WO2022037404A1 - 一种利用红土镍矿生产电池级硫酸镍盐的方法 - Google Patents

一种利用红土镍矿生产电池级硫酸镍盐的方法 Download PDF

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WO2022037404A1
WO2022037404A1 PCT/CN2021/110292 CN2021110292W WO2022037404A1 WO 2022037404 A1 WO2022037404 A1 WO 2022037404A1 CN 2021110292 W CN2021110292 W CN 2021110292W WO 2022037404 A1 WO2022037404 A1 WO 2022037404A1
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ore
nickel
low
nickel sulfate
iron
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French (fr)
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李长东
唐红辉
王春轶
裴新安
李兴对
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广东邦普循环科技有限公司
湖南邦普循环科技有限公司
湖南邦普汽车循环有限公司
宁德邦普循环科技有限公司
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Priority to US18/042,207 priority Critical patent/US11952288B2/en
Priority to EP21857499.4A priority patent/EP4190924A4/en
Publication of WO2022037404A1 publication Critical patent/WO2022037404A1/zh

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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
    • C22B3/06—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
    • C—CHEMISTRY; METALLURGY
    • C01—INORGANIC CHEMISTRY
    • C01G—COMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
    • C01G53/00—Compounds of nickel
    • C01G53/10—Sulfates
    • 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
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • 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
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C22B1/06—Sulfating roasting
    • 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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/005—Preliminary treatment of ores, e.g. by roasting or by the Krupp-Renn process
    • 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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/02—Obtaining nickel or cobalt by dry processes
    • 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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/02—Obtaining nickel or cobalt by dry processes
    • C22B23/025—Obtaining nickel or cobalt by dry processes with formation of a matte or by matte refining or converting into nickel or cobalt, e.g. by the Oxford process
    • 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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0407—Leaching processes
    • C22B23/0415—Leaching processes with acids or salt solutions except ammonium salts solutions
    • C22B23/043—Sulfurated acids or salts thereof
    • 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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
    • 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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
    • 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
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
    • C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
    • C22B3/3842—Phosphinic acid, e.g. H2P(O)(OH)
    • 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
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
    • C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
    • C22B3/3844—Phosphonic acid, e.g. H2P(O)(OH)2
    • 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
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
    • C22B3/384—Pentavalent phosphorus oxyacids, esters thereof
    • C22B3/3846—Phosphoric acid, e.g. (O)P(OH)3
    • 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 belongs to the field of non-ferrous metal metallurgy, and specifically relates to a method for producing battery-grade nickel sulfate by utilizing laterite nickel ore.
  • Laterite ore is a nickel oxide ore formed by large-scale long-term weathering leaching and metamorphism of nickel-containing peridotite in tropical or subtropical regions. Due to differences in geographical location, climatic conditions and weathering degree, the types of laterite ore around the world are not exactly the same. .
  • the weathering process generally produces layered deposits in which complete or most thorough weathering products exist near the surface, graduating to lesser degrees of weathering products with increasing depth, and finally terminating as unweathered rock at some deeper level.
  • a highly regolith layer typically has the majority of its nickel content finely distributed in finely divided goethite grains, the layer is commonly referred to as limonite, which generally contains a high proportion of iron and low proportions of silicon and magnesium.
  • the less weathered layers generally contain more nickel in various magnesium silicate minerals, such as serpentine. There may be many other nickel-containing silicate minerals in incompletely weathered zones. Partially weathered, high-magnesium zones are commonly referred to as saprolites or saprolites. It generally contains low proportions of iron and high proportions of silicon and magnesium. In some deposits there is another zone, usually between limonite and saprolite, containing mainly nontonite clays, called transition minerals. "Low-grade laterite ore” refers to laterite ore without saprolite ore, that is, laterite ore composed of limonite and transition ore.
  • limonite is the main component of laterite nickel ore, accounting for 65-75% of the total laterite ore, saprolite 15-25%, and transition ore 10%.
  • the difficulty in recovering nickel and cobalt from laterite nickel ore is that the useful components of nickel cannot usually be fully enriched by physical means before chemical treatment to separate useful metal components (such as nickel and cobalt), that is, it cannot be carried out by beneficiation technology. enrichment, which makes laterite nickel ore expensive to process. And because of the different mineral and chemical compositions in limonite and saprolite, these ores are generally not suitable for processing using the same processing technology. The search for ways to reduce the cost of processing laterite nickel ore has been ongoing for decades.
  • the well-known metallurgical methods for processing laterite nickel ore include pyrometallurgical process, hydrometallurgy process, and fire-wet combination process.
  • the pyrometallurgical process includes rotary kiln-electric furnace (RKEF) smelting ferronickel, sintering machine-blast furnace smelting ferronickel, briquetting machine-blast furnace smelting ferronickel, pellet-blast furnace smelting low nickel matte, rotary kiln electric furnace smelting low nickel matte, etc.
  • RKEF rotary kiln-electric furnace
  • the pyrometallurgical process is suitable for processing saprolite-type laterite nickel ore with high silicon-magnesium iron and low iron. This process can usually only produce nickel-iron or matte, and cannot recover cobalt. Its application is limited, and in the pyrometallurgical process, in order to obtain high-grade The iron content of the ore needs to be controlled. Although the blast furnace can process high iron ore, the product is generally low-nickel pig iron with poor quality. In order to obtain high-grade nickel-iron, a certain iron-nickel ratio needs to be controlled. In order to make slag by pyrometallurgy, the silicon-magnesium ratio of ore also needs Within a suitable range, the alumina content should not be too high, which limits the use of pyrotechnic ore.
  • leaching agents such as sulfuric acid and hydrochloric acid are generally used to leach valuable metals such as nickel and cobalt in laterite nickel ore, and then purification and electrolysis are used to obtain nickel sulfate or electrolytic nickel products.
  • the hydrometallurgical process is suitable for processing limonite.
  • Hydrometallurgical technologies include high pressure acid leaching and reduction roasting ammonia leaching, as well as atmospheric acid leaching and heap leaching processes that have appeared in recent years.
