碱性加压氧化分离硫化铋精矿中钨钼与铋的研究
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摘要
硫化铋精矿中常伴生有钨、钼、银等有价金属,目前工业上主要应用的是沉淀熔炼方法,但在提取铋的过程中钨、钼等金属会发生分散损失,不利于这些金属的综合回收,并且该方法存在能耗高、污染大、后续回收繁琐等问题。为了解决上述问题,提出本课题的研究。
首先在碱性体系下加压氧化浸出硫化铋精矿中钨、钼、硫,得到含钨钼的加压浸出液和铋的加压氧化渣,实现精矿中钨、钼与铋的分离,同时使硫氧化为硫酸钠,解决了硫的污染问题,然后通过离子交换法从加压浸出液中回收钨、钼,并进行了从加压氧化渣中回收铋、银等有价金属的研究,实现了硫化铋精矿中有价金属的高效分离和提取。在上述研究过程中,得到了以下成果和结论。
通过绘制温度分别为25℃、100℃、150℃、200℃的S-H2O、Bi-S-H2O和Fe-S-H2O的电位-pH图分析铋、铁、硫在加压氧化过程中的行为。在pH值较大、电位较高时,溶液中的硫主要以S042-存在；当pH较低且在有氧化剂存在时,Bi2S3将主要氧化为Bi2(SO4)3,而]pH值较高时,Bi2S3将转化为Bi2O3;溶液中的Fe3+在温度为25℃时易水解为FeO·OH铁,高温下主要转化为Fe2O3;温度的增加,将促使氧化反应所需的平衡电位降低,有助于Bi2S3和FeS2氧化反应的发生,硫将以8O42-离子进入溶液中,铋和铁以金属氧化物存在。
通过比较几种碱性体系处理含钨钼硫化铋精矿的方案,选择氢氧化钠体系加压氧化浸出该精矿,确定其最佳条件为：液固比5:1,温度150℃,氢氧化钠浓度130g/L,填充比0.6,转速1000rpm,时间180min和氧分压0.7MPa。在此条件下,渣率基本稳定在82%左右,硫的浸出率达到96%以上,钨和钼的浸出率分别为93%和95%以上,浸出液中钨、铝的浓度分别为1.31g/L和2.76g/L,浸出渣中铋的含量达39.60%,银的含量为265.3g/t,铅和铜的含量分别为1.25%和2.48%,硫、钨和钼几乎全部进入浸出液中,实现了有价金属的初步分离,有利于后续分别回收加压浸出液中的钨、钼和加压氧化渣中的铋、银等金属。
选择大孔阴离子树脂D363进行钨、钼的静态吸附实验,确定溶液pH在6.6至7.5范围内,树脂对钨的吸附较好,并且在溶液pH为6.66时,钨的吸附量达到236.3mg/g干树脂,钨钼分离因数较好；确定溶液pH在2.2至4.6范围,树脂吸附钼吸附性能较好,并且在pH值为3时,树脂对钼的吸附容量为356.4mg/g干树脂。进行了钨、钼的等温吸附实验研究,结果表明树脂相中离子的平衡吸附量与溶液中的离子浓度更符合Freundlich模型,通过热力学计算确定了树脂对钨、钼的吸附可以白发进行,该吸附过程是吸热过程,也是熵值增加的过程,温度升高有利于树脂对钨钼的吸附。
进行了钨的动态吸附和解吸实验,在溶液pH为6.79、温度为25℃、树脂柱高径比为16:1、吸附接触时间为40min时,树脂对钨的穿透吸附容量为19.3mg/ml湿树脂,并且在流出液体积为416ml(树脂体积的26倍)时,钨的吸附率为85%、钨钼分离因数为7.4,实现了钨的有效回收和钨钼的初步分离；浓度为10%的氨水对钨的解吸率为94.6%;进行了铝的动态吸附和解吸实验,在溶液pH为3.02、温度为25℃、树脂柱高径比为16:1、吸附接触时间为40min时,D363树脂对钼的穿透吸附容量为55.2mg/ml湿树脂,流出液体积为384ml(树脂体积的24倍),树脂对钼的饱和吸附容量为124.6mg/ml湿树脂,实现了钼的有效回收：浓度为10%的氨水对钼的解吸率为87.2%。
进行了氟硅酸体系处理加压氧化渣的实验探索,选择不同加压条件下得到的加压氧化渣进行氟硅酸浸出和通过控制终点pH分步水解两种方法不能有效地解决铋与铁在湿法过程中难以分离的问题,最终确定选择还原熔炼法从加压氧化渣中回收铋等有价金属。进行了CO还原氧化铋的热力学分析,并确定选择以FeO-SiO2-CaO为主的三元系渣型。通过实验确定了熔炼的最佳条件为：还原煤加入量为加压浸出渣重量的7%,渣型为CaO/SiO2=0.5、FeO/SiO2=1.5,熔炼温度1300℃,熔炼时间分别为40min。在上述条件下,铋的回收率为99.6%,银的回收率为99.8%,铜和铅的回收率分别为97.0%和97.3%,实现了铋、银、铅、铜的高效回收。
Bismuth sulfide concentrates, often coexists with tungsten, molybdenum, silver and other valuable metals, is mainly treated by the Precipitation Process in practice. However, the scattered losses of tungsten and molybdenum during the extraction process of bismuth are not conducive to the comprehensive recovery of these metals, and this method has some disadvantages of high energy consumption, severe environment contamination and tedious subsequent recovery process. In order to solve the above problem, this research was posed out.
