[go: up one dir, main page]

WO2020062145A1 - 硫化铜精矿的氧压浸出方法及铜冶炼方法 - Google Patents

硫化铜精矿的氧压浸出方法及铜冶炼方法 Download PDF

Info

Publication number
WO2020062145A1
WO2020062145A1 PCT/CN2018/108652 CN2018108652W WO2020062145A1 WO 2020062145 A1 WO2020062145 A1 WO 2020062145A1 CN 2018108652 W CN2018108652 W CN 2018108652W WO 2020062145 A1 WO2020062145 A1 WO 2020062145A1
Authority
WO
WIPO (PCT)
Prior art keywords
copper
oxygen pressure
pressure leaching
leaching
oxygen
Prior art date
Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
Ceased
Application number
PCT/CN2018/108652
Other languages
English (en)
French (fr)
Inventor
仝一喆
刘自亮
王恒辉
尹泽辉
罗虹霖
冯泽平
杨建平
刘刚
左小红
邓孟俐
谢冰
施耘
张克
陈龙义
吉红
何醒民
鹏苏格·巴图奥奇
阿拉腾苏和·道尔吉贡土布
巴彦巴策仁·恩赫宝鲁德
齐涛
孟凡成
陈德胜
王丽娜
于宏东
林裕安
刘野平
张登凯
徐克华
何磊
Current Assignee (The listed assignees may be inaccurate. Google has not performed a legal analysis and makes no representation or warranty as to the accuracy of the list.)
Danxia Smelter Of Shenzhen Zhongjin Lingnan Nonfemet Co Ltd
Institute of Process Engineering of CAS
CINF Engineering Corp Ltd
Mongolian Ocean Engineering Group LLC
Original Assignee
Danxia Smelter Of Shenzhen Zhongjin Lingnan Nonfemet Co Ltd
Institute of Process Engineering of CAS
CINF Engineering Corp Ltd
Mongolian Ocean Engineering Group LLC
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Application filed by Danxia Smelter Of Shenzhen Zhongjin Lingnan Nonfemet Co Ltd, Institute of Process Engineering of CAS, CINF Engineering Corp Ltd, Mongolian Ocean Engineering Group LLC filed Critical Danxia Smelter Of Shenzhen Zhongjin Lingnan Nonfemet Co Ltd
Priority to CN201880058273.8A priority Critical patent/CN111225988B/zh
Priority to PCT/CN2018/108652 priority patent/WO2020062145A1/zh
Publication of WO2020062145A1 publication Critical patent/WO2020062145A1/zh
Anticipated expiration legal-status Critical
Ceased legal-status Critical Current

Links

Images

Classifications

    • 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
    • C22B15/00—Obtaining copper
    • 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
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
    • 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

