WO2017185946A1 - 一种处理低品位红土镍矿的方法及其选矿方法 - Google Patents

一种处理低品位红土镍矿的方法及其选矿方法 Download PDF

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WO2017185946A1
WO2017185946A1 PCT/CN2017/078993 CN2017078993W WO2017185946A1 WO 2017185946 A1 WO2017185946 A1 WO 2017185946A1 CN 2017078993 W CN2017078993 W CN 2017078993W WO 2017185946 A1 WO2017185946 A1 WO 2017185946A1
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ore
nickel
flotation
selection
laterite nickel
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PCT/CN2017/078993
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English (en)
French (fr)
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米建国
程建国
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上海鑫和镍业科技有限公司
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Publication of WO2017185946A1 publication Critical patent/WO2017185946A1/zh

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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
    • B03B9/00—General arrangement of separating plant, e.g. flow sheets
    • 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
    • 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
    • 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
    • B03D1/014—Organic compounds containing phosphorus
    • 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
    • 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
    • C22B1/26—Cooling of roasted, sintered, or agglomerated ores
    • 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
    • B03D2201/00—Specified effects produced by the flotation agents
    • B03D2201/04—Frothers
    • 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
    • B03D2201/06—Depressants
    • 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
    • B03D2203/025—Precious metal ores

Definitions

  • the invention belongs to the field of non-ferrous metallurgy, and particularly relates to a method for treating low-grade laterite nickel ore and a beneficiation method thereof.
  • Laterite nickel ore is a mineral resource formed by long-term weathering of nickel-bearing peridotite in tropical or subtropical regions, mainly distributed in tropical countries in the equatorial region.
  • the rotary kiln drying pre-reduction-electric furnace smelting ferronickel process is the mainstream process for processing laterite nickel ore at home and abroad (referred to as RKEF process).
  • the ferronickel products produced by this process are mainly used for stainless steel production, and have been produced in dozens of factories abroad.
  • the Srigongshan project invested by MCC For example, the Srigongshan project invested by MCC.
  • the main process of the process is:
  • the calcine is sent to the ore electric furnace, and the smelting is reduced at a high temperature of 1550 to 1600 ° C to produce about 8% of nickel-containing iron;
  • the LF refining furnace is further enriched and produced to produce high-grade ferronickel containing 20-25% of nickel.
  • the main advantages of the ferronickel process are short process, mature process and large-scale production.
  • the high-grade ferronickel produced can be used in the production of medium and high-grade stainless steel; it can handle various types of high-grade laterite nickel ore.
  • the main disadvantage is that the investment is large; the production process requires a large amount of coke and electricity, energy consumption is large, comprehensive energy consumption is high, and the cost is high; the demand for raw ore grade is high, and the original ore grade of >2% is generally required.
  • the main process of the process is:
  • the laterite nickel ore is dried to remove free water and crystal water;
  • the main advantages of the nickel-sulfur process are the low technical difficulty and low market threshold. It is still the main method for handling red earth nickel in small factories in China.
  • the main disadvantage is that the existing equipment manufacturers are unable to produce large-scale blast furnace equipment with low ice nickel, which makes the nickel crucible unable to be mass-produced; the equipment and process produce harmful gases such as sulfur dioxide, and the environmental pollution is serious; it requires a large amount of high-quality coke and has high energy consumption.
  • the small blast furnace reduction smelting process is basically consistent with the traditional blast furnace ironmaking process principle: the crushed laterite nickel ore batch is sintered, the sinter ore is mined into a small blast furnace to produce slag and nickel-containing pig iron. Different from blast furnace ironmaking, the operating conditions and the ratio of raw materials and auxiliary materials are different. The amount of iron slag in the blast furnace is small, and the amount of slag in the laterite nickel ore is large.
  • the main advantage of the small blast furnace reduction smelting process is that the market threshold is low, and the idle blast furnace ironmaking equipment can be utilized, and the limonite-type laterite nickel ore with high iron content is suitable.
  • the main disadvantage is that the blast furnace has low volume utilization rate, requires high-quality coke as fuel, and consumes a large amount of coke, and has high energy consumption.
  • the product contains low nickel, only 3 to 5%, and high impurity content such as sulfur and phosphorus, which can usually only be used for low Quality stainless steel; emits a large amount of harmful fluorine-containing gases during the smelting process.
  • the domestic small blast furnace reduction smelting process has basically been forced to stop production.
  • the reduction roasting ammonia leaching process is also called Caron process, which firstly reduces and calcines the mineral plus reducing agent, and then performs ammonia leaching.
  • the ammonia leaching process uses NH 3 and CO 2 to bring the nickel and cobalt nickel ammonia and cobalt ammonia complexes in the calcined ore into solution.
  • the process mainly deals with laterite nickel ore containing more than 10% MgO, about 1% nickel, and a less complicated state.
