WO2017031949A1 - 一种电氧化浸出低品位氧化铅锌矿石的方法及其浸出设备 - Google Patents

一种电氧化浸出低品位氧化铅锌矿石的方法及其浸出设备 Download PDF

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WO2017031949A1
WO2017031949A1 PCT/CN2016/072471 CN2016072471W WO2017031949A1 WO 2017031949 A1 WO2017031949 A1 WO 2017031949A1 CN 2016072471 W CN2016072471 W CN 2016072471W WO 2017031949 A1 WO2017031949 A1 WO 2017031949A1
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lead
leaching
zinc
low
electrolytic cell
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PCT/CN2016/072471
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French (fr)
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袁崎
廖亦农
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四川华富宇科技有限责任公司
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    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/12—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic alkaline solutions
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B19/00—Obtaining zinc or zinc oxide
    • C22B19/20—Obtaining zinc otherwise than by distilling
    • C—CHEMISTRY; METALLURGY
    • C25—ELECTROLYTIC OR ELECTROPHORETIC PROCESSES; APPARATUS THEREFOR
    • C25C—PROCESSES FOR THE ELECTROLYTIC PRODUCTION, RECOVERY OR REFINING OF METALS; APPARATUS THEREFOR
    • C25C1/00—Electrolytic production, recovery or refining of metals by electrolysis of solutions
    • C25C1/18—Electrolytic production, recovery or refining of metals by electrolysis of solutions of lead
    • Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00—Technologies related to metal processing
    • Y02P10/20—Recycling

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  • the invention belongs to the field of ore separation, and particularly relates to an electrooxidation leaching method and a leaching device for extracting lead and zinc in low-grade lead-zinc ore.
  • Low-grade lead-zinc ore is complex in composition, which generally contains sulfide ore, difficult-to-lead heteropolar ore and silicate metal ore. Due to the complex composition, even in the conventional leaching method, even the addition of chemicals It is good for leaching lead and zinc metal, which is also the reason for the lower recovery rate of the conventional leaching process. In the conventional leaching process, it also needs to consume a large amount of heat energy, which is not conducive to energy saving. In conventional production, it is generally necessary to heat the slurry to about 90 ° C, so that it requires a large amount of fuel such as coal, which greatly increases the cost of production.
  • the existing leaching process is difficult to separate lead and zinc after leaching of lead and zinc metal, and it is necessary to rely on a large amount of chemical precipitating agent to react at a certain temperature for a long time.
  • the existing acid leaching method requires a large amount of chemicals to be consumed, and the separation thereof is complicated, and the separation cost for the low-grade oxidized ore is high.
  • the object of the present invention is to overcome the above problems and to provide a method for effectively simplifying the process flow, significantly reducing energy consumption, pollution and cost, and electrooxidizing and leaching low-grade lead-zinc ore which can significantly improve production efficiency.
  • Another object of the present invention is to provide a leaching apparatus for electrooxidative leaching of low grade lead oxide zinc ore.
  • a method for electrooxidizing leaching of low grade lead oxide zinc ore comprising the steps of:
  • Leaching first adding the slurry to the electrolytic cell, and then adding the alkali solution and the activator to the electrolytic cell to make the mass ratio of solid to liquid in the electrolytic cell 4 to 5:1, forming a solid-liquid mixture; then at normal temperature and pressure Leaching is carried out to obtain leachate and sponge lead;
  • the fineness of the lead-zinc oxide powder in the step A is 200 mesh, and the mass ratio of the lead-zinc ore fine powder in the slurry is 25%.
  • the alkali solution described in the step B is sodium carbonate, sodium hydroxide or potassium hydroxide, and the activator is sodium dodecylbenzenesulfonate, and the concentration of the alkali solution in the solid-liquid mixture formed is 25%, the activator The concentration is 1%.
