采用软锰矿强化辉钼矿的氧化分解及联产钼酸铵与硫酸锰
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摘要
基于软锰矿的强氧化性、辉钼矿的还原性和硫酸锰的稳定性,通过氧化-还原反应的耦合,强化辉钼矿的氧化分解过程,同时实现软锰矿的还原分解,联产钼酸铵和硫酸锰,发展出了辉钼矿与软锰矿联产新工艺。
进行了辉钼矿与软锰矿联合浸出反应的热力学分析,绘制了MoS2-Mn02-H20系的E-pH图,研究结果显示当电势为0.4v-1.0v、pH值小于3,则Mo.Mn可被同时浸出并分别以H2MoO4(MoO3·H20).Mn2+形式存在：
MOS2+9Mn02+16H+→H2MoO4+9Mn2++2HS04-+6H20
考察了物料配比、矿浆液固比L/S、硫酸用量以及反应时间与温度对辉钼矿与软锰矿联合浸出过程的影响,发现反应温度与酸用量的影响较大,在软锰矿用量为理论量的1.2倍(nMnO2:nMos2=10.8).L/S为4、硫酸浓度为500g/L、温度为95℃的条件下反应3h,Mo浸出率可以达到85.8%；继续增大软锰矿用量到理论用量的2倍(nMnO2: nMoS2=18.0),Mo浸出率仅增加了1.2%。单纯依靠增大硫酸、软锰矿的用量不能取得较好的浸出效果,而且酸耗、浸出渣量会随之显著增加,锰回收率急剧下降,导致后续除杂难度与药剂消耗增加。
研究了高锰酸盐、Mn3+/Mn2+氧化还原电对等氧化分解辉钼矿的过程,结果表明NaMn04的氧化效果明显优于软锰矿的,同条件下Mo浸出率可以达到98%以上。尽管Mn3+/Mn2+间接电氧化工艺在强酸介质中可以氧化分解辉钼矿,锰盐可以循环利用,但Mo浸出率只有88%左右。
为了促进两矿联合浸出,采用高锰酸钠协同软锰矿氧化分解辉钼矿,通过正交试验设计考察了各因素的影响规律,结果表明高锰酸钠的加入有效提高了Mo的浸出率,控制物料配比nNaMn04:nMn02: nMoS2=0.36:10.8:1.0.温度为95℃、时间为240min、搅拌速度为400rpm、液固比L/S为4、硫酸浓度为400g/L,Mn.Mo浸出率分别提高到了83.2%与98.8%。浸出液采用N235溶剂萃取技术分离Mo.Mn,分别得到钼酸铵溶液(反萃液)与硫酸锰溶液(萃余液),然后沿用现行冶金工艺流程制得了合格的钼酸铵和硫酸锰产品,Mo.Mn的最终回收率分别达到了92.8%、72.5%(不计加入的高锰酸钠的量)。
高锰酸钠协同软锰矿浸出辉钼矿过程中为了保证Mo的浸出效果,软锰矿用量很大,占到所有固体原料的90%,导致了浸出液中Mo含量与Mn回收率低(计入高锰酸盐加入量,Mn回收率不到70%)。另外,为了强化辉钼矿的氧化效果,加入了大量的高锰酸盐,用量达到辉钼矿用量的1/4倍之多,造成原料成本直线上升。
辉钼矿与软锰矿联合浸出新工艺从实践上证实了软锰矿的强氧化性,以及两矿联产的可行性。为了克服该工艺的上述弊端,采用焙烧方式来实现两矿分解反应的耦合。首先对辉钼矿与软锰矿联合焙烧体系进行了热力学分析,结果表明软锰矿可以通过固-固反应直接氧化分解MoS2、促进MoO2的进一步氧化、吸收转化S02来强化辉钼矿的氧化分解：MoS2+9Mn02=2MnS04+MoO3+7MnO (△rGθ(T)=-0.189T-611.5 kJ-mol-1) Mn02+MoO2=MnO+MoO3 (△,Gθ(T)=-0.104T-1.906 kJ-mol-1) Mn02+S02=MnS04 (△,Gθ(T)=0.184T-246.5 kJ·mol-1)