Definitions

  • the invention relates to an oxygen pressure leaching method of copper sulfide concentrate and a copper smelting method, and belongs to the technical field of non-ferrous metal wet smelting.
  • copper sulfide concentrate is basically smelted by fire. First, copper in the copper sulfide concentrate is smelted into crude copper, and then the crude copper is electrolyzed to obtain cathode copper. The sulfur-containing flue gas is sent to produce acid to produce sulfuric acid. And low-grade copper sulfide ore uses hydrometallurgy, that is, leaching-extraction-electrodeposition process to produce electrodeposited copper, because the raw material contains low sulfur and does not produce sulfuric acid. In some areas where copper sulfide concentrates are produced, due to the weak industrial base in the surroundings, sulfuric acid cannot be transported, stored, and sold, and the method of smelting copper sulfide concentrates by fire cannot be implemented.
  • the present invention is based on the operability of industrial production, and provides an oxygen pressure leaching method of copper sulfide concentrate to control the iron and sulfuric acid content in the leaching solution while ensuring a high leaching rate of copper. Make it meet the requirements of subsequent electrowinning; in addition, the invention also provides a copper smelting method.
  • the oxygen pressure leaching method of copper sulfide concentrate includes the following steps:
  • the first dispersant, the first alum agent, the second stage supernatant and the slurry obtained in S1 are added to an autoclave to obtain a first acid concentration (that is, the concentration of H 2 SO 4 ) of 50 to 70 g / L.
  • a slurry is passed into the autoclave, and oxygen is leached at a temperature of 135 to 145 ° C and a pressure of 1.3 to 1.5 MPa. After the reaction is completed, the temperature and pressure are reduced, and the solid-liquid separation is obtained.
  • the first dispersant is added in an amount of 3 to 5 kg / t-concentrate (that is, after 3 to 5 kg of the first dispersant is added to the corresponding pulp after grinding into a pulp per ton of copper sulfide concentrate), the first
  • the addition amount of the alum agent is 25-30 kg / t-concentrate, preferably, the addition amount of the first dispersant is 3.5-4.5 kg / t-concentrate, and the addition amount of the first alum agent is 26-28 kg / t-concentrate;
  • the concentration of H 2 SO 4 in a section of the supernatant is 25 to 35 g / L
  • the concentration of Cu 2+ is 90 to 100 g / L
  • the concentration of Fe 3+ is 8 to 14 g / L;
  • the concentration of H 2 SO 4 in the waste electricity effluent is 160 to 180 g / L, preferably 165 to 175 g / L; generally, the amount of the second dispersant is 5 to 7 kg / t-concentrate, and the second The amount of alum added is 30-35kg / t-concentrate; preferably, the amount of second dispersant is 5.5-6.5kg / t-concentrate and the amount of second alum is 32-34kg / t -Concentrate;
  • a neutralizing agent is added to a section of the supernatant, and reacted to obtain a neutralizing slurry having a pH value of 2.5-3.5, and then solid-liquid separation is performed to obtain a neutralizing supernatant and a neutralizing residue.
  • the concentration of Cu 2+ in the neutralization supernatant is 90 to 100 g / L
  • the concentration of Fe 3+ is less than or equal to 2 g / L
  • the pH value is 2.5 to 3.5, which can satisfy the composition of the new electrolyte solution in the copper electrodeposition process.
  • two stages of countercurrent oxygen pressure leaching are used, one stage is at intermediate temperature and medium pressure to drop acid, and the second stage is high temperature and high pressure leaching of copper.
  • the second stage supernatant is returned to a stage of oxygen pressure leaching to provide acid for a stage of oxygen pressure leaching reaction and participate in a stage of oxygen.
  • the acid is consumed in the reaction system.
  • the copper leached in the second stage of oxygen pressure leaching is accumulated in the leaching solution and enters the first stage of the supernatant.
  • the temperature and pressure used in the first stage of oxygen pressure leaching are relatively low, and the iron leaching rate is low.
  • alum agents are added to suppress the leaching of iron. In this way, it can not only ensure the high leaching rate of copper, but also effectively control the iron and sulfuric acid content in the leaching solution.
  • the required acid is mainly provided by the waste electricity effusion.
  • the solid content of the pulp is 60-70 wt%, preferably 63-68 wt%.
  • the ore is pulverized by a sand mill.
  • the agitating disc of the sand mill can fully contact the grinding medium with the material.
  • the grinding medium is 0.8 to 1 mm zirconium beads, and the specific surface area is large, which can sufficiently grind the material. fine.
  • the ore particles having a particle size of less than 15 ⁇ m in the pulp account for more than 90% by weight of the total minerals, and preferably, more than 90% by weight of the total minerals. Ensure certain particle size conditions, improve reaction efficiency, and further increase the leaching rate.