  • Heap leaching technology has a low leaching rate and is only suitable for processing laterite ore with high magnesium content.
  • the reduction roasting-ammonia leaching process is rarely used due to high energy consumption and long process flow.
  • Atmospheric pressure acid leaching technology is simple to operate and does not require expensive autoclave, but requires a large amount of acid to completely dissolve the minerals.
  • nickel ore with low magnesium and low iron is selected to reduce acid consumption, and the leaching solution contains various metal ions, which complicates the subsequent leaching and separation process.
  • the high pressure acid leaching process uses sulfuric acid to leach laterite nickel ore at high temperature and pressure. Under high temperature and high pressure conditions, the metal minerals in the ore are almost completely dissolved. The dissolved iron is rapidly hydrolyzed at the high temperature used to precipitate hematite, nickel, cobalt, etc. remain in solution, and after cooling the iron and silicon-containing leaching residue is concentrated by washing in a series of washes, the so-called countercurrent decantation washes It is concentrated in the circuit and separated from the solution containing nickel and cobalt. Therefore, the main purpose of the leaching process is achieved - to separate the nickel from the iron.
  • the advantages of the high-pressure acid leaching process are that the leaching rate of nickel and cobalt is high, the reaction speed is fast, and the reaction time is short.
  • the disadvantages of the high-pressure acid leaching process are also very prominent. It requires complex high-temperature, high-pressure autoclave and related equipment, and its installation and maintenance are very expensive.
  • the magnesium content in the ore is high, and the presence of magnesium will lead to a large increase in the consumption of sulfuric acid.
  • the invention relates to a technology for recovering nickel from laterite nickel ore. After the laterite nickel ore is crushed and ground, carbonaceous reducing agent, compound additives and laterite nickel ore are added in a certain proportion to be mixed and ground, and the ball is formed into balls by a ball egg molding machine. The pellets are 15-20mm, dried at 200-400 °C for 4-6 hours, reduced and roasted in a rotary kiln, and the temperature is controlled at 950-1300 °C.
  • nickel grade of the concentrate is limited by the iron content in the nickel ore. When the iron content is high, the nickel grade is low, and when the iron content is low, the amount of iron used as a trap is small, and the nickel recovery rate is low.
  • Nickel ores of different properties are distributed in layers, there is no obvious boundary between layers, and with the fluctuation of terrain, it is difficult to separate nickel ores of a certain nature in the mining process, resulting in Nickel ores of different properties are mixed with each other, which also brings certain challenges to the stability of the treatment process.
  • the purpose of the present invention is to provide a method for producing battery-grade nickel sulfate by using laterite nickel ore, which can obtain battery-grade nickel sulfate products.
  • a method of utilizing laterite nickel ore to produce battery-grade nickel sulfate comprising the following steps:
  • the difference between the battery-grade nickel sulfate and the electroplating-grade nickel sulfate is that the content of magnetic substances, cobalt, magnesium and silicon is relatively high, Mg content ⁇ 0.002, Si ⁇ 0.001, and magnetic substances ⁇ 0.08 %, Co ⁇ 0.002%.
  • the laterite nickel ore mainly includes the following components by mass percentage: 1.2-2% Ni, 15-40% Fe, 6-20% Mg, 0.03-0.25% Co and 10-40% Co %Si.
  • the lump ore and the silt ore are distributed according to particle size, the particle size of the lump ore is greater than 10 mm, and the particle size of the silt ore is less than 10 mm.
  • the crushing is to crush the lump ore to 1-8 mm.
  • the temperature of the heap leaching is 40°C-60°C, and the time is 30-50 days.
  • step (2) the specific operation of the heap leaching is as follows: placing the crushed lump ore into a heap leaching tank, and then leaching out the nickel in the ore by spraying and soaking sulfuric acid, A crude nickel sulfate solution was obtained.
  • the mass concentration of the dilute sulfuric acid is 1%-5%.
  • the separation is to separate high-chromium ore, low-iron and high-magnesium ore, and high-iron and low-magnesium ore by gravity beneficiation methods such as a groove washer, a spiral chute or a shaking table.
  • the high chromium ore can be directly sold as a finished product.
  • the high-chromium ore contains 30-40% of chromium and 0.1-0.2% of nickel; the low-iron and high-magnesium ore is mainly silicon-magnesium-nickel ore, and contains 1.5-2.1% of nickel, Magnesium 15-25%, iron 8-25%, silicon 35-50%; the high-iron and low-magnesium ore is mainly limonite, containing 0.8-1.3% nickel, 30-50% iron, 0.1-10% magnesium , Silicon 10-30%. Low-iron and high-magnesium ore and high-iron and low-magnesium ore have different particle sizes and properties.
  • the drying is to dry the low-iron high-magnesium ore to a moisture content of 17%-24%.
  • the roasting temperature is 800°C-1000°C, and the time is 1-3h.
  • the reduction temperature is 1400°C-1600°C, and the time is 2-4 h.
  • the reducing agent used in the reduction is at least one of coke, blue charcoal or anthracite.
  • the dosage of the reducing agent that is, the dosage of the reducing agent is 1%-5% of the dry ore
  • the nickel content of ferronickel is controlled, so that the grade of ferronickel reaches 20%-35%, which reduces the content of iron in ferronickel. content, reducing vulcanization and blowing costs.
  • the vulcanization temperature is 1100°C-1400°C, and the vulcanization time is 0.5-3h.
  • the nickel content of the low matte nickel is 15-30%.
  • nickel-sulfur-iron compounds containing more than 50% nickel are called high matte nickel, and below are called low matte nickel.
  • the pressure of the water extraction is 1-5MPa.
  • the temperature of the oxygen pressure leaching is 100°C-250°C, and the pressure is 1-5 Mpa.
  • the temperature of the pressure leaching is 170°C-260°C, and the pressure is 1-5Mpa.
  • the extraction adopts an acidic extractant to take Fe 3+ , Mn 2+ , Co 2+ , Mg 2+ , and Ca 2+ therein to obtain a nickel sulfate solution;
  • the acidic extractant is At least one of P204 (diisooctyl phosphate), P507 (2-ethylhexyl 2-ethylhexyl phosphate) or C272 (bis(2,4,4-trimethylpentyl)).