研究了焙烧温度、时间、物料配比及空气流量等因素对两矿联合焙烧过程中Mo有效分解率(ηMo)与固硫率(ηs)的影响规律,结果显示采用分段升温焙烧方式(450℃×4h+550℃×4h),控制空气流量为20L/h、物料配比nMnO2:nMoS2=4.0,ηMo、ηs可分别达到97.5%和90.6%。软锰矿的加入不仅强化了MoS2的氧化分解,而且获得了较好的固硫效果,同时实现了自身的还原分解。利用热重分析、XRD分析等揭示了两矿联合焙烧过程的反应机理：MoS2+7/202→MoO3+2S02 MoS2+302→MoO2+2S02 9Mn02+MoS2→MnMoO4+6MnO+2MnS04 Mn02+S02→MnS04 Mn02+Mo02→MnMoO4
考察了辉钼矿与软锰矿联合焙烧过程中Re的行为,发现软锰矿的加入强化了Re在焙砂中的富集,两矿联合焙烧工艺中铼在焙砂中的回收率要比传统焙烧工艺的高出33.4%,可以达到64.3%。
采用硫酸浸出含锰钼焙砂,得到的浸出液采用N235萃取回收Mo,负载有机相采用70g/L的硫酸溶液洗涤后再用氨水反萃,Mo的萃取率、反萃率分别达到99.2%与99.1%；反萃液经蒸发、结晶等后续处理可获得八钼酸铵产品,Mo最终回收率为93.9%。含Mn萃余液经黄钠铁钒法除铁、硫化沉淀法去除重金属、净化液蒸发结晶等,制备出了合格的硫酸锰产品(MnSO4·4H2O),Mn的最终回收率达到了83.1%。为了提高Mn产品的附加值和减轻除杂难度,以含Mn萃余液为原料,采用碱式氧化法制备了纯度大于99%的合格化学二氧化锰(CMD)产品,锰转化率大于80%,产品晶型以α-MnO2、γ-MnO2为主。
利用焙烧耦合辉钼矿氧化分解反应与软锰矿还原分解反应,采用酸浸、N235溶剂萃取技术实现了含钼锰焙砂中Mo、Mn的高效浸出与直接分离,获得了合格的Mo、Mn产品,实现了钼酸铵和硫酸锰的联产,发展出了辉钼矿与软锰矿联合焙烧新工艺。对辉钼矿与软锰矿联合新工艺的技术经济进行了分析,与传统焙烧-氨浸工艺相比,新工艺焙烧温度要低50℃以上,焙烧时间也只有传统工艺的一半左右,而Mo的最终回收率提高了3.9%。新工艺在生产一吨钼酸铵的同时多产出硫酸锰(MnSO4·H2O)3.90t,产值要多出14820元,利润空间更大。另外,联合焙烧新工艺中固硫率达到90%以上,相比现有的钼冶金工艺、锰冶金工艺而言,在同时制备出钼酸铵、硫酸锰两种产品的前提下,两矿联产新工艺的流程短得多,生产过程大大降低了SO2、CO2、CO等废气排放,实现了资源、能源的合理化综合利用,达到了减排、降耗的目的。
Based on the strong oxidation of pyrolusite, the reduction of molybdenite and the stability of manganese sulfate, pyrolusite was used to strengthen the oxidation decomposition process of molybdenite concentrates by the coupling of the oxidation-reduction reaction. Meanwhile the reduction decomposition of pyrolusite and the co-production between ammonium molybdate and manganese sulfate were achieved. A novel process of the co-production technology of molybdenite and pyrolusite was developed.
