高铁铟锌精矿无铁渣湿法炼锌提铟及铁源高值化利用工艺与原理研究
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摘要
本论文将湿法炼锌技术与“直接-共沉淀法”制备锰锌软磁铁氧体工艺相结合，开发无铁渣湿法炼锌提铟清洁生产新工艺和锌精矿铁源直接制备锰锌软磁铁氧体的新技术，实现铁源高值化利用及铁渣、二氧化硫零排放。这对湿法炼锌提铟工艺是一次重大进步，而且对锌矿铁资源制备锰锌软磁铁氧体也是首次尝试。以中性浸出渣为原料，研究了高温高酸还原氧化浸出、D2EHPA萃取提铟、硫化法初步净化、复盐深度脱硅，开路锌探索、复盐转化氨浸、电积锌粉或蒸发回收锌、锰锌铁氧体工艺、硫酸铵回收等过程的工艺和高温高酸浸出过程的理论，得出了一些有意义的结论。
首次研究了中浸渣高温高酸浸出动力学，动力学方程符合收缩核模型。试验结果表明浸出过程分为三个阶段进行：第一阶段(≤15min)浸出速度很快，主要为ZnSO_4等易溶物料的溶解，第二阶段(15～60min)主要为ZnO等易溶于酸的锌物种的浸出，其表观活化能为51.792kJ／mol，化学反应为控制步骤。第三阶段(60～300min)主要为ZnFe_2O_4和ZnS的浸出，其表观活化能为20.720kJ／mol，受化学反应和外扩散混合控制。其宏观动力学方程为：
15～60min:
1-(1-a)~(1/3)=5.48[H~+]~(1.08)[Zn~(2+)]~(-1.76)[Fe~(3+)]~(-1.39)d_0~(-1)exp(-51792/RT)t+A_1
60～300min:
1-(1-a)~(1/3)=7.87×10~(-5)[H~+]~(0.40)[Zn~(2+)]~(-0.38)[Fe~(3+)]~(-0.45)d_0~(-1)exp(-20720/RT)t+A_2
A_1，A_2为常数。
研究了中浸渣中的铟提取工艺，结果表明，以硫化锌精矿为还原剂，高温高酸还原-氧化循环浸出高铁富铟中浸渣，锌、铁、锰的浸出率(％)分别为96.25、86.5及96.76，铟的浸出率保持在96％以上，Fe~(3+)的还原率也提高到94.05％以上；铁屑置换除铜方法可行，除铜率为94.61％，铜渣含Cu高达74.20％，铟回收率99.47％；首次用D2EHPA从含有大量Mn~(2+)、Zn~(2+)的硫酸亚铁溶液中直接萃取铟，缩短提铟流程，使铟、锌回收率大幅提高。在30％D2EHPA+70％磺化煤油、室温(298K)、萃取级数为3级、相比(O／A)=1：3、萃取时间为5min的综合条件下，铟的萃取率和反萃率均≥99％。从铟锌精矿到海绵铟，铟的直收率为94.33％，比现有流程提高25％以上，锌的回收率提高8％以上。
选择比较了复盐水洗法、共沉粉氨浸法、锌萃取法和复盐转化氨
With the combination of zinc hydrometallurgy and directed co-precipitation method for preparing of Mn-Zn soft magnetic ferrite materials, a new clean process for hydrometallurgically extracting zinc, indium without ferrite residue emission and directly preparing Mn-Zn soft magnetic ferrite was developed in this dissertation. The ferrite resource in zinc concentrate got value-added utilization and the emission of ferrite residue and sulfur dioxide was avoided simultaneously. It's not only a great progress in hydrometallurgical extracting of zinc and indium, but also a first attempt to directly prepare Mn-Zn soft magnetic ferrite from ferrite resources in zinc ores. Using neutral leaching residue as raw material, systemic studies on the reductive and oxidative hot leaching with high concentration acid, solvent extraction of indium with D2EHPA, sulphide precipitation for primary purification, complex salt precipitation for deep purification, redundant zinc removing, complex salt transformation and ammonia leaching, producing zinc powder by electrowinning or recoverying zinc by vaporizing, Mn-Zn soft magnetic ferrite process and (NH_4)_2SO_4 recovery were carried out in this dissertation. And some meaningful conclusions were obtained.
The kinetics on the hot leaching of neutral residue with high concentration acid was investigated. The results showed the leaching reactions followed the shrinking core model. The leaching process was consisted of the following three stages distinguished by the leaching time. During the first stage (t≤15min), ZnSO_4 dissolved in the solution and the leaching rate was very fast. During the second stage (15～60min), zinc oxide and other zinc dissoluble species dissolved chiefly, the apparent activation energy was 51.792kJ/mol. The leaching rate was controlled by the chemical reaction. While in the third stage (60～300min), zinc ferrite and zinc sulfide were mainly leached out slowly, the apparent activation energy was 20.72kJ/mol. The leaching reaction was simultaneously controlled by the chemical reaction and extemal diffusion. The following expressions were the kinetic equations of leaching reaction: 15～60min:
1-(1-a)~(1/3)=5.48[H~+]~(1.08)[Zn~(2+)]~(-1.76)[Fe~(3+)]~(-1.39)d_0~(-1)exp(-51792/RT)t+Ai 60~300min:
1-(1-a)~(1/3)=7.87×10~(-5)[H~+]~(0.40)[Zn~(2+)]~(-0.38)[Fe~(3+)]~(-0.45)d_0~(-1)exp(-20720/RT)t+A_2 where A1 and A2 are both constants.
