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WO2018161653A1 - 一种包裹型复杂氧化铜矿回收利用的方法 - Google Patents

一种包裹型复杂氧化铜矿回收利用的方法 Download PDF

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Publication number
WO2018161653A1
WO2018161653A1 PCT/CN2017/114280 CN2017114280W WO2018161653A1 WO 2018161653 A1 WO2018161653 A1 WO 2018161653A1 CN 2017114280 W CN2017114280 W CN 2017114280W WO 2018161653 A1 WO2018161653 A1 WO 2018161653A1
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Prior art keywords
copper
copper oxide
leaching
minutes
wrapped
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English (en)
French (fr)
Inventor
文书明
刘媛媛
朱景和
黄草明
刘丹
邓久帅
沈海英
王伊杰
陈瑜
白旭
张谦
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Kunming University of Science and Technology
China Nonferrous Metal Mining Group Co Ltd
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Kunming University of Science and Technology
China Nonferrous Metal Mining Group Co Ltd
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Priority to AU2017403278A priority Critical patent/AU2017403278B2/en
Publication of WO2018161653A1 publication Critical patent/WO2018161653A1/zh
Anticipated expiration legal-status Critical
Ceased legal-status Critical Current

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    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
    • B03B7/00—Combinations of wet processes or apparatus with other processes or apparatus, e.g. for dressing ores or garbage
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D1/00—Flotation
    • B03D1/001—Flotation agents
    • B03D1/004—Organic compounds
    • B03D1/008—Organic compounds containing oxygen
    • 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
    • C22B15/0063—Hydrometallurgy
    • C22B15/0065—Leaching or slurrying
    • C22B15/0067—Leaching or slurrying with acids or salts thereof
    • C22B15/0071—Leaching or slurrying with acids or salts thereof containing sulfur
    • 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
    • C25C1/12—Electrolytic production, recovery or refining of metals by electrolysis of solutions of copper
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D2201/00—Specified effects produced by the flotation agents
    • B03D2201/02—Collectors
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D2203/00—Specified materials treated by the flotation agents; Specified applications
    • B03D2203/02—Ores
    • B03D2203/04—Non-sulfide ores
    • 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 a method for recycling and utilizing a wrapped complex copper oxide ore, belonging to the technical field of ore metallurgy.
  • Copper oxide minerals mainly include malachite, chrysocolla, azurite, cuprite, ferrous copper, imaginary malachite, combined copper, etc., mainly with gangue minerals such as silicate, carbonate and iron oxide.
  • gangue minerals such as silicate, carbonate and iron oxide.
  • For independent oxidized copper ore such as malachite, chrysocolla, azurite, cuprite, epochrite, etc., it can be recovered by flotation of sulphide xanthate, which has high binding rate and dense symbiosis with silicate minerals. Copper ore can be recovered by means of sulfuric acid leaching, solid-liquid separation and extraction electrowinning.
  • ammonia leaching, solid-liquid separation and extraction electrowinning can be used.
  • Method of recycling For the illusion of malachite, part of the copper oxide ore is closely symbiotic with limonite, hematite and biotite, and some copper oxide ore is surrounded by iron and silicate minerals. The method utilized makes this part of the copper oxide ore resources not effectively recycled.
  • Conventional vulcanization-xanthate flotation method is difficult to recover such copper oxide minerals because part of the copper oxide ore is encapsulated by hematite, limonite, biotite, and copper oxide minerals cannot be dissociated by monomers, vulcanizing agents and The collector xanthate is difficult to contact with the surface of the copper oxide mineral and cannot be floated to recover this part of the copper oxide mineral.
  • Conventional ammonia leaching-extraction-electrowinning technology cannot effectively recover this part of copper oxide ore because ammonia cannot damage the structure of iron oxide ore and biotite, and it is difficult to contact with the encapsulated copper oxide mineral, and it is not effective to leach this part of copper oxide. mine.
  • Application method of high-integration rate carbonate gangue type oxygen-sulfur mixed copper with application number 201010178875.2 is for oxygen-sulfur mixed copper ore with high bonding rate and high content of calcium-magnesium carbonate gangue minerals.
