【作者】 夏志华；

【导师】 唐谟堂； 姚维义；

【作者基本信息】 中南大学 ， 有色金属冶金， 2004， 硕士

【摘要】 用无铁渣湿法炼锌新工艺处理广西大厂铟锌精矿，不仅使铟、锌冶炼回收率大幅提高，而且取消除铁过程，消除铁渣对环境的危害，并将精矿中的铁直接加工成高价值的软磁铁氧体材料。本文研究了该方法部分主要过程的工艺和理论问题，得出了一些有意义的结果，对该工艺的完善和应用将具有重要作用。 首先研究了锌焙砂中性浸出渣高温高酸浸出动力学；考察了浸出温度、浸出剂的初始浓度、中浸渣的初始粒度、搅拌强度及金属离子浓度对Zn浸出速率的影响。结果表明，其浸出过程可用收缩核动力学模型描述，该模型较好地说明了浸出机理。 由于该中浸渣除含较多的ZnFe₂O₄外，还富含ZnO等其它锌物种，致使该浸出过程分为三个阶段。第一阶段（≤15min），主要为ZnSO₄及ZnO的溶解，浸出速度很快（浸出率＞35％），第二阶段（15～60min）主要为ZnO等易溶锌物种的浸出，其表观活化能为51.792KJ／mol，属化学反应控制，H⁺、Zn²⁺及Fe³⁺的反应级数分别为1.0771，-1.764和-1.391。第三阶段（60～300min）主要为ZnFe₂O₄及ZnS的溶出，其表观活化能为20.72KJ／mol，属混合控制，H⁺、Zn²⁺及Fe³⁺的反应级数分别为0.3993，-0.3767及-0.4456。 其次研究了硫化锌精矿作还原剂的中浸渣的还原浸出以及以软锰矿作锰源的氧化还原浸出过程，还原浸出优化条件为：硫化锌精矿为理论量的1.03倍，温度为368K，时间为5h，硫化锌精矿分四批加入，始酸酸度为2259/L，规模1009中浸渣/次。氧化浸出优化条件为:时间为sh，温度为368K，软锰矿加入量为259/次，始酸酸度)2509/L。在优化条件下锢、锌、铁的浸出率分别为96%、%.25%及86%，Fe3+的还原率)93%。对铁屑进一步还原Fe3+及置换除铜过程研究结果表明，除铜率为94.61%，产出品位高达74.20%的铜精矿。最后对萃取提锢过程进行了较详细的研究，在有机相组成分为30%PZO4+70%磺化煤油;萃取温度为室温（298K）;萃取级数为3级;相比0/A=3:1;萃取时间为smin及反萃锢条件为:O/A二15:1，6mol.L一’HCI为反萃剂，三级反萃，室温，混合沉清分别为smin的条件下，锢的萃取率和反萃率都为99%。并用锌板置换锢，置换率为99%，从焙砂到海绵锢，锢的总回收率达93%以上，比常规过程提高20%。更多还原

【Abstract】 A new hydrometallurgical process without ferric residue was proposed for treating zinc sulfide concentrate bearing indium from Dachang in Guangxi. Adopting this new process, the recovery ratios of indium and zinc increased greatly. At the same time, this process can also prevent the environmental pollution of ferric residue because of the cancellation of process removing iron. Iron in the zinc sulfide concentrate was used to produce high valuable soft magnetic ferrite materials dierctly. In this paper, the technologies and the theories of the main stages in the new process have been investigated, which is of great significance to the maturity of this new process and its practical application.The first kinetics of leaching the residues from the neutral leaching process of zinc calcine with sulphuric acid of high concentration at high temperature was investigated for examining the effects of reaction temperature, concentration of reagents, the residue granule size, agitation rate and concentration of metal ions on the leaching process. The results show that the leaching process can be simulated with a shrinking core model. Because the residue contains of not only zinc ferrite, but also zinc oxide and other zinc species, the leaching process is composed of three stages distinguished by the leaching time. During the first stage ( 15min),leaching is very fast for solving ZnSO4(s) and part of ZnO. During the second stage (15~60min), zinc oxide and other zinc species are soluted chiefly with the of 51.792KJ/mol. The leaching rate is controlled by the surface chemical reaction .The reaction progression of H+,Zn2+and Fe3+ is 1.0771,-1.764 and -1.391,respectively. While in the third stage (60~300min), zinc ferrite and zinc sulfide were mainly leached out slowly with the activation energy of 20.72KJ/mol, The leaching rate is controlled by the mixed surface chemical reaction and outside diffiision,the reaction progression of H+,Zn2+and Fe3+ is 0.3993, -0.3767,and -0.4456, respectively.On the other hand, The reductive leaching the residue from neutral leaching by using zinc sulfide concentrate as reductive agent has been experimented, to be followed by the oxidative leaching by using pyrolusite as oxidation agent and manganese source. The optimum conditions of the reductive leaching with the scale of 400 g/time are following: the amount of zinc sulfide concentrate to be 1.03 times of the theoretical amount by addition in four batches, temperature of 368K, time of 5 hours, and the begot acidic concentration being 225 g/L. That of the oxidative leaching with the scale of 100 g/time are as: the amount of pyrolusite being 25 g/time, the begot acidic concentration to be more than or equal to 250 g/L, temperature of 368K, and time of 5 hours. Under the optimum conditions, the recovery ratios of indium, zinc and iron were96%, 96.25%, and 86% respectively. The reductive ratio of ferric ions is more than 93%. The results obtained from the experiments of the reduction by iron powder and removing copper show that the removing ratio of copper is 94.61%, the content of copper concentrate produced is higher than 74.20% .At last, the extraction to recover indium was researched in detail. Under the optimum conditions such as the componnent of the organic phase being D2EHPA of 30% and kerosene 70%, extraction temperature of 298K, progressions of 3, ratio of O/A to be 3:1, and extraction time being 5 min, the extractive ratio of indium and its stripping ratio are 99% and 99% respectively. By using zinc plate to cement indium, the displament ratio is 99% . The total recovery ratio of indium is 95%, from zinc calcline to spongy indium, which increased by 25% compared to the routine process.更多还原