  • the main advantage is that the reagents can be recycled, consume a small amount of energy, and can comprehensively recover nickel and cobalt.
  • the main disadvantage is that the nickel and cobalt leaching rate is low, and it is only suitable for the treatment of limonite type laterite nickel ore. It is not suitable for the treatment of silicon magnesium nickel type and laterite nickel ore containing high cobalt. Due to these shortcomings, the development of the process is greatly limited.
  • the high-pressure acid leaching process also known as the HPAL process, is currently the mainstream process of wet processing in the world.
  • the basic process of the process is:
  • the laterite nickel ore enters the autoclave and is leached by high pressure (4 ⁇ 5MPa) and high temperature (230 ⁇ 260°C);
  • the main advantages of this process are low operating costs, low energy consumption, no exhaust emissions, and comprehensive recovery of cobalt.
  • the main disadvantages are high investment, high sulfuric acid consumption, strict requirements on equipment and materials; due to the influence of impurities such as magnesium on acid consumption, and the process economic indicators are mainly affected by the consumption of sulfuric acid, the requirements for magnesium in raw ore are more demanding.
  • the atmospheric pressure acid leaching process is a process proposed for the shortcomings of the high pressure acid leaching process, but due to some technical problems, it is difficult to solve the problem. At present, there is no large and mature plant in the world.
  • the main process is:
  • the raw ore is finely mixed with pulverized coal to form a ball, and the pellet is dried and calcined at a high temperature;
  • the advantage of this process is that the production process is simple and the production cost is low. 85% of the energy is supplied by coal, and 160-180kg of coal consumption per ton of coal, and more than 80% of the energy consumption of the igniting of the fire process electric furnace is provided by electric energy. ⁇ 600kwh, the energy costs of the two are very different.
  • this raw material is rare in mineral deposits, and it has been difficult to obtain in the case where high-quality raw materials have been occupied by the world's major nickel companies. Since the equipment is operated at high temperatures, there is a high demand for equipment manufacturing processes and materials, resulting in a large investment. Since the calcination is carried out under semi-melting, cobalt oxide in the raw material cannot be recovered in the process. Due to the existence of these problems, the Dajiangshan process technology has remained unstable for decades, and the production scale has been staying at an annual output of about 10,000 tons of nickel.
  • the present invention provides a method of treating low grade laterite nickel ore, comprising the steps of:
  • the flotation process is carried out.
  • the flotation process is rough selection, sweeping and selection, and the concentrate is obtained after flotation.
  • the ore dressing agent added by the rough selection has hydrochloric acid or sulfuric acid, copper sulfate and yellow.
  • the selected ore dressing agents are xanthate, black medicine and foaming agent.
  • the invention adopts a combined separation roasting-flotation fire method wet combination process, and can use a conventional ordinary rotary kiln, and can use any type of low grade (about 1% nickel) laterite nickel ore to produce high quality (containing about 30% of nickel). Or higher) nickel-iron alloy.
  • the invention also provides a beneficiation method for treating low-grade laterite nickel ore, comprising the flotation step:
  • the material enters the flotation system for rough selection, sweeping and selection operations, and each flotation operation is separated into two products, namely foam and underflow; wherein the material is first added with a beneficiation agent after stirring Rough selection operation; rough selection foam - coarse concentrate does not add any chemicals into the selection operation, the selected foam is the mineral processing product - concentrate, selected bottom stream - medium mine 1 return coarse selection operation constitutes closed-circuit beneficiation; rough selection underflow addition After the ore dressing agent enters the sweeping operation, sweeping the foam-middle mine 2 back to the rough selection operation to form a closed-circuit beneficiation, and sweeping the bottom stream for the final tailings to be discarded;
  • the ore dressing agents added by the rough selection include hydrochloric acid or sulfuric acid, copper sulfate, xanthate, black medicine and foaming agent, and the selected ore dressing agents are yellow medicine, black medicine and foaming agent.
  • the method of the invention is a production technology with low investment (less than all existing mainstream processes), low production cost (lower than all existing mainstream processes), simple process, and adaptability to various laterite nickel ore raw materials.
  • the method of the present invention for treating low grade laterite nickel ore comprises the following steps:
  • grinding after water quenching preferably grinding fineness of 0.05-0.1 mm accounted for 70-90%, more preferably 0.074mm accounted for 70-90%;
  • the flotation process is carried out.
  • the flotation process is rough selection, sweeping and selection, and the concentrate is obtained after flotation.
  • the ore dressing agent added by the rough selection has hydrochloric acid or sulfuric acid, copper sulfate and yellow.
  • the selected ore dressing agents are xanthate, black medicine and foaming agent.
