WO2010094161A1 - 盐酸全闭路循环法从红土镍矿中提取有价金属的方法 - Google Patents

盐酸全闭路循环法从红土镍矿中提取有价金属的方法 Download PDF

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Publication number
WO2010094161A1
WO2010094161A1 PCT/CN2009/000274 CN2009000274W WO2010094161A1 WO 2010094161 A1 WO2010094161 A1 WO 2010094161A1 CN 2009000274 W CN2009000274 W CN 2009000274W WO 2010094161 A1 WO2010094161 A1 WO 2010094161A1
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Prior art keywords
iron
leaching
nickel
chloride
sulfide
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PCT/CN2009/000274
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English (en)
French (fr)
Inventor
李新海
郭华军
胡启阳
王志兴
彭文杰
张云河
李向群
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中南大学
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Publication of WO2010094161A1 publication Critical patent/WO2010094161A1/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
    • C22B3/10—Hydrochloric acid, other halogenated 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
    • C22B23/0407—Leaching processes
    • C22B23/0415—Leaching processes with acids or salt solutions except ammonium salts solutions
    • C22B23/0423—Halogenated 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
    • C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
    • 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 extracting nickel and cobalt metal from laterite nickel ore and comprehensively utilizing valuable metals such as magnesium and iron, and belongs to the field of nonferrous metallurgy.
  • Laterite nickel ore is a nickel oxide ore that accounts for more than 65% of total nickel storage. Laterite nickel ore is divided into two major categories: limonite type and silicon magnesium nickel ore type.
  • the limonite type is located in the upper part of the deposit. It has high iron and low nickel, and low silicon and magnesium.
  • the cobalt content is relatively high and should be treated by hydrometallurgical process.
  • the silicon-magnesium-nickel ore is located in the lower part of the deposit.
  • the content of silicon and magnesium is relatively high, the iron content is low, and the cobalt content is also low.
  • the nickel content is relatively high and should be treated by pyrometallurgical process.
  • the ore that is in the middle transition can be either pyrometallurgy or wet metallurgy.
  • the traditional hydrometallurgical process of laterite nickel-nickel ore is divided into reduction;) "Ammonia leaching process and sulfuric acid pressure acid leaching process.
  • the former is pre-reduced by the inch ore, and then leached with ammonia; the latter is under pressurized conditions.
  • the leaching of nickel in the ore is carried out using sulfuric acid as a leaching agent.
  • the metallurgical process is also divided into a nickel-iron method and a smelting smelting method, which is mainly adapted to the surface mine.
  • the nickel-iron method is a reduction smelting method
  • the smelting smelting method is a method of smashing in the case of adding a vulcanizing agent.
  • the reduction-ammonia leaching method is rarely used because of environmental problems.
  • the pressurized sulfuric acid method has large investment, low cost, and has problems such as low leaching rate, inability to open magnesium, and high environmental pressure.
  • Most of the laterite nickel ore is high-grade low-grade nickel ore.
  • the process technology and equipment for extracting nickel-cobalt from laterite nickel ore are relatively backward, and the production scale is also small.
  • the environmental protection process is not enough, and it is far from satisfying the development of laterite nickel ore. Demand.
  • Another method for wet treatment of laterite W is to use hydrochloric acid leaching, leaching, neutralization, and sulfide precipitation to obtain a nickel-cobalt-rich sulfide, and the mother liquor obtained after the nickel is properly concentrated, and then calcined at a high temperature to produce hydrogen chloride.
  • the gas regenerates the hydrochloric acid.
  • the method has the characteristics of low investment, high resource utilization and environmental friendliness in the treatment of low-grade laterite nickel ore, and is a wet processing process with great development potential.
  • the object of the present invention is to provide a method for extracting valuable metals from laterite nickel ore by a full closed loop method of hydrochloric acid according to the deficiencies of the prior art, and a method for treating laterite nickel ore by wet chlorination leaching, which can be obtained from laterite. Extraction of valuable metals such as nickel and cobalt. High recovery of nickel and cobalt, closed cycle of water and acid, no waste water discharge, full utilization of nickel, cobalt, iron and magnesium resources, enabling efficient use of resources and clean production.
  • the hydrochloric acid full closed loop method extracts valuable metals from laterite nickel ore by processing the crushed laterite nickel ore as a raw material, and the process steps include chlorination leaching, leaching slag recovery iron, nickel cobalt Extraction, mother liquor atomization drying, roasting (regeneration of hydrochloric acid and comprehensive utilization of iron and magnesium);
  • Chlorination leaching a mixture of hydrochloric acid and metal chloride is used as a leaching agent, and nickel, cobalt and magnesium which are put into the leaching agent are introduced into the leaching agent by chloride, and iron is suppressed in the leaching slag;
  • the leaching is carried out in a pressurized warming leaching tank at a temperature of 60 ° C to 180 ° C and a pressure of 0.1 MPa to 0.5 MPa, and the leaching time is 0.5 hours to 2 hours; wherein, preferably, the temperature is 150 ° C ⁇ 180 ° C, the preferred pressure is 0.30 MPa ⁇ 0.5MPa;
  • the weight ratio of the leaching agent to the ore is 1.3 ⁇ 2.5: 1, the total liquid to solid weight ratio in the hydration control material is 3 ⁇ 4:1 ; the leaching agent has more than 28% HC1 content, and the metal ion in the leaching agent
  • the metal chloride is magnesium chloride or ferric chloride, ferrous chloride.: one or one of i* aluminum, calcium chloride Above, or metal chlorides produced in the present process and mixtures thereof;
  • the leaching result is a leach solution containing nickel, cobalt, magnesium chloride and iron-containing leach residue;
  • the ball is made by mixing the aqueous leaching residue, the conventional composite binder and the reducing agent to form 5 ⁇ 12mm raw spherules;
  • the composite binder may be a mixture of sodium humate, quicklime and sodium carbonate, and the amount of the spheroid is
  • the dry weight of the leaching residue is 2 to 3%;
  • the reducing agent is charcoal or anthracite, bituminous coal or lignite, and the reducing agent is blended in the amount of 6/10% of the dry weight of the leaching residue ;
  • the reduction magnetization roasting is carried out in a rotary kiln.
  • the magnetic iron oxide is obtained by controlling the calcination temperature of 70 (TC ⁇ 780 °C, roasting time 30, 40; the magnetic iron oxide cooked ball obtained after calcination is crushed, ball milled, and ⁇ 200
  • the mesh sieve, the sifted powder magnetic is selected from the weak magnetic separator, the magnetic separation magnetic field strength is 80 Tesla ⁇ 100 Tesla, and the magnetic separation product controls the iron content 55% ⁇ 65% ;
  • the leachate is first dehydrated by one of two methods:
  • the leachate is concentrated, oxidized, and 7 is de-ironed: the leaching solution in the reaction tank is concentrated by the heat in the subsequent atomization drying and roasting process, so that the water and the hydrogen chloride gas in the leachate are largely volatilized, and the ferric chloride and Crystallization of magnesium chloride is carried out to obtain a concentrated leachate for removing the mirror; the generated steam is condensed and returned to the aforementioned "chlorination leaching"step; then, air or oxygen enrichment is introduced into the reaction tank at 60 ° C to 100 ° C.
  • the nickel-cobalt is extracted by the method of precipitating the nickel-cobalt by the composite vulcanizing agent after the iron removal process obtained by the above two methods of removing iron, that is, the composite vulcanizing agent is added to the leach solution after the iron removal by one of the following two methods (vulcanization) Agent A and vulcanizing agent B):
  • vulcanizing agent A and vulcanizing agent B are slowly added or mixed in proportion, and then slowly injected, or partially mixed with vulcanizing agent A or vulcanizing agent B, so that the pH value of the leachate remains in the precipitation process. Between ⁇ 0. 30 of the fixed value;
  • Vulcanizing agent A and vulcanizing agent B are added in a mass ratio of X: (100-x), wherein x is in the range of 1-20, and the vulcanizing agent is added in proportion to: vulcanizing agent A and vulcanizing agent B Slowly added or mixed slowly, or slowly added, or partially mixed with vulcanizing agent A or vulcanizing agent B, so that the pH of the leachate remains unchanged or slowly changes;
  • the vulcanizing agent A is a sulfide containing hydrogen ions, including one or more of hydrogen sulfide or hydrogen sulfide, sodium hydrosulfide, potassium hydrogensulfide, ammonium hydrogen sulfide;
  • the vulcanizing agent B is a sulfide containing no hydrogen ions, including fluidized sodium or potassium sulfide, ammonium sulfide, ferrous sulfide, iron sulfide, magnesium sulfide, and iron sulfide, magnesium sulfide, nickel sulfide, and sulfide formed by precipitation in this process.
  • nickel-cobalt precipitated by adding a composite vulcanizing agent to the leaching solution after iron removal, and a nickel-cobalt sulfide precipitate is obtained by solid-liquid separation, and a magnesium chloride solution or a mixed solution of magnesium chloride and ferric chloride; wherein the nickel-cobalt sulfide is conventional Metallurgical methods for the production of nickel, cobalt metal or compound products;
  • steps (2) and (3) are respectively treating the leachate and the leach residue after the solid-liquid separation, the steps (2) and (3) may be performed sequentially or simultaneously; 3) It can also be carried out before step (2).
  • the atomized dried product, the magnesium chloride solution obtained in the above step (3) or the mixed solution of magnesium chloride and ferric chloride is concentrated and crystallized to obtain magnesium chloride, ferric chloride crystals, or atomized dried product and magnesium chloride obtained by the concentrated crystallization.
  • roasting the mixture of ferric chloride crystals; the calcination can be carried out by one of the following conventional methods:
  • the mass percentage of the atomized dried product in the calcined material is 20% - 100%;
  • the fuel may be various coal, coke, natural gas, 7 gas, liquefied petroleum gas, petroleum products (such as: diesel, heavy oil) , kerosene, etc.); the fuel is added by injection into natural gas, water gas, liquefied petroleum gas, petroleum products, or 10% - 50% (wt) of coal powder or coke powder in the material, the heat is insufficient It can be externally supplemented with heat; the oxidizing atmosphere, weak reducing atmosphere or strong reducing atmosphere can be maintained in the baking furnace according to the process requirements; the roasting equipment can be used in rotary kiln (rotary kiln) and boiling furnace (fluidizing roaster) Or high-temperature metallurgy and chemical equipment such as shaft kiln (vertical kiln).
  • the metal chloride in the material is hydrolyzed to hydrogen chloride and metal oxide at high temperature, and the produced hot hydrogen chloride
  • the body is directly introduced into the laterite nickel ore leaching solution to heat and concentrate the leaching solution, and the hydrogen chloride gas precipitated from the leaching solution is cooled and absorbed to regenerate hydrochloric acid or return the regenerated hot hydrochloric acid to the front end for hot dip of hydrochloric acid for laterite nickel ore (ie, "chlorination" The leaching ") step realizes closed-circuit recycling of hydrochloric acid.
  • the present invention can increase the activity of hydrogen ions in hydrochloric acid by using a mixed solution of hydrochloric acid and metal chloride as a leaching agent, thereby improving the leaching efficiency and the utilization ratio of hydrochloric acid.
  • the metal chloride produced in this process is directly applied to the chlorinated leaching agent, and the efficient use of materials in the system is realized.
  • the results of the leaching of the same laterite nickel ore sample by the technique of the present invention and the conventional hydrochloric acid leaching technique for 3 hours are as shown in Table 1. Due to the presence of metal chlorides, the reactivity of hydrogen ions is enhanced.
  • Nickel-cobalt comprehensive leaching rate /% The present invention 0. lmol / L magnesium chloride +5. 8 mol / L hydrochloric acid solution 95. 8 Conventional technology 6mol / L hydrochloric acid solution 87. 2
  • the invention adopts a mixture solution of hydrochloric acid and metal chloride as a leaching agent, and leaches under a heating pressure cake of 0.3-0.5 MPa and 150-180 ° C, which can ensure high leaching rate of nickel and cobalt while ensuring high leaching rate of nickel and cobalt.
  • the iron leaching rate is controlled to be less than 5%, thereby greatly reducing the acid consumption of the leaching process, greatly reducing the amount of subsequent iron removal work and the loss of nickel and cobalt due to the iron bow.
  • the temperature and pressure used in the present invention are much lower, thereby significantly reducing the requirements for the material of the leaching equipment, and the equipment investment and maintenance costs are greatly reduced.
  • the invention retains most of the iron in the leaching slag by leaching, and then reduces the magnetization roasting by leaching slag, and is subjected to crushing and magnetic separation to enrich the iron for iron making, so that more than 85% of the iron is effectively recovered. With the use of iron.
  • Concentration and crystallization of the leachate reduces or eliminates the expansion of the system in the process and significantly reduces the iron-nickel ratio in the leachate. Significantly reduce the acid consumption of the treatment system and increase the utilization of hydrochloric acid.
  • the invention concentrates and crystallizes the leachate, on the one hand, reduces or eliminates the expansion phenomenon of the system in the process, and improves the utilization efficiency and productivity of the process equipment such as subsequent sulfide precipitation, hydrochloric acid regeneration, etc., because the ferric chloride and magnesium chloride are concentrated.
  • the invention directly uses the ferric oxide magnesia and calcium oxide produced by the regeneration process of hydrochloric acid regeneration as a neutralizing agent, and replaces the traditional method with other basic oxides or compounds, thereby realizing the comprehensive utilization of materials in the system and avoiding The introduction of the impurity compound makes the subsequent treatment of the leachate simpler.
  • the present invention oxidizes ferrous iron to ferric iron by introducing air or oxygen-enriched air during the leachate treatment, and hydrolyzes the precipitate during subsequent neutralization, or extracts iron by organic solvent extraction.
  • the separation of the main lanthanum elemental iron before the precipitation of the sulphide is reduced, the amount of the vulcanizing agent is reduced, and the iron interference during the sulphide precipitation is effectively reduced, and the recovery rate of the nickel-cobalt is improved.
  • the product of hydrolyzed iron, or back-extracted ferric chloride can be used as a raw material for the production of iron products, realizing the comprehensive utilization of laterite nickel ore and increasing the added value of the process.
  • a metal sulfide which has just precipitated in the present process is added to the sulfurized precipitant. Due to the high reactivity of the precipitated iron sulfide, magnesium sulfide, nickel sulfide, etc. in this process, and their excellent surface physicochemical properties, they can be used as seed crystals for the precipitation of nickel sulfide, and deposited on the seed crystal.
  • the sulfide precipitate obtained by growing up has better filtration performance and is easy to re-solid-liquid separation. Therefore, the reduction of the amount of the vulcanizing agent and the improvement of the solid-liquid separation performance can effectively improve the efficiency of the vulcanization and precipitation process.
  • vulcanizing agent A and vulcanizing agent B constitute a buffer system, so that the pH value is basically maintained, the nickel sulfide precipitates crystallinity is good, easy to filter; nickel, cobalt and other precious metals and iron
  • the magnesium separation effect is good, and the content of magnesium and iron in the sulfide is low; the efficiency of the sulfide precipitation is effectively improved, and the amount of the vulcanizing agent is reduced.
  • the composite vulcanizing agent, the hydrogen ion-containing vulcanizing agent A and the hydrogen ion-free vulcanizing agent B constitute a buffer system, and the amount or ratio of the vulcanizing agent A and the vulcanizing agent B in the vulcanization precipitation process can be maintained.