  • the method can also not be used to treat the encapsulated copper oxide ore because the free copper oxide mineral in the encapsulated copper oxide ore is rare, and the sulfide yellow drug flotation has no good effect, while the calcium and magnesium are floating.
  • the selection will cause some of the iron-containing minerals to enter the carbonate minerals, resulting in the loss of copper minerals.
  • a polymer bridge flotation method combining copper impregnation body is a combined copper impregnation body which cannot be recovered by conventional flotation, and adopts polymer bridging agent, copper ion bridge ion and xanthate bridge.
  • the collector is adsorbed by the polymer bridge ion on the surface of the copper-impregnated body, and the copper ion is adsorbed on the bridged adsorbent on the surface, and the collector xanthate anion is adsorbed on the bridged copper ion.
  • the surface of the copper-impregnated body is made hydrophobic to achieve flotation. This method can not be effectively used for the recovery and utilization of the encapsulated copper oxide ore. The reason is that the polymer of the compound of the overseas Chinese cannot be combined with the copper oxide ore which is surrounded by gangue minerals such as hematite, limonite and biotite.
  • the HSC flotation method recovers the copper oxide mineral.
  • Some refractory minerals have achieved good results through warming flotation.
  • Qiu Xianyang et al. studied the kinetics of refining and flocculation of rhombohedrite.
  • Changing the surface properties of the smithsonite, forming a part of the surface of the zinc sulfide on the surface, is beneficial to the adsorption of the amine collector on the surface; heating can accelerate the flotation rate of the renosol, shorten the flotation time, and reduce the dosage of the agent.
  • the dispersion and control of the slime is beneficial to the flotation and the flotation effect.
  • Zhu Congjie studied the effect of slime on the flotation behavior of zinc oxide minerals. The sludge was adsorbed by flotation reagents, and the cover on the surface of the zinc ore and the micro-dissolution affected the uplift of the sphalerite, and the effect was less than 5 micron. For the biggest.
  • the object of the present invention is to provide a method for recycling and utilizing a packaged complex copper oxide ore.
  • a packaged complex copper oxide ore which is difficult to float and recover and has a low direct acid leaching rate, a vulcanization, a yellow drug flotation, a fatty acid combined flotation is used.
  • Free copper oxide, magnetic recovery of copper-bearing iron minerals and biotite minerals, while obtaining low-grade copper-bearing tailings, and copper-containing coarse concentrates are recovered by high temperature pressure leaching to recover copper resources.
  • the copper in the low-grade copper-containing tailings is recovered by the waste heat and residual acid of the high-temperature pressure leaching slurry, and the high-efficiency recycling of the wrapped complex copper oxide ore which cannot be directly floated and directly acid-leached is realized.
  • a method for recycling a packaged complex copper oxide ore is carried out as follows:
  • the distribution ratio of copper in the free copper oxide ore is less than 40%, and the distribution ratio of copper wrapped in limonite, hematite and biotite is 40% to 50%.
  • the ore is firstly crushed and ground. The fineness of the grinding is -0.074mm, and the mass percentage is 75% to 90%.
  • the slurry after grinding enters the mixing tank to adjust the pulp. The concentration of the pulp is 35% ⁇ 40.
  • step (2) recovering the copper-coated iron and biotite minerals from the flotation tailings of step (1) with a strong magnetic field magnetic separator with a magnetic induction strength of 1.0T to 1.6T, and obtaining a low-grade copper-containing concentrate.
  • Magnetic separation tailings
  • step (3) Combining the flotation coarse concentrate of step (1) with the magnetic separation concentrate of step (2) to obtain a copper-containing coarse concentrate, and the copper-containing coarse concentrate is in a liquid-solid ratio of 2 to 3:1.
  • Sulfuric acid leaching in the autoclave the pH value of the leaching slurry is controlled at 1.0 to 1.5, and high temperature and high pressure steam is introduced from the bottom of the autoclave for heating, and the temperature of the slurry is controlled by the amount of steam to be 100 to 130 degrees Celsius, and the leaching time is 90. Minutes to 120 minutes;
  • step (3) The leaching slurry of step (3) is discharged from the top of the pressurized kettle, mixed with the low-grade copper-containing magnetic separation tailings, and the slurry is continuously stirred to leaching the copper oxide for 60 minutes to 90 minutes, the leaching is finished, the solid-liquid separation, the leaching solution
  • the electrowinning copper is obtained by extracting electrowinning.
