WO2022126761A2 - 一种从红土镍矿中综合提取有价金属的方法 - Google Patents

一种从红土镍矿中综合提取有价金属的方法 Download PDF

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WO2022126761A2
WO2022126761A2 PCT/CN2020/141255 CN2020141255W WO2022126761A2 WO 2022126761 A2 WO2022126761 A2 WO 2022126761A2 CN 2020141255 W CN2020141255 W CN 2020141255W WO 2022126761 A2 WO2022126761 A2 WO 2022126761A2
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leaching
extraction
time
ore
pickling
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PCT/CN2020/141255
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English (en)
French (fr)
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许开华
李琴香
王文杰
张坤
朱少文
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荆门市格林美新材料有限公司
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Publication of WO2022126761A2 publication Critical patent/WO2022126761A2/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
    • C22B23/00—Obtaining nickel or cobalt
    • C22B23/04—Obtaining nickel or cobalt by wet processes
    • C22B23/0407—Leaching processes
    • C22B23/0415—Leaching processes with acids or salt solutions except ammonium salts solutions
    • C22B23/043—Sulfurated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00—Obtaining nickel or cobalt
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
    • C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
    • C22B3/08—Sulfuric acid, other sulfurated acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
    • C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
    • C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical 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
    • C22B47/00—Obtaining manganese
    • 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
    • C22B59/00—Obtaining rare earth metals

Definitions

  • the invention relates to the field of hydrometallurgy, in particular to a method for comprehensively extracting valuable metals from laterite nickel ore.
  • Nickel resources mainly include nickel sulfide ore and nickel oxide ore, of which nickel sulfide ore accounts for 40% and nickel oxide ore accounts for 60%.
  • nickel sulfide ore accounts for 40% and nickel oxide ore accounts for 60%.
  • the distribution of laterite-nickel ore layers from top to bottom is limonite layer, transition layer and residual layer, mainly including nickel, cobalt, manganese, scandium and other valuable metals, of which the limonite layer mainly contains 0.9-1.5% nickel, The iron content is between 35% and 45%, and the magnesium content is less than 5%.
  • transition layer laterite nickel ore 5-10% transition layer laterite nickel ore; residual layer is mainly magnesia laterite nickel ore with nickel content of 1.7% to 2.5%, iron content of 5 to 20% and magnesium content of 8 to 30%.
  • the magnesia laterite nickel ore is often used for the production of stainless steel by pyrometallurgy, the laterite nickel ore in the transition layer is often used as the raw material for the production of stainless steel, and the limonite layer is mainly prepared by the wet process or the combined fire-wet process. Nickel intermediates.
  • the hydrometallurgical processes of laterite nickel ore mainly include: pressurized ammonia leaching, pressurized acid leaching, atmospheric leaching and heap leaching, etc.
  • the representative process of pressurized ammonia leaching is the Caron process, that is, the laterite nickel ore is reduced after drying, grinding and other pretreatments, and then ammonia leaching is carried out after reduction.
  • There are mainly disadvantages such as high energy consumption, low nickel and cobalt recovery rate, and can only process laterite nickel ore with iron content greater than 35%.
  • the main flow of the pressurized acid leaching process is ore sample preparation, pressurized acid leaching, CCD washing (continuous countercurrent washing), neutralization and impurity removal, and nickel and cobalt precipitation.
  • Atmospheric leaching is mainly carried out under normal pressure conditions by introducing seawater ball milling combined with reduction leaching for acid leaching.
  • the leaching process has high acid consumption and introduces a large amount of impurities such as iron, aluminum, and chromium.
  • the heap leaching process also has problems such as high acid consumption, long cycle, low metal solution content, and high iron and aluminum impurities.
  • the present invention provides a method for comprehensively extracting valuable metals from laterite nickel ore with low investment cost, low acid consumption and high leaching rate of valuable metals.
  • the present invention is achieved through the following technical solutions.
  • a method for comprehensively extracting valuable metals from laterite nickel ore characterized in that the method comprises:
  • step (2) adding the slurry obtained in step (1) into ore blending and performing pressure leaching to obtain the first leaching solution and the first leaching slag, carrying out the neutralization reaction of the first leaching slag, and performing solid-liquid separation to obtain the second leaching slag and The second leaching solution; wherein, the mass ratio of the added mass of the blended ore to the laterite nickel ore (dry ore) described in step (1) is 0.4-1.5:1, the leaching temperature is 185-210°C, and the leaching time is 1.5- 2h, the pressure is 1.6 ⁇ 2.5Mpa;
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back-extraction and precipitation to obtain scandium hydroxide; the second leaching solution obtained in step (2) is subjected to alkalization to remove impurities, After complexation and precipitation, nickel cobalt manganese hydroxide is obtained.