  • the milled material is passed to a flotation system for rough selection, sweeping, and selection operations, and each flotation operation is separated into two products, namely, a foam and an underflow;
  • the material is first added with a beneficiation agent after being subjected to coarse selection operation; the rough selected foam-coarse concentrate is added to any selected operation without adding any chemicals, and the selected foam is a beneficiation product-concentrate, and the selected underflow-middle mine 1
  • the rough selection operation constitutes closed-circuit beneficiation; after the rough selection of the bottom stream is added to the ore dressing agent, it enters the sweeping operation, and the selected foam-middle mine 2 returns to the rough selection operation to form closed-circuit beneficiation, and the bottom stream is selected as the final tailings to be discarded.
  • the ore dressing agent is: added during rough selection: adjusting agent: hydrochloric acid or sulfuric acid 3000-5000 g/t, activator: copper sulfate 250-500 g/t, collector 1: xanthate 100 ⁇ 300g / t, collector 2: black medicine 50 ⁇ 150g / t, foaming agent: 2 # oil 30 ⁇ 80g / t; add in the sweep: collector 1: xanthate 100 ⁇ 150g / t, Collector 2: black medicine 50 ⁇ 100g / t, foaming agent: 2 # oil 50 ⁇ 100g / t; selected without adding any mineral processing agent; all the mineral processing agents added to the grinding material in the total amount of 3630 ⁇ 6380g / t, based on the total weight of the material after adding the mineral processing agent.
  • the concentrate obtained by flotation is melted to obtain a nickel-iron alloy; or, the concentrate obtained by flotation enters the separation and enrichment process of cobalt, and the concentrate after separation of cobalt is separated. Further, a nickel-iron alloy is obtained by melting, or electrolytic nickel is prepared. Separation and enrichment of cobalt, especially for concentrates obtained from red earth nickel ore raw materials with high cobalt content.
  • the concentrate is melted by a pyrometallurgical process (internal carbon ball roasting and melting process), that is, 8 to 12% of coke and 8 to 16% are blended in the concentrate.
  • Lime after pelletizing, is melted in an electric furnace at a temperature of 1500 to 1600 ° C to separate the nickel iron from the slag to obtain a nickel-iron alloy. According to the iron content of the laterite nickel ore, 40-60% nickel-iron can be produced.
  • the separation and enrichment of cobalt is carried out by a wet process: beneficiation product + sulfuric acid ⁇ (nickel sulfate, cobalt sulfate) + xanthate ⁇ flotation ⁇ (nickel xanthate, cobalt xanthate, Tailings (discarded)) ⁇ concentrate + sodium hydroxide ⁇ (nickel hydroxide, cobalt xanthate, ethyl xanthate (cycle return)) + sulfuric acid ⁇ (nickel sulfate, cobalt xanthate) ⁇ separation ⁇ product 1: nickel sulfate; product 2: cobalt xanthate. Further, nickel sulfate produces electrolytic nickel using a classical nickel electrolysis process, and cobalt xanthate uses a classical cobalt-cobalt process to produce cobalt hydroxide or cobalt carbonate.
  • the composite additive in step 1), consists of a reducing agent, a separating agent and other additives, preferably from coke or pulverized coal as a reducing agent, calcium chloride or chlorine as a separating agent Sodium hydride, additive 1 calcium sulphate and additive 2 calcium carbonate.
  • a reducing agent preferably from coke or pulverized coal as a reducing agent, calcium chloride or chlorine as a separating agent Sodium hydride, additive 1 calcium sulphate and additive 2 calcium carbonate.
  • the composition of the composite additive is: coke or pulverized coal 2 to 5%, calcium chloride or sodium chloride 10 to 25%, calcium sulfate 2 to 5%,
  • the calcium carbonate is 2 to 5%;
  • the total proportion of the composite additive added to the laterite nickel ore raw material is 16 to 40%, which is based on the total weight of the raw materials after the addition of the composite additive.
  • step 1) the resulting mixture can be granulated and then subjected to the treatment of step 2).
  • step 1) the obtained mixture is thoroughly mixed and granulated to have a particle diameter of 5 to 30 mm, and the obtained pellet is dried and then subjected to the treatment of the step 2).
  • the calcination employs a weak reducing atmosphere to a neutral reducing atmosphere.
  • the oxygen content in the calcined tail gas is from 0.5 to 1.5%.
  • Separation reduction roasting is selective composite roasting instead of direct reduction roasting, so various types of laterite nickel ore raw materials can be utilized without having to make excessive demands on the iron-containing titanium content of the raw material and other elements such as magnesium;
  • the invention adopts a combination of pyrolysis and wet separation method of composite separation roasting-flotation at medium and low temperatures, and the process is easy to realize; using conventional ordinary equipment, under the same construction scale, the construction investment is greatly reduced, and the production is lowered. Cost; basically no requirements for raw materials, can handle various types of low-grade raw materials.