  • the pH of the leachate is changed within the range of the set value ⁇ 0.30, or the pH value of the leachate is slowly changed, and the obtained nickel sulfide and cobalt sulfide precipitates have relatively uniform properties, and are favorable for the growth of the particles, and then cooperate with In this process, the precipitated compound is used as a seed crystal, thereby obtaining a compact structure, a large particle, and a uniform sulfide precipitate, which has better filtration performance and is easy to achieve solid-liquid separation.
  • Use a single sodium sulfide as a precipitant pair Compared with the method of the present invention, the sulphide is filtered at a high speed, and the cracking time of the filter cake is about 10 min. It is difficult to filter by the contrast technique, and the cracking time of the filter cake is more than 30 min.
  • a nickel- and cobalt-rich sulfide precipitate having a dense structure and good crystallinity can be obtained, and the adsorption, inclusion of iron sulfide and magnesium sulfide in the precipitation of amorphous nickel sulfide and cobalt sulfide can be reduced.
  • the invention adopts the vulcanizing agent A and the vulcanizing agent B to form a buffer system, the change of the ⁇ ⁇ value of the vulcanization and precipitation process can be adjusted and controlled, and the iron and magnesium are hydrolyzed by the iron hydroxide and the hydric hydroxide due to the excessive ⁇ ⁇ value.
  • Magnesium forms a large amount of sulfide products, which effectively increases the separation rate of nickel, cobalt and iron and magnesium.
  • the content of iron and magnesium compounds in the sulfide is 15% or more lower by the technique of the present invention.
  • Sulfide can provide more precipitant sulfur, which can significantly reduce the amount of vulcanizing agent and improve the separation efficiency of the vulcanization and precipitation process.
  • polysulfide (persulfide) is added to the sulfide precipitant, and since the polysulfide can provide more precipitant sulfur, the amount of the vulcanizing agent can be reduced.
  • the material is subjected to atomization drying before roasting, which improves the drying efficiency and reduces the subsequent; the amount of material burned and the equipment capacity are increased.
  • the present invention provides a hydrochloric acid regeneration method which separates the atomization drying and the roasting operation, reduces the amount of materials in the high-temperature roasting process, saves fuel, improves equipment processing capacity, and reduces equipment investment.
  • the operating cost of equipment is reduced; solid fuels such as coal powder and coke can be used, the adaptability of fuel types is enhanced, the geographical adaptability of industrial production is enhanced, and the fuel cost is greatly reduced.
  • various compounds such as magnesium chloride, iron oxide, magnesium oxide, and ferroferric oxide, and metallic iron and mixtures thereof can be obtained, which can be used for magnetic materials, iron pigments, iron-making raw materials, magnesium-removing materials, Refractory materials and neutralizers for this process leachate are rich in products and can be adapted to different customers and market needs.
  • the invention adds fuel coal powder and coke powder to the calcined material.
  • the coal powder and the coke powder are burned to release a large amount of heat, because the coal powder, the coke powder and the calcined chloride material are uniformly mixed together, thereby Hot profit of pulverized coal and coke powder burning It is very sufficient, and the material in the furnace is evenly heated, which effectively reduces the amount of fuel used and improves fuel utilization. Since the coal powder and coke powder in the material provide heat energy from the inside of the material during the high-temperature roasting process, the chlorination mirror and the ferric chloride are subjected to high-temperature hydrolysis at a lower temperature to generate hydrogen chloride gas to realize hydrochloric acid regeneration.
  • the heat energy is supplied from the inside of the material during the high-temperature roasting process, so that the magnesium chloride and the ferric chloride are fully hydrolyzed at a high temperature, thereby improving the conversion of the chloride. rate.
  • the use of residual heat during the concentration of the mother liquor and the regeneration of hydrochloric acid significantly improves the energy utilization efficiency and reduces energy consumption.
  • the precipitated mother liquid after nickel precipitation is concentrated, and is mixed with the ferric chloride and magnesium chloride crystals generated in the previous concentrated crystallization process, and is calcined together, thereby greatly increasing the chloride ion content in the calcined raw material, thereby significantly increasing the productivity of the high-temperature roasting equipment and improving The efficiency of the roasting process, reducing energy consumption;
  • the hot hydrogen chloride gas generated by the hydrochloric acid regeneration process is directly extracted into the i!A leaching solution to concentrate and crystallize the leaching solution, thereby realizing the comprehensive utilization of the residual heat in the regeneration process.
  • the chloride in the nickel mother liquor and the chloride produced during the concentrated crystallization are regenerated into hydrochloric acid by high temperature hydrolysis, and returned to the process of treating the laterite nickel ore by the chlorination process, thereby realizing the hydrochloric acid treatment of the laterite nickel ore by the chlorination method. Closed loop, for the environment 3 ⁇ 4 ⁇ 4 sub.
  • the method of the invention is used to treat laterite nickel ore, and the high-efficiency and low-cost extraction of nickel and cobalt in the laterite nickel ore, the comprehensive utilization of iron and magnesium resources, and the complete closed circuit of hydrochloric acid have good social and economic benefits.
  • the invention adopts chloride as the leaching agent for warming and pressure leaching; the leaching slag recovers iron treatment; the leaching solution is concentrated, oxidized, neutralized and precipitated by using a composite vulcanizing agent, or extracted and iron removed, neutralized, cobalt; The mother liquor after cobalt is atomized and dried, and the high-efficiency and low-cost extraction of nickel and cobalt in the laterite nickel ore, the comprehensive utilization of iron and magnesium resources, and the complete closed loop of hydrochloric acid have good social and economic benefits. .
  • Table 2 lists the comparison of the technical and economic indicators of the present invention with the comparative technology. Table 2 Comparison of the present invention and comparative technology
  • Vulcanizing agent dosage tons of nickel 8 , 1. 4 1. 4 0. 6
  • Waste water fiber 190 130 - - Advantages and disadvantages of equipment investment is small, the leaching process is obtained by leaching process to obtain leaching equipment investment but - strengthening, but: strengthening, part of the salt is small; iron is dissolved in a large amount; leaching equipment investment acid is regenerated, magnesium nickel cobalt recovery rate magnesium iron is not comprehensive Get used, but: high;
  • magnesium iron is not comprehensively utilized, iron is not comprehensively utilized, comprehensive utilization of magnesium iron, high acid consumption; used; acid consumption;
  • Neutralizer and sulfuric acid consumption Neutralizer and sulfuric acid consumption; High dosage; Neutralizer and vulcanizing agent; High concentration of neutralizing agent and vulcanization environment; High dosage; Environmental pollution; Low; equipment processing capacity, environmental pollution; equipment processing capacity, environmentally friendly; small; equipment processing capacity is small; equipment processing capacity is small;
  • Comparative Technology 1 The characteristics of Comparative Technology 1 are: leaching of hydrochloric acid at atmospheric pressure at 40-95 °C, precipitation of nickel-cobalt by sodium sulfide, and discharge of mother liquor after neutralization;
  • Comparative Technology 2 The characteristics of Comparative Technology 2 are: 200-300 ° C, 3-lOMPa high temperature and high pressure hydrochloric acid leaching, force H sodium sulfide precipitation nickel cobalt, mother liquor neutralization and emission comparison technique 3 characteristics: 200-300 ° C, 3- lOMPa high temperature and high pressure hydrochloric acid leaching, sodium sulfide precipitated with sodium sulfide, mother liquor high temperature atomization roasting regeneration hydrochloric acid
  • the characteristics of the present invention are as follows: 150-165 ⁇ , 0.3-0.5 MPa, using hydrochloric acid and metal chloride as a leaching agent, and leaching; leaching slag to recover iron; 3 ⁇ 4» effluent for concentration, oxidation, neutralization Precipitating with a composite vulcanizing agent, neutralizing the nickel-cobalt after removing iron by extraction; the mother liquor after taking nickel-cobalt is subjected to atomization drying, roasting and regenerating salt, and comprehensively utilizing magnesium iron.
  • the invention adopts a mixed solution of hydrochloric acid and metal chloride as a leaching agent; leaching the laterite nickel ore under appropriate heating and pressure; concentrating and pre-iron removal treatment by concentration of mono-oxidation and neutralization; using a composite vulcanizing agent; Precipitated nickel-cobalt, nickel-cobalt comprehensive recovery rate, can reach more than 95.5%. (See points 1, 2, 4, 6, 7, 8, 15 in the above "Advantages and Positive Effects" of the present invention);
  • the hot hydrogen chloride gas and water vapor generated by the concentration of the solution are returned to the leaching process; the atomized drying, the hot hydrogen chloride gas and the water vapor generated during the calcination process are used for heating, and the solution is concentrated and returned to the leaching process; the chlorine in the nickel mother liquor
  • the chlorides produced during the crystallization and concentration crystallization are regenerated into hydrochloric acid by high temperature hydrolysis, and returned to the process of treatment of laterite nickel ore by chlorination.
  • the closed loop of hydrochloric acid in the treatment of red soil W with water and hydrochloric acid is realized, no waste water is discharged, and the environment is friendly. (points 4, 10, 12, 13, 14, 15 of the advantages of the invention).
  • the recovered iron of the leaching slag is treated as a raw material for iron making; the iron oxide and magnesium oxide produced by the roasting process are neutralizing agents; after calcination, it can be used for magnetic materials, iron pigments, iron making. More raw materials, magnesium-removing materials, and refractory materials
  • the iron and magnesium compound products realize the comprehensive utilization of nickel, cobalt, iron and magnesium resources in laterite nickel ore. (See points 3, 6, 11, and 15 in the above "Advantages and Positive Effects" of the present invention);
  • FIG. 1 is a schematic view showing the process flow of a wet chlorination treatment of laterite nickel ore according to an embodiment of the present invention.
  • the main valuable metal content of one type of laterite nickel ore is: Ni 0.92%; Co 0.07%; Mg 14.21%; Fe 19.36%. Crushed by coarse.
  • the laterite nickel ore is ball milled and passed through a 100 mesh screen at a water to water ratio of 1: 2 (wt).
  • the sum of the iron ion content: the chloride ion content is 0.05: l (wt)), stirred and leached.
  • the laterite nickel ore was leached by chloride, and the nickel-cobalt content (g/L) of the leaching solution was: Ni+Co 2.27; Mg 33.5 Fe2.16 ; the acidity of the leaching solution was: ⁇ " 0.82 ⁇ 1/ ⁇ .
  • the following examples Based on the leachate.
  • the valuable metal component (%) of the leaching residue (dry) is: ⁇ 0.04; Co - 0.01; Mg 0.63; Fe 22.6.
  • the leaching residue contains 16.8% water. The following examples are based on this leaching.
  • the composite binder is 80% sodium humate + 10% quicklime + 10% sodium carbonate, and the blending amount is 3% of the dry amount of the leaching slag; the reducing agent is bituminous coal, and the blending amount is 10% of the dry amount of the leaching slag.
  • Magnetic separation The reduced mature ball is crushed, ball milled, and passed through a 200 mesh sieve; - 200 mesh powder is magnetically selected by a weak magnetic separator, the magnetic separation magnetic field strength is 100 Tesla; the magnetic content of the magnetic product is 58.6%; The iron yield was 86.6%.
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid, so that the volume of the leaching solution is reduced to 20% to 30%, magnesium chloride and ferric chloride are precipitated in the form of crystals, and hydrogen chloride gas and water in the leaching solution are largely precipitated and collected.
  • air is oxidized in a hot concentrated liquid (>60 ° C) for 1 h to oxidize the ferrous iron in the solution to ferric iron.
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquid neutralizes the residual acid of the leachate, and the pH value of the leachate is adjusted to 3 to 5, so that the ferric iron is hydrolyzed to realize iron immersion and the precipitate is filtered.
  • the nickel-cobalt is extracted by sulfide precipitation using hydrogen sulfide (vulcanizing agent A) and sodium sulfide (vulcanizing agent B) as a composite vulcanizing agent. 5 ⁇ The amount of the vulcanizing agent, the theoretical amount of the vulcanizing agent required to precipitate nickel sulfide, cobalt sulfide 1. 5 times control compound The amount of the vulcanizing agent is added, and after solid-liquid separation and washing, a sulfide product rich in nickel and cobalt is obtained, and the filtrate is subjected to atomization drying and calcination to carry out regeneration of hydrochloric acid and comprehensive utilization of magnesium and iron resources.
  • the mother liquor after extracting nickel cobalt is concentrated and atomized and dried.
  • the high-temperature furnace gas obtained by the calcination process is concentrated, and the mother liquid after nickel precipitation is adjusted to a total amount of C1 of 300 g/L, and then atomized and dried at 300 ° C to obtain magnesium chloride containing different crystal water, and different crystal water.
  • the mass percentage of the pulverized coal in the calcined material is 20%
  • the atomized dried product is prepared in a mass percentage of 80% in the calcined material, and is calcined in a rotary kiln.
  • the material moves along the furnace with controlled temperature and atmosphere.
  • the material is first oxidized and burned under 68 TC, oxidizing atmosphere, and then subjected to weak reduction magnetization roasting under 800, weak reducing atmosphere to obtain triiron tetroxide.
  • the hot hydrogen chloride gas generated during the roasting process is passed into a heating jacket to heat and concentrate the laterite nickel ore leaching solution, and the cooled hydrogen chloride gas and the hydrogen chloride gas precipitated from the leachate are cooled and absorbed to obtain regenerated hot hydrochloric acid, and the heat of regeneration is recovered. Hydrochloric acid is returned to the front end for hot dip of hydrochloric acid in laterite ore.
  • the laterite nickel ore is ball milled at a ratio of 1 to 1 (wt) of water and passed through a 100 mesh screen.
  • Pumping -100 mesh material into the pressure reactor according to the liquid-solid ratio of 3.0:1, the ratio of leaching agent to material (weight) is 1.4:1, respectively, pumping water and chloride leaching agent (mixing of hydrochloric acid and mother liquor after extraction of nickel and cobalt) Solution, HC1 content of 28% or more, the sum of magnesium ion and iron ion content: chloride ion content of 0.1: 1), 'stirred leaching.
  • the laterite nickel ore is ball milled through a 100 mesh screen at a water to water ratio of 1: 1 (wt).
  • a mixed solution of iron and hydrochloric acid, the HC1 content is more than 28%, the sum of magnesium ion and iron ion content: the chloride ion content is 03:1), and the mixture is stirred and leached.
  • step (1) The extraction and comprehensive utilization of valuable metals in laterite nickel ore were carried out in the same manner as in Example 1, except that the step (1) was carried out as follows - A. Adding water to the laterite nickel ore according to the ratio of material to water 1: 2 (wt) Ball milled and passed through a 100 mesh screen. Pumping -100 mesh slurry into the pressure reactor, pumping water and hydrochloric acid (a mixture of hydrochloric acid and stripping solution) at a liquid to solid ratio of 4.0:1, a ratio of leaching agent to material (weight) of 1.35:1, respectively.
  • the HC1 content is more than 28%, and the sum of magnesium ion and iron ion content: chloride ion content is 0.02: 1), stirring and leaching.
  • the laterite nickel ore is ball milled and passed through a 100 mesh screen at a water to water ratio of 1: 2 (wt).