  • the time is 0.5h ⁇ 2h, the particle size is 200 meshes ⁇ 300 meshes; dense: the underflow concentration (by mass percentage) is 20% ⁇ 45%, the content of suspended solids in the overflow It is 0 ⁇ 1000ppm; the mass ratio of sulfuric acid (concentrated sulfuric acid) added to laterite nickel ore is 1:2 ⁇ 1.5:1; the pre-dipping temperature is 70 ⁇ 100°C, and the time is 2 ⁇ 6h.
  • step (2) blending ore is limonite layer, transition layer laterite nickel ore or magnesia laterite nickel ore.
  • the pressure is 1.6-2.5Mpa, and the leaching time is 1.5-2h;
  • the leaching time is 1.5-2h, the total pressure is 1.6MPa-2.5MPa, and the oxygen partial pressure is 0-35% of the total pressure.
  • the limonite layer includes (in mass percent): nickel content 0.9-1.5%, iron content 35-45%, magnesium content less than 5%; transition layer laterite nickel ore includes (in mass percent): Nickel content is 1.3-1.7%, iron content is 20-30%, magnesium content is 5-10%; magnesia laterite nickel ore includes (by mass percentage): nickel content 1.7-2.5%, iron content 5-20%, magnesium content 8 to 30%.
  • the first leaching residue adopts calcium carbonate slurry with a concentration of 15-30% for neutralization reaction
  • the process conditions of the neutralization reaction are as follows: the pH of the neutralization reaction is 1-2, and The neutralization reaction temperature is 60-100°C, and the neutralization reaction time is 0.5-2h.
  • the second leaching residue obtained in the step (2) and the sulfuric acid are subjected to pickling at a solid-liquid ratio (mass ratio) of 1:1 to 1:5 to obtain a pickling solution, the acid
  • the technological conditions of the washing are as follows: the pH of the pickling is 0-1, the pickling time is 0.5-4h, the acid-washing temperature is 60-100°C, and the pickling times are 1-5 times.
  • step (3) adding an extractant and a diluent into the pickling solution for extraction to obtain a loaded organic and a raffinate, and the raffinate is incorporated into the second leachate;
  • the extractant is One or more mixtures of P204 (bis(2-ethylhexyl) phosphate), P507 (2-ethylhexyl phosphate mono-2-ethylhexyl phosphate) and TBP (tributyl phosphate), the said The diluent is sulfonated kerosene, the extractant is 1% to 30% of the diluent (by volume percentage), and the ratio of water phase (pickling solution): oil phase (extractant) is 10:1 to 30: 1.
  • the extraction time is 3 to 30 minutes, and the standing time after extraction is 5 to 30 minutes.
  • washing adding a detergent to the loaded organic obtained by extraction and phase separation, the detergent used is sulfuric acid or hydrochloric acid, and the concentration of the detergent is calculated as [H + ] concentration of 1 ⁇ 8mol/L
  • the water phase (hydrochloric acid or sulfuric acid) of the washing process the ratio of the oil phase (extractant) is 1:1 ⁇ 1:30, the washing temperature is 10 ⁇ 90°C, the washing time is 5 ⁇ 30min, and the number of washings is 1 ⁇ 10 time
  • back extraction add back extraction agent to the loaded organic obtained after washing and phase separation, the back extraction agent used is liquid caustic soda, and the concentration of the liquid caustic soda is 1 ⁇ 10mol/L in terms of the concentration of [OH ⁇ ]
  • Back-extraction water phase (back-extraction agent) the ratio of oil phase (loaded organic after washing and phase separation) is 1:1-1:30, the back-extraction time is 3-30min, and the standing time after back-extraction is 5-30min
  • Precipitation add a precipitant to the
  • the process conditions for alkalization and impurity removal in the step (3) are as follows: the alkalization impurity removal agent used is calcium carbonate or sodium hydroxide, the impurity removal temperature is 40 ⁇ 80° C., and the impurity removal pH is 2.8 ⁇ 5.5, The impurity removal time is 2-4h; the process conditions of the complex precipitation reaction are: adding a complexing agent and a precipitating agent to the impurity-removing liquid obtained by alkalization and impurity removal, the complexing agent used is ammonia water, and the amount of ammonia water added It is 10% (by volume percentage) of the second leaching solution, the precipitating agent used is magnesium oxide, the reaction temperature is controlled at 40-80° C., the pH is 7.0-9.0, and the reaction time is 2h-8h.