  • the hematite, limonite, and biotite inclusions contain 1.6 to 2.8% of copper.
  • the encapsulated copper oxide ore has a copper grade of 1.0% to 2.0%.
  • the xanthate is isoamyl and butyl xanthate.
  • the slurry heating in the autoclave is achieved by passing high temperature and high pressure steam to the bottom of the autoclave.
  • the free copper oxide ore and copper-bearing iron mineral can be floated at one time by using a combined collector;
  • the high-temperature pressure leaching slurry is mixed with the low-grade copper-containing tailings, and the leaching is continued, and the waste heat and residual acid of the high-temperature pressure leaching are fully utilized to further recover the copper resources lost in the tailings and improve the recovery rate of copper.
  • Figure 1 is a process flow diagram of the present invention.
  • Embodiment 1 is a diagrammatic representation of Embodiment 1:
  • the combined rate of encapsulated complex copper oxide ore is less than 10%, the copper grade is 2.0%, the distribution of copper in free copper oxide is 35%, and the hematite, limonite and biotite inclusions contain 2.0 to 2.8% of copper.
  • the distribution of copper in hematite, limonite and biotite is 50%, and quartz and kaolinite are the main gangue minerals.
  • the fineness of grinding is -0.074mm, and the mass percentage is 75%.
  • the slurry after grinding enters the mixing tank to adjust the pulp.
  • the concentration of pulp is 35%, and it is dry per ton.
  • the amount of minerals is first added with 800 g of sodium sulfide, 1000 g of inhibitor water glass is added, and the mixture is stirred for 4 minutes to 6 minutes.
  • the collector is 400 g of isoamyl xanthate, 300 g of hydroxamic acid, 500 g of sodium oleate, and stirred for 4 minutes. ⁇ 6 minutes, after a rough selection and a second sweep, the flotation coarse concentrate and flotation tailings are finally produced.
  • step (1) The flotation tailings of step (1) are recovered by a strong magnetic field magnetic separator with a magnetic induction intensity of 1.6T.
  • the unselected iron mineral was selected to obtain a magnetic separation concentrate, and at the same time, a magnetic separation tailing of 0.5% copper was obtained.
  • step (3) Combining the flotation coarse concentrate of step (1) with the step (2) magnetic separation concentrate to obtain a copper-containing coarse concentrate, and the copper-containing coarse concentrate is pressurized under the condition of a liquid-solid ratio of 2:1.
  • the sulfuric acid leaching is carried out in the kettle, and the pH value of the leaching slurry is controlled at 1.0 to 1.2.
  • the high temperature and high pressure steam is introduced from the bottom of the pressure vessel to be heated, and the temperature of the slurry is controlled by the steam addition amount to 120 to 130 degrees Celsius, and the leaching time is 120 minutes.
  • the top of the leaching slurry pressurizer is discharged, mixed with the low-grade copper-containing tailings, and the leaching is continued for 90 minutes.
  • the leaching is completed, the solid-liquid separation is performed, and the leaching solution is obtained by extracting electrowinning.
  • the overall recovery rate of copper is 90%.
  • Embodiment 2 is a diagrammatic representation of Embodiment 1:
  • the combined rate of encapsulated complex copper oxide ore is less than 10%, the copper grade is 1.5%, the distribution of copper in free copper oxide is 30%, hematite, limonite, biotite contains 1.8 to 2.5% copper, and red iron The distribution of copper in ore, limonite and biotite is 47%, and quartz and kaolinite are the main gangue minerals.
  • the fineness of grinding is -0.074mm, and the mass percentage is 80%.
  • the slurry after grinding enters the mixing tank to adjust the pulp.
  • the mass concentration of the pulp is 38%, and it is dry per ton.