  • step (2) was The following way:
  • A. Pelletizing The aqueous leaching residue, the composite binder and the reducing agent are uniformly mixed and granulated by a disc granulator.
  • the composite binder is 80. /.
  • the blending amount is 2.5% of the dry leaching residue; the reducing agent is bituminous coal, and the blending amount is 8% of the dry amount of leaching slag.
  • Magnetic separation The reduced ball is crushed, ball milled, and sieved at 3 ⁇ 4200 mesh; -20 eye powder is magnetically selected by a weak magnetic separator, magnetic field strength is 100 Tesla; magnetic content of iron is 58.5%; iron The yield was 86.4%.
  • Pelletizing Mix the aqueous leaching slag, composite binder and reducing agent evenly, and granulate with a disc granulator.
  • the composite binder is 80% sodium humate + 10% quicklime + 10% sodium carbonate, and the blending amount is 2% of the dry amount of the leaching slag; the reducing agent is bituminous coal, and the blending amount is 6% of the dry amount of the leaching slag.
  • Magnetic separation The reduced mature ball is crushed, ball milled, and 1200 mesh sieve; - 200 mesh powder is magnetically selected by a weak magnetic separator, the magnetic separation magnetic field strength is 80 Tesla; the magnetic separation product iron content is 62.7%; The yield was 86.8%.
  • Example 8 The leachate is oxidized by air; the hydrogen sulfide and the sodium sulfide composite vulcanizing agent are used, and the pH is controlled to be 3. 0-3. 5. The sulfurization precipitant is used to extract nickel and cobalt)
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid to reduce the volume of the leaching solution to about 25%.
  • the magnesium chloride and the ferric chloride are precipitated in the form of crystals, and the hydrogen chloride gas and the water in the leaching solution are largely precipitated, and are collected and introduced into the leaching.
  • air is oxidized in a hot concentrated liquid (>60 ° C) for 1 h to oxidize divalent iron in the solution to ferric iron.
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquid neutralizes the residual acid of the leachate, and the pH value of the leachate is adjusted to 3.0, so that the ferric iron is hydrolyzed to achieve iron immersion and precipitation.
  • Hydrogen sulfide (vulcanizing agent A) and sodium sulfide (vulcanizing agent B) are used for vulcanization precipitation of nickel-cobalt for composite vulcanizing agent.
  • the sulphide is controlled by the amount of the vulcanizing agent, and the theoretical amount of the vulcanizing agent is 1. 1 times controlled composite vulcanization.
  • the amount of the agent added is obtained by solid-liquid separation and washing to obtain a sulfide product rich in nickel and cobalt, and the filtrate is subjected to atomization drying and calcination to carry out regeneration of hydrochloric acid and comprehensive utilization of magnesium and iron resources.
  • Example 9 The leachate is oxidized by oxygen-enriched air; sodium thiohydride, sodium sulfide and nickel sulphide just precipitated in the process are used as a composite vulcanizing agent, and the pH is controlled to carry out the extraction of nickel and cobalt by a sulphide precipitant)
  • step (3) was The following way:
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the salt expansion process, and the volume of the leaching solution is reduced to 20%.
  • the magnesium chloride and the ferric chloride are precipitated in the form of crystals, and the hydrogen chloride gas and the water in the leaching solution are largely precipitated, and are collected and introduced.
  • oxygen-enriched air oxygen: air volume ratio 1: 1 is introduced into the hot concentrate (>60 ° C) to oxidize the ferrous iron in the solution to ferric iron.
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquid neutralizes the residual acid of the leachate, and the pH value of the leachate is adjusted to 3.5, and the ferric iron is hydrolyzed to realize the precipitation and precipitation of the iron, and the iron red is obtained after the heat treatment.
  • the vulcanizing agent A in the composite vulcanizing agent is sodium hydrosulfide
  • the vulcanizing agent B is a mixture of sodium sulfide and the nickel sulfide just precipitated in the present process, and sodium sulfide and hydrogen sulfide are slowly added respectively to make the pH value of the solution during the precipitation process. Maintaining 3. 5 ⁇ 0. 3, according to the theoretical amount of vulcanizing agent required for precipitation of nickel sulfide and cobalt sulfide, 1.08 times, the amount of the composite vulcanizing agent is controlled, and after solid-liquid separation and washing, vulcanization rich in nickel and cobalt is obtained.
  • the product and the filtrate are subjected to atomization drying and simmering treatment to recover hydrochloric acid and comprehensive utilization of magnesium and iron resources.
  • Example 10 (Immersion liquid is oxidized by oxygen-enriched air; and the nickel-cobalt is extracted by a sulfurized precipitant by controlling the pH value).
  • the extraction and comprehensive utilization of the valuable metals in the laterite nickel ore are carried out in the same manner as in the first embodiment. Step (3) of the above is carried out as follows:
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid, so that the volume of the leaching solution is reduced to 30%, magnesium chloride and ferric chloride are precipitated in the form of crystals, and hydrogen chloride gas and water in the leaching solution are largely precipitated, and collected into the leaching agent. in.
  • oxygen-enriched air oxygen: air volume ratio of 1:1
  • the mother liquor is used to atomize the magnesium oxide produced during the drying or roasting process to neutralize the residual acid of the leachate, and the pH value of the leachate is adjusted to 4.0, the ferric iron is hydrolyzed to realize the iron immersion, the precipitate is filtered, and the iron red is obtained after the heat treatment.
  • the vulcanizing agent A in the composite vulcanizing agent is a mixture solution of sodium hydrosulfide, hydrogen sulfide, ammonium hydrogen sulfide, and the vulcanizing agent B is made of potassium sulfide, calcium sulfide, ammonium sulfide, sodium polysulfide, and iron sulfide which has just precipitated in the present process.
  • the mixture of the vulcanizing agent A and the vulcanizing agent B, the pH of the solution is maintained at 4. 0 ⁇ 0. 30, according to the theoretical amount of vulcanizing agent required to precipitate nickel sulfide, cobalt sulfide 1. 06
  • the amount of the composite vulcanizing agent is controlled in multiples.
  • Example 11 passing air oxidation, using sodium hydrosulfide, sodium persulfide, nickel sulfide just precipitated in the present process as a composite vulcanizing agent, controlling the proportion of A and B components, and extracting nickel and cobalt by a sulfurized precipitant) ⁇
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid to reduce the volume of the leaching solution to about 25%.
  • the magnesium chloride and the ferric chloride are precipitated in the form of crystals, and the hydrogen chloride gas and the water in the leaching solution are largely precipitated, and are collected and introduced into the leaching.
  • air is introduced into the hot concentrated liquid (>60 ° C) for oxidation for 1 h to oxidize the ferrous iron in the solution to ferric iron.
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquid neutralizes the residual acid of the leachate, and the pH value of the leachate is adjusted to 2.2, and the ferric iron is hydrolyzed to realize iron immersion, and the precipitate is filtered.
  • the vulcanizing agent A in the composite vulcanizing agent is sodium sulfide
  • the mm B is a mixture of sodium and nickel sulfide which has just precipitated in the present process.
  • Sodium hydrosulfide, sodium persulfide, and nickel sulfide just precipitated in this process are slowly added at a mass ratio of 1:97:2, and the pH of the leachate is slowly changed.
  • the amount of the composite vulcanizing agent is controlled by 1.06 times, and the sulfide product rich in nickel and cobalt is obtained by solid-liquid separation and washing, and the filtrate is atomized and dried.
  • the roasting process performs regeneration of hydrochloric acid and comprehensive utilization of magnesium and iron resources.
  • Example 12 Oxygen-enriched air oxidation, using sodium hydrosulfide, sodium sulfide, nickel sulfide just precipitated in the present process as a composite vulcanizing agent, controlling the proportion of A and B components, and extracting nickel and cobalt by a sulfide precipitant
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid to reduce the volume of the leaching solution to 25%.
  • the magnesium chloride and the ferric chloride are precipitated in the form of crystals, and the hydrogen chloride gas and the water in the leaching solution are largely precipitated, and are collected and introduced into the leaching agent. in.
  • oxygen-enriched air oxygen: air volume ratio of 1:1
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquor neutralizes the residual acid of the leachate, and the pH of the leachate is adjusted to 4.0, and the ferric iron is hydrolyzed to realize iron immersion, and the precipitate is filtered.
  • the vulcanizing agent A in the composite vulcanizing agent is sodium hydrosulfide
  • the vulcanizing agent ⁇ is a mixture of sodium sulfide and nickel sulfide which has just precipitated in the present process. According to the mass ratio of sodium hydrosulfide, sodium sulfide, and nickel sulfide just precipitated in this process, 18:81:1, the pH of the leachate is slowly changed after mixing.
  • Example 13 extraction of iron, sodium hydrosulfide, sodium sulfide, nickel sulfide just precipitated in the present process as a composite vulcanizing agent, controlling the proportion of bismuth and antimony components, and extracting nickel and cobalt by a vulcanization precipitant
  • the iron removal rate was 99.6%, and the stripping solution was obtained as an acidic solution of ferric chloride. After concentration, atomization drying and calcination, hydrochloric acid was regenerated and iron was recovered.
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid, so that the volume of the extraction residual liquid for extracting iron is reduced to 25%, magnesium chloride and ferric chloride are precipitated in the form of crystals, and hydrogen chloride gas and water in the leaching solution are largely precipitated. It is collected and introduced into the leaching agent.
  • the concentrated raffinate is subjected to precipitation of nickel cobalt using a composite vulcanizing agent.
  • the vulcanizing agent in the composite vulcanizing agent is sodium hydrosulfide
  • the vulcanizing agent B is a mixture of sodium sulfide and nickel sulfide which has just precipitated in the present process.
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquor neutralizes the residual acid of the concentrated raffinate leaching solution, and adjusts the pH of the leachate to 3.0, according to sodium hydrosulfide, sodium sulfide, and the newly precipitated nickel sulfide in the present process.
  • the mass ratio is 5: 92: 3 mixed Slowly add after mixing, so that the pH of the leachate changes slowly.
  • the amount of the composite vulcanizing agent is controlled. After the solid-liquid separation and washing, a sulfide product rich in nickel and cobalt is obtained, and the filtrate is atomized and dried. The simmering treatment is used to regenerate hydrochloric acid and comprehensively utilize magnesium and iron resources.
  • the iron removal rate was 99.7%.
  • the stripping solution is obtained as an acidic solution of ferric chloride, and after concentration, atomization drying and calcination, hydrochloric acid is regenerated and iron is recovered.
  • the leaching solution is heated by the hot hydrogen chloride gas generated during the roasting of hydrochloric acid, so that the volume of the extraction residual liquid for extracting iron is reduced to 30%, magnesium chloride and ferric chloride are precipitated in the form of crystals, and hydrogen chloride gas and water in the leaching solution are largely precipitated. It is collected and introduced into the leaching agent.
  • the concentrated raffinate is subjected to precipitation of nickel cobalt using a composite vulcanizing agent.
  • the vulcanizing agent A in the composite vulcanizing agent is sodium hydrosulfide
  • the sulfur agent B is a mixture of sodium sulfide and nickel sulfide which has just precipitated in the present process.
  • the magnesium oxide produced during the atomization drying or calcination of the mother liquor neutralizes the residual acid of the concentrated raffinate leaching solution, and adjusts the pH of the leachate to 3.5, according to sodium hydrosulfide, sodium sulfide, and the newly precipitated nickel sulfide in the present process.
  • the mass ratio is 8:90: 2, and the mixture is slowly added, so that the pH of the leachate changes slowly.
  • the amount of the composite vulcanizing agent is controlled. After solid-liquid separation and washing, a sulfide product rich in nickel and cobalt is obtained, and the filtrate is atomized and dried. The treatment is carried out to regenerate hydrochloric acid and comprehensively utilize magnesium and iron resources.
  • the nickel-cobalt precipitation rate of each of the examples 9-14 was 99.9% or more.
  • Example 15 concentration of mother liquor, atomization drying
  • the mother liquor after nickel deposition is concentrated and atomized and dried.
  • the high-temperature furnace gas obtained by the calcination process is concentrated, and the mother liquid after nickel precipitation is adjusted to a total amount of C1 of 350 g L, and then atomized and dried at 210 ° C to obtain magnesium chloride containing different crystal water and chlorine with different crystal water. Iron solid mixture.
  • the mother liquor after nickel deposition is concentrated and atomized and dried.
  • the high-temperature furnace gas obtained by the calcination process is concentrated, and the mother liquid after nickel precipitation is adjusted to a total amount of C1 of 250 g/L, and then atomized and dried at 380 ° C to obtain magnesium chloride containing different crystal water, and different crystal water.
  • Example 17 (Mother liquid direct atomization drying)
  • the extraction and comprehensive utilization of the valuable metals in the laterite nickel ore are carried out in the same manner as in the first embodiment, except that the step (4) is carried out in the following manner: the mother liquor after nickel sinking is directly subjected to atomization drying, and the high temperature by the calcination process is used. In the furnace gas, the mother liquor after nickel deposition is directly atomized at 280 ° C to obtain a solid mixture of magnesium chloride containing different crystal water and ferric chloride with different crystal water.
  • the mother liquor after nickel immersion is directly atomized and dried, and the mother liquor after nickel precipitation is directly atomized and dried at 370 ° C to obtain chlorination of different crystal water and crystallization of different crystal water by the high temperature furnace gas obtained by the calcination process.
  • Iron solid mixture is obtained.
  • the atomized dried product was sprayed into a high temperature furnace together with natural gas and high pressure air to carry out atomization roasting at 42 CTC.
  • Magnesium chloride and iron oxide, and a small amount of magnesium ferrite are obtained.
  • the solid product after washing with water is iron oxide and magnesium ferrite, and can be used for magnetic materials, iron pigments, iron-making raw materials, and the like.
  • the aqueous solution of magnesium chloride can be further hydrolyzed to recover hydrochloric acid and magnesium oxide, and magnesium oxide can be used for refractory materials.
  • the hot hydrogen chloride generated during the calcination process is directly absorbed by water after heat exchange to obtain regenerated hydrochloric acid, and the regenerated hydrochloric acid is returned to the front end for hydrochloric acid leaching of the laterite ore.
  • Example 20 (low temperature atomization; ⁇ burning)
  • the mass percentage of pulverized coal in the calcined material is 20%
  • the mass percentage of the atomized dried product in the calcined material is 50%
  • the mass percentage of the ferric chloride and magnesium chloride crystal obtained in the leaching solution is 30% in the calcined material.
  • the preparation was carried out, and was sprayed into a high-temperature furnace together with liquefied petroleum gas and high-pressure air to carry out atomization roasting at 48 CTC.
  • Magnesium chloride and iron oxide, a small amount of ferric acid mirror are obtained, and the solid product after washing is iron oxide and magnesium ferrite, and can be used for magnetic materials, iron pigments, iron-making raw materials, and the like.
  • Magnesium chloride aqueous solution can be further hydrolyzed to recover hydrochloric acid and magnesium oxide, and magnesium oxide can be used for refractory materials.
  • the hot hydrogen chloride produced during the calcination process is directly absorbed by water after heat exchange to obtain regenerated hydrochloric acid, and the regenerated hydrochloric acid is returned to the front end for hydrochloric acid leaching of the laterite ore.