Process of the recovery of indium from neutral leaching residue has been investigated. The zinc sulfide concentrate was used as reductive reagent. The leaching ratio of Zn、Fe、Mn was 96.25%、86.5% and 96.76%, respectively, when neutral leaching residue bearing high iron and indium was circularly reductive and oxidative leached with high concentration acid at high temperature. The leaching ratio of
首次研究了中浸渣高温高酸浸出动力学，动力学方程符合收缩核模型。试验结果表明浸出过程分为三个阶段进行：第一阶段(≤15min)浸出速度很快，主要为ZnSO_4等易溶物料的溶解，第二阶段(15～60min)主要为ZnO等易溶于酸的锌物种的浸出，其表观活化能为51.792kJ／mol，化学反应为控制步骤。第三阶段(60～300min)主要为ZnFe_2O_4和ZnS的浸出，其表观活化能为20.720kJ／mol，受化学反应和外扩散混合控制。其宏观动力学方程为：
15～60min:
1-(1-a)~(1/3)=5.48[H~+]~(1.08)[Zn~(2+)]~(-1.76)[Fe~(3+)]~(-1.39)d_0~(-1)exp(-51792/RT)t+A_1
60～300min:
1-(1-a)~(1/3)=7.87×10~(-5)[H~+]~(0.40)[Zn~(2+)]~(-0.38)[Fe~(3+)]~(-0.45)d_0~(-1)exp(-20720/RT)t+A_2
A_1，A_2为常数。
研究了中浸渣中的铟提取工艺，结果表明，以硫化锌精矿为还原剂，高温高酸还原-氧化循环浸出高铁富铟中浸渣，锌、铁、锰的浸出率(％)分别为96.25、86.5及96.76，铟的浸出率保持在96％以上，Fe~(3+)的还原率也提高到94.05％以上；铁屑置换除铜方法可行，除铜率为94.61％，铜渣含Cu高达74.20％，铟回收率99.47％；首次用D2EHPA从含有大量Mn~(2+)、Zn~(2+)的硫酸亚铁溶液中直接萃取铟，缩短提铟流程，使铟、锌回收率大幅提高。在30％D2EHPA+70％磺化煤油、室温(298K)、萃取级数为3级、相比(O／A)=1：3、萃取时间为5min的综合条件下，铟的萃取率和反萃率均≥99％。从铟锌精矿到海绵铟，铟的直收率为94.33％，比现有流程提高25％以上，锌的回收率提高8％以上。
选择比较了复盐水洗法、共沉粉氨浸法、锌萃取法和复盐转化氨
With the combination of zinc hydrometallurgy and directed co-precipitation method for preparing of Mn-Zn soft magnetic ferrite materials, a new clean process for hydrometallurgically extracting zinc, indium without ferrite residue emission and directly preparing Mn-Zn soft magnetic ferrite was developed in this dissertation. The ferrite resource in zinc concentrate got value-added utilization and the emission of ferrite residue and sulfur dioxide was avoided simultaneously. It's not only a great progress in hydrometallurgical extracting of zinc and indium, but also a first attempt to directly prepare Mn-Zn soft magnetic ferrite from ferrite resources in zinc ores. Using neutral leaching residue as raw material, systemic studies on the reductive and oxidative hot leaching with high concentration acid, solvent extraction of indium with D2EHPA, sulphide precipitation for primary purification, complex salt precipitation for deep purification, redundant zinc removing, complex salt transformation and ammonia leaching, producing zinc powder by electrowinning or recoverying zinc by vaporizing, Mn-Zn soft magnetic ferrite process and (NH_4)_2SO_4 recovery were carried out in this dissertation. And some meaningful conclusions were obtained.
The kinetics on the hot leaching of neutral residue with high concentration acid was investigated. The results showed the leaching reactions followed the shrinking core model. The leaching process was consisted of the following three stages distinguished by the leaching time. During the first stage (t≤15min), ZnSO_4 dissolved in the solution and the leaching rate was very fast. During the second stage (15～60min), zinc oxide and other zinc dissoluble species dissolved chiefly, the apparent activation energy was 51.792kJ/mol. The leaching rate was controlled by the chemical reaction. While in the third stage (60～300min), zinc ferrite and zinc sulfide were mainly leached out slowly, the apparent activation energy was 20.72kJ/mol. The leaching reaction was simultaneously controlled by the chemical reaction and extemal diffusion. The following expressions were the kinetic equations of leaching reaction: 15～60min:
1-(1-a)~(1/3)=5.48[H~+]~(1.08)[Zn~(2+)]~(-1.76)[Fe~(3+)]~(-1.39)d_0~(-1)exp(-51792/RT)t+Ai 60~300min:
1-(1-a)~(1/3)=7.87×10~(-5)[H~+]~(0.40)[Zn~(2+)]~(-0.38)[Fe~(3+)]~(-0.45)d_0~(-1)exp(-20720/RT)t+A_2 where A1 and A2 are both constants.
Process of the recovery of indium from neutral leaching residue has been investigated. The zinc sulfide concentrate was used as reductive reagent. The leaching ratio of Zn、Fe、Mn was 96.25%、86.5% and 96.76%, respectively, when neutral leaching residue bearing high iron and indium was circularly reductive and oxidative leached with high concentration acid at high temperature. The leaching ratio of
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