The thermodynamic analysis on the co-leaching reaction between molybdenite and pyrolusite was studied, and an E-pH diagram of MoS2-MnO2-H2O systems was obtained, which showed that electric potential was varied in the range of 0.4v-1.0v, and pH value was less than 3, while Mo and Mn could be leached simultaneously and existed in H2MoO4 (MoO3-H2O)、Mn2+ form respectively:
MoS2+9MnO2+16H+→H2MoO4+ 9Mn2++ 2HSO4-+6H2O
Effects on molybdenite and pyrolusite co-leaching process were investigated, such as the ratio of materials, liquid to solid ratio of slurry (L/S), the amount of sulfuric acid and the reaction time and temperature. It's found that the reaction temperature and the amount of acid had great influences. Under the conditions of nMnO2: nMoS2=10.8, L/S of 4, sulfuric acid concentration of 500g/L, temperature at 95℃and the reaction time of 3h, the leaching rate of molybdenite could reach 85.8%. With the increasing of pyrolusite dosage to nMnO2: nMoS2=18.0, the leaching rate of molybdenite only increased by 1.2%, which indicated that only depended on increasing the amount of acid and pyrolusite a better leaching results could not obtain, and the acid consumption and leaching slag increased significantly, then the recovery of manganese dropped sharply, which results in the removal difficulty of the impurity and an increasing consumption of reagent in the subsequent procedure.
The oxidation susceptibility of NaMnO4 and Mn3+/Mn2+ redox couple for decomposition of molybdenite concentrates were studied, results showed that the oxidation of NaMnO4 was higher than that of pyrolusite at the same condition, and the leaching rate of Mo could achieve more than 98%. Although Mn3+/Mn2+ had ability to oxidation-decompose of molybdenite and most of manganese ion could be recycled in strong acid medium, the leaching rate of Mo could only reach about 88%.
In order to enhance the molybdenite and pyrolusite co-leaching process, NaMnO4 was chosen to co-dissolve the molybdenite with pyrolusite. Some effects had been investigated through Orthogonal test, results showed that the addition of NaMnO4 enhanced the leaching process of Mo efficiently. Under the control of materials ratio nNaMnO4: nMnO2: nMoS2=0.36:10.8:1.0, reaction temperature of 95℃, reaction time of 240 min, stirring speed of 400 rpm, L/S of 4, sulfuric acid concentration of 400g/L, the leaching rate of Mn and Mo increased to 83.2% and 98.8%, respectively. The Mo and Mn were separated by N235 extraction in the leaching solution, and the qualified products of ammonium molybdate and manganese sulfate were generated by existing metallurgical technology. The ultimate recovery of Mo and Mn was 92.8% and 72.5%, respectively (not including the dosage of NaMnO4 added)
For the sake of ensuring the leaching effect of Mo, a large amount of pyrolusite was used in the novel technology of molybdenite and pyrolusite co-leaching process (co-leaching molybdenite by using pyrolusite and sodium permanganate), which reached to 90% by weight, and resulted in the content of Mo and the recovery of Mn dropping in the leaching solution (including potassium permanganate, the recovery of Mn was less than 70%). In addition, in order to enhance the oxidation effect of pyrolusite, a large amount of NaMnO4 were added, the dosage has reached 1/4 times of molybdenite concentrations, which resulted in a soaring cost of raw materials.