The bismuth sulfide concentrates was leached with pressure oxidation method in alkaline medium to get the leaching solution containing tungsten and molybdenum and the oxidation residue containing bismuth, which achieves the separation of tungsten, molybdenum and bismuth, and convert the sulfide into sodium sulfate simultaneously to avoid the sulfur pollution problem. Then the ion-exchange method was used to recover tungsten and molybdenum from the leaching solution, and the research about the recovery of bismuth, silver and other valuable metals from the oxidation residue was also carried out. The main purpose is to achieve the efficient separation and recovery of valuable metals from bismuth sulfide concentrates. The main research content and conclusion are as follows.
The behavior of bismuth, iron and sulfur were analyzed through the potential-pH diagram of S-H2O, Bi-S-H2O and Fe-S-H2O which are drawn under the temperature of25℃,100℃,150℃,200℃. Under the condition of higher pH value and higher potential value, sulfur exists as SO42-. Under the condition of lower pH value and oxidants existing, Bi2S3can be oxided into Bi2(SO4)3, and as the pH value increasing Bi2S3can be converted into Bi2O3. Fe3+in solution can be converted into FeO—OH under25℃through hydrolysis reaction, and under higher temperature it will be changed into Fe2O3. The increasing temperature will reduce the equilibrium potential for oxidation reaction, which is conducive to the oxidation reaction of Bi2O3and Fe2O3, as a result sulfur enter into solution in the form of SO42-while bismuth and iron exist as metal oxides in the leaching residue.
By comparing several alkaline system to treat the bismuth sulfide concentrate containing tungsten and molybdenum, sodium hydroxide system was selected with the optimum leaching conditions were:liquid to solid ratio5:1, temperature150℃, sodium hydroxide concentration of130g/L, fill ratio0.6, stirring speed10000rpm, time of180min, oxygen partial pressure of0.7MPa. Under this condition, the residue ratio is basically stable around82%, sulfur leaching ratio is more than96%, the extraction ratio of tungsten and molybdenum are93%and95%respectively, the concentration of tungsten and molybdenum in leaching solution are1.31g/L and2.76g/L respectively, and in residue the content of bismuth is39.60%and that of silver is265.3g/t, the content of lead and copper are1.25%and2.48%respectively. Nearly all the sulfur, tungsten and molybdenum enter leaching solution, which achieve the preliminary separation of valuable metals and is conducive to the subsequent recovery of tungsten and molybdenum from leaching solution and that of bismuth and silver from the oxide residue.