  • the amount of minerals is first added with 700 g of sodium sulfide, 1200 g of inhibitor water glass is added, stirred for 4 minutes to 6 minutes, collector butyl xanthate 300 g, hydroxamic acid 250 g, oxidized paraffin soap 400 g, stirred for 4 minutes ⁇ After 6 minutes, after a rough selection and a second sweep, the flotation coarse concentrate and flotation tailings are finally produced.
  • the copper-bearing coarse concentrate is leached in the autoclave under the condition of liquid-solid ratio of 2.5:1, and the pH of the slurry is leached.
  • the value is controlled at 1.2 to 1.3, and the high temperature and high pressure steam is introduced from the bottom of the autoclave to be heated, and the temperature of the slurry is controlled by the steam addition amount to be 110 to 120 degrees Celsius, and the leaching time is 100 minutes.
  • the overall recovery rate of copper is 86%.
  • Embodiment 3 is a diagrammatic representation of Embodiment 3
  • the combined rate of encapsulated complex copper oxide ore is less than 10%, the copper grade is 1.0%, the distribution of copper in free copper oxide is 39%, hematite, limonite and biotite contain 1.6-2.0% copper. Mine, limonite, black The distribution of copper in mica is 40%, and quartz and kaolinite are the main gangue minerals.
  • the fineness of grinding is -0.074mm, and the mass percentage is 90%.
  • the slurry after grinding enters the mixing tank to adjust the slurry.
  • the concentration of pulp is 40%, and it is dried per ton.
  • the amount of minerals is first added with 600 g of sodium sulfide, 1000 g of inhibitor water glass is added, and the mixture is stirred for 4 minutes to 6 minutes.