  • the specific extraction operation is as follows: under the conditions of temperature of 50°C-80°C and pH of 2 ⁇ 5, firstly extract Fe 3+ and Mn 2+ with P2O4, Then use one or both of P507 (2-ethylhexyl phosphate) and C272 (bis(2,4,4-trimethylpentyl)) to extract the Co 2+ , Mg 2+ and Ca 2+ to obtain nickel sulfate solution.
  • the present invention also provides the application of the above method in the separation and purification of nickel ore.
  • the method first separates high-magnesium and low-magnesium iron ore and low-magnesium high-iron ore, and then adopts the combined method of RKEF process, pressure leaching process and heap leaching process to treat laterite nickel ore, and can process high-iron and low-magnesium ore and low-magnesium ore at the same time.
  • High-magnesium iron ore and relatively low-grade bulk stone ore generally low-grade bulk stone ore is abandoned in mines), use the characteristics of different ores, process them with appropriate processes, and make full use of fire and wet methods. It can not only reduce the cost of ore processing, but also effectively utilize various ore resources, reduce sewage discharge, and save environmental protection.
  • the present invention makes full use of the advantages of the three technologies of RKEF process, pressure leaching process and heap leaching process, fuses them together, learns from each other, utilizes the characteristics of different ores, and uses suitable processes to process them, resulting in low production cost and comprehensive recovery rate of nickel and cobalt. reach more than 90%.
  • Low-magnesium ore adopts pressure leaching, which has low acid consumption, less magnesium-containing sewage, and is environmentally friendly; large stone ore is broken into 1-10mm, and its particles have good uniform permeability to avoid problems such as osmotic segregation, and heap leaching During the process, nickel is easily leached, the magnesium leaching rate is low, the slag can be comprehensively utilized, and the smelting cost is low.
  • RKEF smelting selects the high-nickel iron process with more than 20% nickel, which can discharge part of the iron in the RKEF process and reduce the cost for the subsequent PS converter nickel-iron sulfidation and low nickel matte blowing process.
  • RKEF process slag can be used as river sand to cast concrete, make cement bricks, etc.
  • heap leaching slag can be used to make cement foam bricks. This technology has low smelting cost, energy saving and environmental protection, and good comprehensive economic benefits. , will cause the increase of smelting costs and poor economic benefits.
  • Fig. 1 is the process flow diagram of utilizing laterite nickel ore to produce battery-grade nickel sulfate in Example 1.
  • the raw materials, reagents or devices used in the following examples can be obtained from conventional commercial channels unless otherwise specified, or can be obtained by existing known methods.
  • a method of utilizing laterite nickel ore to produce battery-grade nickel sulfate comprising the following steps:
  • low matte nickel is blown into high matte nickel with a nickel content of 70%, water is extracted into nickel beans, and then oxygen pressure leaching is carried out at 150 ° C and 2.5 MPa to obtain crude nickel sulfate solution B;
  • a method of utilizing laterite nickel ore to produce battery-grade nickel sulfate comprising the following steps:
  • a method of utilizing laterite nickel ore to produce battery-grade nickel sulfate comprising the following steps:
  • a method of utilizing laterite nickel ore to produce battery-grade nickel sulfate comprising the following steps:
  • the crude nickel sulfate solution in the steps (2) and (3) is first at a temperature of 60 ° C and under the condition of pH 4, using P204 (diisooctyl phosphate) to extract the Fe 3+ , Mn 2 + , and then use P507 (2-ethylhexyl 2-ethylhexyl phosphate) and C272 (bis(2,4,4 trimethylpentyl)) to extract and separate Co 2+ , Mg 2+ , Ca 2+ , The pure nickel sulfate solution is obtained, and then the battery grade nickel sulfate is obtained by evaporation and crystallization.
  • P204 diisooctyl phosphate
  • P507 2-ethylhexyl 2-ethylhexyl phosphate
  • C272 bis(2,4,4 trimethylpentyl)
  • Example 1 The difference between Example 1 and Comparative Example 1 is that in Comparative Example 1, the 0.15-1mm high-magnesium and low-iron ore in the sediment ore was not separated, and all of them were treated by pressure leaching, which would cause a large amount of magnesium to react with acid, acid
  • the consumption can reach 70t/t-Ni, and the smelting wastewater contains a lot of Mg ions, the sewage treatment cost is high, and the environment is polluted.
  • a method of utilizing laterite nickel ore to produce battery-grade nickel sulfate comprising the following steps:
  • the crude nickel sulfate solution in the steps (2) and (4) is first at a temperature of 60 ° C and under the condition of pH 4, using P2O4 (diisooctyl phosphate) to extract the Fe 3+ , Mn 2 + , and then use P507 (2-ethylhexyl 2-ethylhexyl phosphate) and C272 (bis(2,4,4 trimethylpentyl)) to extract and separate Co 2+ , Mg 2+ , Ca 2+ , The pure nickel sulfate solution is obtained, and then the battery grade nickel sulfate is obtained by evaporation and crystallization.
  • P2O4 diisooctyl phosphate
  • Example 2 The difference between Example 2 and Comparative Example 2 is that in Comparative Example 2, the high iron and low magnesium ore of 0.045-0.15mm was not separated, but the sediment ore would enter the RKEF production line uniformly to produce ferronickel, resulting in high iron content in the ore, It becomes difficult to produce high-grade ferronickel, and the discharge of a large amount of iron in the electric furnace slag (reduction) will take away part of the nickel, and the recovery rate of nickel will decrease. In the process of vulcanization and blowing of low-grade ferronickel, a large amount of iron needs to be discharged through slag making, the amount of nickel carried away by the slag increases at the same time, and the recovery rate of nickel decreases.