  • Example 21 (Medium temperature oxidizing simmering)
  • the mass percentage of the coke powder in the calcined material is 10%
  • the mass percentage of the atomized dried product in the calcined material is 70%
  • the ferric chloride and magnesium crystal obtained after the nickel liquid is concentrated in the calcined material are prepared.
  • the mass percentage is 20%
  • the preparation was carried out, and it was sprayed into a high-temperature furnace together with high-pressure air at 800 ° C for calcination. A solid mixture of magnesium oxide and ferric oxide acid is obtained, and the mixture is partially returned to the residual acid in the front end and the leachate, and the remainder is used for iron making and magnesium extraction raw materials.
  • the hot hydrogen chloride gas generated during the roasting process is directly introduced into the laterite nickel ore leaching solution to heat and concentrate the leachate, and the hydrogen chloride gas precipitated from the leachate is cooled and absorbed to obtain regenerated hot hydrochloric acid; the regenerated hot hydrochloric acid is returned to the front end for the laterite hydrochloric acid. Hot dip.
  • Example 22 (Medium temperature oxidizing simmering)
  • the atomized dried product was sprayed into a high temperature furnace together with diesel oil and high pressure air for atomization roasting at 700.
  • the obtained solid products of magnesium chloride, iron oxide and magnesium ferrite are washed with iron oxide and magnesium ferrite, and can be used for magnetic materials, iron pigments, iron-making raw materials and the like.
  • the chlorinated mirror aqueous solution can be further hydrolyzed to recover hydrochloric acid and magnesium oxide, and the magnesium oxide can be used for the refractory material.
  • the hot hydrogen chloride produced during the calcination process is directly absorbed by water after heat exchange to obtain regenerated hydrochloric acid, and the regenerated hydrochloric acid is returned to the front end for hydrochloric acid leaching of the laterite ore.
  • the atomized dried product is sprayed into a high temperature furnace together with high pressure air to carry out atomization at 1000 ° C; Magnesium oxide and iron oxide are obtained, and the solid product is subjected to strong magnetic separation.
  • the magnetic products are iron oxide and magnesium ferrite, which can be used for magnetic materials, fabrics, ironmaking raw materials and the like.
  • Non-magnetic magnesium oxide can be used for refractory materials.
  • the hot hydrogen chloride produced during the calcination process is directly absorbed by water after heat exchange to obtain regenerated hydrochloric acid, and the regenerated hydrochloric acid is returned to the front end for hydrochloric acid leaching of the laterite ore.
  • Example 24 (Medium temperature oxidation; ⁇ burning, weak reduction magnetization roasting)
  • the mass percentage of pulverized coal in the calcined material is 25%
  • the mass percentage of the atomized dried product in the calcined material is
  • the hot hydrogen chloride gas generated during the roasting process is passed into a heating jacket to heat and concentrate the laterite nickel ore leaching solution, and the cooled hydrogen chloride gas and the hydrogen chloride gas precipitated from the leachate are cooled and absorbed to obtain regenerated hot hydrochloric acid, and the heat of regeneration is recovered. Hydrochloric acid is returned to the front end for hot dip of hydrochloric acid in laterite ore.
  • Example 25 (Medium temperature oxidizing sinter, weak reduction magnetization; ⁇ burning)
  • the mass percentage of pulverized coal in the calcined material is 30%, the mass percentage of the atomized dried product in the calcined material is 20%, and the quality of the ferric chloride and magnesium chloride crystal obtained in the leachate treatment;
  • the percentage was 50% for stock preparation and roasting in a rotary kiln.
  • the material moves along the furnace with controllable temperature and atmosphere.
  • the material is first oxidized and calcined at 700 ° C under an oxidizing atmosphere, and then subjected to weak reduction magnetization roasting at 830 ° C under a weak reducing atmosphere to obtain triiron tetroxide.
  • the hot hydrogen chloride gas generated during the roasting process is directly introduced into the laterite nickel ore leaching system for the hot dip of hydrochloric acid in the laterite ore.
  • the mass percentage of coke in the calcined material is 50%
  • the atomized dried product is prepared in a mass percentage of 50% in the calcined material, and is calcined in a rotary kiln.
  • the material moves along the furnace with controllable temperature and atmosphere.
  • the material is first oxidized and calcined under 1000 V under oxidizing atmosphere, and then subjected to reduction magnetization roasting at 1350 ° C under strong reducing atmosphere to obtain a mixture of metallic iron and magnesium oxide.
  • Magnetically selected iron is used as a raw material for iron making
  • magnesium oxide is used as a raw material for refractory materials or as other products.
  • the hot hydrogen chloride gas generated during the roasting process is directly introduced into the laterite nickel ore leaching system for the hot dip of hydrochloric acid in the laterite ore.
  • 20% of the mass of the atomized dried product in the calcined material is 60%, and calcination is carried out in a rotary kiln.
  • the material moves along a furnace with a controlled temperature and atmosphere, starting at 1100.
  • C. Oxidation roasting under an oxidizing atmosphere, followed by reduction magnetization roasting under a strong reducing atmosphere at 1250 to obtain a mixture of metallic iron and magnesium oxide, magnetically selecting iron for use in ironmaking raw materials, - magnesium oxide for refractory materials or as Raw materials for other products.
  • the hot hydrogen chloride gas generated during the roasting process is passed into a heating jacket to heat and concentrate the laterite nickel ore leaching solution, and the cooled hydrogen chloride gas and the hydrogen chloride gas precipitated from the leachate are cooled and absorbed to obtain regenerated hot hydrochloric acid, and the heat of regeneration is recovered. Hydrochloric acid is returned to the front end for hot dip of hydrochloric acid in laterite ore.

