditions, such as 1 h stirring at a temperature around 25 ℃ with a mixing speed of 500 r/min, an H2SO4 concentration of 1.0 mol/L and a mass ratio of the ore-slurry-liquid to solid (mL/mS) of 3. The overall yield of refined ore after flotation and leaching is over 89.18% of the copper, which is much better than sole flotation or leaching. A copper product containing more than 99.9% copper was obtained by using the process: flotation-agitation leaching-solvent extraction-electro-winning.     

Molybdenum removal from copper ore concentrate by sodium hypochlorite leaching

The removal of molybdenum from a copper ore concentrate by sodium hypochlorite leaching was investigated. The results show that leaching time, liquid to solid ratio, leaching ternperature, agitation speed, and sodium hypochlorite and sodium hydroxide concentrations all have a significant effect on the removal of molybdenum. The optimum process operating parameters were found to be: time, 4 h: sodium hydroxide concentration, 10%; sodium hypochlorite concentration, 8%; liquid to solid ratio, 10:1; temperature, 50℃; and,agitation speed, 500 r/min. Under these conditions the extraction of molybdenum is greater than 99.9% and the extraction of copper is less than 0.01%. A shrinking particle model could be used to describe the leaching process. The apparent activation energy of the dissolution reaction was found to be approximately 8.8 kJ/mol.     

置换—转化法综合治理含铜废水的研究

本文在研究含铜废水综合治理方法时着重研究了铁的置换作用,发现在置换过程中存在着一个有意义的转变点。在该转变点前铜的去除率随置换时间增加迅速增加,以后则增加缓慢。同时还表明、在废水pH值为2～5、含铁总量<1000mg/L时进行置换处理,对铜的去除率没有多大影响。单一采用铁屑置换法除铜,出水铜含量往往偏高;用置换—转化法综合治理,废水中大部分铜在置换中以海绵铜回收,小部分铜在转化中以硫化铜回收,总回收率>99％。出。水铜含量<0.5mg/L,可以回用。     

REDUCTION OF FERRIC IRON IN THE TITANIUM SULFATE SOLUTION BY THE ION EXCHANGE MEMBRANE PRIMARY CELL METHOD *

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浸铜渣中提银的研究

用严充酸和硫化硫酸盐的溶液浸出浸铜渣回收银,研究了影响浸出过程的主要因素,讨论了堆浸的作用问题。对于含银２００ｇ／ｔ的浸铜渣,银的回收率达８０％以上。     

Reduction of frerric iron in the titanium sulfate solution by the ion-exchange membrane primary cell method

In the production process of titanium dioxide with sulfuric acid, the contamination of the titanium sulfate solution (the ilmenite leaching solution) in the Fe³⁺ reduction stage by iron scraps is a practical problem because it is difficult to guarantee the quality of the iron scraps. In this research, a new method, called the ion-exchange membrane primary cell method, for reduction of Fe³⁺ in the titanium sulfate solution has been advanced. The positive compartment of the primary cell consists of lead (copper) electrode and the titanium sulfate solution, and the negative compartment consists of iron electrode and acidic FeSO₄ solution. The anion ion-exchange membrane is used as the diaphragm between two compartments. Fe³⁺ in the titanium sulfate solution is reduced by the electric discharge of the primary cell. The effects of temperature, stirring strength of the solution and membrane area on the reduction rate have been investigated. The experimental result shows that the optimum current density can be higher than 100 A/m².     

难处理高结合率氧化铜矿选冶联合工艺研究

采用选冶联合工艺对含铜1.53％、氧化率47.06％、结合率21.57％的高结合率氧化铜矿进行回收.原矿的砂光片分析结果表明,矿石中大部分铜矿物嵌布粒度极细,多呈星点状和不均匀浸染状分布,与硅、钙、镁、铝等脉石共生严重,导致浮游性较差.针对该矿石的特点,研究了工艺参数及流程结构对指标的影响,确定了“三次粗选—粗精矿再磨—三次精选”的硫化浮选工艺流程,获得了含铜品位为23.43％、回收率为53.72％的铜精矿.对尾矿的形貌及矿物组成表征发现：铜矿物呈细粒浸染状或被硅酸盐矿物包裹,导致这部分铜损失在尾矿中.在最佳的酸浸工艺条件下,对浮选尾矿进行酸浸试验,获得了相对原矿的浸出率为33.21％的试验指标;铜综合回收率为86.93％.     

Investigation on steelmaking dust recycling and iron oxide red preparing

To investigate the physical and chemical properties of the steelmaking dust, wet sieve separation, XRD, SEM, EDS, and traditional chemical analysis were carded out to obtain the particle size distribution, mineralogy, morphology, and the chemical composition of the dust. The dust with a total Fe content of 64.08wt% has coarse metallic iron, magnetite and hematite grains, while free clay minerals with a size of 〈38 μm are mainly iosidefite, calcium silicate, and calcite, which are conglomerated to each other. By following the procedures of wet magnetic separation, acid leaching, and oxidization calcination, magnetic materials were recycled and further prepared as iron oxide red with a productivity of 0.54 ton per unit ton of the dust. Middle iron concentrate with an Fe content of 65.92wt% can be reused as feeding material in the ironmaking industry. Additionally, washed water from acid leaching with an Fe^3＋ ion content of less than 5 g·L^-1 was recovered as feeding water in the wet magnetic separation procedure. 2008 University of Science and Technology Beijing. All fights reserved.