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Abstract

本发明公开了一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：将红土镍矿分选，得到块矿和泥沙矿；将块矿破碎，再进行堆浸处理，得到粗硫酸镍溶液A；将泥沙矿分离，得到高铬矿、低铁高镁矿、高铁低镁矿，将低铁高镁矿干燥、焙烧、还原、硫化，得到低冰镍；将低冰镍进行吹炼，水萃，再进行氧压浸出，得到粗硫酸镍溶液B；将高铁低镁矿进行压力浸出，得到粗硫酸镍溶液C；将上述粗硫酸镍溶液A、B、C进行萃取，再蒸发结晶即得电池级硫酸镍盐。本发明充分利用RKEF工艺、压力浸出工艺、堆浸工艺三种技术的优势，融合到一起，取长补短，利用不同矿石自身的特点，用合适的工艺处理，生产成本低，镍钴综合回收率达到90％以上。

Description

一种利用红土镍矿生产电池级硫酸镍盐的方法 技术领域
本发明属于有色金属冶金领域，具体涉及到一种利用红土镍矿生产电池级硫酸镍盐的方法。
背景技术
红土矿是由含镍橄榄岩在热带或亚热带地区经过大规模长期风化淋滤变质而成的氧化镍矿，由于存在地理位置、气候条件以及风化程度的差异，世界各地的红土矿类型不完全相同。风化过程一般产生层状沉积，其中在表面附近存在着完全的或最彻底的风化产物，随着深度增加渐变为程度较轻的风化产物，最后在某个更深处终止为未风化的岩石。高度风化层通常将其含有的大部分镍细微分布在细碎的针铁矿颗粒中，该层通常称为褐铁矿，它一般含有高比例的铁和低比例的硅和镁。风化较轻的层所含的镍一般更多地包含于各种硅酸镁矿物中，例如蛇纹石。不完全风化带中可能有很多其他含有镍的硅酸盐矿物。部分风化的高含镁带通常称为腐泥土或硅镁镍矿。它一般含有低比例的铁和高比例的硅和镁。在一些矿床中还有另一种通常处于褐铁矿和腐泥土之间的主要含有绿脱石粘土的带，称为过渡矿。“低品位红土矿”是指没有腐泥土矿的红土矿，也就是由褐铁矿和过渡矿组成的红土矿。通常情况下，褐铁矿为红土镍矿的主要组成部分，占红土矿总量的65-75％，腐泥土占15-25％，过渡矿占10％。从红土镍矿中回收镍、钴的困难之处在于，在进行化学处理分离金属有用成分(如镍和钴)之前通常不能通过物理方式充分富集镍的有用成分，即无法用选矿的技术进行富集，这使得红土镍矿的处理成本很高。并且由于褐铁矿和腐泥土矿中不同的矿物和化学组成，这些矿石通常不适于使用同一处理技术进行处理。几十年来一直在寻找降低处理红土镍矿的成本的方法。
公知的处理红土镍矿的冶金方法有火法冶炼工艺、湿法冶炼工艺，火湿结合工艺。
1、火法冶炼工艺
主要采用焦炭或兰炭等碳质材料在冶金炉内，进行还原或者硫化，产出镍铁或冰镍的冶金方法。火法冶炼工艺包括回转窑-电炉(RKEF)冶炼镍铁、烧结机-高炉冶炼镍铁、压球机-鼓风炉冶炼镍铁、球团-鼓风炉冶炼低冰镍、回转窑电炉冶炼低冰镍等技术方案。
火法冶炼工艺适合处理硅镁高铁低的腐泥土型红土镍矿，该工艺通常只能生产镍铁或冰镍，不能回收钴，其应用受到限制，且在火法冶炼过程中为了获得高品位的镍铁，需要控制 矿石中铁的含量。高炉虽然可以处理高铁矿石，但是产品一般是质量差的低镍生铁，要想获得高品位的镍铁，需要控制一定的铁镍比例，为了火法冶炼造渣需要，矿石的硅镁比也需要在合适范围之内，氧化铝含量不能太高，这限制了火法工艺矿石的使用范围。
2、湿法冶炼工艺
湿法冶炼工艺一般采用硫酸、盐酸等浸出剂将红土镍矿中的镍、钴等有价金属浸出出来，再用净化、电解等方法，获得硫酸镍或者电解镍产品。湿法冶金工艺适合处理褐铁矿。
湿法冶金技术包括高压酸浸和还原焙烧氨浸以及近年来出现的如常压酸浸、堆浸工艺等。堆浸技术浸出率较低，只适用于处理高镁含量的红土矿。还原焙烧一氨浸工艺由于能耗较高，工艺流程长而较少被采用。常压酸浸技术操作简单，不需使用昂贵的高压釜，但要使矿物完全溶解则所需酸耗量较大，一般选择低镁低铁的镍矿以降低酸耗，且浸出液中含有各种金属离子，使后续浸化分离工序变得复杂。高压酸浸工艺使用硫酸在高温和高压下浸出红土镍矿。在高温、高压条件下，矿石中的金属矿物几乎完全溶解。溶解的铁在所采用的高温下迅速水解为赤铁矿沉淀，镍、钴等留在溶液中，在冷却之后含铁和硅的浸出残渣通过在一系列洗涤浓缩，即所谓的逆流倾析洗涤回路中浓缩而从含镍、钴的溶液中分离。因此达到了浸出工艺的主要目的——将镍与铁分离。高压酸浸出工艺的优点是镍、钴浸出率高反应速度快、反应时间短铁在酸浸过程中理论上不消耗硫酸且水解产物为赤铁矿沉淀。但高压酸浸出工艺的缺点也很突出，它需要复杂的高温、高压的高压釜以及相关的设备，其安装与维护都很昂贵，高压浸出工艺只限于处理褐铁矿类的原料，因为腐泥土矿中镁含量较高，镁的存在会导致硫酸消耗量的大量增加。
3、火湿结合工艺
火法与湿法相结合的工艺处理氧化镍的工厂，目前世界上只有日本冶金公司的大江山冶炼厂，其主要工艺过程为原矿磨细与粉煤混合制团，团矿经干燥和高温还原焙烧，焙烧矿团再磨细，矿浆进行选矿重选和磁选分离得到镍铁合金产品。该工艺的最大特点是能耗中的能源由煤提供，因而生产成本低。而火法工艺电炉熔炼的能耗以上由电能提供，两者能耗成本差价很大。但是该工艺存在的问题还比较多，大江山冶炼厂虽经多次改进，工艺技术仍不够稳定，经过几十年其生产规模仍停留在年产镍万吨左右。