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Description

盐酸全闭路循环法从红土 W中提取有价金属的方法
技术领域
本发明涉及从红土镍矿中提取镍、 钴金属及综合利用镁、 铁等有价金属的方法， 属有色冶金 领域。
背景技术
红土镍矿是一种氧化镍矿， 占总镍贮量的 65%以上。 红土镍矿分为褐铁矿型和硅镁镍矿型两大 类。褐铁矿类型位于矿床的上部， 铁高、镍低， 硅、镁也较低， 但钴含量比较高， 宜采用湿法冶金 工艺处理。硅镁镍矿位于矿床下部， 硅、镁的含量比较高， 铁含量较低， 钴含量也较低， 但镍的含 量比较高， 宜采用火法冶金工艺处理。而处于中间过渡的矿石可以釆用火法冶金， 也可以采用湿法 冶金工艺。
红土镍镍矿的传统湿法冶金工艺又分为还原;) "氨浸工艺和硫酸加压酸浸工艺。前者通 寸 矿石进行预还原处理， 再用氨浸出镍； 后者是在加压条件下， 用硫酸作浸出剂对矿石中的镍进行浸 取。近年来镍红土矿的湿法冶金技术有了很大的发展， 特别是加压浸出技术和各种组合的溶剂萃取 工艺。火法冶金工艺也分为镍铁法和造锍熔炼法，主要适应于麵矿。镍铁法是一种还原熔炼方法， 而造锍熔炼是在外加硫化剂的情况下造锍的方法。
目前红土矿处理工艺中，还原一氨浸法因为环保问题极少使用。加压硫酸法因投资大、成本髙， 并且存在浸出率低、镁不能开路、 环保压力大等问题。 红土镍矿中大部分为高铁低品位镍矿， 目前 从红土镍矿中提镍钴生产的工艺技术及装备较落后， 生产规模也小， 特别是环保工艺不过关， 远不 能满足开发红土镍矿的需求。
另一种湿法处理红土 W的方法是采用盐酸浸出， 经浸出、 中和、硫化沉淀后得到富集了镍钴 的硫化物， 而沉镍后所得母液经适当浓缩， 然后经高温焙烧产生氯化氢气体使盐酸得到再生。该方 法在处理低品位红土镍矿方面具有投资省、 资源利用率高、对环境友好的特点， 是一种极具发展潜 力的湿法处理工艺。但是， 目前盐酸再生过程中一般是将沉镍后的母液稍作浓缩后直接进行高温焙 烧， 存在能耗大、设备产能低等问题， 从而使得盐酸再生成本很高。 开发经济高效的盐酸再生技术 对于低品位红土 W的开发利用具有重要意义。
因此， 目前世界范围内迫切需要开发适用范围宽、 投资省、 成本低、 资源利用率高、 X寸环境友 好的红土镍矿提取与综合利用生产技术。
发明内容
本发明的目的在于， 针对现有技术存在的不足， 提供一种盐酸全闭路循环法从红土镍矿中提 取有价金属的方法， 为湿法氯化浸出处理红土镍矿方法， 可从红土 中提取镍、钴等有价金属，. 镍钴回收率高， 水、 酸全封闭循环， 无废水排放， 镍、 钴、 铁、 镁资源全部利用， 实现资源高效 利用与清洁生产。 本发明的技术方案是， 所述盐酸全闭路循环法从红土镍矿中提取有价金属的方法以加工破碎 的红土镍矿矿石为原料， 工艺步骤包括氯化浸出、 浸出渣回收铁、 镍钴提取、母液雾化干燥、 焙 烧（再生盐酸与综合利用铁镁)； 其中：
(1)氯化浸出： 以盐酸与金属氯化物的混合 ¾作为浸出剂，将投入所述浸出剂的矿石中镍、 钴、 镁以氯化物进入浸出剂中， 将铁抑制在浸出渣中；
所述浸出在加压加温的浸出槽中，并于温度为 60°C~180°C、压力 0.1 lMPa~0.5MPa的条件下进行， 浸出时间为 0.5小时〜 2小时； 其中， 优选温度为 150°C~180°C， 优选压力为 0.30 MPa~0.5MPa;
所述浸出剂与矿石的重量比为 1.3~2.5 : 1， 补水控制物料中总的液体与固体的重量比 3~4: 1 ; 所述浸出剂中 HC1含量 28%以上，浸出剂中金属离子含量之和：氯离子含量为 0.02~0.3： 1 (wt); 所述金属氯化物为氯化镁或氯化铁、 氯化亚铁.: i *化铝、 氯化钙中的一种或一种以上， 或者为 本流程产生的金属氯化物及它们的混合物；
浸出所得为含镍、 钴、 镁氯化物的浸出液和含铁的浸出渣；
(2)从浸出渣中回收铁：所述浸出渣与浸出液经固液分离后，将浸出渣造球、还原磁化; ^烧， 再经破碎、 磁选富集铁， 用于炼铁； 其中：
造球是将含水浸出渣、常规复合粘结剂、还原剂混合后制造成 5~12mm生球粒； 所述复合粘结 剂可以是腐植酸钠、 生石灰、碳酸钠的混合物， 配入量为浸出渣干重的 2~3%； 还原剂为木炭或无 烟煤、 烟煤或褐煤， 所述还原剂在造球时配入量为浸出渣干重的 6/^10%;
还原磁化焙烧是在回转窑中进行， 通过控制焙烧温度 70(TC~780°C， 焙烧时间 30 中40 中 得到磁性氧化铁； 焙烧后得到的磁性氧化铁熟球经破碎、 球磨， 并^200目筛， 过筛的粉料磁选用 弱磁选机， 磁选磁场强度 80特斯拉〜 100特斯拉， 磁选产物控制铁含量 55%~65%;
(3)镍钴提取：
先将所述浸出液采用以下两种方法之一进行除饮处理：
a.所述浸出液进行浓缩、氧化、 7解除铁：利用后续雾化干燥及焙烧过程中的热量对反应罐中 的浸出液进行浓缩处理， 使浸出液中的水分及氯化氢气体大量挥发， 氯化铁与氯化镁结晶析出， 得 到除去部 镜的浓缩浸出液； 所产生的蒸汽冷凝后返回前述"氯化浸出"步骤； 然后在 60°C~100 °C条件下向所述反应罐中通入空气或富氧空气或 气， 使所述浓縮浸出液中的二价铁氧化成三价 铁； 再以后续雾化干燥或焙烧过程中产生的氧化镁、氧化铁或它们的混合物为中和剂， 将所述含有 三价铁浓缩浸出液的 pH值调整到 2. 2~4. 0， 使三价铁水解实现沉铁， 过滤沉淀， 经热处理后得到铁 红； 滤液为除铁后浸出液；
b. 采用萃取方法对所述浸出液进行除铁处理： 浸出 ¾¾用萃取剂萃取得到含铁的有机相及镍 钴、 镁氯化物水溶液， 萃取除铁后的萃取余液经浓缩后为除铁后浸出液； 从萃取含铁的有机相中 反萃得到的氯化铁溶液； 其中： 萃取铁用萃取剂可以是中性有机膦或胺类化合物， 萃取铁萃取剂 优选为 20%N235+10%丙三醇 +70%煤油 (wt); 萃取操作用箱式萃取设备或立式萃取设备； 采用逆 流萃取和逆流反萃取、 常温下操作； 有效萃取级数 3~5级， 萃取相比 0/Α=1 :3~5; 有效反萃取级 数 3〜5级， 反萃取相比 0/A=3~4: 1；
对经上述两种除铁方法所得除铁后浸出液， 采用复合硫化剂沉淀镍钴的方法提取镍钴， 即采 用下述两种方法之一在所述除铁后浸出液中加入复合硫化剂（硫化剂 A与硫化剂 B):
(i)硫化剂 A与硫化剂 B分别缓慢加入或按比例混合后缓動口入， 或者部分混合后与硫化剂 A或硫化剂 B分别加入， 使得浸出液的 pH值在沉淀过程中保持在设定值的 ±0. 30之间； ：
(ii)硫化剂 A与硫化剂 B按质量比为 X： (100-x)加入， 其中 x在 1-20范围内取值， 硫化剂 按比例加入的方式为： 硫化剂 A与硫化剂 B分别缓慢加入或混合后缓慢加入， 或者部分混合后与 硫化剂 A或硫化剂 B分别加入， 使浸出液的 pH值保持不变或缓慢变化；
所述硫化剂 A为含氢离子的硫化物， 包括硫化氢或过硫化氢、 硫氢化钠、 硫氢化钾、 硫氢化 铵中的一种或一种以上；
所属硫化剂 B为不含氢离子的硫化物， 包括流化钠或硫化钾、 硫化铵、 硫化亚铁、硫化铁、 硫化镁以及本流程中沉淀产生的硫化铁、 硫化镁、 硫化镍、 硫化钴、 多硫化钠、 多硫化铵中的一 种或一种以上；
所述除铁后浸出液中加入复合硫化剂沉淀镍钴后， 经固液分离得到镍钴硫化物沉淀， 以及氯 化镁溶液或氯化镁与氯化铁的混合溶液； 其中， 所述镍钴硫化物经传统冶金方法生产镍、 钴金属 或化合物产品；
由于步骤 (2)和步骤 (3 )是分别对固液分离'后的浸出液和浸出渣进行处理，故所述步骤 (2) 和步骤 (3)可依序进行， 也可同步进行； 步骤 (3)还可先于步骤 (2)进行。
(4)雾化干燥： 利用焙烧工序来的高温炉气， 将萃取分离铁的反萃氯化铁溶液、 提取镍钴后 的氯化镁溶液或氯化镁与氯化铁的混合溶液， 控制 C1的总量 >250 g L， 再在 200°C→00°C下雾化 干燥， 得到含不同结晶水的氯化铁、 或含不同结晶水的氯化镁、 或它们的固体混合物， 即雾化干 燥产物； 雾化干燥的余热用于向前述 "氯化浸出":或 "镍钴提取"中 "浸出液的浓缩、 氧化"供 热； 雾化方式可采用溶液加压离心雾化、 高压气体携带雾化、 超声辅助雾化； 雾化方向可以有卧 式雾化与立式雾化。
(5) ¾
将所述雾化干燥的产物， 上述步骤 (3 )所得氯化镁溶液或氯化镁与氯化铁的混合溶液浓缩结 晶得到的氯化镁、 氯化铁晶体， 或雾化干燥产物与所述浓缩结晶得到的氯化镁、 氯化铁晶体的混 合物进行焙烧； 所述焙烧可采用下述常规 方法之一进行：
(a)低温 (420-500°C )焙烧，用于氯化铁、或氯化铁与氯化镜的混合物的焙烧，得到铁的氧化 物， 或铁的氧化物与氯化镁的混合物， zK洗混合 ¾ /后的固体产物为氧化铁等铁的氧化物， 可以用 于磁性材料、 铁颜料、 炼铁原料等； 水洗液为氯化镁溶液， 可以再高温水解回收盐酸与氧化镁， 氧化镁可以用于耐火材料；
(b)中温（600°C-1000°C)氧化焙烧， 用于氯化镁、 或氯化铁与氯化镁的混合物的焙烧， 得到 氧化镁、或氧化镜与氧化铁（铁的氧化物）的固体混合物， 混合物部分返回前端步骤 (1 )中和浸 出液中的残余酸， 其余用于炼铁及提缓原料；
(c)中温（60(TC-850°C )分段氧化焙烧与弱还原磁性化焙烧，用于氯化铁、或氯化铁与氯化镁 的混合物的焙烧， 得到以四氧化三铁、或四氧化三铁与氧化镁混合物， 混合物磁选后的四氧化三 铁用于炼铁原料或其它铁质原料， 氧化 于耐火材料或作为其它产品的原料；
(d)高温 ( 1000°C-1500°C )分段氧化焙烧与 还原焙烧， 用于氯化铁、 或氯化铁与氯化镁的 混合物的焙烧， 得到金属铁、 或金属铁与氧化镁混合物， 混合物磁选出铁用于炼铁原料， 氧化镁 用于耐火材料或作为其它产品的原料；
上述的焙烧过程中， 烧物料中雾化干燥产物的质量百分数为 20%— 100%; 燃料可以是各种 煤炭、 焦炭、 天然气、 7煤气、 液化石油气、 石油类产品 （如： 柴油、 重油、 煤油等）等； 燃料 加入方式为焙烧时喷入天然气、水煤气、液化石油气、石油类产品，或者在物料中配入 10%—50% (wt) 的煤粉或焦炭粉， 热量不足部分可以通过外， ^热补充； 焙烧炉内按工艺要求分区保持氧 化性气氛、 弱还原性气氛或强还原性气氛； 焙烧设备可釆用回转炉（回转窑)、沸腾炉（流态化焙 烧炉）或立窑 (竖窑)等高温冶金、化工设备。
高温; ^烧时物料中的金属氯化物在高温下水解为氯化氢和金属氧化物，所产生的炙热氯化氢气 体直接通入红土镍矿浸出液中对浸出液进行加热浓缩， 从浸出液中析出的氯化氢气体经冷却吸收 后再生盐酸或将再生的热盐酸返回前端用于红土镍矿的盐酸热浸（即 "氯化浸出" ）步骤， 实现 了盐酸的闭路循环利用。
本发明具有以下的优点与积极效果：
(1) 采用氯化物（盐酸与金属氯化物）浸出提高了浸出效率与盐酸利用率。
本发明通过采用盐酸与金属氯化物的混合溶液作为浸出剂， 可以提高盐酸中氢离子的活性， 从而提髙浸出效率与盐酸的利用率。 并且在稳定生产过程中， 直接采用本流程中产生的金属氯化 物应用于氯化浸出剂， 实现了系统内物料的高效利用。 在其它条件（液固比、浸出温度、 时间、 压力、搅拌等）相同时， 釆用本发明技术与传统的盐酸浸出技术对相同红土镍矿样浸出 3h后的结 果比较如表 1所列。 由于金属氯化物的存在， 氢离子的反应活性增强， 本发明中采用了较低的盐 酸浓度， 但可以在较短的时间内获得较好的镍钴综 ¾r浸出率。 表 1 釆用本发明技术与传统技术将红土镍矿浸泡 lh后的镍、 钴综合浸出率
项目 浸出剂组成 镍钴综合浸出率 /% 本发明 0. lmol/L氯化镁 +5. 8 mol/L盐酸溶液 95. 8 传统技术 6mol/L盐酸溶液 87. 2
(2) 采用氯化物浸出剂与适当的加温加压浸出 法， 可以有效抑制浸出过程中铁的溶出， 提 高浸出过程的效率， 与传统的高温高压浸出比较 则可显著降低对设备的要求与大幅减少一次性 投资。
本发明采用盐酸与金属氯化物的混合物溶液为浸出剂， 并在 0.3~0.5MPa, 150~180°C的加温 加压餅下进行浸出， 可以在确保高的镍、钴浸出率的同时， 使铁的浸出率控制在 5%以下， 从而 大大降低了浸出过程的酸耗， 大幅减少了后续除铁的工作量以及由于铁弓 I起的镍钴损失。 同传统 的高温髙压 (200-300Ό, 3-10MPa)浸出比较， 本发明中采用的温度与压力要低得多， 从而显著降低 了对浸出设备材料的要求， 设备投资与维护费用大幅减少。
(3) 对浸出渣的回收铁处理， 实现了红土镍矿中铁资源的回收利用， 提高矿产资源的综合利 用价值。
本发明通过浸出时将绝大部分铁留在浸出渣中， 然后经浸出渣造球还原磁化焙烧， 经破碎、 磁选富集铁， 用于炼铁， 使 85%以上的铁得到了有效回收与利用铁。
(4) 釆用对浸出液浓缩结晶， 减少或消除了流程中的体系膨胀现象， 并显著降低浸出液中的 铁镍比。 显著减少处理系统的酸耗， 提高了盐酸利用率。 本发明通过对浸出液进行浓縮结晶， 一方面， 减少或消除流程中体系膨胀现象， 提高后续硫 化沉淀、 盐酸再生等工序设备利用效率与产能； 一 7了面， 由于氯化铁、 氯化镁在浓缩过程中结 晶析出， 从而使浸出液中的镍得到富集， 铁镍比 1;镁镍比降低， 减少了后续硫化沉淀所需的沉淀 剂， 显著提高硫化沉淀中所得中间产物中的镍含量。
本发明在盐酸中和前进行加热浓缩结晶时， 使浸出液中大部分剩余的盐酸挥发并返回到前面 的浸出过程，从而显著减少调 pH所需加入中和的碱性氧化物或化合物的量， 因此显著减少了中和 过程弓 I起的酸耗， 提高盐酸的利用率。
(5)釆用本流程中产生的氧化铁、 氧化镁为中和剂， 实现系统内物料的综合利用。
本发明直接釆用盐酸再生的焙烧过程产生的氧 铁 氧化镁、氧化钙为中和剂，取代传统方法 夕卜加其它碱性氧化物或化合物， 从而实现了系统内物料的综合利用， 并避免杂质化合物的引入， 使浸出液的后续处理更加简单。
(6) 摄出液进 fim化处理、 水解沉铁、或有机溶剂萃取分离铁， 实现大部 、或全部铁 在硫化沉淀之前的低成本分离与综合利用， 简化了工艺流程。
本发明通过在浸出液处理过程中通入空气或富氧空气使二价铁氧化成三价铁，并在后续中和过 程水解沉淀， 或通过有机溶剂萃取分离铁。 实现硫化沉淀之前主要贱金属元素铁的分离， 降低了 •硫化剂的用量， 同时有效减小了硫化沉淀时铁的千扰，·提高了镍钴的回收率。水解沉铁的产物、 或反萃取氯化铁， 可作为铁产品生产原料， 实现了红土镍矿的综合利用， 增加了工艺的附加值
(7) 硫化沉淀时加入本流程中刚沉淀的金属硫化物， 有效提高了硫化沉淀的效率， 降低了硫 化剂的用量。
本发明采用硫化沉淀剂中加入本流程中刚沉淀的金属硫化物。 由于本流程中刚沉淀的硫化铁、 硫化镁、硫化镍等具有较高的反应活性， 并且其恃殊的表面物化性质使得它们可以作为硫化镍析 出的晶种， 而且在这种晶种上面沉积长大得到的硫化物沉淀具有更好的过滤性能， 易于实再固液 分离。 因此， 硫化剂用量的减少， 固液分离性能的改善， 均有效提高硫化沉淀过程的效率。