  • the nickel-iron alloy produced by the invention has high grade, containing about 30% or more of nickel, and can contain 40-60% of nickel by melting of the concentrate; in particular, it can treat the laterite nickel ore with high cobalt content, and obtain nickel products efficiently. And cobalt products.
  • the process flow is: raw material processing ⁇ pellet drying (or powder) ⁇ separation roasting ⁇ water quenching ⁇ grinding ⁇ flotation ⁇ cobalt separation enrichment ⁇ concentration ⁇ melting ⁇ high quality nickel iron alloy.

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Abstract

一种处理低品位红土镍矿的方法，包括：1)向红土镍矿原料中加入复合添加剂，得混合物并干燥；2)干燥后进行还原焙烧；焙烧温度为850～1100℃，焙烧时间为30～90min；3)焙烧后的焙砂进行水淬；4)水淬后进行磨矿；5)磨矿后进入浮选流程，浮选过程为粗选、扫选和精选，浮选后得精矿。还提供了一种处理低品位红土镍矿的选矿方法，其包括浮选步骤。该处理低品位红土镍矿的方法采用复合离析焙烧-浮选的火法湿法结合工艺，可使用常规普通回转窑，可利用任何类型的低品位(含镍1％左右)红土镍矿生产高品质(含镍30％左右或更高)的镍铁合金。

Description

一种处理低品位红土镍矿的方法及其选矿方法 技术领域
本发明属于有色金属冶金领域，具体涉及一种处理低品位红土镍矿的方法及其选矿方法。
背景技术
红土镍矿是含镍橄榄岩在热带或亚热带地区经长期风化而成的一种矿产资源，主要分布在赤道地区的热带国家。
世界上红土镍矿的处理工艺归纳起来有三类：火法工艺、湿法工艺和火湿法结合工艺，现分别介绍如下：
1.火法工艺
1)回转窑干燥预还原-电炉熔炼镍铁工艺
回转窑干燥预还原-电炉熔炼镍铁工艺是目前国内外处理红土镍矿的主流工艺(简称RKEF工艺)，该工艺生产的镍铁产品主要用于不锈钢生产，在国外已有数十家工厂生产，例如中冶投资的缅甸达贡山项目。
该工艺的主体流程为：
(1)将红土镍矿破碎筛分到50～150mm，然后送干燥窑干燥，使矿石既不粘结也不太粉化；
(2)添加还原剂配料，送入煅烧回转窑，在700℃温度下干燥和预还原；
(3)还原后焙砂送入矿热电炉，在1550～1600℃高温下还原熔炼产出含镍约8％的粗镍铁；
(4)再经LF精炼炉吹炼富集产出含镍20～25％的高品位镍铁。
镍铁工艺的主要优点是流程短，工艺成熟，可实现大规模生产；生产的高品位镍铁可用于中高档不锈钢生产；可处理各种类型高品位的红土镍矿。主要缺点是投资大；生产过程中需大量的焦炭和电，能源消耗大，综合能耗高，成本高；对原矿品位要求高，一般要求＞2％的原矿品位。
2)镍锍工艺
该工艺的主体流程为：
(1)先将红土镍矿经过干燥脱去游离水和结晶水；
(2)然后添加硫磺、硫精矿或石膏等硫化剂，再加还原剂、助熔剂等辅助原 料，在鼓风炉内1600℃高温下，红土镍矿与硫化剂发生硫化反应，产出含镍约12％左右的低冰镍；
(3)再经过转炉锤炼成高冰镍。
镍硫工艺的主要优点是技术难度小，市场门槛低，目前仍是国内小型工厂处理红土镍的主要方法。主要缺点是现有设备生产商无法生产低冰镍的大型鼓风炉设备，导致镍锍无法大规模生产；该设备及工艺流程产生二氧化硫等有害气体，环境污染严重；需要大量优质焦炭，能耗高。
3)小高炉还原熔炼工艺
小高炉还原熔炼工艺与传统高炉炼铁工艺原理基本一致：将破碎后的红土镍矿配料烧结，烧结矿配矿进小高炉熔炼，产出炉渣和含镍生铁。与高炉炼铁不同点是操作条件和原辅料配比不同，高炉炼铁渣量少，红土镍矿小高炉渣量大。