汪云华等人在中国专利文献中提出了不同类型红土镍矿的还原一磨选处理方法的技术方案(申请号200610163831.6)。本发明涉及一种从红土镍矿中回收镍的技术，红土镍矿经破碎和磨细、按一定比例，加入碳质还原剂、复合添加剂与红土镍矿混磨，用球蛋成型机制成球团15-20mm，在200-400℃干燥4-6小时，采用回转窑还原焙烧，温度控制在950-1300℃。还原焙烧后，进行粗破，然后按一定矿浆配比，进行湿法球磨后，采用摇床进行重选，重选 获得的镍精矿采用3000-5000高斯的磁选机再进行选别，便得到高品位的镍铁混合精矿，其含镍可达到7-15％。该技术方案具有原料适应性强、工艺流程短、环境友好，以煤作为主要能源，不用昂贵的电力作为能源等特点，为处理不同类型的红土镍矿提供了一种新的方法，具有良好的应用和推广前景。但是该方案在后续的应用中存在着以下有待改进的问题：(1)由于红土镍矿成分的波动，回转窑在焙烧过程中结圈情况严重。(2)由于整个还原过程在熔融条件下进行的，镍分布在硅酸盐晶格中且较为分散，冶金动力学条件导致部分还原的镍不能有效迁移汇聚并回收，镍的回收率极低。(3)精矿镍品位受限于镍矿中铁的含量。当铁含量高时，镍品位低，当铁含量低时，铁作为捕集剂量少，镍回收率低。
也有人提出采用RKEF和常压浸出联合法处理红土镍矿，但是因为常压浸出无法处理褐铁矿，只能选择铁低镁低的过渡层矿石，这样的矿石较难匹配。且常压浸出存在酸耗高、镍回收率低，冶炼污水量大难以处理，冶炼成本高等问题无法有效实施。
不同性质的镍矿虽然呈层状分布，但是层与层之间没有明显的界限，且随着地形的起伏，在开采过程中很难做到将某一性质的镍矿单独分离开来，导致不同性质的镍矿相互混杂在一起，也给处理工艺稳定性带来一定的挑战。
发明内容
本发明目的是提供一种利用红土镍矿生产电池级硫酸镍盐的方法，可以得到电池级硫酸镍盐产品，该工艺充分利用火法、湿法、堆浸这几种技术方案的优势，融合到一起，取长补短，生产成本低，环境友好，镍钴回收率高，有效利用镍矿资源，应用和推广前景广阔的特点。
为了实现上述目的，本发明采取以下技术方案：
一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：
(1)将红土镍矿进行分选，得到块矿和泥沙矿；
(2)将块矿破碎，再进行堆浸处理，得到粗硫酸镍溶液A；
(3)将泥沙矿分选，得到高铬矿、低铁高镁矿、高铁低镁矿，将低铁高镁矿干燥、焙烧、还原、硫化，得到低冰镍；
(4)将低冰镍进行吹炼，水萃，再进行氧压浸出，得到粗硫酸镍溶液B；
(5)将高铁低镁矿进行压力浸出，得到粗硫酸镍溶液C；
(6)将粗硫酸镍溶液A、粗硫酸镍溶液B和粗硫酸镍溶液C进行萃取，再蒸发结晶即得所述电池级硫酸镍盐。
优选地，步骤(6)中，所述电池级硫酸镍与电镀级硫酸镍的区别是对磁性物质、钴、镁、硅含量要求比较高，Mg含量<0.002，Si<0.001，磁性物质<0.08％，Co<0.002％。
优选地，步骤(1)中，所述红土镍矿主要包括以下质量百分比的组分：1.2-2％Ni、15-40％Fe、6-20％Mg、0.03-0.25％Co和10-40％Si。
优选地，步骤(1)中，所述块矿和的泥沙矿是按粒径分配的，块矿的粒径大于10mm，泥沙矿的粒径小于10mm。
优选地，步骤(2)中，所述破碎是将块矿破碎至1-8mm。
优选地，步骤(2)中，所述堆浸的温度为40℃-60℃，时间为30-50天。
优选地，步骤(2)中，所述堆浸具体的操作为：将粉碎后的块矿，放进堆浸池中，再采用喷淋、浸泡硫酸的方式，将矿石中的镍浸出来，得到粗硫酸镍溶液。
更优选地，所述稀硫酸的质量浓度为1％-5％。
优选地，步骤(3)中，所述分选是用槽式洗矿机、螺旋溜槽或摇床等重力选矿的方法，分离出高铬矿、低铁高镁矿、高铁低镁矿。所述高铬矿可以作为成品直接出售。
优选地，步骤(3)中，所述高铬矿含有铬30-40％，镍0.1-0.2％；所述低铁高镁矿是以硅镁镍矿为主，含有镍1.5-2.1％，镁15-25％，铁8-25％，硅35-50％；所述高铁低镁矿是以褐铁矿为主，含有镍0.8-1.3％，铁30-50％，镁0.1-10％，硅10-30％。低铁高镁矿和高铁低镁矿两种矿在粒度，性质也不一样。
优选地，步骤(3)中，所述干燥是将低铁高镁矿干燥至含水率为17％-24％。
优选地，步骤(3)中，所述焙烧的温度为800℃-1000℃，时间为1-3h。
优选地，步骤(3)中，所述还原的温度为1400℃-1600℃，时间为2-4h。
优选地，步骤(3)中，所述还原所用的还原剂为焦炭、兰炭或无烟煤中的至少一种。
利用降低还原剂的配入量，即还原剂的加入量为干矿的1％-5％，控制镍铁的含镍量，使得镍铁品位达到20％-35％，降低了镍铁中铁的含量，降低硫化和吹炼成本。
优选地，步骤(3)中，所述硫化的温度为1100℃-1400℃，硫化的时间为0.5-3h。
优选地，步骤(3)中，所述低冰镍的含镍量为15-30％。相对来说，含镍在50％以上的镍硫铁化合物称为高冰镍，以下的称为低冰镍。
优选地，步骤(4)中，所述水萃的压力为1-5MPa。