  • minerals can be understood as copper sulfide concentrates.
  • neutralizing slag is added, minerals can be understood as a mixture of neutralizing slag and copper sulfide concentrates.
  • reaction time for a period of oxygen pressure leaching is 1.5-2.5h, preferably 1.8-2.3h.
  • the oxygen partial pressure in the autoclave is 0.9-1.1 MPa, and preferably 0.95-1.05 MPa.
  • reaction time of the two-stage oxygen pressure leaching is 2.5-3.5 h, preferably 2.8-3.3 h.
  • the oxygen partial pressure in the autoclave is 0.8-1.0 MPa, and preferably 0.85-0.95 MPa.
  • the reaction time Through reasonable control of the reaction time, the high completion of the leaching reaction can be ensured, and the leaching time can be shortened, thereby avoiding time wastage.
  • the control of the oxygen partial pressure Through the control of the oxygen partial pressure, the requirements for the reaction conditions of the oxygen pressure leaching reaction can be guaranteed, and the oxygen pressure leaching reaction can be performed efficiently.
  • the neutralizing agent is limestone and / or copper roasting sand.
  • the neutralizing agent is copper roasting sand.
  • the neutralization slag obtained by the neutralization reaction is sent to S1 for grinding and pulping, and further
  • the mass ratio of neutralization slag and copper sulfide concentrate can be set as required.
  • the neutralization slag and copper sulfide concentrate can be mixed at any ratio.
  • the copper content is 18-25 wt%
  • the iron content is 25-35% by weight
  • the sulfur content is 4-6 wt%
  • the remaining components are gangue.
  • the neutralizing agent is limestone, and the underflow generated can be neutralized slag by pressure filtration, and can be processed by storage and the like.
  • the waste electricity accumulation liquid is a copper electricity accumulation waste liquid, wherein the Cu 2+ concentration is not higher than 40 g / L, and the Fe 3+ concentration is not higher than 2 g / L.
  • both the dispersant and the alum agent can be selected from the dispersants and alum agents commonly used in the non-ferrous metal wet smelting and leaching process.
  • the dispersant may be a lignin sulfonate
  • the alum agent may be It is an ammonium salt or a base.
  • the solid-liquid separation process can be performed by a thickener.
  • the concentration of oxygen introduced into the autoclave is not less than 99%.
  • the oxygen pressure leaching method of the present invention can be applied to leaching treatment of copper sulfide concentrates of different grades.
  • the present invention also provides a copper smelting method.
  • the copper sulfide concentrate is subjected to leaching treatment according to the oxygen pressure leaching method described above to obtain a neutralized supernatant; and then, the neutralized supernatant is used. It is used to collect electric liquid, and to obtain electrolytic copper and copper electrolytic waste liquid.
  • control current density is 160 to 180 A / m 2
  • cycle of the electric product is 7 days.
  • A is selected from Na + , K + , NH 4 + ; reactions (1), (2), (3), and (4) mainly occur during the oxygen pressure leaching process.
  • Two-stage countercurrent oxygen pressure leaching can increase the leaching rate of copper.
  • the total leaching rate of copper in two-stage oxygen pressure leaching can reach more than 90%, and the supernatant of one stage is low in acid and iron, and the neutralizing dose used is small. , Less slag, less copper loss. For the copper that has not been leached in the second-stage leaching slag, flotation can be sent to recover copper and sulfur.
  • the smelting waste electricity effluent can be effectively used, which can reduce production costs and realize the reuse of resources.
  • FIG. 1 is a flowchart of a copper smelting method according to the present invention.
  • the section of supernatant contained Cu 90.78g / L, Fe10.81g / L, and H 2 SO 4 28g / L.
  • the generated one-stage underflow was sent to a two-stage autoclave. Waste electric effusion, 0.6g of dispersant, and 3g of alum were added.
  • the liquid-solid ratio was controlled to 4: 1, oxygen was passed in, the temperature was controlled to 165 ° C, and the pressure was 1.4MPa.
  • a two-stage oxygen pressure leaching was performed, and the reaction was performed for 2.5 hours to produce 70 g of two-stage leaching slag (containing Cu3.1wt%, Fe33.7wt%), the total copper leaching rate of two stages was 90.49%, and the total iron leaching rate was 13.34%.
  • the section of the supernatant contained 92.34 g / L of Cu, 13.84 g / L of Fe, and 30 g / L of H 2 SO 4 .
  • the generated one-stage underflow was sent to a two-stage autoclave, and waste electric effusion, 0.7g of dispersant and 3.5g of alum were added, the liquid-solid ratio was controlled to 4: 1, oxygen was passed in, the temperature was controlled at 170 ° C, and the pressure was 1.5MPa With a reaction time of 3 h, two-stage oxygen pressure leaching was performed to produce 68 g of two-stage leaching slag (containing Cu2.6wt%, Fe33.1wt%), the total copper leaching rate of two stages was 92.25%, and the total iron leaching rate was 17.31%.

