含钴铜转炉渣的工艺矿物学

研究了铜转炉渣的化学与矿物学特征,用光学显微镜查明了炉渣各主要渣相为铁橄榄石,磁铁矿和无定形玻璃体,铜主要以冰铜相裹夹于渣相中,铜滴的尺度与渣的冷却历史有关。用X射线衍射谱,扫描电子显微镜,X射线能谱和X射线波谱及化学物相分析,对铜和钴的赋存状态及在各相中的分布进行了表征和量化。结果表明,钴主要以类质同象形式取代铁橄榄石和磁铁矿晶格中部分Fe^2 ,作为氧化物富集其中,二者约占钴总量的95％,渣中钴的提取需要以铁橄榄石和磁铁矿的分解为前提,使钴的氧化的游离出来而酸溶。     

Leaching of blended copper slag in microwave oven

Leaching of blended slag (BS) was investigated in a microwave oven using hydrogen peroxide and acetic acid. The BS was a mixture of converter and flash furnace slag containing 51% Fe₂O₃, 3.8% CuO, and 3.2% ZnO. The important variables that influence the metal extraction yield were leaching time, liquid-solid ratio, H₂O₂ and CH₃COOH concentrations. The preferred leaching conditions were as follows: CH₃COOH concentration 4 mol/L; H₂O₂ concentration 4 mol/L; microwave power 900 W; leaching time 30 min; liquid-solid ratio 25 mL/g BS; leaching temperature 100 °C. Under these conditions, the metal extractions of 95% Cu, 1.6% Fe, and 30% Zn were obtained. The results were compared with the traditional leaching results. It is evident that microwave heating causes a reduction in the leaching time. Also, the extraction yield results indicate that selective leaching of BS can be achieved under the preferred conditions. The dissolution kinetic of BS in hydrogen peroxide with acetic acid is controlled by a shrinking unreacted core model equation. The apparent activation energy and reaction order were found to be 16.64 kJ/mol and 1.09, respectively.     

Extraction of arsenic from arsenic-containing cobalt and nickel slag and preparation of arsenic-bearing compounds

The arsenic extraction from the arsenic-containing cobalt and nickel slag, which came from the purification process of zinc sulfate solution in a zinc smelting factory, was investigated. The alkaline leaching method was proposed according to the mode of occurrence of arsenic in the slag and its amphoteric characteristic. The leaching experiments were conducted in the alkaline aqueous medium, with bubbling of oxygen into the solution, and the optimal conditions for leaching arsenic were determined. The results showed that the extraction rate of arsenic was maximized at 99.10% under the optimal conditions of temperature 140 °C, NaOH concentration 150 g/L, oxygen partial pressure 0.5 MPa, and a liquid-to-solid ratio 5:1. Based on the solubilities of As₂O₅, ZnO and PbO in NaOH solution at 25 °C, a method for the separation of As in the form of sodium arsenate salt from the arsenic-rich leachate via cooling crystallization was established, and the reaction medium could be fully recycled. The crystallization rate was confirmed to reach 88.9% (calculated on the basis of Na₃AsO₄) upon a direct cooling of the hot leachate down to room temperature. On the basis of redox potentials, the sodium arsenate solution could be further reduced by sulfur dioxide (SO₂) gas to arsenite, at a reduction yield of 92% under the suitable conditions. Arsenic trioxide with regular octahedron shape could be prepared successfully from the reduced solution, and further recycled to the purification process to purify the zinc sulfate solution. Also, sodium arsenite solution obtained after the reduction of arsenate could be directly used to purify the zinc sulfate solution. Therefore, the technical scheme of alkaline leaching with pressured oxygen, cooling crystallization, arsenate reduction by SO₂ gas, and arsenic trioxide preparation, provides an attractive approach to realize the resource utilization of arsenic-containing cobalt and nickel slag.     