小高炉还原熔炼工艺的主要优点是市场门槛低，可以利用闲置的现有高炉炼铁的淘汰设备，适宜处理含铁高的褐铁矿型的红土镍矿。主要缺点是高炉体积利用率低，需用优质焦炭作燃料，且焦炭消耗量大，能耗高；产品含镍低，只有3～5％，硫磷等杂质含量高，通常只能用于低品质不锈钢；在冶炼过程中排放大量有害含氟气体。随着国家环保政策的落实，目前国内的小高炉还原熔炼工艺已基本强制停产。
2.湿法工艺
1)还原焙烧氨浸工艺
还原焙烧氨浸工艺又称Caron工艺，该工艺先将矿物加还原剂进行还原焙烧，然后进行氨浸。其中氨浸过程为采用NH3及CO2使焙烧矿中的镍和钴镍氨和钴氨络合物进入溶液。该工艺主要处理含MgO大于10％，含镍1％左右的且赋存状态不太复杂的红土镍矿。主要优点是试剂可循环使用，消耗量小，能综合回收镍和钴。主要缺点是镍、钴浸出率低，只适合处理表层的褐铁矿型红土镍矿，对于处理硅镁镍型以及含钴高的红土镍矿均不适用。正是由于这些缺点，该工艺的发展受到很大限制。
2)高压酸浸工艺
高压酸浸工艺又称HPAL工艺，是目前国际上湿法处理的主流工艺，该工艺的基本过程为：
(1)红土镍矿经过破碎制浆后进入高压釜经高压(4～5MPa)、高温(230～260℃)硫酸浸出；
(2)浸出后进行固液分离，对浸出液进行中和、除铁；
(3)除铁后通过萃取进行镍钴分离，再根据不同需要进一步冶炼得到不同的镍钴产品。
该工艺的主要优点是操作成本低，能耗小，无废气排放，可实现钴的综合回收。主要缺点是投资高，硫酸消耗大，对设备和材料的要求比较严格；由于镁等杂质对于耗酸影响较大，而工艺经济指标主要受硫酸消耗量的影响，所以对原矿含镁的要求较高，一般适应处理含镁小于5％的红土镍矿。
3)常压酸浸工艺
常压酸浸工艺是针对高压酸浸工艺的缺点提出的一种工艺，但由于一些技术问题难于解决，目前国际上尚无大型成熟的工厂实际运行。
3.火法湿法结合工艺
目前世界上采用火法湿法结合工艺处理红土镍的工厂只有日本冶金公司的大江山冶炼厂。
主要工艺过程为：
(1)原矿磨细与煤粉混合造球，球团经干燥和高温还原焙烧；
(2)焙砂再磨细，矿浆进行选矿(重选和磁选)分离得到镍铁合金产品。
该工艺优点是生产过程简单，生产成本低，能源中的85％由煤提供，吨矿耗煤160～180kg，而火法工艺电炉熔炼的能耗80％以上由电能提供，吨矿耗电560～600kwh，两者能耗成本差别很大。
但是，该工艺存在很多问题。在回转窑内1250～1400℃高温下焙烧，物料在此温度下容易烧结而结窑，一旦结窑则生产过程无法顺利进行，所以运行大江山法要求很高的工艺技术水平。由于该法本质上是直接还原，在还原氧化镍的同时铁也被还原，若要得到高品位的镍铁，则对原料有很高的要求，即要求原料含镍＞2％，含铁10％左右，这种原料在矿藏中极少，在优质原料已经被世界各大镍业公司占据的情形下，已经很难得到。由于设备在高温下运行，所以对设备制造工艺和材质有很高的要求，由此造成投资较大。由于在半熔融下焙烧，所以原料中的氧化钴在该工艺中不能回收。由于这些问题的存在，几十年来大江山法工艺技术仍不稳定，生产规模一直停留在年产镍1万吨左右。
发明内容
本发明提供一种处理低品位红土镍矿的方法，包括以下步骤：
1)向红土镍矿原料中加入复合添加剂，得混合物并干燥；
2)干燥后进行还原焙烧；焙烧温度为850～1100℃，焙烧时间为30～90min；
3)焙烧后的焙砂进行水淬；
4)水淬后进行磨矿；
5)磨矿后进入浮选流程，浮选过程为粗选、扫选和精选，浮选后得精矿；浮选过程中，粗选添加的选矿药剂有盐酸或硫酸、硫酸铜、黄药、黑药和起泡剂，扫选添加的选矿药剂有黄药、黑药和起泡剂。
本发明采用复合离析焙烧-浮选的火法湿法结合工艺，可使用常规普通回转窑，可利用任何类型的低品位(含镍1％左右)红土镍矿生产高品质(含镍30％左右或更高)的镍铁合金。
本发明还提供一种处理低品位红土镍矿的选矿方法，包括浮选步骤：
使磨矿后物料进入浮选系统，进行粗选、扫选和精选作业，每一浮选作业分离为二个产品，即泡沫和底流；其中，所述物料先添加选矿药剂经搅拌后进行粗选作业；粗选泡沫-粗精矿不添加任何药剂进入精选作业，精选泡沫即为选矿产品-精矿，精选底流-中矿1返回粗选作业构成闭路选矿；粗选底流添加选矿药剂后进入扫选作业，扫选泡沫-中矿2返回粗选作业构成闭路选矿，扫选底流为最终尾矿予以抛弃；
浮选过程中，粗选添加的选矿药剂有盐酸或硫酸、硫酸铜、黄药、黑药和起泡剂，扫选添加的选矿药剂有黄药、黑药和起泡剂。
本发明的方法是一种低投资(低于现有全部主流工艺)、低生产成本(低于现有全部主流工艺)、工艺简单、能适应各种红土镍矿原料的生产技术。