Landscapes

  • Chemical & Material Sciences (AREA)
  • Engineering & Computer Science (AREA)
  • Organic Chemistry (AREA)
  • Metallurgy (AREA)
  • Mechanical Engineering (AREA)
  • Materials Engineering (AREA)
  • Manufacturing & Machinery (AREA)
  • Environmental & Geological Engineering (AREA)
  • Geology (AREA)
  • Geochemistry & Mineralogy (AREA)
  • General Life Sciences & Earth Sciences (AREA)
  • Life Sciences & Earth Sciences (AREA)
  • Inorganic Chemistry (AREA)
  • Manufacture And Refinement Of Metals (AREA)

Abstract

一种硫化铜精矿的氧压浸出方法及铜冶炼方法，先将硫化铜精矿加水磨制成矿浆；再将第一分散剂、第一沉矾剂、二段上清液和矿浆加入到高压釜中，进行一段氧压浸出，获得一段底流和一段上清液；然后将一段底流、废电积液、第二分散剂和第二沉矾剂加入到高压釜中，进行二段氧压浸出，获得二段上清液和二段浸出渣；向一段上清液中加入中和剂，获得中和上清液和中和渣，使用中和上清液电积铜。本方法在保证铜的高浸出率同时，控制浸出液中的铁及硫酸含量。