加压浸出脱除钴白合金中金属铜和钴的研究

阐述了采用加压浸出方法对金属铜和钴进行有效脱除。实验结果表明：加压浸出过程的较佳反应条件为反应温度150℃、硫酸浓度30 g/L、反应压力1.5 MPa、反应时间6 h、液固比6：1。在此反应条件下,金属Cu、Co的浸出效率分别达到了90％和88％。     

含Ti高炉渣的加压酸解

含Ti高炉渣中的TiO2 含量很高 ,质量分数约为 2 4%左右 ,是宝贵的二次资源。用稀硫酸加压酸解含Ti高炉渣提取其中的Ti,并分析物料粒度、酸浓度、渣酸比、反应温度、反应时间等因素对TiO2 酸解率的影响 ,得出了加压酸解的适宜条件。TiO2 酸解率达到 90 %以上。酸解液水解获得的TiO2 产物纯度为 90 %以上。为开发利用废稀酸提取含Ti高炉渣中的Ti提供了依据。     

Recovery of valuable metals from spent lithium ion batteries by smelting reduction process based on FeO–SiO2–Al2O3 slag system

A novel smelting reduction process based on FeO–SiO₂–Al₂O₃ slag system for spent lithium ion batteries with Al cans was developed, while using copper slag as the only slag former. The feasibility of the process and the mechanism of copper loss in slag were investigated. 98.83% Co, 98.39% Ni and 93.57% Cu were recovered under the optimum conditions of slag former/battery mass ratio of 4.0:1, smelting temperature of 1723 K, and smelting mass ratio of time of 30 min. The FeO–SiO₂–Al₂O₃ slag system for the smelting process is appropriate under the conditions of m(FeO):m(SiO₂)=0.58:1–1.03:1, and 17.19%–21.52% Al₂O₃ content. The obtained alloy was mainly composed of Fe–Co–Cu–Ni solid solution including small amounts of matte. The obtained slag mainly consisted of fayalite and hercynite. Meanwhile, the mechanism of copper loss is the mechanical entrainment from strip-like fayalite particles in the main form of copper sulfide and metallic copper.     

Study of spent battery material leaching process

The recovery of Ni, Co and Mn from spent battery material is very important to environment protection, utilization of resources and cost reduction of the material. The dissolution rates of Ni, Co and Mn with hydrochloric acid as leachant are all over 95% under the optimal conditions of initial hydrochloric acid of 6 mol/L, particle size of 120 μm for the exhausted scraps, molar ratio of H2O2 to MeS of 2, leaching temperature about 60 ℃, ratio of liquid to solid of 8, and leaching time of 2 h. The NixCoyMnz precursor for cathode material prepared from the purified leaching solution, can meet the demand of precursor by pure chemicals. The process is economic and feasible for base metals from spent battery material.     

An investigation of oxygen pressure acid leaching of Gacun complex Cu-Pb bulk concentrate

The treatment of the Gacun complex Cu-Pb bulk concentrate with high Zn,Ag,etc.,by oxygen pressure acid leaching was studied.The primary copper and lead minerals in the concentrate are tetrahedrite and galena.The treatment of tetrahedrite was rarely studied,and most of silver occurred in the mineral too.The optimum operating parameters of oxygen pressure acid leaching were established by conditional tests.Under these parameters,the result of pilot scale test showed that the leaching percentages of copper and zinc were separately as high as 98.9 wt.％ and 94.9 wt.％,while lead and silver were transformed into sulfate and sulfide precipitations,respectively.The copper and zinc in lixivium were reclaimed by extraction-electrowinning and purification-electrowinning,respectively,and the lead and silver in the residue were reclaimed separately by carbonate transformation-silicofluoric acid leaching and thiourea leaching.     