  • the collector is 200 g of isoamyl xanthate, 200 g of hydroxamic acid, 300 g of oxidized paraffin soap, and stirred for 4 minutes. ⁇ 6 minutes, after a rough selection and a second sweep, the flotation coarse concentrate and flotation tailings are finally produced.
  • the leaching slurry is pressurized and discharged from the top of the kettle, mixed with the low-grade copper-containing tailings, and the leaching is continued for 60 minutes.
  • the leaching is completed, and the solid-liquid separation is performed.
  • the leaching solution is obtained by extracting electrowinning.
  • the overall recovery rate of copper is 80%.

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Abstract

一种包裹型复杂氧化铜矿回收利用的方法，对浮选难以回收、直接酸浸浸出率低的包裹型氧化铜矿，采用硫化，黄药、脂肪酸盐和羟肟酸联合浮选游离氧化铜，磁选回收矿石中的包裹铜矿物的铁质和黑云母矿物，分离出大量低品位含铜尾矿。含包裹型铜的粗精矿通过高温加压搅拌浸出强化回收其中的铜矿资源，高温浸出矿浆与低品位含铜尾矿混合继续进行余热余酸浸出尾矿中的铜矿物，使不能直接浮选回收和直接酸浸回收的包裹型氧化铜矿得以高效回收利用。经济有效地解决了包裹型复杂氧化铜矿资源加工利用的技术难题。

Description

一种包裹型复杂氧化铜矿回收利用的方法 技术领域
本发明涉及一种包裹型复杂氧化铜矿回收利用的方法，属于选矿冶金技术领域。
背景技术
氧化铜矿物主要有孔雀石、硅孔雀石、蓝铜矿、赤铜矿、黑铜矿、假象孔雀石、结合铜等，主要与硅酸盐、碳酸盐、氧化铁等脉石矿物共伴生。对于独立的孔雀石、硅孔雀石、蓝铜矿、赤铜矿、黑铜矿等氧化铜矿，可以通过硫化黄药浮选法回收，对于结合率高，与硅酸盐矿物致密共生的氧化铜矿，可以采用硫酸浸出、固液分离、萃取电积的方式回收利用，而对于结合率低，与碳酸盐矿物致密共生的氧化铜矿，可以采用氨浸、固液分离、萃取电积的方法回收。但对于以假象孔雀石为主，部分氧化铜矿与褐铁矿、赤铁矿、黑云母致密共生，部分氧化铜矿被铁质和硅酸盐矿物包裹的氧化铜矿，至今还没有高效回收利用的方法，致使这部分氧化铜矿资源没有得到有效回收利用。
常规的硫化—黄药浮选法难以回收这种氧化铜矿物，原因在于部分氧化铜矿被赤铁矿、褐铁矿、黑云母包裹，氧化铜矿物不能单体解离，硫化剂和捕收剂黄药类难以与氧化铜矿物表面接触，不能浮选回收这部分氧化铜矿物。常规的氨浸—萃取—电积技术不能有效回收这部分氧化铜矿，原因在于氨不能破坏氧化铁矿、黑云母的结构，难以与包裹的氧化铜矿物接触，不能有效浸出这部分氧化铜矿。常规的酸浸技术用于处理这种氧化铜矿，由于需要通过矿浆加温，促进赤铁矿、黑云母矿物晶格变形，硫酸才能扩散进入赤铁矿、褐铁矿、黑云母内部，浸出被包裹的氧化铜矿物，所以，由于原矿品位低，矿量大，矿浆加温至摄氏60度甚至更高时，浸出成本高，没有经济效益。