  • the voltage in the electrolytic cell is controlled in the range of 2.5 to 3.5 V, and the current density of the anode is maintained at 100 to 200 A/m 2 .
  • the leaching time is 4 to 6 hours.
  • the "separating and extracting the leachate" described in the step C is specifically to obtain the crude filtrate of the leachate by means of precipitation or centrifugal filtration, and then obtain the zinc powder by electrolyzing the crude filtrate.
  • the regeneration treatment described in the step C includes two steps of chemical neutralization precipitation and calcium hydroxide heating and impurity removal; after the filtrate after the regeneration treatment is returned to the electrolytic tank, the alkali solution and the activator need to be gradually added to the electrolytic tank, so that The concentration of the lye in the slurry was maintained at 25% and the concentration of the activator was maintained at 1%.
  • the "recovering sponge lead into ingot” as described in step D specifically means: collecting and washing the sponge lead formed on the cathode in the electrolytic cell, then compacting the washed sponge lead, and finally passing through the melting furnace. The compacted sponge is fired into a lead ingot.
  • step E The waste residue in step E is washed and filtered until it reaches the standard, and is piled up; after the production wastewater is mixed with the waste residue treatment wastewater, the impurities are precipitated by the neutralization precipitation method, and the wastewater is recovered and reused by adjusting the pH value of the wastewater to 7.8 with sulfuric acid.
  • a leaching apparatus for electrooxidizing a method for leaching low-grade lead-zinc ore is The electrolytic cell has a "ten" shape, the four ends of the electrolytic cell are a cathode region, the center is an anode region, the anode region and the cathode region are spaced apart by a diaphragm, and a stirrer is disposed at a central position of the anode region.
  • a cathode plate and an anode plate are further disposed in the electrolytic cell, a cathode bus bar is connected to the cathode plate, and an anode bus bar is connected to the anode plate.
  • anode plate is disposed at a position adjacent to the diaphragm in the anode region
  • the cathode plate is disposed adjacent to the diaphragm in the cathode region
  • two pairs of cathode plates are disposed in pairs in each cathode region.
  • the process of the present invention is relatively simple, and the lead and zinc can be extracted quickly and easily, the production efficiency is improved, the production cost is lowered, and the leaching rate can be improved by the method of the present invention.
  • the invention does not need to heat the slurry, greatly saves heat energy, reduces energy consumption in the production process, and further saves production cost.
  • the invention can recycle the chemical liquid, thereby reducing the consumption of the chemical liquid in the production process, and further reducing the cost of the production input.
  • the current density of the invention in the electrolysis process is much lower than the conventional 500-2000 A/m 2 , which greatly reduces the cost of equipment manufacturing and the power consumption in the production process, thereby further saving the equipment. Manufacturing and use costs.
  • the invention is based on its rapid efficiency and good effect, has broad industry prospects, can be well promoted and used in the industry, and thus better promotes the progress of the industry and the development of the enterprise.
  • FIG. 1 is a block diagram showing the steps of the present invention.
  • Figure 2 is a plan view of the electrical cell of the present invention.
  • Figure 3 is a side cross-sectional view of the electrical cell of the present invention.
  • the present invention includes the following steps:
  • the fineness of the lead-zinc oxide powder in the step A is 200 mesh, and the mass ratio of the lead-zinc ore fine powder in the slurry is 25%.
  • Leaching first adding the slurry to the electrolytic cell, and then adding the alkali solution and the activator to the electrolytic cell to make the mass ratio of solid to liquid in the electrolytic cell 4 to 5:1, forming a solid-liquid mixture; then at normal temperature and pressure Leaching is carried out to obtain leachate and sponge lead;
  • the alkali solution described in the step B is sodium carbonate, sodium hydroxide or potassium hydroxide, and the activator is sodium dodecylbenzenesulfonate, and the concentration of the alkali solution in the solid-liquid mixture formed is 25%, the activator The concentration is 1%.