The novel technology of molybdenite and pyrolusite co-leaching process has approved the strong oxidation of pyrolusite as well as the feasibility of the two mineral co-production technologies in practice. In order to overcome its disadvantages, a roasting process has been used to achieve the coupling of the decomposition reaction of molybdenite and pyrolusite. Thermodynamic analysis on the molybdenite and pyrolusite co-roasting system was studied firstly, results indicated that pyrolusite had the ability to enhance the oxidation decomposition of molybdenite through promoting the further oxidation of MoO2 and absorbing SO2 which generated already. Meanwhile the oxidation of MoS2 could realize directly in a solid-solid reaction: MoS2+9MnO2=2MnSO4+MoO3+7MnO (△rGθ(T)=-0.189T-611.5 kJ·mol-1) MnO2+MoO2=MnO+MoO3 (△rGθ(T)=-0.104T-1.906 kJ-mol-1) MnO2+SO2=MnSO4 (△rGθ(T)=0.1847-246.5 KJ·mol-1)
Effects such as roasting temperature, roasting time, material ratio and air flow etc. were investigated to characterize the decomposition rate (ηMo) and the the sulfur-retained ratio (ηs) of molybdenite. Under conditions of subsection calefactive mode (450℃x4h+550℃x4h), material ratio nMnO2: nMoS2= 4:1,ηMo andηs could reach 97.5% and 90.6%, respectively. The results indicated that the addition of pyrolusite in molybdenite roasting process, not only enhanced oxidation of molybdenite decomposition, but also could obtain a better sulfur-retained effect. Meanwhile the reduction decomposition of pyrolusite was achieved, and the resources and energy were used reasonablely in this process. The mechanism of molybdenite and pyrolusite co-roasting process was studied by the TGA analysis and the XRD analysis: MoS2+7/2O2→MoO3+2SO2 MoS2+3O2→MoO2+2SO2 9MnO2+MoS2→MnMoO4+6MnO+2MnSO4 MnO2+SO2→MnSO4 MnO2+MoO2→MnMoO4
Effects on recovery of Re in molybdenite and pyrolusite co-roasting process were investigated. It was found that the addition of pyrolusite strengthened the enrichment of Re in the calcine, and the recovery of Re in the molybdenite and pyrolusite co-roasting process was 64.3%, which was 33.4% higher than that of the traditional roasting process with the same conditions.
The calcine which contained Mo and Mn was leached with sulfuric acid, then Mo was extracted by N235 from the leaching solution. And 70g/L of sulfuric acid solution was used to washing the as-product on organic phase load before ammonia stripping, the extraction rate of Mo and the stripping rate can reach 99.2% and 99.1%, respectively. The ammonium octamolybdate products could be extracted after evaporation, crystallization and other processes from the stripping solution, the overall recovery rate of Mo could achieve to 93.9% in the whole co-roasting process. A research on the properties of raffinate which contained Mn had been studied, a qualified product manganese sulfate (MnSO4·4H2O) was prepared after removaling iron by methods of carphosiderite and natrojarosi, removing heavy metals by sulfide precipitation and purifying the evaporation and crystallization processes, the overall recovery rate of Mn could reach 83.1% in the whole process. In order to increase the value of the Mn as-products and reduce the removal difficulty of the impurity, a chemical manganese dioxide (CMD) was prepared by alkaline oxidation of Mn raffinate, and the ratio of CMD conversion was more than 80%, the main crystalline form in the product wereα-MnO2 andγ-MnO2.
Coupling the processes of the roasting oxidation reduction reaction of molybdenite and the decomposition reaction of pyrolusite, the acid leaching technology and N235 solvent extraction technology were adopt to obtain the efficient leaching and a direct separation of Mo and Mn, which existed in the two mineral co-roasting process. The qualified products of ammonium molybdate and manganese sulfate were gained in this method, and a novel process of the co-roasting technology of molybdenite and pyrolusite was developed. The technical and economic analysis of the new molybdenite and pyrolusite co-leaching process were carried out. Compared to the traditional process of roasting-ammonia leaching, the roasting temperature of the new process was lower 50℃and the roasting time was only about half of the traditional process's, while the overall recovery of Mo increased by 3.9%. In the new process, when one ton of ammonium sulfate was produced, another 3.90 tons of manganese sulfate (MnSO4H2O) yielded at the same time, and the output value was 14,820 Yuan more than the traditional process, which indicated that there has greater profit space. In addition, the sulfur-retained ratio of the new co-roasting process was higher than 90%. Compared to the existing molybdenum and manganese metallurgical process, ammonium molybdate and manganese sulfate were prepared simultaneously on this condition. The procedure of the new process was much shorter than others. The process also reduced the emission of the waste gas SO2, CO2 and CO significantly, which finally made a reasonable use of resources and energy and achieved the purpose of the reduction of emission and consumption.