优选地，步骤(4)中，所述氧压浸出的温度为100℃-250℃，压力为1-5Mpa。
优选地，步骤(5)中，所述压力浸出的温度为170℃-260℃，压力为1-5Mpa。
优选地，步骤(6)中，所述萃取采用酸性萃取剂采取其中的Fe 3+、Mn 2+、Co 2+、Mg 2+、Ca 2+，得到硫酸镍溶液；所述酸性萃取剂为P204(磷酸二异辛酯)、P507(2-乙基己基磷酸2-乙基己基酯)或C272(二(2,4,4-三甲基戊基))中的至少一种。
更优选地，步骤(6)中，所述萃取具体的操作为：在温度为50℃-80℃和pH为2～5的条件下，先用P204萃取其中的Fe 3+、Mn 2+，再用P507(2-乙基己基磷酸2-乙基己基酯)、C272(二(2,4,4-三甲基戊基))中的一种或两种，萃取其中的Co 2+、Mg 2+、Ca 2+，得到硫酸镍溶液。
本发明还提供上述方法在镍矿分离提纯中的应用。
目前在矿石的选矿时，粒度参数的选定很重要，对于矿石的选矿，还要考虑矿石选矿后的性能，否则不能分离高镁低铁矿和低镁高铁矿，也就发挥不出各种工艺的优点，合适的粒度分界点，不仅需要大量的实验论证，由于每个矿山的矿石风化程度不一样，还要根据发明人的经验和对工艺的改进，重新做实验摸索参数，因此粒度分界点不是一个绝对的定值。本发明先将高镁低铁矿和低镁高铁矿分离，再采用RKEF工艺、压力浸出工艺、堆浸工艺三种技术联合的方法处理红红土镍矿，可以同时处理高铁低镁矿和低镁高铁矿以及品位比较低的大块石头矿(一般低品位大块石头矿在矿山都是废弃不用)，利用不同矿石自身的特点，用合适的工艺处理，充分利用和火法和湿法工艺各自优点，扬长避短，不但可以降低矿石处理成本，有效利用各种矿石资源，降低污水排放量，节约环保。
有益效果
1.本发明充分利用RKEF工艺、压力浸出工艺、堆浸工艺三种技术的优势，融合到一起，取长补短，利用不同矿石自身的特点，用合适的工艺处理，生产成本低，镍钴综合回收率达到90％以上。
2.低镁矿采用压力浸出，酸耗低，含镁污水量少，环境友好；大块石头矿破碎成1-10mm，利用其颗粒均匀渗透性好，避免造成渗透偏析等问题，且堆浸过程中镍容易浸出，镁浸出率低，渣可综合利用，冶炼成本低；高镁矿石采用RKEF处理，避免用湿法处理镁与酸反应造成酸耗高，产生大量含镁废水等问题。
3.RKEF冶炼选择含镍20％以上的高镍铁工艺，可以在RKEF流程中排出部分铁，给后面PS转炉镍铁硫化及低冰镍吹炼过程降低成本。
4.RKEF工艺炉渣可以充当河沙浇筑混凝土、做水泥砖等，堆浸渣可以做水泥泡沫砖，该技术冶炼成本低，节能环保，综合经济效益好，如果采用任何单一的方法处理多种矿石，都会造成冶炼成本的上升，经济效益差。
附图说明
图1为实施例1利用红土镍矿生产电池级硫酸镍盐的工艺流程图。
具体实施方式
为了让本领域技术人员更加清楚明白本发明所述技术方案，现列举以下实施例进行说明。需要指出的是，以下实施例对本发明要求的保护范围不构成限制作用。
以下实施例中所用的原料、试剂或装置如无特殊说明，均可从常规商业途径得到，或者可以通过现有已知方法得到。
实施例1
一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：
(1)将主要含1.75％Ni、18％Fe、15％Mg、0.03Co、30％Si的红土镍矿用颚式破碎机破碎后，再用反击破碎机配合振动筛，得到粒径>10mm块矿和粒径<10mm泥沙矿；
(2)将块矿破碎至1-8mm，堆入堆浸池进行喷淋质量浓度为2％的稀硫酸水溶液，在温度为50℃下进行浸出40天，得到粗硫酸镍溶液A；
(3)将泥沙矿分离，得到高铬矿、0.15-1mm的低铁高镁矿、0.045-0.15mm高铁低镁矿，将低铁高镁矿干燥至含水率为20％-24％、先在900℃下焙烧、再在1500℃下还原(加入焦炭、兰炭进行还原)、最后在1200℃下硫化1.5h，得到含镍量为18％的低冰镍；
(4)将低冰镍进行吹炼成含镍量为70％的高冰镍，水萃成镍豆，再在150℃和2.5MPa下进行氧压浸出，得到粗硫酸镍溶液B；
(5)将高铁低镁矿在200℃和2.5MPa下进行高压浸出，得到粗硫酸镍溶液C；
(6)将粗硫酸镍溶液A、粗硫酸镍溶液B和粗硫酸镍溶液C先在温度为60℃和pH为4的条件下，采用P204(磷酸二异辛酯)萃取其中的Fe 3+，Mn 2+，再用P507(2—乙基己基磷酸单2—乙基己基脂)和C272【二(2,4,4三甲基戊基)】萃取其中的Co 2+、Mg 2+，Ca 2+，得到纯硫酸镍溶液，再蒸发结晶即得电池级硫酸镍盐。
实施例1的硫酸镍产品中杂质量为：钴(Co)＝0.0003％，铁(Fe)＝0.0002％，镁(Mg)＝0.0001％，锰(Mn)＝0.0001％，锌(Zn)＝0.0002％，镍矿综合回收率92％(镍回收率＝产品中的镍量/投入矿石中镍量*100％)，硫酸镍冶炼成本10000美元/吨基镍。
实施例2
一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：
(1)将主要含1.2％Ni、40％Fe、6％Mg、0.2％Co、25％Si的红土镍矿用颚式破碎机破碎后，再用反击破碎机配合振动筛，得到粒径为>10mm块矿和粒径<10mm泥沙矿；