  • the voltage in the electrolytic cell is controlled in the range of 2.5 to 3.5 V, and the current density of the anode is maintained at 100 to 200 A/m 2 .
  • the leaching time is 4 to 6 hours.
  • the lead leaching equation is: PbO+2H + ⁇ Pb 2+ +H 2 O; the reaction equation of lead in the electrolysis process is: Pb 2+ +2e - ⁇ Pb; when the alkaline agent is sodium hydroxide, the pulp
  • the "separating and extracting the leachate" described in the step C is specifically to obtain the crude filtrate of the leachate by means of precipitation or centrifugal filtration, and then obtain the zinc powder by electrolyzing the crude filtrate.
  • the regeneration treatment described in the step C includes two steps of chemical neutralization precipitation and calcium hydroxide heating and impurity removal; after the filtrate after the regeneration treatment is returned to the electrolytic tank, the alkali solution and the activator need to be gradually added to the electrolytic tank, so that The concentration of the lye in the slurry was maintained at 25% and the concentration of the activator was maintained at 1%.
  • the process of electrolyzing zinc powder In the cathode device and the anode plate of the electrolysis device, 316 stainless steel plates are used.
  • the "recovering sponge lead into ingot” as described in step D specifically means: collecting and washing the sponge lead formed on the cathode in the electrolytic cell, then compacting the washed sponge lead, and finally passing through the melting furnace. The compacted sponge is fired into a lead ingot.
  • step E The waste residue in step E is washed and filtered until it reaches the standard, and is piled up; after the production wastewater is mixed with the waste residue treatment wastewater, the impurities are precipitated by the neutralization precipitation method, and the wastewater is recovered and reused by adjusting the pH value of the wastewater to 7.8 with sulfuric acid.
  • the neutralization precipitation method lime is used to precipitate with a concentration of 10 to 15% of lime milk. The equation in the precipitation process is:
  • the low-grade lead-zinc ore is crushed and ball-milled to 200 mesh 25% pulp, and the solid-liquid mass ratio is adjusted to 4:1.
  • a voltage of 2.9-3.5 V is applied to the electrolytic cell while controlling.
  • the current density of the anode is 100 A/m 2
  • the speed of the stirrer per minute is controlled in the range of 300 to 500 rpm, and the operation is performed at normal temperature and normal pressure for 4 hours.
  • the final lead leaching rate was 92%, the lead ingot after the ingot was 87.5 to 93.5%, and the zinc powder yield was 95%.
  • the low-grade lead-zinc ore is crushed and ball-milled to 200 mesh 25% pulp, and the solid-liquid mass ratio is adjusted to 4:1.
  • a voltage of 2.9-3.5 V is applied to the electrolytic cell while controlling.
  • the current density of the anode is 100 A/m 2
  • the speed of the stirrer per minute is controlled in the range of 300 to 500 rpm, and the operation is performed at normal temperature and normal pressure for 6 hours.
  • the final lead leaching rate was 94.3%, while the zinc powder yield was 86.7%.
  • a leaching apparatus for electrooxidizing a method for leaching low-grade lead-zinc ore is an electrolytic cell, the electrolytic cell is formed in a "ten" shape, the four ends of the electrolytic cell are a cathode region 3, the center is an anode region 4, and the anode region 4 and the cathode region 3 are spaced apart by a separator 6,
  • a stirrer 5 is disposed at a central position of the anode region 4, and a cathode plate 8 and an anode plate 7 are further disposed in the electrolytic cell, a cathode bus bar 1 is connected to the cathode plate 8, and an anode bus bar 2 is connected to the anode plate 7.
  • the anode plate 7 is disposed at a position adjacent to the separator 6 in the anode region 4, and the cathode plate 8 is disposed at a position adjacent to the separator 6 in the cathode region 3, and two other pairs are disposed in each of the cathode regions 3
  • the cathode plate 8 and the anode plate 7 are made of titanium plate
  • the anode plate is made of titanium plate or 316 stainless steel plate.