申请号为201010178875.2的一种高结合率碳酸盐脉石型氧硫混合铜的选冶方法，是针对结合率高、钙镁碳酸盐脉石矿物含量高的氧硫混合铜矿,先通过浮选回收其中的硫化铜矿物和游离氧化铜矿物,浮选尾矿用脂肪酸反浮选其中的 钙镁碳酸盐矿物,得到含钙镁碳酸盐矿物低,含结合铜的中矿,再添加硫酸搅拌浸出结合铜,固液分离后的含铜溶液通过冶金方法获得铜产品。该方法也不能用于处理这种包裹型氧化铜矿，其原因在于，该包裹型氧化铜矿中的游离氧化铜矿物很少，硫化黄药浮选没有好的效果，而钙镁反浮选将使部分含铁矿物进入碳酸盐矿物之中，导致铜矿物的损失。
申请号为201210201306.4的一种结合铜浸染体的高分子桥联浮选方法，是针对常规浮选不能回收的结合铜浸染体，采用高分子桥联剂、铜离子桥联离子、黄药桥联捕收剂，通过高分子桥联剂离子在结合铜浸染体表面发生多原子吸附，铜离子在表面上吸附的桥联剂上再吸附，捕收剂黄药阴离子在桥联铜离子上吸附，造成结合铜浸染体表面疏水而实现浮选。该方法也不能有效用于该包裹型氧化铜矿的回收利用，原因在于高分子侨联剂分子不能与被赤铁矿、褐铁矿、黑云母等脉石矿物包裹的氧化铜矿作用，难以通过侨联浮选方式回收该氧化铜矿物。
有些难选的矿物通过加温浮选，取得了较好的效果，比如：邱显扬等人对菱锌矿加温硫化浮选动力学进行了研究，得出菱锌矿加温硫化过程可以直接改变菱锌矿表面的性质，在其表面形成部分硫化锌表面，有利于胺类捕收剂在其表面吸附；加温能使菱锌矿硫化浮选速率加快，缩短浮选时间，药剂用量减少，可浮性增加，回收率上升【邱显扬等，菱锌矿加温硫化浮选动力学研究，有色金属(选矿部分)，2007(1)：24～26】。过建光等人对湖南柿竹园钨矿加温浮选工艺进行了改造，实践表明，新的加温浮选方法应用后，流程结构更趋合理，工艺操作更加简洁，取消了脱药和脱硫工序，每年可节约大量的电耗，彻底解决了脱药时造成的金属流失，白钨精矿质量、回收率均得到提高【过建光等，柿竹园钨加温浮选工艺改造实践，有色金属(选矿部分),2002(6)：13～14】。
矿泥的分散和控制有利于浮选的进行和提高浮选效果。朱从杰研究了矿泥对氧化锌矿物浮选行为的影响，矿泥通过吸附浮选药剂，在菱锌矿表面的罩盖以及微量溶解影响菱锌矿的上浮，而且以小于5微米矿泥的影响为最大。同时得出添加少量六偏磷酸钠和水玻璃以及使用超声波处理可降低矿泥的影响【朱从杰，矿泥对氧化锌矿物浮选行为的影响，矿产综合利用，2005(2):7～11】。丰奇成等人对新疆泥质难选氧化铜矿进行了浮选试验，研究得出：通过添加高效组合矿泥抑制剂CHO+A22有效地抑制了矿泥在浮选过程中的上浮，解决了浮选过程泡沫多 且矿浆粘性大的问题，使整个浮选工艺顺畅进行，最终获得了铜品位18.18％，铜回收率为75.04％的指标【丰奇成等，矿产综合利用，2011(6):21～24，49】。
但对于本发明所涉及的包裹型氧化铜矿，极其难选，单一的加温，或者是单一的分散和抑制矿泥，均不能获得令人满意的效果。
发明内容
本发明的目的提供一种包裹型复杂氧化铜矿回收利用的方法，对难以浮选回收、直接酸浸浸出率低的包裹型复杂氧化铜矿，采用硫化，黄药浮选、脂肪酸联合浮选游离氧化铜物，磁选回收含铜铁质矿物和黑云母矿物，同时获得低品位含铜尾矿，包裹型含铜粗精矿通过高温加压浸出回收其中的铜矿资源。低品位含铜尾矿中的铜通过高温加压浸出矿浆的余热和余酸继续浸出得以回收，实现不能直接浮选回收和直接酸浸回收的包裹型复杂氧化铜矿的高效回收利用。
本发明通过以下技术方案来实现：
一种包裹型复杂氧化铜矿回收利用的方法按以下步骤进行：
(1)对于结合率小于10％，游离氧化铜矿中铜的分布率小于40％，褐铁矿、赤铁矿、黑云母中包裹铜的分布率40％～50％的包裹型复杂氧化铜矿石，首先进行碎矿和磨矿，磨矿细度为-0.074mm质量百分含量占75％～90％，磨矿后的矿浆进入搅拌桶调浆，矿浆质量百分浓度35％～40％，按照每吨干矿量加入硫化钠600克～800克、加入抑制剂水玻璃1000克～1500克，搅拌4分钟～6分钟，加入捕收剂黄药200克～400克、羟肟酸200克～300克脂肪酸类捕收剂油酸钠或氧化石蜡皂300克～500克，搅拌4分钟～6分钟，经过一次粗选、二次扫选，产出浮选粗精矿和浮选尾矿；
(2)将步骤(1)的浮选尾矿用磁感应强度为1.0T～1.6T的强磁场磁选机回收包裹铜的铁质和黑云母矿物获得磁选精矿，同时获得低品位含铜磁选尾矿；