(2)将块矿破碎至1-8mm，堆入堆浸池进行喷淋质量浓度为2％的稀硫酸水溶液，在温度为50℃下进行浸出40天，得到粗硫酸镍溶液；
(3)将泥沙矿分离，得到高铬矿、0.15-1mm的低铁高镁矿、0.045-0.15mm高铁低镁矿，将低铁高镁矿干燥至含水率为20％-24％、先在900℃下焙烧、再在1500℃下还原、最后在1200℃下硫化1.5h，得到15％低冰镍；
(4)将低冰镍进行吹炼、水萃，再在150℃和2.5MPa下进行氧压浸出，得到粗硫酸镍溶液；
(5)将高铁低镁矿在200℃和2.5MPa下进行高压浸出，得到粗硫酸镍溶液；
(6)将粗硫酸镍溶液A、粗硫酸镍溶液B和粗硫酸镍溶液C先在温度为60℃和pH为4的条件下，采用P204(磷酸二异辛酯)萃取其中的Fe 3+，Mn 2+，再用P507(2-乙基己基磷酸2-乙基己基酯)和C272(二(2,4,4三甲基戊基))萃取分离Co 2+、Mg 2+，Ca 2+，得到纯硫酸镍溶液，再蒸发结晶即得电池级硫酸镍盐。
实施例2的硫酸镍产品中杂质量为：钴(Co)＝0.0005％，铁(Fe)＝0.0006％，镁(Mg)＝0.0001％，锰(Mn)＝0.0002％，锌(Zn)＝0.0002％，镍矿综合回收率94％，硫酸镍冶炼成本10500美元/吨基镍。
实施例3
一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：
(1)将主要含2.0％Ni、18％Fe、15％Mg、0.05％Co、35％Si的红土镍矿用颚式破碎机破碎后，再用反击破碎机配合振动筛，得到>10mm块矿和粒径<10mm泥沙矿；
(2)将块矿破碎至1-8mm，堆入堆浸池进行喷淋质量浓度为2％的稀硫酸水溶液，在温度为50℃下进行浸出40天，得到粗硫酸镍溶液；
(3)将泥沙矿分离，得到高铬矿、0.15-1mm的低铁高镁矿、0.045-0.15mm高铁低镁矿，将低铁高镁矿干燥至含水率为20％-24％、先在900℃下焙烧、再在1500℃下还原、最后在1200℃下硫化1.5h，得到28％低冰镍；
(4)将低冰镍进行吹炼、水萃，再在150℃和2.5MPa下进行氧压浸出，得到粗硫酸镍溶液；
(5)将高铁低镁矿在200℃和2.5MPa下进行高压浸出，得到粗硫酸镍溶液；
(6)将粗硫酸镍溶液A、粗硫酸镍溶液B和粗硫酸镍溶液C先在温度为60℃和pH为4的条件下，采用P204(磷酸二异辛酯)萃取其中的Fe 3+，Mn 2+，再用P507(2-乙基己基磷 酸2-乙基己基酯)和C272(二(2,4,4三甲基戊基))萃取分离Co 2+、Mg 2+，Ca 2+，得到纯硫酸镍溶液，再蒸发结晶即得电池级硫酸镍盐。
实施例3的硫酸镍产品中杂质量为：钴(Co)＝0.0003％，铁(Fe)＝0.0004％，镁(Mg)＝0.0006％，锰(Mn)＝0.0002％，锌(Zn)＝0.0001％，镍矿综合回收率94％，硫酸镍冶炼成本9850美元/吨基镍。
对比例1
一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：
(1)将主要含1.75％Ni、18％Fe、15％Mg、0.03Co、30％Si的红土镍矿用颚式破碎机破碎后，再用反击破碎机配合振动筛，得到粒径>10mm的块矿和粒径<10mm的泥沙矿；
(2)将块矿破碎，堆入堆浸池进行喷淋质量浓度为2％的稀硫酸水溶液，在温度为50℃下进行浸出40天，得到粗硫酸镍溶液；
(3)将泥沙矿在150℃和2.5MPa下进行氧压浸出，得到粗硫酸镍溶液；
(4)将步骤(2)和(3)中的粗硫酸镍溶液先在温度为60℃和pH为4的条件下，采用P204(磷酸二异辛酯)萃取其中的Fe 3+，Mn 2+，再用P507(2-乙基己基磷酸2-乙基己基酯)和C272(二(2,4,4三甲基戊基))萃取分离Co 2+、Mg 2+，Ca 2+，得到纯硫酸镍溶液，再蒸发结晶即得电池级硫酸镍盐。
对比例1的硫酸镍产品中杂质量为：钴(Co)＝0.0003％，铁(Fe)＝0.0004％，镁(Mg)＝0.0006％，锰(Mn)＝0.0002％，锌(Zn)＝0.0001％，镍矿综合回收率90％，硫酸镍冶炼成本12500美元/吨基镍。
实施例1与对比例1的区别在于，对比例1没有将泥沙矿中的0.15-1mm的高镁低铁矿石分离出来，全部采用压力浸出处理，会导致大量的镁与酸反应，酸耗可达70t/t-Ni，且冶炼废水中含有大量Mg离子，污水处理成本高，污染环境。
对比例2
一种利用红土镍矿生产电池级硫酸镍盐的方法，包括以下步骤：
(1)将主要含1.2％Ni、40％Fe、6％Mg、0.2％Co、25％Si的红土镍矿用颚式破碎机破碎后，再用反击破碎机配合振动筛，得到粒径>10mm的块矿和粒径<10mm的泥沙矿；
(2)将块矿破碎，堆入堆浸池进行喷淋质量浓度为2％的稀硫酸水溶液，在温度为50℃下进行浸出40天，得到粗硫酸镍溶液；
(3)将泥沙矿干燥至含水率为20％-24％、先在900℃下焙烧、再在1500℃下还原、最后在1200℃下硫化1.5h，得到10％低冰镍；
(4)将低冰镍进行吹炼、水萃，再在150℃和2.5MPa下进行氧压浸出，得到粗硫酸镍溶液；