进行了辉钼矿与软锰矿联合浸出反应的热力学分析,绘制了MoS2-Mn02-H20系的E-pH图,研究结果显示当电势为0.4v-1.0v、pH值小于3,则Mo.Mn可被同时浸出并分别以H2MoO4(MoO3·H20).Mn2+形式存在：
MOS2+9Mn02+16H+→H2MoO4+9Mn2++2HS04-+6H20
考察了物料配比、矿浆液固比L/S、硫酸用量以及反应时间与温度对辉钼矿与软锰矿联合浸出过程的影响,发现反应温度与酸用量的影响较大,在软锰矿用量为理论量的1.2倍(nMnO2:nMos2=10.8).L/S为4、硫酸浓度为500g/L、温度为95℃的条件下反应3h,Mo浸出率可以达到85.8%；继续增大软锰矿用量到理论用量的2倍(nMnO2: nMoS2=18.0),Mo浸出率仅增加了1.2%。单纯依靠增大硫酸、软锰矿的用量不能取得较好的浸出效果,而且酸耗、浸出渣量会随之显著增加,锰回收率急剧下降,导致后续除杂难度与药剂消耗增加。
研究了高锰酸盐、Mn3+/Mn2+氧化还原电对等氧化分解辉钼矿的过程,结果表明NaMn04的氧化效果明显优于软锰矿的,同条件下Mo浸出率可以达到98%以上。尽管Mn3+/Mn2+间接电氧化工艺在强酸介质中可以氧化分解辉钼矿,锰盐可以循环利用,但Mo浸出率只有88%左右。
为了促进两矿联合浸出,采用高锰酸钠协同软锰矿氧化分解辉钼矿,通过正交试验设计考察了各因素的影响规律,结果表明高锰酸钠的加入有效提高了Mo的浸出率,控制物料配比nNaMn04:nMn02: nMoS2=0.36:10.8:1.0.温度为95℃、时间为240min、搅拌速度为400rpm、液固比L/S为4、硫酸浓度为400g/L,Mn.Mo浸出率分别提高到了83.2%与98.8%。浸出液采用N235溶剂萃取技术分离Mo.Mn,分别得到钼酸铵溶液(反萃液)与硫酸锰溶液(萃余液),然后沿用现行冶金工艺流程制得了合格的钼酸铵和硫酸锰产品,Mo.Mn的最终回收率分别达到了92.8%、72.5%(不计加入的高锰酸钠的量)。
高锰酸钠协同软锰矿浸出辉钼矿过程中为了保证Mo的浸出效果,软锰矿用量很大,占到所有固体原料的90%,导致了浸出液中Mo含量与Mn回收率低(计入高锰酸盐加入量,Mn回收率不到70%)。另外,为了强化辉钼矿的氧化效果,加入了大量的高锰酸盐,用量达到辉钼矿用量的1/4倍之多,造成原料成本直线上升。
辉钼矿与软锰矿联合浸出新工艺从实践上证实了软锰矿的强氧化性,以及两矿联产的可行性。为了克服该工艺的上述弊端,采用焙烧方式来实现两矿分解反应的耦合。首先对辉钼矿与软锰矿联合焙烧体系进行了热力学分析,结果表明软锰矿可以通过固-固反应直接氧化分解MoS2、促进MoO2的进一步氧化、吸收转化S02来强化辉钼矿的氧化分解：MoS2+9Mn02=2MnS04+MoO3+7MnO (△rGθ(T)=-0.189T-611.5 kJ-mol-1) Mn02+MoO2=MnO+MoO3 (△,Gθ(T)=-0.104T-1.906 kJ-mol-1) Mn02+S02=MnS04 (△,Gθ(T)=0.184T-246.5 kJ·mol-1)