  • the present invention can be achieved well by the above method.

Abstract

一种电氧化浸出低品位氧化铅锌矿石的方法，包括以下步骤：A、制浆，将低品位氧化铅锌矿破碎并球磨后加水混合制成矿浆；B、浸出，将矿浆加入电解槽后添加碱液与活化剂形成固液混合物；在常温常压下进行浸出，得到浸出液和海绵铅；C、对浸出液进行分离与提取，得到锌粉，滤液进行再生处理后返回电解槽中循环利用；D、对海绵铅进行回收；E、对废水与废渣进行处理。上述电解槽包括阴极区（3）、阳极区（4）、搅拌器（5）、阴极板（8）以及阳极板（7）。

Description

一种电氧化浸出低品位氧化铅锌矿石的方法及其浸出设备 技术领域
本发明属于矿石分离领域，特别涉及一种用于提取低品位氧化铅锌矿石中铅锌的电氧化浸出方法与浸出设备。
背景技术
低品位氧化铅锌矿石的成份复杂，其中一般都含有硫化矿石、难浸出的异极矿以及硅酸金属矿石等成份，由于含有复杂的成份，在常规的浸出方法中，即使是添加药剂也不能很好的使得其中的铅与锌金属浸出，这也是使得常规的浸出工艺回收率较为低下的原因，而在常规的浸出工艺中还需要消耗较大的热能，不利于能源的节省。在常规的生产中，一般都需要将矿浆加热到90℃左右，如此便需要耗费大量的煤炭等燃料，大大提高生产的成本。同时，现有的浸出工艺在铅锌金属浸出后，其中的铅锌分离难度较高，需要依靠大量的化学沉淀剂在一定的温度下经过较长时间的反应才能完成。而现有的酸浸方法需要消耗大量的药剂，且其分离复杂，用于低品位氧化矿的分离成本较高。
综上所述，现在迫切的需要一种新的方法来降低低品位氧化铅锌矿石的分离难度与成本，提高分离的效率。
发明内容
本发明的目的在于克服上述问题，提供一种能有效的简化工艺流程，显著降低能耗、污染和成本，以及能显著提高生产效率的电氧化浸出低品位氧化铅锌矿石的方法。
本发明的另一目的是提供一种电氧化浸出低品位氧化铅锌矿石的浸出设备。
为了实现上述目的，本发明采用以下技术方案实现：
一种电氧化浸出低品位氧化铅锌矿石的方法，包括以下步骤：
A、制浆，将低品位氧化铅锌矿破碎并球磨成氧化铅锌矿粉，再将该氧化铅 锌矿粉与水混合制成矿浆；
B、浸出，先将矿浆加入电解槽，再向电解槽中添加碱液与活化剂使电解槽中的固体与液体的质量比为4～5：1，形成固液混合物；然后在常温常压下进行浸出，得到浸出液和海绵铅；
C、对浸出液进行分离与提取，得到锌粉；对分离锌粉后的滤液进行再生处理后再次返回电解槽中循环利用；
D、对海绵铅进行回收成锭；
E、对生产过程中生成的废水与废渣进行处理。
步骤A中的氧化铅锌矿粉的细度为200目，矿浆中的氧化铅锌矿粉的质量占比为25％。
步骤B中所述的碱液为碳酸钠、氢氧化钠或氢氧化钾，活化剂为十二烷基苯磺酸钠，而形成的固液混合物中的碱液的浓度为25％、活化剂的浓度为1％。所述的“在常温常压下进行浸出，得到浸出液和海绵铅”的过程中，电解槽中的电压控制在2.5～3.5V的范围内，阳极的电流密度保持在100～200A/m2，浸出的时间为4～6小时。
步骤C中所述所述的“对浸出液进行分离与提取”，具体为先通过沉淀或者离心过滤的方式取得浸出液的粗滤液，接着再通过电解粗滤液的方式得到锌粉。
步骤C中所述的再生处理包括化学中和沉淀与氢氧化钙加热除杂两步；在再生处理后的滤液返回电解槽后，还需要再向电解槽中逐渐加入碱液与活化剂，使得矿浆中碱液的浓度保持在25％、活化剂的浓度保持在1％。