(8)釆用复合硫化剂进行硫化沉淀，硫化剂 A与硫化剂 B构成一种缓冲体系，使 pH值基本维 持不变， 硫化镍沉淀结晶性好， 易过滤； 镍、 钴等贵金属与铁、 镁分离效果好， 硫化物中镁、铁 含量低； 有效提高了硫化沉淀的效率， 降低了硫化剂的用量。
本发明釆用复合硫化剂，含氢离子的硫化剂 A与不含氢离子的硫化剂 B构成了缓冲体系，通过 控制硫化沉淀过程中硫化剂 A与硫化剂 B的加入量或比例， 可以保持浸出液的 pH在设定值 ±0. 30 的范围内变化，或者使浸出液的 pH值缓慢变化，得到的硫化镍、硫化钴沉淀具有较为均一的性质， 并有利于颗粒的长大， 再配合以本流程中刚沉淀的 ^化物作晶种， 从而得到结构密实、颗粒较大、 均匀的硫化物沉淀， 具有更好的过滤性能， 易于实现固液分离。 与采用单一的硫化钠作沉淀剂对 比， 采用本发明技术得到的硫化物抽滤速度快， 滤饼出现开裂的时间为 lOmin左右， 采用对比技 术时很难抽滤， 滤饼出现开裂的时间为 30min以上。
采用本发明技术可以得到结构密实、结晶性较好的富含镍、钴的硫化物沉淀，减少了无定形硫 化镍、硫化钴沉淀中吸附、 夹杂的硫化铁、硫化镁。并且由于本发明采用硫化剂 A与硫化剂 B构 成一种缓冲体系， 可以调节、 控制硫化沉淀过程 ΐ ρΗ值的变化， 防止由于 ρΗ值过高而导致铁、 镁水解以氢氧化铁、氢氧化镁的形式大量进入硫化物产品， 从而有效提高了镍、钴与铁、镁的分 离率。与釆用单一的硫化钠作沉淀剂对比，采用本发明技术时硫化物中铁、镁化合物的含量低 15% 以上。
由于稳定的硫化沉淀环境及硫化镍、硫化钴沉淀良好的物化特性， 可以使硫化沉淀过程中更 好地将镍、 钴富集到硫化物中， 减少铁、镁等贱金属的析出， 而且多硫化物可以提供更多沉淀剂 硫， 从而可以显著降低硫化剂用量， 提高硫化沉淀过程的分离效率。 与釆用单一的硫化钠作沉淀 剂对比，采用本发明技术时硫化剂的中硫：（镍 +钴）（摩尔比）为 1.05时，沉镍率即可达到 99.9%， 而对比技术中硫化剂的中硫： （镍 +钴）（摩尔比）为 2—6。
(9) 硫化沉淀时加入多硫化物， 降低了硫化剂的用量。
本发明釆用硫化沉淀剂中加入多硫化物（过硫化物)， 由于多硫化物可以提供更多沉淀剂硫， 从而可以降低硫化剂用量。
(10)在焙烧之前对物料进行雾化方式干燥， 提高了干燥效率， 并减少了后续;^烧物料量， 提高 设备产能。
与传统的雾化焙烧相比， 本发明提供了一种将雾化干燥与焙烧分开操作的盐酸再生方法， 减 少了高温焙烧过程的物料量， 节约了燃料， 设备处理能力提高， 设备投资减少， 设备运行费用降 低； 可以使用煤粉、 焦炭等固体燃料， 燃料种类的适应性增强， 工业生产的地域适应性增强， 燃 料成本大幅降低。
(11 )提出了雾化干燥后的物料的多种焙烧方式，得到多种铁、续化合物产品， 实现铁、镁产 品的多样化， 提升产品的市场适应性与竞争能力。
通过采用不同的谙烧方式， 得到氯化镁、氧化铁、氧化镁、四氧化三铁等多种化合物与金属铁 以及它们的混合物， 可以用于磁性材料、 铁颜料、 炼铁原料、 提镁材料、 耐火材料及用于本流程 浸出液的中和剂等， 产品丰富， 可以适应不同的客户及市场需求。
(12)焙烧物料中加入燃料煤粉、焦炭粉，减少了外加燃料用量； 可以在较低温度下实现盐酸 再生； 提髙了氯化物的转化率；
本发明在焙烧物料中加入燃料煤粉、焦炭粉，在高温 烧过程中， 煤粉、焦炭粉燃烧释放出大 量的热， 由于煤粉、 焦炭粉与焙烧的氯化物物料均匀混合在一起， 从而煤粉、焦炭粉燃烧的热利 用非常充分， 并使炉内物料受热均匀， 有效减少了夕卜加燃料用量及提高燃料利用率。 由于物料中的煤粉、焦炭粉在高温焙烧过程中从物料内部提供热能，从而使氯化镜、氯化铁在 较低温度下进行高温水解， 产生氯化氢气体而实现盐酸再生。
由于物料中的煤粉、焦炭粉与氯化铁、氯化镁^高 炉中均匀混合， 高温焙烧过程中从物料内 部提供热能， 从而使氯化镁、 氯化铁在高温下充分水解， 提高了氯化物的转化率。
( 13 )沉淀母液浓缩、盐酸再生过程中余热的利用，显著提高了能量综合利用效率，降低能耗。 将沉镍后的沉淀母液浓缩，并与前面浓缩结晶过程中产生的氯化铁、氯化镁晶体混合后一起焙 烧， 大大提高了焙烧原料中氯离子含量， 从而显著增加了高温焙烧设备的产能， 提高了焙烧过程 的效率， 减少能耗；
将盐酸再生过程产生的炙热氯化氢气体直接 i!A浸出液，使浸出液浓缩结晶，实现了再生过程 中余热的综合利用。
( 14)实现了氯化法处理红土镍矿时盐酸的闭路循环， X寸环境友好；
本发明中沉镍母液中的氯化物及浓缩结晶时产生的氯化物均通过高温水解再生为盐酸，返回到 氯化法处理红土镍矿过程中， 实现了氯化法处理红土镍矿时盐酸的闭路循环， 对环境 ¾¾子。
( 15 )采用本发明方法处理红土镍矿， 实现了红土镍矿中镍、钴的高效、低成本提取， 铁、镁 资源的综合利用， 盐酸全闭路循环， 具有很好的社会经济效益。
本发明采用氯化物为浸出剂加温加压浸出； 浸出渣回收铁处理； 对浸出液进行浓缩、氧化、中 和后采用复合硫化剂沉淀， 或采用萃取除铁后中和、 钴； 对提取镍钴后的母液进行雾化干燥、 焙烧， 实现了红土镍矿中镍、 钴的高效、 低成本提取， 铁、镁资源的综合利用， 盐酸全闭路循环, 对环境友好具有很好的社会经济效益。表 2列出了本发明与对比技术的有关技术经济指标的比较。 表 2 本发明与对比技术的比较
项目 对比技术 1 对比技术 2 对比技术 3 本发明 镍钴回收率 (%) 85-90% 88-92% 88-92% >95. 5% 铁回收率 (%) - - >85% 镁回收率 (0/0) ― ― 90% >93% 酸耗 (35%盐酸）咖屯镍） 200 140 5 5 外加中和剂 （氧化钙等） 7 4 4 -
(n¾/吨镍）
硫化剂用量 (吨麵镍） 8 , 1. 4 1. 4 0. 6 废水纖 (m3/吨镍） 190 130 - - 优缺点 设备投资小， 浸出过程得到 浸出过程得到 浸出设备投资 但- 强化， 但： 强化， 部分盐 较小； 铁大量溶出； 浸出设备投资 酸被再生， 镁 镍钴回收率 镁铁未综合利 大； 得到利用，但： 高；
用； 镁铁未综合利 铁未综合利用 镁铁综合利用 酸耗大； 用； 酸耗大； 好；
中和剂与硫化 酸耗大； 中和剂与硫化 酸耗小； 剂用量高； 中和剂与硫化 剂用量高； 中和剂与硫化 环境污染大； 剂用量高； 环境污染大； 剂用量低； 设备处理能力 环境污染大； 设备处理能力 环境友好； 小； 设备处理能力 小； 设备处理能力 小； 大；
注： 对比技术 1的特征为： 40- 95 °C下常压盐酸浸出， 力口硫化钠沉淀镍钴， 母液中和后排放;
对比技术 2的特征为： 200-300°C, 3-lOMPa下高温高压盐酸浸出， 力 H硫化钠沉淀镍钴， 母液中和后排放 对比技术 3的特征为： 200-300°C, 3-lOMPa下高温高压盐酸浸出， 加硫化钠沉淀镍钴， 母液高温雾化焙烧再 生盐酸
本发明技术的特征为： 150-165Ό, 0.3-0.5MPa下采用盐酸与金属氯化物为浸出剂加 ί¾]口压浸出； 浸出渣回收 铁处理； ¾»出液进行浓缩、 氧化、 中和后采用复合硫化剂沉淀， 用萃取除铁后中和沉镍钴； 取镍 钴后的母液进行雾化干燥、 焙烧再生盐艱综合利用镁铁。
综上所述， 本发明的特点有：
本发明通过釆用盐酸与金属氯化物混合溶液作为浸出剂； 在适当的加温与压力的 下对红土 镍矿进行浸出； 釆用浓缩一氧化一中和进行预先除铁处理； 采用复合硫化剂沉淀镍钴， 镍钴综合 回收率髙， 可以达到 95.5%以上。 （参见本发明上述"优点与积极效果"中的第 1、 2、 4、 6、 7、 8、 15点）；
通过将溶液浓缩时产生的热氯化氢气体与水汽返回到浸出过程； 雾化干燥、 与焙烧过程中产生 的炙热氯化氢气体与水汽用于加热、 浓缩溶液后返回到浸出过程； 沉镍母液中的氯化物及浓缩结 晶时产生的氯化物均通过高温水解再生为盐酸， 返回到氯化法处理红土镍矿过程中， 实现了水与 盐酸在处理红土 W时盐酸的闭路循环， 无废水排放， 对环境友好。 （发明优点中的第 4、 10、 12、 13、 14、 15点）。
在高效提取镍钴的基础上，通过对浸出渣的回收铁处理作为炼铁原料；焙烧过程产生的氧化铁、 氧化镁为中和剂； 焙烧后得到可以用于磁性材料、 铁颜料、 炼铁原料、 提镁材料、 耐火材料的多 种铁、镁化合物产品， 实现了红土镍矿中镍、钴、铁、镁资源的的综合利用。（参见本发明上述"优 点与积极效果"中的第 3、 6、 11、 15点）；
通过釆用盐酸与金属氯化物混合溶液作为浸出^; 在适当的加温与压力的^ ί牛下对红土 W进 行浸出； 浸出渣的回收铁处理； 采用浓缩一氧化一中和进行预先除铁处理； 采用复合硫化剂沉淀 镍钴； 沉镍后母液雾化干燥、 焙烧； 过程中余热与残余酸的回收利用， 实现资源高效利用与清洁 生产（参见本发明上述 "优点与积极效果"中的第 1一 点)。
附图说明
图 1 本发明一种实施例的湿法氯化处理红土镍矿的工艺流程示意图。
具体实施方式：
下面结合具体实施对本发明做进一步描述。 本发明可以按发明内容的任一方式实施， 但这些 实施例的给出决不是限制本发明。
以下实施例釆用的红土镍矿， 其中一种红土镍矿的主要有价金属含量分别为： Ni 0.92%； Co 0.07%； Mg 14.21%； Fe 19.36%。 经粗破碎。
实施例 1
(1 )浸出
A.按料水比 1 :2 (wt)加水将红土镍矿矿石球磨并过 100目筛。取 -100目浆料泵入压力反应器中； 按液固比 4.0: 1、浸出剂与料（重量）比为 1.4:1分别泵入水和氯化物浸出剂 (HC1含量 28%以上， 镁离子与铁离子含量之和:氯离子的含量为 0.05 : l(wt))， 搅拌浸出。
B. 浸出条件控制： 150°C (最大压强控制为 0.3MPa)下搅拌浸出 2小时。
C. 趁热过滤、 对残渣进行洗涤。 有价金属综合浸出率分别为： Ni+Co 95.8%； Mg 95.0%； Fe 4.6%。
经氯化物加压浸出红土镍矿，浸出液镍钴含量（g/L)分别为： Ni+Co 2.27； Mg 33.5 Fe2.16; 浸出液酸度为： Η" 0.82πιο1/ί。 下述实施例以此浸出液为基础。
浸出渣（干）有价金属成分 (%)为： Μ 0.04； Co -0.01； Mg 0.63； Fe 22.6。浸出渣含水 16.8%。 下述实施例以此浸出 为基础。
(2)浸出渣回收铁
A. 造球： X寸含水浸出渣、 复合粘结剂、还原剂混合均匀， 并用园盘式造粒机造粒。 复合粘结剂为 80%腐植酸钠 +10%生石灰 +10%碳酸钠， 配入量为浸出渣干量的 3%; 还原剂为烟煤，配入量为 浸出渣干量的 10%。
B. 磁化焙烧： 生球入还原焙烧回转窑中， 焙烧温度 700°C， 焙烧时间 40分钟。
C. 磁选： 还原熟球经破碎、 球磨， 并过 200目筛； -200目粉料用弱磁选机进行磁选， 磁选磁场强 度 100特斯拉； 磁选产物铁含量 58.6%; 铁收率 86.6%。
ΐϋ (3)镍钴提取
利用盐酸焙烧过程中产生的炙热氯化氢气体加热浸出液， 使浸出液体积减小为原来的 20%~30%， 氯化镁、 氯化铁以晶体形式析出， 浸出液中的氯化氢气体与水大量析出， 经收集引入 到浸出剂中。 固液分离后， 往热的浓缩液 (>60°C ) 中通入空气进行氧化处理 lh,使溶液中的二 价铁氧化成三价铁。 然后以母液雾化干燥或焙烧过程中产生的氧化镁中和浸出液的残酸， 将浸出 液的 pH值调整到 3~5， 使三价铁水解实现沉铁， 过滤沉淀。
采用硫化氢（硫化剂 A)及硫化钠 （硫化剂 B)为复合硫化剂进行硫化沉淀提取镍钴。 分别缓 慢加入硫化钠及通入硫化氢， 使得沉淀过程中溶液的 pH值保持 3. 7-4. 3之间， 按照沉淀硫化镍、 硫化钴所需硫化剂理论量的 1. 05倍控制复合硫化剂的加入量， 经固液分离、 洗涤后得到富含镍、 钴的硫化物产品， 滤液经雾化干燥、 焙烧处理进行盐酸的再生与镁、 铁资源的综合利用。
(4)提镍钴后母液雾化干燥
提取镍钴后的母液进行浓缩与雾化干燥。 利用焙烧工序来的高温炉气， 进行浓缩， 将沉镍后 的母液调整到 C1总量为 300g/L， 再在 300°C下雾化干燥得到含不同结晶水的氯化镁、 与不同结晶 水的氯化铁固体混合物。
(5) 烧回收盐酸与综合利用镁铁
按照煤粉在焙烧物料中的质量百分数为 20%， 雾化干燥产物在焙烧物料中的质量百分数为 80%进行备料，在回转窑中进行焙烧。物料沿着具有可控温度与气氛的炉膛移动，物料先在 68(TC、 氧化气氛下进行氧化烧烧， 然后在 800 、弱还原气氛下进行弱还原磁性化焙烧， 得到以四氧化三 铁、 氧化镁、 铁酸镁混合物， 经磁选后， 磁性的四氧化三铁、 铁酸镁用于炼铁原料或其它铁质原 料， 非磁性的氧化镁用于耐火材料或作为其它产品的原料。
焙烧过程中产生的炙热氯化氢气体通入通入加热套管中对红土镍矿浸出液进行加热浓缩， 冷 却后的氯化氢气体及从浸出液中析出的氯化氢气体经冷却吸收后得到再生热盐酸， 再生的热盐酸 返回前端用于红土矿的盐酸热浸。
实施例 2
按实施例 1相同的方式进行红土镍矿中有价金属 提取与综合利用， 只是其中的步骤 (1)按以 下方式进行：
Α. 按料水比 1 : 1 (wt)加水将红土镍矿球磨并过 100目筛。取 -100目 料泵入压力反应器中， 按 液固比 3.0: 1、 浸出剂与料（重量） 比为 1.4:1分别泵入水和氯化物浸出剂 (盐酸与提取镍钴后 母液的混合溶液， HC1含量 28%以上， 镁离子与铁离子含量之和:氯离子的含量为 0.1 : 1 )， '搅 拌浸出。
B. 浸出条件控制： 165°C (最大压强控制为 0.45MPa)下搅拌浸出 1小时。
C. 趁热过滤、 对残渣进行洗涤。 有价金属综合浸 率分别为： Ni+Co 96.2%； Mg 95.5%； Fe 3·4%。
实施例 3
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (1 )按以 下方式进行：
D. 按料水比 1 : 1 (wt)加水将红土镍矿球磨并过 100目筛。取 -100目浆料泵入压力反应器中， 按 液固比 3.5 : 1、 浸出剂与料（重量） 比为 1.35:1分别泵入水和氯化物浸出剂 (浓缩结晶时得到的 氯化镁、 氯化铁与盐酸的混合溶液， HC1含量 28%以上， 镁离子与铁离子含量之和:氯离子的 含量为 03 : 1 )， 搅拌浸出。
E. 浸出条件控制： 180°C (最大压强控制为 0.5MPa)下搅拌浸出 0.5小时。
F. 趁热过滤、 对残渣进行洗涤。 有价金属综合浸出率分别为： Ni+Co 96.1%； Mg 95.6%； Fe 3.1%。
实施例 4
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (1 )按以 下方式进行- A. 按料水比 1 :2 (wt)加水将红土镍矿球磨并过 100目筛。 取 -100目浆料泵入压力反应器中， 按 液固比 4.0: 1、浸出剂与料（重量）比为 1.35:1分别泵入水和盐酸 (盐酸与反萃得到的溶液的混合 液， HC1含量 28%以上， 镁离子与铁离子含量之和:氯离子的含量为 0.02: 1 )， 搅拌浸出。
B. 浸出^ ί空制： 160°C (最大压强控制为 0.40MPa)下搅拌浸出 1.5小时。