浮选药剂对嗜酸氧化亚铁硫杆菌浸出低品位铜尾矿的影响

以分离于湖北某铜矿的嗜酸氧化亚铁硫杆菌LD-1菌株(At.f LD-1)为研究对象,研究了4种不同碳链长度黄药、丁胺黑药、2号油、11号油单一浮选药剂和由这些药剂组成的5种组合药剂对At.f LD-1菌株生长、氧化活性以及浸出尾矿效率的影响规律。结果表明:单一浮选药剂均会影响At.f LD-1菌的氧化活性,影响程度有所差异,起泡剂对菌种的影响最小,捕收剂的影响较大;5种组合浮选药剂对At.f LD-1菌株生长、氧化活性和浸铜效率的抑制作用从小到大顺序为丁基黄药+丁胺黑药+2号油、异丙基黄药+丁胺黑药+2号油、异戊基黄药+丁胺黑药+2号油、乙基黄药+异戊基黄药+2号油、乙基黄药+丁胺黑药+2号油。     

铜在富FeO熔渣中的氧化溶解行为

在1873 K条件下对铜在富FeO熔渣中的氧化溶解行为进行了研究.结果表明:铜氧化物在富FeO熔渣中以Cu2O的形式存在;熔渣中CaO的含量将对铜的氧化溶解产生影响,随着渣中CaO含量的增加,熔渣中的铜含量逐渐下降,但在不同温度下,这种影响的作用各不相同;当温度较高时,渣中CaO含量对铜的氧化溶解影响较大,且随着温度的降低,其影响作用明显变缓;反应体系的氧分压对铜的氧化溶解影响较大,在熔渣中γ(CuO0.5)不变的情况下,铜氧化物的溶解与体系氧分压的四分之一次方之间呈正线性关系.     

选择性还原-磁选回收镍渣中的有价金属

采用选择性还原-磁选工艺富集某镍渣中的镍、铜,通过控制还原过程参数实现选择性还原。结果表明：添加熔剂并适当提高渣料的碱度（CaO与SiO2质量比）有助于镍、铜的富集;对碱度0.15、还原温度1200℃、还原时间20 min、内配煤量5%（质量分数）的优化条件下得到的还原样品,通过磨矿-磁选获得镍、铜、铁品位分别为3.25%、1.20%、75.26%的精矿,镍、铜、铁的回收率分别为82.20%、80.00%、42.17%,实现了镍、铜相对于铁的选择性富集;选择性还原-磁选没有显著降低S、P的含量,两者在工艺过程中的行为需要进一步研究。     

添加剂氧化钙对镍渣强化还原的影响

针对铁以铁橄榄石形式存在而难以直接还原磁选回收的问题,采用在镍渣中配加CaO提升镍渣碳热还原速率及铁金属化率的技术思路。将添加剂与镍渣按比例混合并配加适量还原剂后,在高温炉中进行还原,结合热力学计算及产物特征分析结果,研究添加剂强化还原的机理及还原效果。结果表明:CaO的加入有效促进了硅酸铁的解离,强化了还原过程,还原速率加快,金属相在渣相中的形态结构发生改变,金属铁的聚集长大明显,有利于后续分离回收。随着CaO添加量由0增加到10%(质量分数),还原产物的金属化率和Fe颗粒平均粒径呈先增大后减小的趋势,FeO含量和残碳量变化规律相反。在CaO添加量为8%时,还原产物中铁的金属化率和粒径达到峰值,分别为87.25%和41.3μm。     

低品位高钛渣中钛组分制备金红石(英文)

介绍了以70%～75%TiO2的低品位高钛渣为原料制备人造金红石的分离工艺。低品位高钛渣中MgO,FeO,CaO,Al2O3和SiO2进入杂质相,钛组分进入金红石相,金红石相中TiO2品位达到90%～95%,可满足流态化氯化对杂质的要求。1050℃的低温预氧化与1510℃的高温热处理促使渣中分散于各矿物相的钛组分选择性转移并富集于金红石相,金红石相析出与长大,用稀硫酸和稀盐酸实现金红石相的分离。实验结果表明,金红石矿物相平均晶粒度可以达到25μm,通过稀酸选择性浸出改性渣,可以获得95%TiO2品位的人造金红石。     

Electrochemistry of oxygen ion transport in slag

1　ＩＮＴＲＯＤＵＣＴＩＯＮＴｈｅｒｅａｃｔｉｏｎｂｅｔｗｅｅｎａｍｅｔａｌｄｒｏｐｌｅｔａｎｄｓｌａｇｉｓａｖｅｒｙｉｎｔｅｒｅｓｔｉｎｇｓｕｂｊｅｃｔｉｎｍｅｔａｌｌｕｒｇｉｃａｌｒｅｓｅａｒｃｈ ,ｓｉｎｃｅｉｔｉｓｄｉｒｅｃｔｌｙｒｅｌａｔｅｄｔｏｉ     