(3)将步骤(1)的浮选粗精矿和步骤(2)的磁选精矿合并获得含铜粗精矿，含铜粗精矿在液固比为2～3:1的条件下在加压釜中进行硫酸浸出，浸出矿浆pH值控制在1.0～1.5，从加压釜底部通入高温高压蒸汽进行加温，通过蒸汽加入量控制矿浆温度摄氏100度～130度，浸出时间90分钟～120分钟；
(4)将步骤(3)的浸出矿浆从加压釜顶部放出，与低品位含铜磁选尾矿混合，继续搅拌矿浆浸出氧化铜60分钟～90分钟，浸出结束，固液分离，浸出液 采用萃取电积获得电积铜。
所述的赤铁矿、褐铁矿、黑云母包裹体含铜1.6～2.8％。
所述的包裹型氧化铜矿石含铜品位为1.0％～2.0％。
所述的黄药为异戊基和丁基黄药。
加压釜中的矿浆加热通过在加压釜底部通入高温高压蒸汽来实现。
本发明具有以下优点和积极效果：
(1)采用组合捕收剂可以一次浮选游离氧化铜矿和含铜铁质矿物；
(2)采用强磁选可以回收部分包裹铜的弱磁性铁矿物和黑云母；
(3)浮选和磁选联合可以分离出大量低品位含铜尾矿，减少进入高温加压酸浸流程的矿石量，降低硫酸消耗和降低加温成本；
(4)高温加压浸出矿浆与低品位含铜尾矿混合，继续浸出，充分利用高温加压浸出的余热和余酸，进一步回收损失于尾矿中的铜资源，提高铜的回收率。
附图说明
图1为本发明的工艺流程图。
具体实施方式
本领域技术人员将会理解，下列实施例仅用于说明本发明，而不应视为限定本发明的范围。实施例中未注明具体技术或条件者，按照本领域内的文献所描述的技术或条件或者按照产品说明书进行。所用试剂或仪器未注明生产厂商者，均为可以通过购买获得的常规产品。
实施例一：
包裹型复杂氧化铜矿结合率小于10％，含铜品位2.0％，游离氧化铜中的铜的分布率为35％，赤铁矿、褐铁矿、黑云母包裹体含铜2.0～2.8％，赤铁矿、褐铁矿、黑云母中铜的分布率50％，石英和高岭石为主要脉石矿物。
(1)首先进行碎矿和磨矿，磨矿细度为-0.074mm质量百分含量占75％，磨矿后的矿浆进入搅拌桶调浆，矿浆质量百分浓度35％，按照每吨干矿量先加入硫化钠800克，加入抑制剂水玻璃1000克，搅拌4分钟～6分钟，捕收剂异戊基黄药400克，羟肟酸300克，油酸钠500克，搅拌4分钟～6分钟，经过一次粗选、二次扫选，最终产出浮选粗精矿和浮选尾矿。
(2)将步骤(1)的浮选尾矿用磁感应强度为1.6T的强磁场磁选机回收浮 选没有回收的铁质矿物获得磁选精矿，同时获得含铜0.5％的磁选尾矿。
(3)将步骤(1)的浮选粗精矿和步骤(2)磁选精矿合并获得含铜粗精矿，含铜粗精矿在液固比为2:1的条件下在加压釜中进行硫酸浸出，浸出矿浆pH值控制在1.0～1.2，从加压釜底部通入高温高压蒸汽进行加温，通过蒸汽加入量控制矿浆温度摄氏120度～130度，浸出时间120分钟。
(4)将浸出矿浆加压釜顶部放出，与低品位含铜尾矿混合，继续搅拌浸出90分钟，浸出结束，固液分离，浸出液采用萃取电积获得电积铜。
铜的综合回收率为90％。
实施例二：
包裹型复杂氧化铜矿结合率小于10％，含铜品位1.5％，游离氧化铜中的铜的分布率为30％，赤铁矿、褐铁矿、黑云母含铜1.8～2.5％，赤铁矿、褐铁矿、黑云母中铜的分布率47％，石英和高岭石为主要脉石矿物。
(1)首先进行碎矿和磨矿，磨矿细度为-0.074mm质量百分含量占80％，磨矿后的矿浆进入搅拌桶调浆，矿浆质量百分浓度38％，按照每吨干矿量先加入硫化钠700克，加入抑制剂水玻璃1200克，搅拌4分钟～6分钟，捕收剂丁基黄药300克，羟肟酸250克，氧化石蜡皂400克，搅拌4分钟～6分钟，经过一次粗选、二次扫选，最终产出浮选粗精矿和浮选尾矿。
(2)将浮选尾矿用磁感应强度为1.4T的强磁场磁选机回收铁质矿物获得磁选精矿，同时获得含铜0.4％的磁选尾矿。
(3)将浮选粗精矿和磁选精矿合并获得含铜粗精矿，含铜粗精矿在液固比为2.5:1的条件下在加压釜中进行硫酸浸出，浸出矿浆pH值控制在1.2～1.3，从加压釜底部通入高温高压蒸汽进行加温，通过蒸汽加入量控制矿浆温度摄氏110度～120度，浸出时间100分钟。