C 趁热过滤、 对残渣进行洗漆。 有价金属综合浸出率分别为： Ni+Co 95.9%； Mg 95.0%； Fe 4.6%。 实施例 5
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (1 )按以 下方式进行：
A. 按料水比 1 :2 (wt)加水将红土镍矿球磨并过 100目筛。 取 -100目浆料泵入压力反应器中， 按 液固比 4.0: 1、 浸出剂与料（重量） 比为 2.4: 1分别泵入水和盐酸 (HC1含量 28%以上， 镁离子与 铁离子含量之和:氯离子的含量为 0.08 : 1 )， 搅拌浸出。
B. 浸出条件控制： 60°C (最大压强控制为 0.12MPa)下搅拌浸出 2小时。
C. 趁热过滤、 对残渣进行洗漆。 有价金属综合浸出率分别为： Ni+Co 92.3%； Mg 93.80%； Fe 47%。
实施例 6
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤（2)按以 下方式进行：
A. 造球： 对含水浸出渣、 复合粘结剂、 还原剂混合均匀， 并用园盘式造粒机造粒。 复合粘结剂为 80。/。腐植酸钠 +10%生石灰 +10%碳酸钠 配入量为；浸 渣干量的 2.5%; 还原剂为烟煤， 配入量为浸 出渣干量的 8%。
B. 磁化焙烧： 生球入还原焙烧回转窑中， 焙烧温度 750°C, 焙烧时间 40分钟。
C. 磁选： 还原熟球经破碎、 球磨， 并 ¾200目筛； -20ϋ目粉料用弱磁选机进行磁选， 磁选磁场强度 100特斯拉； 磁选产物铁含量 58.5%; 铁收率 86.4%。
实施例 7
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (2)按以 下方式进行：
Α. 造球： 对含水浸出渣、 复合粘结剂、 还原剂混合均匀， 并用园盘式造粒机造粒。 复合粘结剂为 80%腐植酸钠 +10%生石灰 +10%碳酸钠， 配入量为浸出渣干量的 2%; 还原剂为烟煤， 配入量为浸出 渣干量的 6%。
Β. 磁化焙烧： 生球入还原; ^烧回转窑中， 焙烧温度 780°C， 焙烧时间 30分钟。
C. 磁选： 还原熟球经破碎、 球磨， 并1200目筛； -200目粉料用弱磁选机进行磁选， 磁选磁场强度 80特斯拉； 磁选产物铁含量 62.7%; 铁收率 86.8%。
实施例 8 (浸出液通空气氧化； 采用硫化氢及硫化钠复合硫化剂， 控制 pH值为 3. 0-3. 5进行硫 化沉淀剂提取镍钴）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤（3)按 以下方式进行：
利用盐酸焙烧过程中产生的炙热氯化氢气体加热浸出液， 使浸出液体积减小为原来的 25%左 右， 氯化镁、 氯化铁以晶体形式析出， 浸出液中的氯化氢气体与水大量析出， 经收集引入到浸出 剂中。 固液分离后， 往热的浓缩液 ( >60°C ) 中通入空气进行氧化处理 lh,使溶液中的二价铁氧 化成三价铁。 然后以母液雾化干燥或焙烧过程中产生的氧化镁中和浸出液的残酸， 将浸出液的 pH 值调整到 3.0， 使三价铁水解实现沉铁， 过 沉淀,，
釆用硫化氢 （硫化剂 A)及硫化钠 （硫化剂 B) 为复合硫化剂进行硫化沉淀提取镍钴。 分别缓 慢加入硫化钠及通入硫化氢， 使得沉淀过程中溶液的 值保持 3. 0—3. 5之间， 按照沉淀硫化镍、 硫化钴所需硫化剂理论量的 1. 1倍控制复合硫化剂的加入量， 经固液分离、 洗绦后得到富含镍、 钴的硫化物产品， 滤液经雾化干燥、 烧处理进行盐酸的再生与镁、 铁资源的综合利用。
实施例 9 (浸出液通富氧空气氧化； 采用硫氢化钠， 硫化钠与本流程中刚沉淀的硫化镍作为 复合硫化剂， 控制 pH值进行硫化沉淀剂提取镍钴）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (3)按以 下方式进行：
利用盐膨咅烧过程中产生的炙热氯化氢气体加热浸出液， 使浸出液体积减小为原来的 20%， 氯化镁、 氯化铁以晶体形式析出， 浸出液中的氯化氢气体与水大量析出， 经收集引入到浸出剂中。 固液分离后， 往热的浓缩液 (>60°C) 中通入富氧空气（氧气： 空气体积比为 1: 1 ) ,使溶液中 的二价铁氧化成三价铁。 然后以母液雾化干燥或焙烧过程中产生的氧化镁中和浸出液的残酸， 将 浸出液的 pH值调整到 3.5， 使三价铁水解实现沉铁' 过滤沉淀， 经热处理后得到铁红。
复合硫化剂中的硫化剂 A采用硫氢化钠，硫化剂 B为硫化钠与本流程中刚沉淀的硫化镍的混合 物， 分别缓慢加入硫化钠及通入硫化氢， 使得沉淀过程中溶液的 pH值保持 3. 5±0. 3， 按照沉淀硫 化镍、 硫化钴所需硫化剂理论量的 1. 08倍控制复合硫化剂的加入量， 经固液分离、 洗涤后得到富 含镍、 钴的硫化物产品， 滤液经雾化干燥、 谙烧处理进行盐酸的再生与镁、 铁资源的综合利用。
实施例 10 (浸出液通富氧空气氧化；采用复贫硫化剂、控制 pH值进行硫化沉淀剂提取镍钴） 按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (3)按以 下方式进行：
利用盐酸焙烧过程中产生的炙热氯化氢气体加热浸出液， 使浸出液体积减小为原来的 30%， 氯化镁、 氯化铁以晶体形式析出， 浸出液中的氯化氢气体与水大量析出， 经收集引入到浸出剂中。 固液分离后， 往热的浓缩液（>60°C ) 中通入富氧空气 （氧气： 空气体积比为 1 : 1 ) ,使溶液中 的二价铁氧化成三价铁。 然后以母液雾化千燥或焙烧过程中产生的氧化镁中和浸出液的残酸， 将 浸出液的 pH值调整到 4.0， 使三价铁水解实现沉铁， 过滤沉淀， 经热处理后得到铁红。
复合硫化剂中的硫化剂 A釆用硫氢化钠、过硫化氢、硫氢化铵的混合物溶液， 硫化剂 B采用硫 化钾、 硫化钙、 硫化铵、 多硫化钠、 本流程中刚沉淀的硫化铁的混合物， 分别缓慢加入上述硫化 剂 A与硫化剂 B， 使得沉淀过程中溶液的 pH值保持为 4. 0±0. 30， 按照沉淀硫化镍、 硫化钴所需 硫化剂理论量的 1. 06倍控制复合硫化剂的加入量， 经固液分离、 洗漆后得到富含镍、 钴的硫化物 产品， 滤液经雾化干燥、 焙烧处理进行盐酸的再生与镁、 铁资源的综合利用。
实施例 11 (通空气氧化， 采用硫氢化钠、过硫化钠、本流程中刚沉淀的硫化镍为复合硫化剂、 控制 A、 B组分比例进行硫化沉淀剂提取镍钴） ■
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (3)按 以下方式进行：
利用盐酸焙烧过程中产生的炙热氯化氢气体加热浸出液， 使浸出液体积减小为原来的 25%左 右， 氯化镁、 氯化铁以晶体形式析出， 浸出液中的氯化氢气体与水大量析出， 经收集引入到浸出 剂中。 固液分离后， 往热的浓缩液 (>60°C ) 中通入空气进行氧化处理 lh,使溶液中的二价铁氧 化成三价铁。 然后以母液雾化干燥或焙烧过程中产生的氧化镁中和浸出液的残酸， 将浸出液的 pH 值调整到 2.2， 使三价铁水解实现沉铁， 过滤沉淀。 复合硫化剂中的硫化剂 A采用硫氛化钠， mm B为 化钠与本流程中刚沉淀的硫化镍的混 合物。 硫氢化钠、 过硫化钠、 本流程中刚沉淀的硫化镍按质量比为 1: 97： 2混合后缓慢加入， 使 浸出液的 pH值缓慢变化。 按照沉淀硫化镍、 硫化钴所需硫化剂理论量的 1. 06倍控制复合硫化剂 的加入量， 经固液分离、 洗涤后得到富含镍、 钴的硫化物产品， 滤液经雾化干燥、 焙烧处理进行 盐酸的再生与镁、 铁资源的综合利用。
实施例 12 (通富氧空气氧化， 采用硫氢化钠、 硫化钠、 本流程中刚沉淀的硫化镍为复合硫化 剂、 控制 A、 B组分比例进行硫化沉淀剂提取镍钴）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (3)按 以下方式进行：
利用盐酸焙烧过程中产生的炙热氯化氢气体加热浸出液， 使浸出液体积减小为原来的 25%， 氯化镁、 氯化铁以晶体形式析出， 浸出液中的氯化氢气体与水大量析出， 经收集引入到浸出剂中。 固液分离后， 往热的浓縮液 ( >60°C ) 中通入富氧空气 （氧气： 空气体积比为 1 : 1 ) ,使溶液中 的二价铁氧化成三价铁。 然后以母液雾化干燥或焙烧过程中产生的氧化镁中和浸出液的残酸， 将 浸出液的 pH值调整到 4.0， 使三价铁水解实现沉铁， 过滤沉淀。
复合硫化剂中的硫化剂 A采用硫氢化钠， 硫化剂 β采用硫化钠、 本流程中刚沉淀的硫化镍的 混合物。 按照硫氢化钠、 硫化钠、 本流程中刚沉淀的硫化镍的质量比为 18: 81： 1混合后缓動口 入， 使浸出液的 pH值缓慢变化。 按照沉淀硫化镍、 硫化钴所需硫化剂理论量的 1. 1倍控制复合硫 化剂的加入量， 经固液分离、 洗涤后得到富含镍、 钴的硫化物产品， 滤液经雾化干燥、 焙烧处理 进行盐酸的再生与镁、 铁资源的综合利用。
实施例 13 (萃取除铁， 采用硫氢化钠、 硫化钠、 本流程中刚沉淀的硫化镍为复合硫化剂、 控 制 Α、 Β组分比例进行硫化沉淀剂提取镍钴） ·
按实施例 1相同的方式进行红土镍矿中有价金属.的 取与综合利用， 只是其中的步骤 (3)按以 下方式进行：
浸出液用 20%Ν235+10%丙三醇 +煤油萃取除铁： 用 3级萃取、 5级反萃取串联， 常温下操， 萃 取相比 0/Α=1:3， 反萃取相比 0/Α=5:1。 铁去除率 99.6%， 得到反萃取液为氯化铁酸性溶液， 经浓 缩、 雾化干燥与焙烧后再生盐酸并回收利用铁。 利用盐酸焙烧过程中产生的炙热氯化氢气体加热 浸出液， 使萃取除铁的萃取余液体积减小为原来的 25%, 氯化镁、 氯化铁以晶体形式析出， 浸出 液中的氯化氢气体与水大量析出， 经收集引入到浸出剂中。 浓缩后萃余液进行采用复合硫化剂沉 淀镍钴。
复合硫化剂中的硫化剂 Α釆用硫氢化钠， 硫化剂 B釆用硫化钠、 本流程中刚沉淀的硫化镍的 混合物。 以母液雾化干燥或焙烧过程中产生的氧化镁中和浓缩后萃余液浸出液的残酸， 将浸出液 的 pH值调整到 3.0， 按照硫氢化钠、 硫化钠、 本流程中刚沉淀的硫化镍的质量比为 5: 92： 3混 合后缓慢加入， 使浸出液的 PH值缓慢变化。 按照沉淀硫化镍、 硫化钴所需硫化剂理论量的 1. 06 倍控制复合硫化剂的加入量， 经固液分离、 洗漆后得到富含镍、 钴的硫化物产品， 滤液经雾化干 燥、 燎烧处理进行盐酸的再生与镁、 铁资源的综合利用。
实施例 14 (萃取除铁， 中和沉镍钴）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (3)按以 下方式进行：
浸出液用 20%N235+10%丙三醇 +煤油萃取除铁： 用 5级萃取、 3级反萃取串联， 常温下操， 萃 取相比 0/Α=1:4， 反萃取相比 0/Α=4:1。 铁去除率 99.7%。 得到反萃取液为氯化铁酸性溶液， 经浓 缩、 雾化干燥与焙烧后再生盐酸并回收利用铁。 利用盐酸焙烧过程中产生的炙热氯化氢气体加热 浸出液， 使萃取除铁的萃取余液体积减小为原来的 30%， 氯化镁、 氯化铁以晶体形式析出， 浸出 液中的氯化氢气体与水大量析出， 经收集引入到浸出剂中。 浓缩后萃余液进行采用复合硫化剂沉 淀镍钴。
复合硫化剂中的硫化剂 A采用硫氢化钠， 硫 剂 B采用硫化钠、 本流程中刚沉淀的硫化镍的 混合物。 以母液雾化干燥或焙烧过程中产生的氧化镁中和浓缩后萃余液浸出液的残酸， 将浸出液 的 pH值调整到 3.5， 按照硫氢化钠、 硫化钠、 本流程中刚沉淀的硫化镍的质量比为 8: 90： 2裩 合后缓慢加入， 使浸出液的 pH值缓慢变化。 按照沉淀硫化镍、 硫化钴所需硫化剂理论量的 1 06 倍控制复合硫化剂的加入量， 经固液分离、 洗涤后得到富含镍、 钴的硫化物产品， 滤液经雾化干 燥、 焙烧处理进行盐酸的再生与镁、 铁资源的综合利用。
经分析， 实施例 9-14中镍钴沉淀率均为 99. 9%以上。
实施例 15 (母液浓缩、 雾化干燥）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (4)按以 下方式进行：
沉镍后的母液进行浓缩与雾化干燥。 利用焙烧工序来的高温炉气， 进行浓缩， 将沉镍后的母 液调整到 C1总量为 350g L， 再在 210°C下雾化干燥得到含不同结晶水的氯化镁、 与不同结晶水的 氯化铁固体混合物。
实施例 16 (母液浓缩、 雾化干燥）
按实施例 1相同的方式进行红土镍矿中有价金属'; ¾^取与综合利用， 只是其中的步骤（4)按以 下方式进行：
沉镍后的母液进行浓缩与雾化干燥。 利用焙烧工序来的高温炉气， 进行浓縮， 将沉镍后的母 液调整到 C1总量为 250g/L， 再在 380°C下雾化干燥得到含不同结晶水的氯化镁、 与不同结晶水的 氯化铁固体混合物。
实施例 17 (母液直接雾化干燥） 按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (4)按 以下方式进行- 沉镍后的母液直接进行雾化干燥， 利用焙烧工序来的高温炉气， 将沉镍后的母液直接在 280 °C下雾化干燥得到含不同结晶水的氯化镁、 与不同结晶水的氯化铁固体混合物。
实施例 18 (母液直接雾化干燥）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (4)按 以下方式进行：
沉镍后的母液直接进行雾化干燥， 利用焙烧工序来的高温炉气， 将沉镍后的母液直接在 370 °C下雾化干燥得到含不同结晶水的氯化镁、 与不同结晶水的氯化铁固体混合物。
实施例 19 (低温雾化焙烧）
按实施例 1相同的方式进行红土镍矿中有价金 的提取与综合利用， 只是其中的步骤 (5)按 以下方式进行：
将雾化干燥产物与天燃气、高压空气一起喷入高温炉中在 42CTC下进行雾化焙烧。得到氯化镁 及氧化铁、 少量铁酸镁， 水洗后的固体产物为氧化铁、 铁酸镁， 可以用于磁性材料、 铁颜料、 炼 铁原料等。 氯化镁水溶液可以再高温水解回收盐酸与氧化镁， 氧化镁可以用于耐火材料。 ； 焙烧过程中产生的炙热氯化氢经换热后直接经水吸收后获得再生盐酸， 再生的盐酸返回前端 用于红土矿的盐酸浸出。
实施例 20 (低温雾化; ^烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (5)按 以下方式进行：
按照煤粉在焙烧物料中的质量百分数为 20%， 雾化干燥产物在焙烧物料中的质量百分数为 50%, 浸出液浓缩时得到的氯化铁、 氯化镁晶体在焙烧物料中的质量百分数为 30%进行备料， 与 液化石油气、高压空气一起喷入高温炉中在 48CTC下进行雾化焙烧。得到氯化镁及氧化铁、少量铁 酸镜， 7洗后的固体产物为氧化铁、 铁酸镁， 可以用于磁性材料、 铁颜料、 炼铁原料等。 氯化镁 水溶液可以再高温水解回收盐酸与氧化镁， 氧化镁可以用于耐火材料。
焙烧过程中产生的炙热氯化氢经换热后直接经水吸收后获得再生盐酸， 再生的盐酸返回前端 用于红土矿的盐酸浸出。