Description

硫化铜精矿的氧压浸出方法及铜冶炼方法 技术领域
本发明涉及硫化铜精矿的氧压浸出方法及铜冶炼方法，属于有色金属湿法冶炼技术领域。
背景技术
目前，硫化铜精矿基本都是采用火法冶炼，首先将硫化铜精矿中铜冶炼成粗铜，然后将粗铜电解得到阴极铜，含硫烟气送制酸生产硫酸，而氧化铜矿以及低品位硫化铜矿则采用湿法冶金，即浸出—萃取—电积工艺生产电积铜，因原料含硫低不产出硫酸。在一些生产硫化铜精矿的地区，由于周边工业基础薄弱，硫酸无法运输贮存及销售，也就无法实施火法冶炼硫化铜精矿的工艺方法。因此，研究类似硫化锌精矿氧压浸出生产硫磺的方法应用于硫化铜精矿的湿法冶金方法具有十分重要的现实意义。虽然，硫化锌精矿氧压浸出工艺用于工业生产已有三十余年实践经验，但硫化铜精矿氧压浸出工艺还只是处于试验研究阶段，目前大多数研究工作内容及方法路线与工业化生产还有很大差距。硫化铜精矿氧压浸出工艺最大难点是铜浸出需要高温，温度升高，铜浸出率高，但随之进入浸出液中的铁及硫酸含量增加，后续工序采用中和剂如石灰石降酸除铁，将产出大量中和渣导致铜损失，由于浸出液中铁为Fe 3+，铁渣过滤性能也差。2007年在美国Arizona建成一套6.6万t/a铜氧浸工厂，该厂采用一段高温高压氧浸工艺，产出的浸出液经脱硅后进行铜电积，产出的大部分电积液送往氧化铜矿堆浸降酸，堆浸液经萃取净化后返回氧压浸出，即进入高压釜的混合电积液酸度低。因此，该厂完全是利用硫化铜精矿氧压浸出产出的多余硫酸用于氧化铜矿堆浸。
然而，在没有足够氧化铜矿来消耗多余硫酸的地区，在保证铜浸出率的同时，如何控制进入浸出液中的铁及硫酸含量以满足后续电积铜工序的工况要求也具有很大的难度。
发明内容
针对现有技术的不足，本发明立足于工业化生产的可操作性，提供一种硫化铜精矿的氧压浸出方法，以在保证铜的高浸出率同时，控制浸出液中的铁及硫酸含量，使之满足后续电积要求；另外，本发明还提供一种铜冶炼方法。
为了解决上述技术问题，本发明的技术方案如下：
硫化铜精矿的氧压浸出方法，包括如下步骤：
S1、将硫化铜精矿加水磨制成矿浆；
S2、将第一分散剂、第一沉矾剂、二段上清液和S1中获得的矿浆加入到高压釜中，获得始酸浓度(即H 2SO 4的浓度)为50～70g/L的第一料浆，再向高压釜中通入氧气，在温度为135～145℃、压力为1.3～1.5MPa的条件下，进行一段氧压浸出，反应完毕后，降温降压，固液分离，获得一段底流和一段上清液；
其中，一般地，第一分散剂的添加量为3～5kg/t-精矿(即每吨硫化铜精矿磨成矿浆后，在相应矿浆配入3～5kg第一分散剂)，第一沉矾剂的添加量为25～30kg/t-精矿，优选地，第一分散剂的添加量为3.5～4.5kg/t-精矿，第一沉矾剂的添加量为26～28kg/t-精矿；一段上清液中H 2SO 4浓度为25～35g/L，Cu 2+浓度为90～100g/L，Fe 3+浓度为8～14g/L；
S3、将一段底流、废电积液、第二分散剂、第二沉矾剂加入到高压釜中，获得始酸浓度为100～130g/L的第二料浆，再向高压釜中通入氧气，在温度为165～175℃、压力为1.4～1.6MPa的条件下，进行二段氧压浸出，反应完毕后，降温降压，固液分离，获得二段上清液和二段浸出渣；
其中，废电积液中H 2SO 4浓度为160～180g/L，优选为165-175g/L；一般地，第二分散剂的添加量为5～7kg/t-精矿，第二沉矾剂的添加量为30～35kg/t-精矿；优选地，第二分散剂的添加量为5.5～6.5kg/t-精矿，第二沉矾剂的添加量为32～34kg/t-精矿；
向一段上清液中加入中和剂，反应，获得pH值为2.5-3.5的中和料浆，然后固液分离，获得中和上清液和中和渣。
本发明中，中和上清液中Cu 2+浓度为90～100g/L，Fe 3+浓度≤2g/L，pH值为2.5～3.5，可满足铜电积过程对电积新液的成分要求。
本发明中，采用两段逆流氧压浸出，一段中温中压降酸，二段高温高压浸铜，其中，二段上清液返回至一段氧压浸出，为一段氧压浸出反应提供酸，参与一段氧压浸出，反应系统中酸得到消耗，二段氧压浸出中浸出的铜在浸出液中得到累积，并进入一段上清液中，一段氧压浸出采用的温度、压 力相对较低，铁浸出率低；同时，两段氧压浸出过程中，均加入沉矾剂，抑制铁的浸出。如此，既可保证铜的高浸出率，又可有效控制浸出液中的铁及硫酸含量。另外，两段氧压浸出反应过程中，所需的酸主要由废电积液提供。
S1中，矿浆的固含量为60-70wt％，优选为63-68wt％。
进一步地，S1中，通过砂磨机进行磨矿，砂磨机的搅拌磨盘能使研磨介质与物料充分接触，进一步地，研磨介质为0.8～1mm锆珠，比表面积大，能将物料充分磨细。
S1中，矿浆中粒度小于15μm的矿粒占矿物总量的90wt％以上，优选地，占矿物总量的90wt％以上。保证一定的粒度条件，提升反应效率，进一步地提高浸出率。一般而言，矿物可理解为硫化铜精矿，当加入中和渣时，矿物可理解为中和渣与硫化铜精矿的混合物。
S2中，一段氧压浸出反应时间为1.5-2.5h，优选为1.8-2.3h。
S2中，一段氧压浸出反应时，高压釜中氧分压为0.9-1.1MPa，优选为0.95-1.05MPa。
S3中，二段氧压浸出反应时间为2.5-3.5h，优选为2.8-3.3h。
S3中，二段氧压浸出反应时，高压釜中氧分压为0.8-1.0MPa，优选为0.85-0.95MPa。
通过反应时间的合理调控，即可保证浸出反应的高完成度，又可缩短浸出耗时，避免时间浪费。通过氧分压的控制，可保证氧压浸出反应过程对反应条件的要求，保证氧压浸出反应高效进行。
S3中，所述中和剂为石灰石和/或铜焙砂，优选地，所述中和剂为铜焙砂，将中和反应获得的中和渣送入S1中磨矿制浆，进一步地，中和渣与硫化铜精矿的质量比可根据需要设定，一般而言，中和渣、硫化铜精矿可为任一比例混合。进一步地，所述铜焙砂中，铜含量为18-25wt％，铁含量为25-35wt％，硫含量为4-6wt％，其余成分为脉石。
所述中和剂为石灰石，产生的底流经压滤可得中和渣，可通过堆存等方式处理。