  • the precipitation liquid produced by the complex precipitation in the step (3) enters the ammonia recovery system for ammonia water recovery, and the evaporation temperature is controlled to be 80-150° C., and the recovered ammonia water is recycled for use.
  • the beneficial technical effect of the present invention is that the present invention provides a method for comprehensively extracting valuable metals from laterite nickel ore, which has low investment cost, good working conditions, low acid consumption, low maintenance cost, high nickel-cobalt-manganese leaching rate, and applicable range of raw materials. wide.
  • the recovery rate of scandium in valuable metals is >90%; the leaching rate of nickel is ⁇ 95%, the recovery rate of the whole process is ⁇ 92%; the leaching rate of cobalt is ⁇ 95%, the recovery rate is ⁇ 92%; the leaching rate of manganese is ⁇ 95%, Recovery rate ⁇ 92%.
  • Fig. 1 is a process flow diagram of the present invention.
  • a method for comprehensively extracting valuable metals from laterite nickel ore includes:
  • Concentration (by mass percentage) obtained after ball milling and thickening the laterite nickel ore (the laterite nickel ore is one or more of the limonite layer, the transition layer laterite nickel ore, and the magnesia laterite nickel ore) 20% to 45% of the underflow (latterite nickel ore pulp), preferably 20% to 35%, preferably 35% to 40%, is added with sulfuric acid for pre-impregnation to obtain the slurry; wherein, the ball milling time is 0.5h to 2h, preferably 0.5 to 40%.
  • the particle size requirement is 200 meshes ⁇ 300 meshes; the content of suspended solids in the overflow obtained densely is 0 ⁇ 1000ppm; the added mass of sulfuric acid (concentrated sulfuric acid) and the mass ratio of laterite nickel ore are 1: 2 ⁇ 1.5:1, preferably 1:3 ⁇ 1.5:1, most preferably 1.2:1; the prepreg temperature is 70°C ⁇ 100°C, preferably 85°C ⁇ 100°C, and the time is 2 ⁇ 6h.
  • the ore blending is a limonite layer, a transition layer laterite nickel ore or a magnesia laterite nickel ore;
  • the limonite layer includes (pressing By mass percentage): nickel content 0.9-1.5%, iron content 35-45%, magnesium content less than 5%;
  • transition layer laterite nickel ore includes (by mass percentage): nickel content 1.3-1.7%, iron content 20-30% %, the magnesium content is 5-10%;
  • the magnesia laterite nickel ore includes (by mass percentage): nickel content 1.7-2.5%, iron content 5-20%, magnesium content 8-30%;
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back-extraction and precipitation to obtain scandium hydroxide; the second leaching solution obtained in step (2) is subjected to alkalization to remove impurities, After complex precipitation, nickel cobalt manganese hydroxide is obtained; the precipitated liquid produced by complex precipitation enters the ammonia recovery system for ammonia water recovery, and the evaporation temperature is controlled to be 80°C to 150°C, and the recovered ammonia water is recycled.
  • pickling pickling the second leaching residue obtained in step (2) and 98% sulfuric acid according to a solid-liquid ratio (mass ratio) of 1:1 to 1:5 (preferably 1:3 to 1:5) , the pickling solution is obtained, and the process conditions of the pickling are: the pH of the pickling is 0 ⁇ 1, the pickling time is 0.5 ⁇ 4h, the pickling temperature is 60°C ⁇ 100°C, and the pickling times are 1 ⁇ 5 times.
  • Extraction be that extractant, diluent are added in pickling solution to extract, phase separation obtains load organic and raffinate, and the raffinate obtained is incorporated into the second leachate to reclaim cobalt, nickel and manganese; wherein, extractant is P2O4 ( One or more mixtures of bis(2-ethylhexyl) phosphate), P507 (2-ethylhexyl phosphate mono-2-ethylhexyl phosphate), TBP (tributyl phosphate), and the diluent is sulfonic acid kerosene, the extraction agent is 1% to 30% of the diluent (by volume percentage), preferably 5% to 20%, and the ratio of water phase (pickling solution): oil phase (extractant) is 10:1 to 30 : 1, preferably 15:1 to 25:1, the extraction time is 3 to 30 minutes, and the standing time after extraction is 5 to 30 minutes.