(4)将浸出矿浆加压釜顶部放出，与低品位含铜尾矿混合，继续搅拌浸出80分钟，浸出结束，固液分离，浸出液采用萃取电积获得电积铜。
铜的综合回收率为86％。
实施例三：
包裹型复杂氧化铜矿结合率小于10％，含铜品位1.0％，游离氧化铜中的铜的分布率为39％，赤铁矿、褐铁矿、黑云母含铜1.6～2.0％，赤铁矿、褐铁矿、黑 云母中铜的分布率40％，石英和高岭石为主要脉石矿物。
(1)首先进行碎矿和磨矿，磨矿细度为-0.074mm质量百分含量占90％，磨矿后的矿浆进入搅拌桶调浆，矿浆质量百分浓度40％，按照每吨干矿量先加入硫化钠600克，加入抑制剂水玻璃1000克，搅拌4分钟～6分钟，捕收剂异戊基黄药200克，羟肟酸200克，氧化石蜡皂300克，搅拌4分钟～6分钟，经过一次粗选、二次扫选，最终产出浮选粗精矿和浮选尾矿。
(2)将浮选尾矿用磁感应强度为1.0T的强磁场磁选机回收浮选没有回收的铁质矿物获得磁选精矿，同时获得含铜品位0.3％磁选尾矿。
(3)将浮选粗精矿和磁选精矿合并获得含铜粗精矿，含铜粗精矿在液固比为3:1的条件下在加压釜中进行硫酸浸出，浸出矿浆pH值控制在1.3～1.5，从加压釜底部通入高温高压蒸汽进行加温，通过蒸汽加入量控制矿浆温度摄氏100度～110度，浸出时间90分钟。
(4)将浸出矿浆加压从釜顶部放出，与低品位含铜尾矿混合，继续搅拌浸出60分钟，浸出结束，固液分离，浸出液采用萃取电积获得电积铜。
铜的综合回收率为80％。
以上显示和描述了本发明的基本原理、主要特征和本发明的优点。本行业的技术人员应该了解，本发明不受上述实施例的限制，上述实施例和说明书中描述的只是说明本发明的原理，在不脱离本发明精神和范围的前提下，本发明还会有各种变化和改进，这些变化和改进都落入要求保护的本发明范围内。本发明要求保护范围由所附的权利要求书及其等效物界定。

Claims (5)

  1. 一种包裹型复杂氧化铜矿回收利用的方法，其特征在于按以下步骤进行：
    (1)对于结合率小于10％，游离氧化铜矿中铜的分布率小于40％，褐铁矿、赤铁矿、黑云母中包裹铜的分布率40％～50％的包裹型复杂氧化铜矿石，首先进行碎矿和磨矿，磨矿细度为-0.074mm质量百分含量占75％～90％，磨矿后的矿浆进入搅拌桶调浆，矿浆质量百分浓度35％～40％，按照每吨干矿量加入硫化钠600克～800克，加入抑制剂水玻璃1000克～1500克，搅拌4分钟～6分钟，捕收剂黄药200克～400克，羟肟酸200克～300克，脂肪酸类捕收剂300克～500克，搅拌4分钟～6分钟，经过一次粗选、二次扫选，产出浮选粗精矿和浮选尾矿；
    (2)将步骤(1)的浮选尾矿用磁感应强度为1.0T～1.6T的强磁场磁选机回收包裹铜的铁质和黑云母矿物获得磁选精矿，同时获得低品位含铜磁选尾矿；
    (3)将步骤(1)的浮选粗精矿和步骤(2)的磁选精矿合并获得含铜粗精矿，含铜粗精矿在液固比为2～3:1的条件下在加压釜中进行硫酸浸出，浸出矿浆pH值控制在1.0～1.5，矿浆温度摄氏100度～130度，浸出时间90分钟～120分钟；
    (4)将步骤(3)的浸出矿浆从加压釜顶部放出，与低品位含铜尾矿混合搅拌浸出60分钟～90分钟，浸出结束，固液分离，浸出液采用萃取电积获得电积铜。
  2. 根据权利要求1所述的包裹型复杂氧化铜矿回收利用的方法，其特征在于，所述的赤铁矿、褐铁矿、黑云母包裹体含铜1.6～2.8％。
  3. 根据权利要求1所述的包裹型复杂氧化铜矿回收利用的方法，其特征在于，所述的包裹型氧化铜矿石含铜品位为1.0％～2.0％。
  4. 根据权利要求1所述的包裹型复杂氧化铜矿回收利用的方法，其特征在于，所述的黄药为异戊基和丁基黄药。
  5. 根据权利要求1所述的包裹型复杂氧化铜矿回收利用的方法，其特征在于，加压釜中的矿浆加热通过在加压釜底部通入高温高压蒸汽来实现。
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