实施例 21 (中温氧化燎烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (5)按 以下方式进行：
按照焦炭粉在焙烧物料中的质量百分数为 10%, 雾化干燥产物在焙烧物料中的质量百分数为 70%进行备料， 提镍后母液浓缩时得到的氯化铁、 化镁晶体在焙烧物料中的质量百分数为 20% 进行备料，与高压空气一起喷入高温炉中 800°C下 化 烧。得到氧化镁、氧化 铁酸镜固体混 合物， 该混合物部分返回前端中和浸出液中的残余酸， 其余用于炼铁及提镁原料。
焙烧过程中产生的炙热氯化氢气体直接通入红土镍矿浸出液中对浸出液进行加热浓缩， 从浸 出液中析出的氯化氢气体经冷却吸收后得到再生热盐酸； 再生的热盐酸返回前端用于红土矿的盐 酸热浸。
实施例 22 (中温氧化愔烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (5)按 以下方式进行：
将雾化干燥产物与柴油、高压空气一起喷入高温炉中在 700 下进行雾化焙烧。得到氯化镁及 氧化铁、 铁酸镁， 7洗后的固体产物为氧化铁、 铁酸镁， 可以用于磁性材料、 铁颜料、 炼铁原料 等。氯化镜水溶液可以再高温水解回收盐酸与氧化镁， 氧化镁可以用于耐火材料。
焙烧过程中产生的炙热氯化氢经换热后直接经水吸收后获得再生盐酸， 再生的盐酸返回前端 用于红土矿的盐酸浸出。
实施例 23 (中温氧化焙烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤（5)按 以下方式进行：
将雾化干燥产物与高压空气一起喷入高温炉中在 1000'C下进行雾化;^烧。 得到氧化镁及氧化 铁， 固体产物经强磁选， 磁性产物为氧化铁、铁酸镁， 可以用于磁性材料、麵料、炼铁原料等。 非磁性氧化镁可以用于耐火材料。
焙烧过程中产生的炙热氯化氢经换热后直接经水吸收后获得再生盐酸， 再生的盐酸返回前端 用于红土矿的盐酸浸出。
实施例 24 (中温氧化; ^烧、 弱还原磁化焙烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (5)按 以下方式进行：
按照煤粉在焙烧物料中的质量百分数为 25%, 雾化干燥产物在焙烧物料中的质量百分数为
75%进行备料，在回转窑中进行焙烧。物料沿着具有可控温度与气氛的炉膛移动，物料先在 600°C、 氧化气氛下进行氧化焙烧， 然后在 750 、弱还原气氛下进行弱还原磁性化焙烧，得到以四氧化三 铁、 氧化镁、 铁酸镁混合物， 经磁选后， 磁性的四氧化三铁、 铁酸镁用于炼铁原料或其它铁质原 料， 非磁性的氧化镁用于耐火材料或作为其它产品的原料。
焙烧过程中产生的炙热氯化氢气体通入通入加热套管中对红土镍矿浸出液进行加热浓缩， 冷 却后的氯化氢气体及从浸出液中析出的氯化氢气体经冷却吸收后得到再生热盐酸， 再生的热盐酸 返回前端用于红土矿的盐酸热浸。 实施例 25 (中温氧化谙烧、 弱还原磁化; ^烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (5)按 以下方式进行：
按照煤粉在焙烧物料中的质量百分数为 30%， 雾化干燥产物在焙烧物料中的质量百分数为 20%进行备料， 浸出液处理时得到的氯化铁、 氯化镁晶体在;^烧物料中的质量百分数为 50%进行 备料，在回转窑中进行焙烧。物料沿着具有可控温 与气氛的炉膛移动， 物料先在 700°C、氧化气 氛下进行氧化焙烧， 然后在 830°C、弱还原气氛下进行弱还原磁性化焙烧， 得到以四氧化三铁、氧 化镁、铁酸镁混合物， 经磁选后， 磁性的四氧化三铁、 铁酸镁用于炼铁原料或其它铁质原料， 非 磁性的氧化 于耐火材料或作为其它产品的原料。
焙烧过程中产生的炙热氯化氢气体直接通入红土镍矿浸出体系用于红土矿的盐酸热浸。
实施例 26 (高温氧化焙烧、 强还原磁化 烧）
按实施例 1相同的方式进行红土镍矿中有价金属的提取与综合利用， 只是其中的步骤 (5)按 以下方式进 fi1:
按照焦炭在焙烧物料中的质量百分数为 50%， 雾化干燥产物在焙烧物料中的质量百分数为 50%进行备料， 在回转窑中进行焙烧。物料沿着具有可控温度与气氛的炉膛移动， 物料先在 1000 V、氧化气氛下进行氧化焙烧， 然后在 1350°C、 强还原气氛下进行还原磁性化焙烧， 得到金属铁 与氧化镁混合物， 磁选出铁用于炼铁原料， 氧化镁用于耐火材料或作为其它产品的原料。
焙烧过程中产生的炙热氯化氢气体直接通入红土镍矿浸出体系用于红土矿的盐酸热浸。
实施例 27 (高温氧化焙烧、 强还原磁化焙烧）
按实施例 1相同的方式进行红土镍矿中有价金 的提取与综合利用，只是其中的步骤 (5)按以 下方式进行：
按照煤粉与焦炭在焙烧物料中的质量百分数各占 20%， 雾化干燥产物在焙烧物料中的质量百 分数为 60%进行备料， 在回转窑中进行焙烧。物料沿着具有可控温度与气氛的炉膛移动， 物料先 在 1100。C、 氧化气氛下进行氧化焙烧， 然后在 1250 、 强还原气氛下进行还原磁性化焙烧， 得到 金属铁与氧化镁混合物， 磁选出铁用于炼铁原料， -氧化镁用于耐火材料或作为其它产品的原料。
焙烧过程中产生的炙热氯化氢气体通入通入加热套管中对红土镍矿浸出液进行加热浓缩， 冷 却后的氯化氢气体及从浸出液中析出的氯化氢气杯 冷却吸收后得到再生热盐酸， 再生的热盐酸 返回前端用于红土矿的盐酸热浸。

Claims

权 利 要 求
1、 一种盐酸全闭路循环法从红土镍矿中提取有价金属的方法， 它以加工破碎的红土镍矿 矿石为原料， 其特征是， 该方法包括氯化浸出、 出渣回收铁、 镍钴提取、 母液雾化干燥、 焙烧； 其中：
( 1 )氯化浸出： 以盐酸与金属氯化物的混合溶液作为浸出剂， 将投入所述浸出剂的矿石 中镍、 钴、 镁以氯化物进入浸出剂中， 将铁抑制在浸出渣中；
所述浸出在加压加温的浸出槽中， 并于温度为 60°C~180°C、 压力 0.11MPa~0.5MPa的条件 下进行， 浸出时间为 0.5小时〜 2小时；
所述浸出剂与矿石的重量比为 1.3~2.5 : 1， 补氷控制物料中总的液体与固体的重量比 3-4 : 1；
所述浸出剂中 HC1含量 28%以上， 浸出剂中金属离子含量之和：氯离子含量为 0.02~0.3 : 1
(wt) ;
所述金属氯化物为氯化镁或氯化铁、 氯化亚铁、 氯化铝、 氯化钙中的一种或一种以上； 浸出所得为含镍、 钴、 镁氯化物的浸出液和含铁的浸出渣；
(2) 从浸出渣中回收铁： 所述浸出渣与浸出液经固液分离后， 将浸出渣造球、 还原磁 化焙烧， 再经破碎、 磁选富集铁， 用于炼铁； 其中： '
造球是将含水浸出渣、 常规复合粘结剂、 还原剂混合后制造成 5~12mm生球粒;'所述复 合粘结剂配入量为浸出渣干重的 2~3%； 还原剂为木炭或无烟煤、 烟煤或褐煤， 所述还原剂在 造球时配入量为浸出渣干重的 6%~10%；
还原磁化焙烧是在回转窑中进行， 通过控制焙烧温度 700°C~78(TC， 焙烧时间 30分钟〜 40 分钟得到磁性氧化铁， 焙烧后得到的熟球经破碎、 球磨， 并过 200目筛， 过筛的粉料磁选用弱 磁选机， 磁选磁场强度 80特斯拉〜 100特斯拉， 磁选产物控制铁含量 55%~65%；
(3 ) 镍钴提取：
先将所述浸出液采用以下两种方法之一进行除铁处理： '
a.所述浸出液进行浓缩、 氧化、 水解除铁； 除铁后的滤液为除铁后浸出液；
b.采用萃取方法对所述浸出液进行除铁处理，萃取除铁后的萃取余液经浓缩后为除铁后浸 出液； 从萃取含铁的有机相中反萃得到的氯化铁溶液；
对经上述两种除铁方法所得除铁后浸出液， 采用复合硫化剂沉淀镍钴的方法提取镍钴， 即采用下述两种方法之一在所述除铁后浸出液中加入复合硫化剂： ( i )硫化剂 A与硫化剂 B分别缓慢加入或按比例混合后缓慢加入， 或者部分混合后与硫 化剂 A或硫化剂 B分别加入， 使得浸出液的 pH值在沉淀过程中保持在设定值的 ±0. 30之间；
(ii)硫化剂 A与硫化剂 B按质量比为 X： ( 100— X )加入， 其中 x在 1-20范围内取值， 硫 化剂按比例加入的方式为： 硫化剂 A与硫化剂 B分别缓慢加入或混合后缓慢加入， 或者部分 混合后与硫化剂 A或硫化剂 B分别加入， 使浸出液的 pH值保持不变或缓慢变化；
所述硫化剂 A为含氢离子的硫化物， 包括硫化氢或过硫化氢、 硫氢化钠、 硫氢化钾、 硫 氢化铵中的一种或一种以上；
所属硫化剂 B为不含氢离子的硫化物， 包括硫化钠或硫化钾、 硫化铵、 硫化亚铁、 硫化 铁、 硫化镁以及本流程中沉淀产生的硫化铁、 硫化镁、 硫化镍、 硫化钴、 多硫化钠、 多硫化 铵中的一种或一种以上；
所述除铁后浸出液中加入复合硫化剂沉淀镍钴后， 经固液分离得到镍钴硫化物沉淀， 以 及氯化镁溶液或氯化镁与氯化铁的混合溶液； 其中， 所述镍钴硫化物经传统冶金方法生产镍、 钴金属或化合物产品；
(4) 雾化干燥： 利用焙烧工序来的高温炉 ^萃取分离铁的反萃氯化铁溶液、 提取镍 钴后的氯化镁溶液或氯化镁与氯化铁的混合溶液， 控制 C1的总量 >250 g/L， 再在 200°C— 400 °C下雾化干燥， 得到含不同结晶水的氯化铁、 或含不同结晶水的氯化镁、 或它们的固体混合 物， 即雾化干燥产物； 雾化干燥的余热用于向前述 "氯化浸出"或 "镍钴提取" 中 : "浸出液 的浓缩、 氧化"供热； ·
(5)赔烧：
将所述雾化干燥的产物， 上述步骤 （3) 中所得氯化镁溶液或氯化镁与氯化铁的混合溶液 浓缩结晶得到的氯化镁、 氯化铁晶体， 或雾化干 与所述浓缩结晶得到的氯化镁、 氯化 铁晶体的混合物进行焙烧； 所述焙烧可采用下述常规焙烧方法之一进行：
(a)低温 （420-500°C ) 焙烧， 用于氯化铁、 或氯化铁与氯化镁的混合物的焙烧， 得到铁 的氧化物， 或铁的氧化物与氯化镁的混合物， 水洗混合物后的固体产物为氧化铁等铁的氧化 物；
(b) 中温 （600°C-1000°C ) 氧化焙烧， 用于氯化镁、 或氯化铁与氯化镁的混合物的焙烧， 得到氧化镁、 或氧化镁与铁的氧化物的固体混合物， 所述混合物部分返回前端步骤（1 ) 中和 浸出液中的残余酸；
(c) 中温（600°C 85CTC ) 分段氧化焙烧与弱还原磁性化焙烧， 用于氯化铁、 或氯化铁与 氯化镁的混合物的焙烧， 得到以四氧化三铁、 或四氧化三铁与氧化镁混合物；
(d) 高温 （1000°C-150(TC ) 分段氧化焙烧与强还原焙烧， 用于氯化铁、 或氯化铁与氯化 镁的混合物的焙烧， 得到金属铁、 或金属铁与氧化镁混合物；
上述的焙烧过程中， 焙烧物料中雾化干燥产物的质量百分数为 20%— 100%; 燃料为煤炭 或焦炭、 天然气、 水煤气、 液化石油气、 石油类产品， 所述石油类产品为柴油或重油、 煤油； 焙烧焙烧炉内按工艺要求分区保持氧化性气氛、 弱还原性气氛或强还原性气氛；
高温焙烧时产生的炙热的氯化氢气体直接通入红土镍矿浸出液中对浸出液进行加热浓 缩， 从浸出液中溢出的氯化氢气体经冷却吸收后再生盐酸或将再生的热盐酸返回前述红土镍 矿的 "氯化浸出" 步骤。
2、根据权利要求 1所述盐酸全闭路循环法从红土镍矿中提取有价金属的方法，其特征是， 所述复合粘结剂为腐植酸钠、 生石灰、 碳酸钠的混合物。
3、根据权利要求 1所述盐酸全闭路循环法从 土镍矿中提取有价金属的方法，其特征是， 所述浸出液进行浓缩、 氧化、 水解除铁的方法为： 利用后续雾化干燥及焙烧过程中的热量对 反应罐中的浸出液进行浓缩处理， 使浸出液中的水分及氯化氢气体大量挥发， 氯化铁与氯化 镁结晶析出， 得到除去部分铁镁的浓缩浸出液； 所产生的蒸汽冷凝后返回前述 "氯化浸出" 工段； 然后在 60°C~100°C条件下向所述反应罐中通入空气或富氧空气或氧气， 使所述浓缩浸 出液中的二价铁氧化成三价铁； 再以后续雾化千燥或焙烧过程中产生的氧化镁、 氧化铁或它 们的混合物为中和剂， 将所述含有三价铁浓缩浸出液的 pH值调整到 2. 2~4. 0， 使三价铁水解实 现沉铁， 过滤沉淀， 经热处理后得到铁红； ·
4、根据权利要求 1所述盐酸全闭路循环法从红土镍矿中提取有价金属的方法，其特征是， 所述采用萃取方法对所述浸出液进行除铁处理为， 萃取铁用萃取剂为中性有机膦或胺类化 合物， 萃取铁萃取剂优选为 20%N235+10%丙三醇 +70%煤油 （wt); 萃取操作用箱式萃取设备 或立式萃取设备； 采用逆流萃取和逆流反萃取、 常温下操作； 有效萃取级数 3~5级， 萃取相比 0/Α=1 : 3~5 ; 有效反萃取级数 3~5级， 反萃取相比 0/Α=3~4: 1。
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CN111187922A (zh) * 2020-02-18 2020-05-22 云南锡业研究院有限公司 一种常压下从高镍铜锍中选择性浸出镍的方法
CN112520766A (zh) * 2020-09-21 2021-03-19 河南省睿博环境工程技术有限公司 一种石棉尾矿低成本综合利用方法
CN112646973A (zh) * 2020-12-09 2021-04-13 金川集团股份有限公司 一种三段选择性浸出红土矿的方法
CN112646973B (zh) * 2020-12-09 2022-11-25 金川集团股份有限公司 一种三段选择性浸出红土矿的方法
CN114686698A (zh) * 2020-12-30 2022-07-01 北京博萃循环科技有限公司 一种红土镍矿的处理方法
CN114686698B (zh) * 2020-12-30 2023-10-13 北京博萃循环科技有限公司 一种红土镍矿的处理方法
CN113582213A (zh) * 2021-07-26 2021-11-02 四川顺应动力电池材料有限公司 一种粉煤灰综合利用的方法
CN114737068A (zh) * 2022-03-17 2022-07-12 贵州金瑞新材料有限责任公司 一种高品位软锰矿的高效浸出方法
CN114715853A (zh) * 2022-03-22 2022-07-08 湖南世纪垠天新材料有限责任公司 一种锌冶炼过程中除钴渣综合利用的方法
CN114715853B (zh) * 2022-03-22 2023-07-07 湖南世纪垠天新材料有限责任公司 一种锌冶炼过程中除钴渣综合利用的方法
CN115058597A (zh) * 2022-06-30 2022-09-16 盛隆资源再生（无锡）有限公司 一种含钙、铁、钴、镍的电镀污泥的回收处理方法
CN115286021A (zh) * 2022-08-11 2022-11-04 衢州华友钴新材料有限公司 一种镍钴中间品浸出溶液中回收氧化镁的方法
CN115286021B (zh) * 2022-08-11 2024-05-03 衢州华友钴新材料有限公司 一种镍钴中间品浸出溶液中回收氧化镁的方法
CN116287683A (zh) * 2022-12-31 2023-06-23 广西中伟新能源科技有限公司 一种硫化矿物的浸出方法

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