  • extractant is P2O4 ( One
  • Washing add detergent to the loaded organic obtained by extraction and phase separation, the detergent used is sulfuric acid or hydrochloric acid, the concentration of the detergent is 1-8 mol/L in terms of [H + ] concentration, and the aqueous phase (hydrochloric acid) of the washing process is used. or sulfuric acid): the oil phase (extractant) ratio is 1:1 to 1:30, preferably 1:15 to 1:20, the washing temperature is 10°C to 90°C, the washing time is 5 to 30 minutes, and the number of washings is 1 to 10 times. ;
  • Back extraction add back extraction agent to the loaded organic obtained after washing and phase separation, the back extraction agent used is liquid caustic soda, and the concentration of liquid caustic soda is 1 ⁇ 10mol/L in terms of the concentration of [OH ⁇ ], preferably 5 ⁇ 8mol /L
  • back extraction water phase (reverse extraction agent) oil phase (loaded organic after washing and phase separation) ratio is 1:1 ⁇ 1:30, preferably 1:15 ⁇ 1:20, back extraction time 3 ⁇ 30min , the standing time after back extraction is 5 ⁇ 30min;
  • Precipitation add a precipitant to the water phase obtained by back extraction, the used precipitant is sulfuric acid or hydrochloric acid, the pH is controlled to be 8 ⁇ 11, and the reaction time is 0.5 ⁇ 4h.
  • the alkalization impurity removal agent used is calcium carbonate or sodium hydroxide
  • the impurity removal temperature is 40 °C ⁇ 80 °C
  • the impurity removal pH is 2.8 ⁇ 5.5, preferably 3.2 ⁇ 5.0
  • the impurity removal time is 40 °C ⁇ 80 °C.
  • the process conditions of the complex precipitation reaction are: adding a complexing agent and a precipitating agent to the impurity-removed liquid obtained by alkalization and impurity removal, the complexing agent used is ammonia water, and the addition amount of ammonia water is the second 10% of the leaching solution (by volume percentage), the precipitant used is magnesium oxide, the reaction temperature is controlled to be 40°C to 80°C, the pH is 7.0 to 9.0, and the reaction time is 2h to 8h.
  • step (2) adding the slurry obtained in step (1) into ore blending and performing pressure leaching to obtain the first leaching solution and the first leaching slag, carrying out the neutralization reaction of the first leaching slag, and performing solid-liquid separation to obtain the second leaching slag and The second leaching solution;
  • the ore blending is a limonite layer (including 1.5% nickel content, 45% iron content, 4% magnesium content), and the mass ratio of the added mass of the blending ore to the laterite nickel ore of step (1) is 1.4 : 1, the leaching pressure is 2.5Mpa, the temperature is 210°C, and the leaching time is 1.5 hours; neutralization reaction: use calcium carbonate slurry with a concentration of 15% for neutralization reaction, the pH of the neutralization reaction is 1, and the neutralization reaction temperature 80°C, neutralization reaction time 2h.
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back extraction, and precipitation to obtain scandium hydroxide;
  • pickling the second leaching residue and sulfuric acid are pickled at a mass ratio of 1:1 to obtain a pickling solution, the pH of the pickling is 1, the pickling time is 0.5h, the pickling temperature is 80°C, and the pickling times are 2;
  • the extractant and the diluent are added to the pickling solution for extraction, and the obtained raffinate is incorporated into the second leaching solution; wherein, the extractant is P204, the diluent is sulfonated kerosene, and the extractant is 1% of the diluent , the water phase: the oil phase ratio is 10:1, the extraction time is 3min, and the standing time after extraction is 30min.