S3中，所述废电积液为铜电积废液，其中，Cu 2+浓度不高于40g/L，Fe 3+浓度不高于2g/L。
本发明中，分散剂、沉矾剂均可选用有色金属湿法冶炼浸出处理过程中常用的分散剂、沉矾剂，进一步地，所述分散剂可以为木质素磺酸盐，沉矾剂可以为铵盐或碱。
本发明中，所述固液分离过程均可采用浓密机进行。
本发明中，通入高压釜中的氧气的浓度不低于99％。
本发明的氧压浸出方法可适用于不同品位的硫化铜精矿的浸出处理。
基于同一发明构思，本发明还提供一种铜冶炼方法，先按如上所述的氧压浸出方法对硫化铜精矿进行浸出处理，获得中和上清液；然后以所述中和上清液为电积液，电积，获得电积铜和铜电积废液。
进一步地，电积过程中，控制电流密度为160～180A/m 2，电积周期为7天。
本发明中涉及的主要反应如下：
CuFeS 2+O 2+2H 2SO 4→CuSO 4+FeSO 4+2S 0+2H 2O        (1)
2FeSO 4+0.5O 2+H 2SO 4→Fe 2(SO 4) 3+H 2O                (2)
Fe 2(SO 4) 3+3H 2O→Fe 2O 3+3H 2SO 4                      (3)
3Fe 2(SO 4) 3+2(A)OH+10H 2O→2(A)Fe 3(SO 4) 2(OH) 6+5H 2SO 4(4)
其中，A选自Na +、K +、NH 4 +；反应(1)、(2)、(3)、(4)主要发生于氧压浸出过程。
与现有技术相比，本发明的有益效果如下：
(1)将硫化铜精矿等物料充分磨细成矿浆，提升浸出效率和浸出率。
(2)采用两段逆流氧压浸出，二段氧压浸出采用废电积液直接浸出一段底流，二段上清液经一段氧压浸出降酸后，产出一段上清液成分(H 2SO 4浓度为25～35g/L，Cu 2+浓度为90～100g/L，Fe 3+浓度为8～14g/L)，一段上清液经中和处理后，获得的中和上清液可满足传统铜电积新液的成分要求(Cu 2+浓度为90～100g/L，Fe 3+浓度为2g/L，pH为2.5～3.5)。
(3)在高温高酸条件下进行除铁，直接将铁以矾渣形式沉积在高压釜内，解决了铜氧压浸出上清液铁含量高的难题。
(4)对一段上清液采用中和剂中和，除中和硫酸外，还可除去溶液中部分杂质如铁、砷等，不需溶剂萃取，进一步降低杂质含量，可满足电积铜工况要求，简化了工艺流程及设备配置。
(5)两段逆流氧压浸出可提高铜的浸出率，两段氧压浸出中铜的总浸出率可达90％以上，且一段上清液酸低、铁低，使用的中和剂量少，渣量少，铜损失量小。对二段浸出渣中未被浸出的铜，可送浮选回收铜及硫。
(6)冶炼产生的废电积液可得到有效利用，可降低生产成本，并实现资源的再利用。
附图说明
图1是本发明的一种铜冶炼方法的流程图。
具体实施方式
以下将结合实施例来详细说明本发明。
实施例1
取100g硫化铜精矿(含Cu22.81wt％，Fe27.22wt％)，磨矿至90％精矿粒度小于15μm，获得矿浆。将二段上清液、分散剂0.4g、沉矾剂2.5g与矿浆加入高压釜中，控制液固比为4:1，，通入氧气，控制温度为135℃，压力1.3MPa，进行一段氧压浸出，反应2h，获得一段上清液370mL、一段底流，该一段上清液含Cu 90.78g/L、Fe10.81g/L、H 2SO 4 28g/L。将产生的一段底流送入二段高压釜，加入废电积液、分散剂0.6g及沉矾剂3g，控制液固比为4:1，通入氧气，控制温度为165℃，压力1.4MPa，进行二段氧压浸出，反应2.5h，产生二段浸出渣70g(含Cu3.1wt％，Fe33.7wt％)，两段铜总浸出率90.49％，铁总浸出率13.34％。
实施例2
取100g硫化铜精矿(含Cu22.81wt％，Fe27.22wt％)，磨矿至90％精矿粒度小于15μm，获得矿浆。将二段上清液、分散剂0.5g及沉矾剂3g与矿浆加入高压釜中，控制液固比为4:1，通入氧气，控制温度140℃，压力1.4MPa，进行一段氧压浸出，反应2.5h，产生一段上清液367mL、一段底流，该一段上清液含Cu 92.34g/L、Fe13.84g/L、H 2SO 4 30g/L。将产生的一段底流送入二段高压釜，加入废电积液、分散剂0.7g及沉矾剂3.5g，控制液固比为4:1，通入氧气，控制温度170℃，压力1.5MPa，反应时间3h，进行二段氧压浸出，产生二段浸出渣68g(含Cu2.6wt％，Fe33.1wt％)，两段铜总浸出率92.25％，铁总浸出率17.31％。
对比例1
取100g硫化铜精矿(含Cu22.81wt％，Fe27.22wt％)，磨矿至90％精矿粒度小于15μm，获得矿浆。将矿浆与浸出剂混合(硫酸浓度为160g/L、铜含量为40g/L、铁含量为2g/L)，添加适量分散剂， 分两组，进行一段氧压浸出，具体试验条件及结果如表1所示。
表1对比例1一段氧压浸出参数及结果情况
Figure PCTCN2018108652-appb-000001
由表1可知，一段氧压浸出的浸出液含酸、含铁高，且温度越高，浸出液中酸、铁含量越高，对比可知，对比例中铁的浸出率远远超出本申请中铁的浸出率。
对比例2
取100g硫化铜精矿(含Cu22.81wt％，Fe27.22wt％)，磨矿至90％精矿粒度小于15μm，获得矿浆。将矿浆与浸出剂混合(硫酸浓度为40g/L、铜含量为70g/L、铁含量为35g/L)，添加适量分散剂、沉矾剂，分两组，进行一段氧压浸出，具体试验条件及结果如表2所示。
表2对比例2一段氧压浸出参数及结果情况
Figure PCTCN2018108652-appb-000002
由表2可知，沉矾剂的添加，可有效降低浸出液中铁含量，但是铜的浸出率也有所降低，不利于获得高的铜浸出率。而本申请通过两段氧压浸出相关工艺参数的调控，可在降低铁浸出率的同时，获得高的铜浸出率，有效解决这一矛盾。
上述实施例阐明的内容应当理解为这些实施例仅用于更清楚地说明本发明，而不用于限制本发明的范围，在阅读了本发明之后，本领域技术人员对本发明的各种等价形式的修改均落入本申请所附权利要求所限定的范围。