具体实施方式
以下对本发明的技术方案作进一步的说明。在本发明中，如无特别说明，则所有操作均在室温、常压条件下实施；所有百分比为重量百分比。
本发明处理低品位红土镍矿的方法，包括以下步骤：
1)向红土镍矿原料中加入复合添加剂，得混合物并干燥；
2)干燥后进行还原焙烧；焙烧温度为850～1100℃，焙烧时间为30～90min；
3)焙烧后的焙砂进行水淬，水淬可防止焙砂氧化；
4)水淬后进行磨矿，优选磨矿细度为0.05-0.1mm的占70～90％，更优选0.074mm的占70～90％；
5)磨矿后进入浮选流程，浮选过程为粗选、扫选和精选，浮选后得精矿；浮选过程中，粗选添加的选矿药剂有盐酸或硫酸、硫酸铜、黄药、黑药和起泡剂，扫选添加的选矿药剂有黄药、黑药和起泡剂。
在具体的实施方案中，在步骤5)中，使磨矿后物料进入浮选系统，进行粗选、扫选和精选作业，每一浮选作业分离为二个产品，即泡沫和底流；其中，所述物料先添加选矿药剂经搅拌后进行粗选作业；粗选泡沫-粗精矿不添加任何药剂进入精选作业，精选泡沫即为选矿产品-精矿，精选底流-中矿1返回粗选作业构成闭路选矿；粗选底流添加选矿药剂后进入扫选作业，扫选泡沫-中矿2返回粗选作业构成闭路选矿，扫选底流为最终尾矿予以抛弃。
在具体的实施方案中，所述选矿药剂为：在粗选时添加：调整剂：盐酸或硫酸3000～5000g/t，活化剂：硫酸铜250～500g/t，捕收剂1：黄药100～300g/t，捕收剂2：黑药50～150g/t，起泡剂：2#油30～80g/t；在扫选时添加：捕收剂1：黄药100～150g/t，捕收剂2：黑药50～100g/t，起泡剂：2#油50～100g/t；精选不添加任何选矿药剂；所有加入磨矿后物料中的选矿药剂的总量为3630～6380g/t，均按加入选矿药剂后物料的总重计。
在具体的实施方案中，在步骤5)后，将浮选所得精矿进行熔分得镍铁合金；或者，将浮选所得精矿进入钴的分选富集过程，分离出钴后的精矿再进行熔分得镍铁合金，或制备得到电解镍。钴的分选富集，特别针对含钴量高的红土镍矿原料获得的精矿进行。
在更具体的实施方案中，所述精矿的熔分采用火法工艺(内配炭造球焙烧熔分工艺)，即在精矿中配入8～12％的焦炭和8～16％的石灰，造球后在电炉中在1500～1600℃的温度下熔分，使镍铁与渣分离，得镍铁合金。根据红土镍原矿含铁量，可产出含镍40～60％的镍铁。
在更具体的实施方案中，钴的分选富集采用湿法工艺：选矿产品+硫酸→(硫酸镍，硫酸钴)+乙黄药→浮选→(黄原酸镍，黄原酸钴，尾矿(抛弃))→精矿+氢氧化钠→(氢氧化镍，黄原酸钴↓，乙黄药(循环返回))+硫酸→(硫酸镍，黄原酸钴↓)→分离→产品1：硫酸镍；产品2：黄原酸钴。进一步，硫酸镍采用经典镍电解工艺产出电解镍，黄原酸钴采用经典沉钴工艺产出氢氧化钴或碳酸钴。
在具体的实施方案中，在步骤1)中，所述复合添加剂由还原剂、离析剂和其他添加剂组成，优选地，由作为还原剂的焦炭或煤粉、作为离析剂的氯化钙或氯化钠、添加剂1硫酸钙和添加剂2碳酸钙组成。离析剂可在工艺过程中循环使 用，理论上并未消耗。
在更具体的实施方案中，在步骤1)中，所述复合添加剂的组成为：焦炭或煤粉2～5％，氯化钙或氯化钠10～25％，硫酸钙2～5％，碳酸钙2～5％；加入红土镍矿原料中的复合添加剂的总比例为16～40％，均按加入复合添加剂后原料的总重计。
在具体的实施方案中，在步骤1)中，所得混合物可造粒后再进行步骤2)的处理。在更具体的实施方案中，在步骤1)中，将所得混合物充分混合后造粒，粒径为5～30mm，将所得球粒干燥后再进行步骤2)的处理。
在具体的实施方案中，在步骤2)中，焙烧采用弱还原气氛至中性还原气氛。例如使焙烧尾气中氧含量为0.5～1.5％。
本发明的优点是：
1)在850～1100℃中低温下还原焙烧而不是1250～1400℃下焙烧，将不会产生烧结而出现的结窑现象，使生产过程更容易控制；
2)离析还原焙烧为选择性复合焙烧而不是直接还原焙烧，所以可以利用各种类型的红土镍矿原料，而不必对原料含铁含钛以及其他元素如镁等提出过多要求；
3)采用中低温焙烧，在达到相同技术指标时，可以采用常规普通的回转窑焙烧而不必采用高品质材质和高技术水平制造的焙烧设备；
4)让矿物在正常状态下而不是半熔融状态下焙烧，所以在后续的工艺中可以回收原料中所含的氧化钴；
5)采用浮选而不是采用重选和磁选，有更高的分选富集效率。