A macroporous anion resin D363was selected to make static adsorption of tungsten and molybdenum. The results show that when the pH value of solution is6.6-7.5, D363has a good adsorption property for tungsten and the adsorption capacity reaches to236.3mg/g dry resin when pH is6.66; when the pH value of solution is2.2-4.6, D363shows great adsorption property for molybdenum and the adsorption capacity reaches356.4mg/g dry resin when pH is3. The results of adsorption isotherm experiments demonstrate that the adsorption process is close to the Freundich model, and the process is endothermic, higher temperature is conducive to the adsorption of tungsten and molybdenum by D363resin. Results of the dynamic adsorption and desorption experiment for tungsten show that the penetration adsorption capacity of tungsten by D363resin can reach19.3mg/ml wet resin when the pH of solution is6.79, temperature is25℃, the resin column ratio of height to diameter is16:1, the time for adsorption is40min and the volume of raffinate is416ml(26 times the volume of wet resin). The adsorption of tungsten reaches85%and the distribution ratio of tungsten and molybdenum is7.4, which achieve the goal of effective recovery of tungsten and preliminary separation of tungsten and molybdenum. Tungsten loaded on D363resin can be effectively desorbed by10%ammonia solution with the desorption ratio of94.6%. Results of the dynamic adsorption and desorption experiment for molybdenum show that the penetration adsorption capacity of molybdenum by D363resin can reach55.2mg/ml wet resin when the pH of solution is3.02, temperature is25℃, the resin column ratio of height to diameter is16:1, the time for adsorption is40min, and the volume of raffinate is384ml(24times the volume of wet resin), which make the effective recovery of molybdenum. Molybdenum loaded on D363resin can effectively desorbed by10%ammonia solution with the desorption ratio of87.2%.
Fluosilicic acid system was used to treat the oxidative residue and the optimum condition was determined according to a series of experiment. However, the problem of difficult separation of bismuth and iron during wet process cannot be effectively solved neither by changing the leaching condition to change the composition of oxide residue nor by controlling the final pH value to make hydrolysis reactions. In order to resolve this difficult, the reducing smelting method is selected to recover bismuth and other valuable metals from tne oxide residue. The ternary slag type FeO-SiO2-CaO is selected according to the the compostions of oxidative residue. The corresponding optimum condition is determined as follows:the amount of coal added is7%of the weight of leaching residue, in slag the ratio of CaO to SiO2is0.5and FeO to SiO2is1.5, temperature is1300℃and time is40min. Under this condition the recovery ratio of bismuth, silver, copper and lead are99.6%,99.8%,97.0%and97.3%respectively, which achieve the efficient recovery of bismuth, silver, lead and copper.