Claims (10)

  1. 硫化铜精矿的氧压浸出方法，其特征在于，包括如下步骤：
    S1、将硫化铜精矿加水磨制成矿浆；
    S2、将第一分散剂、第一沉矾剂、二段上清液和S1中获得的矿浆加入到高压釜中，获得始酸浓度为50～70g/L的第一料浆，再向高压釜中通入氧气，在温度为135～145℃、压力为1.3～1.5MPa的条件下，进行一段氧压浸出，反应完毕后，降温降压，固液分离，获得一段底流和一段上清液；
    其中，一段上清液中H 2SO 4浓度为25～35g/L，Cu 2+浓度为90～100g/L，Fe 3+浓度为8～14g/L；
    S3、将一段底流、废电积液、第二分散剂、第二沉矾剂加入到高压釜中，获得始酸浓度为100～130g/L的第二料浆，再向高压釜中通入氧气，在温度为165～175℃、压力为1.4～1.6MPa的条件下，进行二段氧压浸出，反应完毕后，降温降压，固液分离，获得二段上清液和二段浸出渣；
    其中，废电积液中H 2SO 4浓度为160～180g/L；
    向一段上清液中加入中和剂，反应，获得pH值为2.5-3.5的中和料浆，然后固液分离，获得中和上清液和中和渣。
  2. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S1中，矿浆的固含量为60-70wt％；优选地，S1中，矿浆中粒度小于15μm的矿粒占矿物总量的90wt％以上。
  3. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S2中，一段氧压浸出反应时间为1.5-2.5h。
  4. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S2中，一段氧压浸出反应时，高压釜中氧分压为0.9-1.1MPa。
  5. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，第一分散剂的添加量为3～5kg/t-精矿，第一沉矾剂的添加量为25～30kg/t-精矿；第二分散剂的添加量为5～7kg/t-精矿，第二沉矾剂的添加量为30～35kg/t-精矿。
  6. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S3中，二段氧压浸出反应时间为2.5-3.5h。
  7. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S3中，二段氧压浸出反应时，高压釜中氧分压为0.8-1.0MPa。
  8. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S3中，所述中和剂为石灰石和/或铜焙砂，优选地，所述中和剂为铜焙砂，将中和反应获得的中和渣送入S1中磨矿制浆。
  9. 根据权利要求1所述的硫化铜精矿的氧压浸出方法，其特征在于，S3中，所述废电积液为铜电积废液，其中，Cu 2+浓度不高于40g/L，Fe 3+浓度不高于2g/L。
  10. 铜冶炼方法，其特征在于，先按如权利要求1-9任一项所述的氧压浸出方法对硫化铜精矿进行浸出处理，获得中和上清液；然后以所述中和上清液为电积液，电积，获得电积铜和铜电积废液。
PCT/CN2018/108652 2018-09-29 2018-09-29 硫化铜精矿的氧压浸出方法及铜冶炼方法 Ceased WO2020062145A1 (zh)

Priority Applications (2)

Application Number Priority Date Filing Date Title
CN201880058273.8A CN111225988B (zh) 2018-09-29 2018-09-29 硫化铜精矿的氧压浸出方法及铜冶炼方法
PCT/CN2018/108652 WO2020062145A1 (zh) 2018-09-29 2018-09-29 硫化铜精矿的氧压浸出方法及铜冶炼方法

Applications Claiming Priority (1)

Application Number Priority Date Filing Date Title
PCT/CN2018/108652 WO2020062145A1 (zh) 2018-09-29 2018-09-29 硫化铜精矿的氧压浸出方法及铜冶炼方法

Publications (1)

Publication Number Publication Date
WO2020062145A1 true WO2020062145A1 (zh) 2020-04-02

Family

ID=69952811

Family Applications (1)

Application Number Title Priority Date Filing Date
PCT/CN2018/108652 Ceased WO2020062145A1 (zh) 2018-09-29 2018-09-29 硫化铜精矿的氧压浸出方法及铜冶炼方法

Country Status (2)

Country Link
CN (1) CN111225988B (zh)
WO (1) WO2020062145A1 (zh)

Cited By (3)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN116040772A (zh) * 2023-03-27 2023-05-02 长春黄金研究院有限公司 高铜酸性水治理产渣无害化的处理方法
CN116200595A (zh) * 2023-02-23 2023-06-02 四川四环锌锗科技有限公司 一种从高硫含铜物料中综合回收有价金属的方法
CN116463497A (zh) * 2023-04-20 2023-07-21 云南驰宏锌锗股份有限公司 一种高铜高铅硫化锌精矿氧压浸出方法

Families Citing this family (2)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN113215398B (zh) * 2021-05-06 2022-03-29 长沙有色冶金设计研究院有限公司 一种硫化镍精矿的氧压浸出方法
CN113215399B (zh) * 2021-05-06 2022-04-05 长沙有色冶金设计研究院有限公司 一种硫化镍精矿的氧压浸出方法

Citations (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CA2363969A1 (en) * 2001-11-26 2003-05-26 Walter Curlook Process for the recovery of residual metal values from smelter waste slags, and from converter slags
CN101643857A (zh) * 2009-08-25 2010-02-10 西部矿业股份有限公司 一种复杂铜铅锌银多金属硫化矿综合回收方法
CN101775502A (zh) * 2010-02-04 2010-07-14 云南澜沧铅矿有限公司 湿法炼锌流程提取铜工艺
CN107099669A (zh) * 2017-04-21 2017-08-29 昆明冶金研究院 一种含铜渣中高效清洁除砷的方法