步骤D中所述的“对海绵铅进行回收成锭”具体是指：先对电解槽中阴极上形成的海绵铅进行收集与洗涤，接着再将洗净的海绵铅压实，最后通过熔炼炉将压实后的海绵铅火炼成铅锭。
步骤E中的废渣经过洗涤与过滤直至达标后进行堆放；生产废水与废渣处理废水混合后采用中和沉淀法进行杂质沉淀，并利用硫酸调节废水的PH值至7.8后将废水进行回收再利用。
一种电氧化浸出低品位氧化铅锌矿石的方法的浸出设备，所述浸出设备为 电解槽，该电解槽成“十”字状，该电解槽的四端为阴极区，中央为阳极区，阳极区与阴极区通过隔膜间隔开来，在阳极区的中央位置设置有搅拌器，在电解槽中还设置有阴极板与阳极板，在阴极板上连接有阴极母线，在阳极板上连接有阳极母线。
进一步的，所述阳极板设置在阳极区内与隔膜相邻的位置，阴极板设置在阴极区内与隔膜相邻的位置，在每个阴极区内还成对设置有另外两对阴极板与阳极板。
本发明较现有技术相比，具有以下优点及有益效果：
(1)本发明的工艺方法较为简单，能够简单快捷的提取出铅与锌，提高了生产效率，降低了生产成本，且采用本发明的方法能够提高浸出率。
(2)本发明无需对矿浆进行加热，大大节省了热能，降低了生产过程中的能耗，进而节省了生产成本。
(3)本发明能够对药液进行循环使用，进而降低了生成过程中药液的消耗，进一步降低了生产所需投入的成本。
(4)本发明在电解过程中的电流密度大大低于常规的500～2000A/m2，很好的降低了设备制造的成本与生产过程中的耗电量，进而更好的节省了设备的制造与使用成本。
(5)本发明基于其快捷的效率与良好的效果，具有广泛的行业前景，能够很好的在行业中进行推广与使用，进而更好的促进了行业的进步与企业的发展。
附图说明
图1为本发明的步骤框图。
图2为本发明的电积槽的俯视图。
图3为本发明的电积槽的侧视剖面图。
附图标记说明：1、阴极母线；2、阳极母线；3、阴极区；4、阳极区；5、搅拌器；6、隔膜；7、阳极板；8、阴极板。
具体实施方式
下面结合实施例对本发明作进一步地详细说明，但本发明的实施方式不限 于此。
实施例1
如图1所示，本发明包括以下步骤：
A、制浆，将低品位氧化铅锌矿破碎并球磨成氧化铅锌矿粉，再将该氧化铅锌矿粉与水混合制成矿浆；
步骤A中的氧化铅锌矿粉的细度为200目，矿浆中的氧化铅锌矿粉的质量占比为25％。
B、浸出，先将矿浆加入电解槽，再向电解槽中添加碱液与活化剂使电解槽中的固体与液体的质量比为4～5：1，形成固液混合物；然后在常温常压下进行浸出，得到浸出液和海绵铅；
步骤B中所述的碱液为碳酸钠、氢氧化钠或氢氧化钾，活化剂为十二烷基苯磺酸钠，而形成的固液混合物中的碱液的浓度为25％、活化剂的浓度为1％。所述的“在常温常压下进行浸出，得到浸出液和海绵铅”的过程中，电解槽中的电压控制在2.5～3.5V的范围内，阳极的电流密度保持在100～200A/m2，浸出的时间为4～6小时。
在浸出的过程中，锌停留在浸出液中，铅则浸出在阴极板上，并随着铅的堆积加厚最终形成海绵铅附着在浸出液的表面上。该铅的浸出方程式为：PbO+2H+→Pb2++H2O；铅在电解过程中的反应方程式为：Pb2++2e-→Pb；当碱性药剂为氢氧化钠时，矿浆中的氧化锌与氢氧化钠的反应方程式为：2NaOH+ZnO＝Na2ZnO2+H2O。
C、对浸出液进行分离与提取，得到锌粉；对分离锌粉后的滤液进行再生处理后再次返回电解槽中循环利用；