研究了焙烧温度、时间、物料配比及空气流量等因素对两矿联合焙烧过程中Mo有效分解率(ηMo)与固硫率(ηs)的影响规律,结果显示采用分段升温焙烧方式(450℃×4h+550℃×4h),控制空气流量为20L/h、物料配比nMnO2:nMoS2=4.0,ηMo、ηs可分别达到97.5%和90.6%。软锰矿的加入不仅强化了MoS2的氧化分解,而且获得了较好的固硫效果,同时实现了自身的还原分解。利用热重分析、XRD分析等揭示了两矿联合焙烧过程的反应机理：MoS2+7/202→MoO3+2S02 MoS2+302→MoO2+2S02 9Mn02+MoS2→MnMoO4+6MnO+2MnS04 Mn02+S02→MnS04 Mn02+Mo02→MnMoO4
考察了辉钼矿与软锰矿联合焙烧过程中Re的行为,发现软锰矿的加入强化了Re在焙砂中的富集,两矿联合焙烧工艺中铼在焙砂中的回收率要比传统焙烧工艺的高出33.4%,可以达到64.3%。
采用硫酸浸出含锰钼焙砂,得到的浸出液采用N235萃取回收Mo,负载有机相采用70g/L的硫酸溶液洗涤后再用氨水反萃,Mo的萃取率、反萃率分别达到99.2%与99.1%；反萃液经蒸发、结晶等后续处理可获得八钼酸铵产品,Mo最终回收率为93.9%。含Mn萃余液经黄钠铁钒法除铁、硫化沉淀法去除重金属、净化液蒸发结晶等,制备出了合格的硫酸锰产品(MnSO4·4H2O),Mn的最终回收率达到了83.1%。为了提高Mn产品的附加值和减轻除杂难度,以含Mn萃余液为原料,采用碱式氧化法制备了纯度大于99%的合格化学二氧化锰(CMD)产品,锰转化率大于80%,产品晶型以α-MnO2、γ-MnO2为主。
利用焙烧耦合辉钼矿氧化分解反应与软锰矿还原分解反应,采用酸浸、N235溶剂萃取技术实现了含钼锰焙砂中Mo、Mn的高效浸出与直接分离,获得了合格的Mo、Mn产品,实现了钼酸铵和硫酸锰的联产,发展出了辉钼矿与软锰矿联合焙烧新工艺。对辉钼矿与软锰矿联合新工艺的技术经济进行了分析,与传统焙烧-氨浸工艺相比,新工艺焙烧温度要低50℃以上,焙烧时间也只有传统工艺的一半左右,而Mo的最终回收率提高了3.9%。新工艺在生产一吨钼酸铵的同时多产出硫酸锰(MnSO4·H2O)3.90t,产值要多出14820元,利润空间更大。另外,联合焙烧新工艺中固硫率达到90%以上,相比现有的钼冶金工艺、锰冶金工艺而言,在同时制备出钼酸铵、硫酸锰两种产品的前提下,两矿联产新工艺的流程短得多,生产过程大大降低了SO2、CO2、CO等废气排放,实现了资源、能源的合理化综合利用,达到了减排、降耗的目的。
Based on the strong oxidation of pyrolusite, the reduction of molybdenite and the stability of manganese sulfate, pyrolusite was used to strengthen the oxidation decomposition process of molybdenite concentrates by the coupling of the oxidation-reduction reaction. Meanwhile the reduction decomposition of pyrolusite and the co-production between ammonium molybdate and manganese sulfate were achieved. A novel process of the co-production technology of molybdenite and pyrolusite was developed.