  • Washing add detergent to the loaded organic obtained by extraction and phase separation, the detergent used is sulfuric acid, the concentration of the detergent is 8mol/L in terms of [H + ] concentration, and the water phase of the washing process: the oil phase is compared to 1:1, washing temperature 10°C, washing time 30min, washing times 5 times;
  • the process conditions of back extraction add back extraction agent to the loaded organic obtained by washing and phase separation, the back extraction agent used is liquid caustic soda, and the concentration of liquid alkali is calculated as [OH - ] concentration as 2mol/L, and back extraction water phase:
  • the oil phase ratio is 1:1, the stripping time is 3min, and the standing time after stripping is 30min;
  • Precipitation add a precipitant to the water phase obtained by back extraction and phase separation, the precipitant used is sulfuric acid, the pH is controlled to be 11, and the reaction time is 0.5h.
  • the second leaching solution obtained in step (2) is subjected to alkalization, impurity removal and complex precipitation to obtain nickel cobalt manganese hydroxide, and the precipitation liquid produced by complex precipitation enters the ammonia recovery system for ammonia recovery, and the evaporation temperature is controlled to be 150 ° C , the recovered ammonia water is recycled; wherein, alkalization and impurity removal: the alkalization impurity removal agent adopted is calcium carbonate, the impurity removal temperature is 40 °C, the impurity removal pH is 3.2, and the impurity removal time is 2 hours; complex precipitation: After removing impurities, a complexing agent and a precipitating agent are added to the liquid, and the complexing agent used is ammonia water, and the addition amount of the ammonia water is 10% of the second leaching solution.
  • the precipitant used was magnesium oxide, the reaction temperature was controlled at 40°C, the pH was 7.0, and the reaction time was 8 hours.
  • the recovery rate of scandium in valuable metals is 91%; the leaching rate of nickel is 95%, and the recovery rate of the whole process is 92%; the leaching rate of cobalt is 95%, and the recovery rate is 92%; the leaching rate of manganese is 95%, and the recovery rate is 92%.
  • ball milling time is 2h, particle size is 300 mesh; dense process conditions: underflow concentration is 35%, and the content of suspended solids in the overflow is 10ppm; the mass ratio of sulfuric acid added to laterite nickel ore is 1.5:1; The pre-dipping temperature is 100°C and the time is 6h.
  • step (2) adding the slurry obtained in step (1) into ore blending and performing pressure leaching to obtain the first leaching solution and the first leaching slag, carrying out the neutralization reaction of the first leaching slag, and performing solid-liquid separation to obtain the second leaching slag and the second leachate;
  • the ore blending is transition layer laterite nickel ore (including nickel content of 1.5%, iron content of 25%, magnesium content of 8%), and the mass ratio of the added mass of the blended ore to the laterite nickel ore of step (1) is 1.2:1;
  • Oxygen pressure leaching was adopted, the temperature was 200°C, the leaching time was 2h, the total pressure was 2.0Mpa, and the oxygen partial pressure was 35% of the total pressure;
  • the process conditions of the neutralization reaction are as follows: the neutralization reaction is carried out with calcium carbonate slurry with a concentration of 30%, the pH of the neutralization reaction is 2, the neutralization reaction temperature is 100° C., and the neutralization reaction time is 1h.
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back extraction, and precipitation to obtain scandium hydroxide;
  • the process conditions of pickling are as follows: the second leaching residue and sulfuric acid are pickled at a mass ratio of 1:5 to obtain a pickling solution, the pH of the pickling is 0.5, the pickling time is 2h, the pickling temperature is 60°C, and the number of pickling times 4 times;
  • the technological conditions of extraction the extractant and the diluent are added to the pickling solution for extraction, and the obtained raffinate is incorporated into the second leachate; wherein, the extractant is P507, the diluent is sulfonated kerosene, and the extractant is a diluent
  • the ratio of water phase: oil phase is 30:1, the extraction time is 15min, and the standing time after extraction is 5min.
  • the detergent used is hydrochloric acid
  • the concentration of the detergent is 1mol/L in terms of [H + ] concentration
  • the ratio of water phase: oil phase in the washing process is 1:30
  • the washing temperature is 90°C
  • the washing time 20min washing times 1 time;
  • the back-extraction agent used is liquid caustic soda, the concentration is 10mol/L in terms of the concentration of [OH - ], the back-extraction water phase: the oil phase is 1:30, the back-extraction time is 30min, and the back-extraction phase is 1:30.
  • the post-resting time is 5min;
  • the precipitant used is hydrochloric acid
  • the pH is controlled to be 8
  • the reaction time is 4h.