步骤C中所述所述的“对浸出液进行分离与提取”，具体为先通过沉淀或者离心过滤的方式取得浸出液的粗滤液，接着再通过电解粗滤液的方式得到锌粉。步骤C中所述的再生处理包括化学中和沉淀与氢氧化钙加热除杂两步；在再生处理后的滤液返回电解槽后，还需要再向电解槽中逐渐加入碱液与活化剂，使得矿浆中碱液的浓度保持在25％、活化剂的浓度保持在1％。在电解锌粉的过程 中，该电解装置的阴极板与阳极板均采用316不锈钢板。
D、对海绵铅进行回收成锭；
步骤D中所述的“对海绵铅进行回收成锭”具体是指：先对电解槽中阴极上形成的海绵铅进行收集与洗涤，接着再将洗净的海绵铅压实，最后通过熔炼炉将压实后的海绵铅火炼成铅锭。
E、对生产过程中生成的废水与废渣进行处理。
步骤E中的废渣经过洗涤与过滤直至达标后进行堆放；生产废水与废渣处理废水混合后采用中和沉淀法进行杂质沉淀，并利用硫酸调节废水的PH值至7.8后将废水进行回收再利用。在中和沉淀法中，采用石灰调成的浓度为10～15％的石灰乳进行沉淀，该沉淀过程中的方程式为：
Pb2++Ca(OH)2→Pb(OH)2；
Zn2++Ca(OH)2→Zn(OH)2；
Cd2++Ca(OH)2→Cd(OH)2；
Fe2++Ca(OH)2→Fe(OH)2。
实施例2
本实施例将低品位氧化铅锌矿石破碎后球磨到200目25％的矿浆，并调整固液质量比为4：1，加入相应的药剂后对电解槽施加2.9～3.5V的电压，同时控制阳极的电流密度在100A/m2，其搅拌器每分钟的转速控制在300～500转的范围中，常温常压中工作4小时。最终铅的浸出率为92％，成锭后的铅锭品味87.5～93.5％；而锌粉的产出率为95％。
实施例3
本实施例将低品位氧化铅锌矿石破碎后球磨到200目25％的矿浆，并调整固液质量比为4：1，加入相应的药剂后对电解槽施加2.9～3.5V的电压，同时控制阳极的电流密度在100A/m2，其搅拌器每分钟的转速控制在300～500转的范围中，常温常压中工作6小时。最终铅的浸出率为94.3％，而锌粉的产出率为86.7％。
实施例4
如图2、3所示，一种电氧化浸出低品位氧化铅锌矿石的方法的浸出设备， 所述浸出设备为电解槽，该电解槽成“十”字状，该电解槽的四端为阴极区3，中央为阳极区4，阳极区4与阴极区3通过隔膜6间隔开来，在阳极区4的中央位置设置有搅拌器5，在电解槽中还设置有阴极板8与阳极板7，在阴极板8上连接有阴极母线1，在阳极板7上连接有阳极母线2。
所述阳极板7设置在阳极区4内与隔膜6相邻的位置，阴极板8设置在阴极区3内与隔膜6相邻的位置，在每个阴极区3内还成对设置有另外两对阴极板8与阳极板7。其中，阴极板采用钛板，阳极板采用钛板或者316不锈钢板。
通过上述方法，便能很好的实现本发明。

Claims (10)

  1. 一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，包括以下步骤：
    A、制浆，将低品位氧化铅锌矿破碎并球磨成氧化铅锌矿粉，再将该氧化铅锌矿粉与水混合制成矿浆；
    B、浸出，先将矿浆加入电解槽，再向电解槽中添加碱液与活化剂使电解槽中的固体与液体的质量比为4～5：1，形成固液混合物；然后在常温常压下进行浸出，得到浸出液和海绵铅；
    C、对浸出液进行分离与提取，得到锌粉；对分离锌粉后的滤液进行再生处理后再次返回电解槽中循环利用；
    D、对海绵铅进行回收成锭；
    E、对生产过程中生成的废水与废渣进行处理。