The thermodynamic analysis on the co-leaching reaction between molybdenite and pyrolusite was studied, and an E-pH diagram of MoS2-MnO2-H2O systems was obtained, which showed that electric potential was varied in the range of 0.4v-1.0v, and pH value was less than 3, while Mo and Mn could be leached simultaneously and existed in H2MoO4 (MoO3-H2O)、Mn2+ form respectively:
MoS2+9MnO2+16H+→H2MoO4+ 9Mn2++ 2HSO4-+6H2O
Effects on molybdenite and pyrolusite co-leaching process were investigated, such as the ratio of materials, liquid to solid ratio of slurry (L/S), the amount of sulfuric acid and the reaction time and temperature. It's found that the reaction temperature and the amount of acid had great influences. Under the conditions of nMnO2: nMoS2=10.8, L/S of 4, sulfuric acid concentration of 500g/L, temperature at 95℃and the reaction time of 3h, the leaching rate of molybdenite could reach 85.8%. With the increasing of pyrolusite dosage to nMnO2: nMoS2=18.0, the leaching rate of molybdenite only increased by 1.2%, which indicated that only depended on increasing the amount of acid and pyrolusite a better leaching results could not obtain, and the acid consumption and leaching slag increased significantly, then the recovery of manganese dropped sharply, which results in the removal difficulty of the impurity and an increasing consumption of reagent in the subsequent procedure.
The oxidation susceptibility of NaMnO4 and Mn3+/Mn2+ redox couple for decomposition of molybdenite concentrates were studied, results showed that the oxidation of NaMnO4 was higher than that of pyrolusite at the same condition, and the leaching rate of Mo could achieve more than 98%. Although Mn3+/Mn2+ had ability to oxidation-decompose of molybdenite and most of manganese ion could be recycled in strong acid medium, the leaching rate of Mo could only reach about 88%.
In order to enhance the molybdenite and pyrolusite co-leaching process, NaMnO4 was chosen to co-dissolve the molybdenite with pyrolusite. Some effects had been investigated through Orthogonal test, results showed that the addition of NaMnO4 enhanced the leaching process of Mo efficiently. Under the control of materials ratio nNaMnO4: nMnO2: nMoS2=0.36:10.8:1.0, reaction temperature of 95℃, reaction time of 240 min, stirring speed of 400 rpm, L/S of 4, sulfuric acid concentration of 400g/L, the leaching rate of Mn and Mo increased to 83.2% and 98.8%, respectively. The Mo and Mn were separated by N235 extraction in the leaching solution, and the qualified products of ammonium molybdate and manganese sulfate were generated by existing metallurgical technology. The ultimate recovery of Mo and Mn was 92.8% and 72.5%, respectively (not including the dosage of NaMnO4 added)
For the sake of ensuring the leaching effect of Mo, a large amount of pyrolusite was used in the novel technology of molybdenite and pyrolusite co-leaching process (co-leaching molybdenite by using pyrolusite and sodium permanganate), which reached to 90% by weight, and resulted in the content of Mo and the recovery of Mn dropping in the leaching solution (including potassium permanganate, the recovery of Mn was less than 70%). In addition, in order to enhance the oxidation effect of pyrolusite, a large amount of NaMnO4 were added, the dosage has reached 1/4 times of molybdenite concentrations, which resulted in a soaring cost of raw materials.