还原熔炼法从谦比希铜冶炼厂转炉渣中回收钴(英文)

研究从赞比亚谦比希铜冶炼厂转炉渣中回收钴的还原熔炼过程。实验考察还原剂用量、熔炼温度、保温时间及渣型改善剂CaO和TiO2的添加对还原熔炼金属回收率的影响。采用X射线衍射、扫描电子显微镜及能谱分析对所得贫化渣和含钴合金进行表征。结果表明,在优化条件下,转炉渣中钴、铜、铁的回收率分别为94.02%,95.76%和小于18%;贫化渣的主要物相组成为铁橄榄石和铁尖晶石,含钴合金中主要含有金属铜、含钴铜的铁合金和少量的硫化物。     

机械活化强化钒渣钙化提钒工艺

针对转炉钒渣钙化焙烧酸浸工艺中存在的钒转浸率低的问题,采用高能球磨对钒渣进行活化预处理,以期强化其提钒效果。采用激光粒度分析仪、BET比表面积测定仪和XRD对活化前后钒渣进行了粒度、比表面积及物相结构分析;采用浸出实验研究了机械活化对钙化焙烧和浸出的影响规律。结果表明:机械活化法增大了钒渣的比表面积,增加了晶格畸变与微观应力,使含钒物相充分解离,由此可改善钒渣钙化焙烧的动力学条件。在浸出20 min条件下,机械活化80 min可将钒浸出率提高10%,最佳焙烧温度降低100℃。     

氟盐对炼镁还原渣中氧化铝浸出率的影响研究

以白云石和菱镁石为原料以铝粉为还原剂真空热还原炼镁过程中添加氟化钙可使镁还原率提高5%以上,还原温度降低50℃,还原后还原渣的主要物相为CaO.2Al2O3,加入的氟化钙在还原过程会参与反应生成氟铝酸钙。在实验室以氢氧化钠和碳酸钠的混合碱液对该含氟盐还原渣中氧化铝的浸出进行了研究,研究结果表明:经碱液浸出后还原渣中的CaO.2Al2O3全部被分解,还原渣中的氧化铝浸出率在70%以上,浸出渣的主要物相为CaCO3。含氟盐炼镁还原中氧化铝的浸出率比不含氟盐的氧化铝浸出率低10%以上,在还原过程中生成的氟铝酸钙和浸出过程中生成的水合铝酸钙是导致氧化铝损失增加的主要原因。     

微波作用下铝酸钙炉渣非等温浸出动力学

研究了微波作用下铝酸钙炉渣非等温浸出动力学，考察了不同微波辐射功率对炉渣中氧化铝浸出率的影响以及相应微波功率下反应体系温度随时间的变化规律。结果表明：该浸出动力学过程的控制步骤为界面化学反应控制，表观反应活化能为40kJ／mol左右；微波作用可改变反应的频率因子，对反应活化能影响不明显；随着微波辐射功率的增大，频率因子增加，宏观表现出微波功率的增加对反应有促进作用。     

炼镍转炉溅渣护炉的水模试验

利用1/4水力学模型试验研究炼镍转炉溅渣工艺参数对炉衬各部位溅渣量的影响。结果表明:溅渣时间和炉体倾角是影响溅渣总量的显著性因素,且与之成正比关系。溅渣量分布受炉体角度和初始熔池深度影响较大,当炉体角度由-10°增至-30°或初始熔池深度(h/D)由0.078增至0.172时,风口对面的溅渣量比例由80%急剧降为5%左右,风口面和端墙面溅渣量相应增大。溅渣高度随着炉体角度和初始熔池深度增加而降低。溅渣模式分为喷溅、渣涌或两者共存。溅渣过程通过调整炉体倾角,可以实现较大的溅渣总量和均匀的分布。工业溅渣试验验证了水模型的研究结果,风口粘结过多等问题得到了解决。