(5)将步骤(2)和(4)中的粗硫酸镍溶液先在温度为60℃和pH为4的条件下，采用P204(磷酸二异辛酯)萃取其中的Fe 3+，Mn 2+，再用P507(2-乙基己基磷酸2-乙基己基酯)和C272(二(2,4,4三甲基戊基))萃取分离Co 2+、Mg 2+，Ca 2+，得到纯硫酸镍溶液，再蒸发结晶即得电池级硫酸镍盐。
对比例2的硫酸镍产品中杂质量为：硫酸镍产品中杂质量为：钴(Co)＝0.0005％，铁(Fe)＝0.0006％，镁(Mg)＝0.0001％，锰(Mn)＝0.0002％，锌(Zn)＝0.0002％，镍矿综合回收率87％，硫酸镍冶炼成本11500美元/吨基镍。
实施例2与对比例2的区别在于，对比例2没有将0.045-0.15mm的高铁低镁的矿石分离出来，而是泥沙矿将统一的进入RKEF生产线生产镍铁，导致矿石中铁含量高，生产高品位镍铁变得困难，且电炉渣(还原)中排出大量铁会带走部分镍，镍的回收率降低。低品位镍铁在硫化和吹炼过程中，需要通过造渣排出大量铁，渣带走的镍量同时增加，镍回收率降低。
以上对本发明提供的一种利用红土镍矿生产电池级硫酸镍盐的方法进行了详细的介绍，本文中应用了具体实施例对本发明的原理及实施方式进行了阐述，以上实施例的说明只是用于帮助理解本发明的方法及其核心思想，包括最佳方式，并且也使得本领域的任何技术人员都能够实践本发明，包括制造和使用任何装置或系统，和实施任何结合的方法。应当指出，对于本技术领域的普通技术人员来说，在不脱离本发明原理的前提下，还可以对本发明进行若干改进和修饰，这些改进和修饰也落入本发明权利要求的保护范围内。本发明专利保护的范围通过权利要求来限定，并可包括本领域技术人员能够想到的其他实施例。如果这些其他实施例具有不是不同于权利要求文字表述的结构要素，或者如果它们包括与权利要求的文字表述无实质差异的等同结构要素，那么这些其他实施例也应包含在权利要求的范围内。

Claims (10)

  1. 一种利用红土镍矿生产电池级硫酸镍盐的方法，其特征在于，包括以下步骤：
    (1)将红土镍矿进行分选，得到块矿和泥沙矿；
    (2)将块矿破碎，再进行堆浸处理，得到粗硫酸镍溶液A；
    (3)将泥沙矿分离，得到高铬矿、低铁高镁矿、高铁低镁矿，将低铁高镁矿干燥、焙烧、还原、硫化，得到低冰镍；
    (4)将低冰镍进行吹炼、水萃，再进行氧压浸出，得到粗硫酸镍溶液B；
    (5)将高铁低镁矿进行压力浸出，得到粗硫酸镍溶液C；
    (6)将粗硫酸镍溶液A、粗硫酸镍溶液B和粗硫酸镍溶液C进行萃取，再蒸发结晶即得所述电池级硫酸镍盐。
  2. 根据权利要求1所述的方法，其特征在于，步骤(1)中，所述红土镍矿主要包括以下质量百分比的组分：1.2-2％Ni、15-40％Fe、6-20％Mg、0.03-0.25％Co和10-40％Si。
  3. 根据权利要求1所述的方法，其特征在于，步骤(2)中，所述堆浸具体的操作为：将粉碎后的块矿，放进堆浸池中，再采用喷淋、浸泡硫酸的方式，将矿石中的镍浸出来，得到粗硫酸镍溶液；步骤(2)中，所述堆浸的温度为40℃-60℃，时间为30-50天。
  4. 根据权利要求1所述的方法，其特征在于，步骤(3)中，所述高铬矿含有铬30-40％、镍0.1-0.2％；所述低铁高镁矿是以硅镁镍矿为主，含有镍1.5-2.1％，镁15-25％，铁8-25％，硅35-50％；所述高铁低镁矿是以褐铁矿为主，含有镍0.8-1.3％，铁30-50％，镁0.1-10％，硅10-30％；步骤(6)中，所述电池级硫酸镍中Mg的含量<0.002％，Si<0.001％，磁性物质<0.08％，Co<0.002％。
  5. 根据权利要求1所述的方法，其特征在于，步骤(3)中，所述焙烧的温度为800℃-1000℃，时间为1-3h；步骤(3)中，所述还原的温度为1400℃-1600℃，时间为2-4h。
  6. 根据权利要求1所述的方法，其特征在于，步骤(3)中，所述还原所用的还原剂为焦炭、兰炭或无烟煤中的至少一种。
  7. 根据权利要求1所述的方法，其特征在于，步骤(3)中，所述硫化的温度为1100℃-1400℃，硫化的时间为0.5-3h；步骤(4)中，所述氧压浸出的温度为100℃-250℃，压力为1-5Mpa；步骤(5)中，所述压力浸出的温度为170℃-260℃，压力为1-5Mpa。
  8. 根据权利要求1所述的方法，其特征在于，步骤(6)中，所述萃取采用酸性萃取剂萃取其中的Fe 3+、Mn 2+、Co 2+、Mg 2+、Ca 2+，得到硫酸镍溶液；所述酸性萃取剂为磷酸二异辛酯、2-乙基己基磷酸2-乙基己基酯或二(2,4,4-三甲基戊基)中的至少一种。
  9. 根据权利要求8所述的方法，其特征在于，步骤(6)中，所述萃取具体的操作为：在温度为50℃-80℃和pH为2～5的条件下，先用磷酸二异辛酯萃取其中的Fe 3+、Mn 2+，再用2-乙基己基磷酸2-乙基己基酯、二(2,4,4-三甲基戊基)中的一种或两种，萃取其中的Co 2+、Mg 2+、Ca 2+，得到硫酸镍溶液。
  10. 权利要求1-9中任一项所述方法在镍矿分离提纯中的应用。
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