The novel technology of molybdenite and pyrolusite co-leaching process has approved the strong oxidation of pyrolusite as well as the feasibility of the two mineral co-production technologies in practice. In order to overcome its disadvantages, a roasting process has been used to achieve the coupling of the decomposition reaction of molybdenite and pyrolusite. Thermodynamic analysis on the molybdenite and pyrolusite co-roasting system was studied firstly, results indicated that pyrolusite had the ability to enhance the oxidation decomposition of molybdenite through promoting the further oxidation of MoO2 and absorbing SO2 which generated already. Meanwhile the oxidation of MoS2 could realize directly in a solid-solid reaction: MoS2+9MnO2=2MnSO4+MoO3+7MnO (△rGθ(T)=-0.189T-611.5 kJ·mol-1) MnO2+MoO2=MnO+MoO3 (△rGθ(T)=-0.104T-1.906 kJ-mol-1) MnO2+SO2=MnSO4 (△rGθ(T)=0.1847-246.5 KJ·mol-1)
Effects such as roasting temperature, roasting time, material ratio and air flow etc. were investigated to characterize the decomposition rate (ηMo) and the the sulfur-retained ratio (ηs) of molybdenite. Under conditions of subsection calefactive mode (450℃x4h+550℃x4h), material ratio nMnO2: nMoS2= 4:1,ηMo andηs could reach 97.5% and 90.6%, respectively. The results indicated that the addition of pyrolusite in molybdenite roasting process, not only enhanced oxidation of molybdenite decomposition, but also could obtain a better sulfur-retained effect. Meanwhile the reduction decomposition of pyrolusite was achieved, and the resources and energy were used reasonablely in this process. The mechanism of molybdenite and pyrolusite co-roasting process was studied by the TGA analysis and the XRD analysis: MoS2+7/2O2→MoO3+2SO2 MoS2+3O2→MoO2+2SO2 9MnO2+MoS2→MnMoO4+6MnO+2MnSO4 MnO2+SO2→MnSO4 MnO2+MoO2→MnMoO4
Effects on recovery of Re in molybdenite and pyrolusite co-roasting process were investigated. It was found that the addition of pyrolusite strengthened the enrichment of Re in the calcine, and the recovery of Re in the molybdenite and pyrolusite co-roasting process was 64.3%, which was 33.4% higher than that of the traditional roasting process with the same conditions.
The calcine which contained Mo and Mn was leached with sulfuric acid, then Mo was extracted by N235 from the leaching solution. And 70g/L of sulfuric acid solution was used to washing the as-product on organic phase load before ammonia stripping, the extraction rate of Mo and the stripping rate can reach 99.2% and 99.1%, respectively. The ammonium octamolybdate products could be extracted after evaporation, crystallization and other processes from the stripping solution, the overall recovery rate of Mo could achieve to 93.9% in the whole co-roasting process. A research on the properties of raffinate which contained Mn had been studied, a qualified product manganese sulfate (MnSO4·4H2O) was prepared after removaling iron by methods of carphosiderite and natrojarosi, removing heavy metals by sulfide precipitation and purifying the evaporation and crystallization processes, the overall recovery rate of Mn could reach 83.1% in the whole process. In order to increase the value of the Mn as-products and reduce the removal difficulty of the impurity, a chemical manganese dioxide (CMD) was prepared by alkaline oxidation of Mn raffinate, and the ratio of CMD conversion was more than 80%, the main crystalline form in the product wereα-MnO2 andγ-MnO2.
Coupling the processes of the roasting oxidation reduction reaction of molybdenite and the decomposition reaction of pyrolusite, the acid leaching technology and N235 solvent extraction technology were adopt to obtain the efficient leaching and a direct separation of Mo and Mn, which existed in the two mineral co-roasting process. The qualified products of ammonium molybdate and manganese sulfate were gained in this method, and a novel process of the co-roasting technology of molybdenite and pyrolusite was developed. The technical and economic analysis of the new molybdenite and pyrolusite co-leaching process were carried out. Compared to the traditional process of roasting-ammonia leaching, the roasting temperature of the new process was lower 50℃and the roasting time was only about half of the traditional process's, while the overall recovery of Mo increased by 3.9%. In the new process, when one ton of ammonium sulfate was produced, another 3.90 tons of manganese sulfate (MnSO4H2O) yielded at the same time, and the output value was 14,820 Yuan more than the traditional process, which indicated that there has greater profit space. In addition, the sulfur-retained ratio of the new co-roasting process was higher than 90%. Compared to the existing molybdenum and manganese metallurgical process, ammonium molybdate and manganese sulfate were prepared simultaneously on this condition. The procedure of the new process was much shorter than others. The process also reduced the emission of the waste gas SO2, CO2 and CO significantly, which finally made a reasonable use of resources and energy and achieved the purpose of the reduction of emission and consumption.
引文
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