Family Cites Families (5)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
AU749257B2 (en) * 1998-07-24 2002-06-20 Intec Hellyer Metals Pty Ltd Processing copper sulphide ores
FI120315B (fi) * 2007-11-27 2009-09-15 Outotec Oyj Menetelmä pyriittisen, kultaa, kuparia ja arseenia sisältävän rikasteen käsittelemiseksi
CN101225476B (zh) * 2008-02-22 2010-10-13 昆明理工大学 从铅冰铜中回收铜的工艺
CN100554452C (zh) * 2008-07-17 2009-10-28 北京矿冶研究总院 一种含铜硫化矿湿法提取铜的方法
CN101440434A (zh) * 2008-12-17 2009-05-27 昆明金湖冶金有限公司 一种高钙镁硫化铜湿法提取铜的工艺

Patent Citations (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CA2363969A1 (en) * 2001-11-26 2003-05-26 Walter Curlook Process for the recovery of residual metal values from smelter waste slags, and from converter slags
CN101643857A (zh) * 2009-08-25 2010-02-10 西部矿业股份有限公司 一种复杂铜铅锌银多金属硫化矿综合回收方法
CN101775502A (zh) * 2010-02-04 2010-07-14 云南澜沧铅矿有限公司 湿法炼锌流程提取铜工艺
CN107099669A (zh) * 2017-04-21 2017-08-29 昆明冶金研究院 一种含铜渣中高效清洁除砷的方法

Cited By (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN116200595A (zh) * 2023-02-23 2023-06-02 四川四环锌锗科技有限公司 一种从高硫含铜物料中综合回收有价金属的方法
CN116040772A (zh) * 2023-03-27 2023-05-02 长春黄金研究院有限公司 高铜酸性水治理产渣无害化的处理方法
CN116040772B (zh) * 2023-03-27 2023-08-15 长春黄金研究院有限公司 高铜酸性水治理产渣无害化的处理方法
CN116463497A (zh) * 2023-04-20 2023-07-21 云南驰宏锌锗股份有限公司 一种高铜高铅硫化锌精矿氧压浸出方法

Also Published As

Publication number Publication date
CN111225988A (zh) 2020-06-02
CN111225988B (zh) 2022-04-29

Similar Documents

Publication Publication Date Title
CN109234522B (zh) 一种钴硫精矿综合回收处理方法
CN101225476B (zh) 从铅冰铜中回收铜的工艺
WO2020062145A1 (zh) 硫化铜精矿的氧压浸出方法及铜冶炼方法
CN102994747B (zh) 一种从高铅铜锍中回收金属铜的工艺
CN113897491B (zh) 一种综合高效处理锌浸出渣的方法
CN104531988B (zh) 一种难处理复杂多金属矿的回收工艺
CN110184482B (zh) 一种含锗次氧化锌粉浸出工艺
CN113846214B (zh) 一种湿法炼锌生产中含锌物料的处理方法
CN107630146A (zh) 镍回收方法
CN113215399B (zh) 一种硫化镍精矿的氧压浸出方法
CN111235404A (zh) 一种铜萃余液生产氢氧化钴的除杂方法
CN113416843A (zh) 一种硫化镍精矿的超细磨-氧压浸出工艺
CN116568831A (zh) 一种从高冰镍中提取硫酸镍的方法
CN103952562A (zh) 一种铁矾渣综合利用方法
CN114990351B (zh) 一种锌精矿与铜白烟尘协同处理的方法
CN115029562B (zh) 一种湿法炼锌过程分离铜锗的方法
CN107099669A (zh) 一种含铜渣中高效清洁除砷的方法
CN105110300A (zh) 一种含硫化锰的复合锰矿提取锰及硫的方法
CN105399132B (zh) 一种用黄铜炉渣和含锌烟道灰制备碱式氯化铜及碱式氯化锌的工艺
CN104947145B (zh) 一种高铅铜锍氧压浸出电积工艺平衡酸的方法
CN103993170A (zh) 从铜铅锌砷锑混合精矿中回收金属的方法
CN109913647B (zh) 一种回收铋中矿中铜、锌的湿法处理方法
CN117926027A (zh) 一种红土镍矿石的综合利用方法
CN115011810B (zh) 一种提升锌焙烧矿中铜回收率的浸出工艺
CN105886797A (zh) 一种从多金属硫化物原料中制备海绵铟的方法

Legal Events

Date Code Title Description
121 Ep: the epo has been informed by wipo that ep was designated in this application

Ref document number: 18934803

Country of ref document: EP

Kind code of ref document: A1

NENP Non-entry into the national phase

Ref country code: DE

122 Ep: pct application non-entry in european phase

Ref document number: 18934803

Country of ref document: EP

Kind code of ref document: A1