  2. 根据权利要求1所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，步骤A中的氧化铅锌矿粉的细度为200目，矿浆中的氧化铅锌矿粉的质量占比为25％。
  3. 根据权利要求2所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，步骤B中所述的碱液为碳酸钠、氢氧化钠或氢氧化钾，活化剂为十二烷基苯磺酸钠，而形成的固液混合物中的碱液的浓度为25％、活化剂的浓度为1％。
  4. 根据权利要求3所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，在步骤B中的“在常温常压下进行浸出，得到浸出液和海绵铅”的过程中，该电解槽中的电压控制在2.5～3.5V的范围内，阳极的电流密度保持在100～200A/m2，浸出的时间为4～6小时。
  5. 根据权利要求4所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，步骤C中所述的“对浸出液进行分离与提取”，具体为先通过沉淀或者离心过滤的方式取得浸出液的粗滤液，接着再通过电解粗滤液的方式得到锌粉。
  6. 根据权利要求5所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其 特征在于，步骤C中所述的再生处理包括化学中和沉淀与氢氧化钙加热除杂两步；在再生处理后的滤液返回电解槽后，还需要再向电解槽中逐渐加入碱液与活化剂，使得矿浆中碱液的浓度保持在25％、活化剂的浓度保持在1％。
  7. 根据权利要求6所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，步骤D中所述的“对海绵铅进行回收成锭”具体是指：先对电解槽中阴极上形成的海绵铅进行收集与洗涤，接着再将洗净的海绵铅压实，最后通过熔炼炉将压实后的海绵铅火炼成铅锭。
  8. 根据权利要求7所述的一种电氧化浸出低品位氧化铅锌矿石的方法，其特征在于，步骤E中的废渣经过洗涤与过滤直至达标后进行堆放；生产废水与废渣处理废水混合后采用中和沉淀法进行杂质沉淀，并利用硫酸调节废水的PH值至7.8后将废水进行回收再利用。
  9. 根据权利要求1-8任意一项所述的一种电氧化浸出低品位氧化铅锌矿石的方法的浸出设备，其特征在于，所述浸出设备为电解槽，该电解槽成“十”字状，该电解槽的四端为阴极区(3)，中央为阳极区(4)，阳极区(4)与阴极区(3)通过隔膜(6)间隔开来，在阳极区(4)的中央位置设置有搅拌器(5)，在电解槽中还设置有阴极板(8)与阳极板(7)，在阴极板(8)上连接有阴极母线(1)，在阳极板(7)上连接有阳极母线(2)。
  10. 根据权利要求9所述的一种电氧化浸出低品位氧化铅锌矿石的方法的浸出设备，其特征在于，所述阳极板(7)设置在阳极区(4)内与隔膜(6)相邻的位置，阴极板(8)设置在阴极区(3)内与隔膜(6)相邻的位置，在每个阴极区(3)内还成对设置有另外两对阴极板(8)与阳极板(7)。
PCT/CN2016/072471 2015-08-25 2016-01-28 一种电氧化浸出低品位氧化铅锌矿石的方法及其浸出设备 WO2017031949A1 (zh)

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