首先在碱性体系下加压氧化浸出硫化铋精矿中钨、钼、硫,得到含钨钼的加压浸出液和铋的加压氧化渣,实现精矿中钨、钼与铋的分离,同时使硫氧化为硫酸钠,解决了硫的污染问题,然后通过离子交换法从加压浸出液中回收钨、钼,并进行了从加压氧化渣中回收铋、银等有价金属的研究,实现了硫化铋精矿中有价金属的高效分离和提取。在上述研究过程中,得到了以下成果和结论。
通过绘制温度分别为25℃、100℃、150℃、200℃的S-H2O、Bi-S-H2O和Fe-S-H2O的电位-pH图分析铋、铁、硫在加压氧化过程中的行为。在pH值较大、电位较高时,溶液中的硫主要以S042-存在；当pH较低且在有氧化剂存在时,Bi2S3将主要氧化为Bi2(SO4)3,而]pH值较高时,Bi2S3将转化为Bi2O3;溶液中的Fe3+在温度为25℃时易水解为FeO·OH铁,高温下主要转化为Fe2O3;温度的增加,将促使氧化反应所需的平衡电位降低,有助于Bi2S3和FeS2氧化反应的发生,硫将以8O42-离子进入溶液中,铋和铁以金属氧化物存在。
通过比较几种碱性体系处理含钨钼硫化铋精矿的方案,选择氢氧化钠体系加压氧化浸出该精矿,确定其最佳条件为：液固比5:1,温度150℃,氢氧化钠浓度130g/L,填充比0.6,转速1000rpm,时间180min和氧分压0.7MPa。在此条件下,渣率基本稳定在82%左右,硫的浸出率达到96%以上,钨和钼的浸出率分别为93%和95%以上,浸出液中钨、铝的浓度分别为1.31g/L和2.76g/L,浸出渣中铋的含量达39.60%,银的含量为265.3g/t,铅和铜的含量分别为1.25%和2.48%,硫、钨和钼几乎全部进入浸出液中,实现了有价金属的初步分离,有利于后续分别回收加压浸出液中的钨、钼和加压氧化渣中的铋、银等金属。
选择大孔阴离子树脂D363进行钨、钼的静态吸附实验,确定溶液pH在6.6至7.5范围内,树脂对钨的吸附较好,并且在溶液pH为6.66时,钨的吸附量达到236.3mg/g干树脂,钨钼分离因数较好；确定溶液pH在2.2至4.6范围,树脂吸附钼吸附性能较好,并且在pH值为3时,树脂对钼的吸附容量为356.4mg/g干树脂。进行了钨、钼的等温吸附实验研究,结果表明树脂相中离子的平衡吸附量与溶液中的离子浓度更符合Freundlich模型,通过热力学计算确定了树脂对钨、钼的吸附可以白发进行,该吸附过程是吸热过程,也是熵值增加的过程,温度升高有利于树脂对钨钼的吸附。
进行了钨的动态吸附和解吸实验,在溶液pH为6.79、温度为25℃、树脂柱高径比为16:1、吸附接触时间为40min时,树脂对钨的穿透吸附容量为19.3mg/ml湿树脂,并且在流出液体积为416ml(树脂体积的26倍)时,钨的吸附率为85%、钨钼分离因数为7.4,实现了钨的有效回收和钨钼的初步分离；浓度为10%的氨水对钨的解吸率为94.6%;进行了铝的动态吸附和解吸实验,在溶液pH为3.02、温度为25℃、树脂柱高径比为16:1、吸附接触时间为40min时,D363树脂对钼的穿透吸附容量为55.2mg/ml湿树脂,流出液体积为384ml(树脂体积的24倍),树脂对钼的饱和吸附容量为124.6mg/ml湿树脂,实现了钼的有效回收：浓度为10%的氨水对钼的解吸率为87.2%。
进行了氟硅酸体系处理加压氧化渣的实验探索,选择不同加压条件下得到的加压氧化渣进行氟硅酸浸出和通过控制终点pH分步水解两种方法不能有效地解决铋与铁在湿法过程中难以分离的问题,最终确定选择还原熔炼法从加压氧化渣中回收铋等有价金属。进行了CO还原氧化铋的热力学分析,并确定选择以FeO-SiO2-CaO为主的三元系渣型。通过实验确定了熔炼的最佳条件为：还原煤加入量为加压浸出渣重量的7%,渣型为CaO/SiO2=0.5、FeO/SiO2=1.5,熔炼温度1300℃,熔炼时间分别为40min。在上述条件下,铋的回收率为99.6%,银的回收率为99.8%,铜和铅的回收率分别为97.0%和97.3%,实现了铋、银、铅、铜的高效回收。
Bismuth sulfide concentrates, often coexists with tungsten, molybdenum, silver and other valuable metals, is mainly treated by the Precipitation Process in practice. However, the scattered losses of tungsten and molybdenum during the extraction process of bismuth are not conducive to the comprehensive recovery of these metals, and this method has some disadvantages of high energy consumption, severe environment contamination and tedious subsequent recovery process. In order to solve the above problem, this research was posed out.
The bismuth sulfide concentrates was leached with pressure oxidation method in alkaline medium to get the leaching solution containing tungsten and molybdenum and the oxidation residue containing bismuth, which achieves the separation of tungsten, molybdenum and bismuth, and convert the sulfide into sodium sulfate simultaneously to avoid the sulfur pollution problem. Then the ion-exchange method was used to recover tungsten and molybdenum from the leaching solution, and the research about the recovery of bismuth, silver and other valuable metals from the oxidation residue was also carried out. The main purpose is to achieve the efficient separation and recovery of valuable metals from bismuth sulfide concentrates. The main research content and conclusion are as follows.