综上，本发明采用中低温度下复合离析焙烧-浮选的火法湿法结合的工艺，工艺过程易于实现；采用常规普通设备，在同等建设规模下，大幅降低了建设投资，降低了生产成本；对原料基本没有要求，可以处理各种类型低品位原料。本发明产出的镍铁合金品位高，含镍30％左右或更高，通过精矿熔分含镍可达40～60％；尤其能处理含钴高的红土镍矿，且高效的得到镍产品和钴产品。
在一个具体的实施例中，工艺流程为：原料处理→造球干燥(或粉料)→离析焙烧→水淬→磨矿→浮选→钴分离富集→精矿→熔分→高品质镍铁合金。
最后说明的是，以上实施方案仅用以说明本发明的技术方案而非限制。本领域技术人员在不脱离本发明技术方案的宗旨和范围的情况下，对本发明的技术方案进行的修改或者等同替换，均应涵盖在本发明的权利要求范围当中。

Claims (10)

  1. 一种处理低品位红土镍矿的方法，其特征在于，包括以下步骤：
    1)向红土镍矿原料中加入复合添加剂，得混合物并干燥；
    2)干燥后进行还原焙烧；焙烧温度为850～1100℃，焙烧时间为30～90min；
    3)焙烧后的焙砂进行水淬；
    4)水淬后进行磨矿；
    5)磨矿后进入浮选流程，浮选过程为粗选、扫选和精选，浮选后得精矿；浮选过程中，粗选添加的选矿药剂有盐酸或硫酸、硫酸铜、黄药、黑药和起泡剂，扫选添加的选矿药剂有黄药、黑药和起泡剂。
  2. 根据权利要求1所述的处理低品位红土镍矿的方法，其特征在于，在步骤5)中，使磨矿后物料进入浮选系统，进行粗选、扫选和精选作业，每一浮选作业分离为二个产品，即泡沫和底流；其中，所述物料先添加选矿药剂经搅拌后进行粗选作业；粗选泡沫-粗精矿不添加任何药剂进入精选作业，精选泡沫即为选矿产品-精矿，精选底流-中矿1返回粗选作业构成闭路选矿；粗选底流添加选矿药剂后进入扫选作业，扫选泡沫-中矿2返回粗选作业构成闭路选矿，扫选底流为最终尾矿予以抛弃。
  3. 根据权利要求1或2所述的处理低品位红土镍矿的方法，其特征在于，所述选矿药剂为：
    在粗选时添加：调整剂：盐酸或硫酸3000～5000g/t，活化剂：硫酸铜250～500g/t，捕收剂1：黄药100～300g/t，捕收剂2：黑药50～150g/t，起泡剂：2#油30～80g/t；
    在扫选时添加：捕收剂1：黄药100～150g/t，捕收剂2：黑药50～100g/t，起泡剂：2#油50～100g/t；
    精选不添加任何选矿药剂；
    所有加入磨矿后物料中的选矿药剂的总重为3630～6380g/t，均按加入选矿药剂后物料的总重计。
  4. 根据权利要求1至3之一所述的处理低品位红土镍矿的方法，其特征在于，在步骤5)后，将浮选所得精矿进行熔分得镍铁合金；或者，将浮选所得精矿进入钴的分选富集过程，分离出钴后的精矿再进行熔分得镍铁合金，或制备得到电解镍。
  5. 根据权利要求4所述的处理低品位红土镍矿的方法，其特征在于，所述精 矿的熔分采用火法工艺，即在精矿中配入8～12％的焦炭和8～16％的石灰，造球后在电炉中在1500～1600℃的温度下熔分，使镍铁与渣分离，得镍铁合金。
  6. 根据权利要求1至5之一所述的处理低品位红土镍矿的方法，其特征在于，在步骤1)中，所述复合添加剂的组成为：
    焦炭或煤粉2～5％，
    氯化钙或氯化钠10～25％，
    硫酸钙2～5％，
    碳酸钙2～5％；
    加入红土镍矿原料中的复合添加剂的总比例为16～40％，均按加入复合添加剂后原料的总重计。
  7. 根据权利要求1至6之一所述的处理低品位红土镍矿的方法，其特征在于，在步骤1)中，将所得混合物充分混合后造粒，粒径为5～30mm，将所得球粒干燥后再进行步骤2)的处理。
  8. 根据权利要求1至7之一所述的处理低品位红土镍矿的方法，其特征在于，在步骤2)中，焙烧采用弱还原气氛至中性还原气氛。
  9. 一种处理低品位红土镍矿的选矿方法，其特征在于，包括浮选步骤：
    使磨矿后物料进入浮选系统，进行粗选、扫选和精选作业，每一浮选作业分离为二个产品，即泡沫和底流；其中，所述物料先添加选矿药剂经搅拌后进行粗选作业；粗选泡沫-粗精矿不添加任何药剂进入精选作业，精选泡沫即为选矿产品-精矿，精选底流-中矿1返回粗选作业构成闭路选矿；粗选底流添加选矿药剂后进入扫选作业，扫选泡沫-中矿2返回粗选作业构成闭路选矿，扫选底流为最终尾矿予以抛弃；
    浮选过程中，粗选添加的选矿药剂有盐酸或硫酸、硫酸铜、黄药、黑药和起泡剂，扫选添加的选矿药剂有黄药、黑药和起泡剂。
  10. 根据权利要求9所述的处理低品位红土镍矿的选矿方法，其特征在于，所述选矿药剂为：
    在粗选时添加：调整剂：盐酸或硫酸3000～5000g/t，活化剂：硫酸铜250～500g/t，捕收剂1：黄药100～300g/t，捕收剂2：黑药50～150g/t，起泡剂：2#油30～80g/t；
    在扫选时添加：捕收剂1：黄药100～150g/t，捕收剂2：黑药50～100g/t，起泡剂：2#油50～100g/t；
    精选不添加任何选矿药剂；
    所有加入磨矿后物料中的选矿药剂的总重为3630～6380g/t，按加入选矿药剂后物料的总重计。
PCT/CN2017/078993 2016-04-26 2017-03-31 一种处理低品位红土镍矿的方法及其选矿方法 WO2017185946A1 (zh)

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CN111921699A (zh) * 2020-06-29 2020-11-13 金堆城钼业汝阳有限责任公司 一种选钼尾矿的高效选铁新工艺
CN112226612A (zh) * 2020-10-29 2021-01-15 矿冶科技集团有限公司 一种低硫铜精矿的焙烧收尘工艺
CN113145304A (zh) * 2021-04-26 2021-07-23 昆明理工大学 一种电镀污泥原位还原硫化浮磁电联合梯度回收铜镍锌铁铬的方法
CN114100842A (zh) * 2021-11-02 2022-03-01 湖南有色郴州氟化学有限公司 一种低品位萤石降碳降硅提质方法
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CN115821058A (zh) * 2023-02-15 2023-03-21 矿冶科技集团有限公司 红土镍矿氯化挥发提取镍钴的方法
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CN108554646A (zh) * 2018-06-07 2018-09-21 北矿机电科技有限责任公司 一种浮选机的差异化配置系统
CN111921699A (zh) * 2020-06-29 2020-11-13 金堆城钼业汝阳有限责任公司 一种选钼尾矿的高效选铁新工艺
CN112226612A (zh) * 2020-10-29 2021-01-15 矿冶科技集团有限公司 一种低硫铜精矿的焙烧收尘工艺
CN113145304A (zh) * 2021-04-26 2021-07-23 昆明理工大学 一种电镀污泥原位还原硫化浮磁电联合梯度回收铜镍锌铁铬的方法
CN113145304B (zh) * 2021-04-26 2023-03-21 昆明理工大学 一种电镀污泥原位还原硫化浮磁电联合梯度回收铜镍锌铁铬的方法
US20240034639A1 (en) * 2021-07-20 2024-02-01 Guangdong Brunp Recycling Technology Co., Ltd. Nickel-iron wet treatment method and application thereof
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CN114100842A (zh) * 2021-11-02 2022-03-01 湖南有色郴州氟化学有限公司 一种低品位萤石降碳降硅提质方法
CN114100842B (zh) * 2021-11-02 2023-10-17 湖南有色郴州氟化学有限公司 一种低品位萤石降碳降硅提质方法
CN115055278A (zh) * 2022-05-25 2022-09-16 西北矿冶研究院 一种从混合型炉渣中综合回收铜及伴生金属的选矿方法
CN115055278B (zh) * 2022-05-25 2023-11-14 西北矿冶研究院 一种从混合型炉渣中综合回收铜及伴生金属的选矿方法
CN115821058A (zh) * 2023-02-15 2023-03-21 矿冶科技集团有限公司 红土镍矿氯化挥发提取镍钴的方法

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