The behavior of bismuth, iron and sulfur were analyzed through the potential-pH diagram of S-H2O, Bi-S-H2O and Fe-S-H2O which are drawn under the temperature of25℃,100℃,150℃,200℃. Under the condition of higher pH value and higher potential value, sulfur exists as SO42-. Under the condition of lower pH value and oxidants existing, Bi2S3can be oxided into Bi2(SO4)3, and as the pH value increasing Bi2S3can be converted into Bi2O3. Fe3+in solution can be converted into FeO—OH under25℃through hydrolysis reaction, and under higher temperature it will be changed into Fe2O3. The increasing temperature will reduce the equilibrium potential for oxidation reaction, which is conducive to the oxidation reaction of Bi2O3and Fe2O3, as a result sulfur enter into solution in the form of SO42-while bismuth and iron exist as metal oxides in the leaching residue.
By comparing several alkaline system to treat the bismuth sulfide concentrate containing tungsten and molybdenum, sodium hydroxide system was selected with the optimum leaching conditions were:liquid to solid ratio5:1, temperature150℃, sodium hydroxide concentration of130g/L, fill ratio0.6, stirring speed10000rpm, time of180min, oxygen partial pressure of0.7MPa. Under this condition, the residue ratio is basically stable around82%, sulfur leaching ratio is more than96%, the extraction ratio of tungsten and molybdenum are93%and95%respectively, the concentration of tungsten and molybdenum in leaching solution are1.31g/L and2.76g/L respectively, and in residue the content of bismuth is39.60%and that of silver is265.3g/t, the content of lead and copper are1.25%and2.48%respectively. Nearly all the sulfur, tungsten and molybdenum enter leaching solution, which achieve the preliminary separation of valuable metals and is conducive to the subsequent recovery of tungsten and molybdenum from leaching solution and that of bismuth and silver from the oxide residue.
A macroporous anion resin D363was selected to make static adsorption of tungsten and molybdenum. The results show that when the pH value of solution is6.6-7.5, D363has a good adsorption property for tungsten and the adsorption capacity reaches to236.3mg/g dry resin when pH is6.66; when the pH value of solution is2.2-4.6, D363shows great adsorption property for molybdenum and the adsorption capacity reaches356.4mg/g dry resin when pH is3. The results of adsorption isotherm experiments demonstrate that the adsorption process is close to the Freundich model, and the process is endothermic, higher temperature is conducive to the adsorption of tungsten and molybdenum by D363resin. Results of the dynamic adsorption and desorption experiment for tungsten show that the penetration adsorption capacity of tungsten by D363resin can reach19.3mg/ml wet resin when the pH of solution is6.79, temperature is25℃, the resin column ratio of height to diameter is16:1, the time for adsorption is40min and the volume of raffinate is416ml(26 times the volume of wet resin). The adsorption of tungsten reaches85%and the distribution ratio of tungsten and molybdenum is7.4, which achieve the goal of effective recovery of tungsten and preliminary separation of tungsten and molybdenum. Tungsten loaded on D363resin can be effectively desorbed by10%ammonia solution with the desorption ratio of94.6%. Results of the dynamic adsorption and desorption experiment for molybdenum show that the penetration adsorption capacity of molybdenum by D363resin can reach55.2mg/ml wet resin when the pH of solution is3.02, temperature is25℃, the resin column ratio of height to diameter is16:1, the time for adsorption is40min, and the volume of raffinate is384ml(24times the volume of wet resin), which make the effective recovery of molybdenum. Molybdenum loaded on D363resin can effectively desorbed by10%ammonia solution with the desorption ratio of87.2%.
Fluosilicic acid system was used to treat the oxidative residue and the optimum condition was determined according to a series of experiment. However, the problem of difficult separation of bismuth and iron during wet process cannot be effectively solved neither by changing the leaching condition to change the composition of oxide residue nor by controlling the final pH value to make hydrolysis reactions. In order to resolve this difficult, the reducing smelting method is selected to recover bismuth and other valuable metals from tne oxide residue. The ternary slag type FeO-SiO2-CaO is selected according to the the compostions of oxidative residue. The corresponding optimum condition is determined as follows:the amount of coal added is7%of the weight of leaching residue, in slag the ratio of CaO to SiO2is0.5and FeO to SiO2is1.5, temperature is1300℃and time is40min. Under this condition the recovery ratio of bismuth, silver, copper and lead are99.6%,99.8%,97.0%and97.3%respectively, which achieve the efficient recovery of bismuth, silver, lead and copper.
引文
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