  • the second leaching solution obtained in step (2) is subjected to alkalization, impurity removal and complex precipitation to obtain nickel cobalt manganese hydroxide, and the precipitation liquid produced by complex precipitation enters the ammonia recovery system for ammonia recovery, and the evaporation temperature is controlled to be 100 ° C , the recovered ammonia water is recycled;
  • the process conditions of alkalization and impurity removal are as follows: the alkalization impurity removal agent used is sodium hydroxide, the impurity removal temperature is 80° C., the impurity removal pH is 5.5, and the impurity removal time is 4 hours; the process conditions of complex precipitation are: : The complexing agent used is ammonia water, the addition amount of ammonia water is 10% of the second leaching solution, the precipitating agent used is magnesium oxide, the reaction temperature is controlled at 80°C, the pH is 9.0, and the reaction time is 2 hours.
  • the recovery rate of scandium in valuable metals is 92%; the leaching rate of nickel is 96%, and the recovery rate of the whole process is 93%; the leaching rate of cobalt is 97%, and the recovery rate is 94%; the leaching rate of manganese is 96%, and the recovery rate is 93%.
  • the process conditions of ball milling the time is 1h, the particle size is 250 mesh; the dense process conditions: the concentration of the underflow is 25%, the content of suspended solids in the overflow is 100ppm; the mass ratio of the added sulfuric acid to the laterite nickel ore is 1.3 : 1; the pre-soak temperature is 85°C, and the time is 3h.
  • step (2) adding the slurry obtained in step (1) into ore blending and performing pressure leaching to obtain the first leaching solution and the first leaching slag, carrying out the neutralization reaction of the first leaching slag, and performing solid-liquid separation to obtain the second leaching slag and the second leachate;
  • the ore blending is magnesia laterite nickel ore (including 1.7% nickel content, 20% iron content, 15% magnesium content), and the mass ratio of the added mass of the ore blending to the laterite nickel ore of step (1) is 0.8:1;
  • oxygen pressure leaching the temperature is 190 °C, the leaching time is 1.8h, the total pressure is 1.8Mpa, and the oxygen partial pressure is 30% of the total pressure;
  • the process conditions of the neutralization reaction are as follows: the neutralization reaction is carried out by using calcium carbonate slurry with a concentration of 20%, the pH of the neutralization reaction is 1.5, the neutralization reaction temperature is 80°C, and the neutralization reaction time is 0.5h.
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back extraction, and precipitation to obtain scandium hydroxide;
  • the process conditions of pickling are as follows: the second leaching residue and the concentration of 98% sulfuric acid are pickled at a mass ratio of 1:3 to obtain a pickling solution, the pH of the pickling is 0.2, the pickling time is 2.5h, and the pickling temperature 85°C, pickling times 3 times;
  • the technological conditions of extraction adding extractant and diluent to the pickling solution for extraction, and the obtained raffinate is incorporated into the second leachate; wherein, the extractant is TBP, the diluent is sulfonated kerosene, and the extractant is a diluent
  • the ratio of water phase: oil phase is 15:1
  • the extraction time is 30min
  • the standing time after extraction is 15min.
  • the detergent used is sulfuric acid
  • the concentration is 4mol/L in terms of [H + ] concentration
  • the water phase oil phase ratio of the washing process is 1:15
  • the washing temperature is 50 ° C
  • the washing time is 5 min
  • the back-extraction agent adopted is liquid caustic soda
  • the concentration of liquid caustic soda [OH - ] is calculated as 5mol/L
  • the ratio of back-extraction water phase: oil phase is 1:15
  • the back-extraction time is 15min
  • the back-extraction time is 15min.
  • the post-resting time is 15min;
  • Precipitation process conditions the used precipitant is sulfuric acid, the pH is controlled to be 10, and the reaction time is 2h.
  • the second leaching solution obtained in step (2) is subjected to alkalization, impurity removal, and complex precipitation to obtain nickel cobalt manganese hydroxide.
  • the precipitation liquid produced by complex precipitation enters the ammonia recovery system for ammonia recovery, and the evaporation temperature is controlled to be 130 ° C. , the recovered ammonia water is recycled;
  • the process conditions of alkalization and impurity removal are: the alkalization impurity removal agent used is calcium carbonate, the impurity removal temperature is 60 ° C, the impurity removal pH is 4, and the impurity removal time is 3 hours; the process conditions of the complex precipitation reaction are: : The used complexing agent is ammonia water, the added amount is 10% of the second leaching solution, the used precipitating agent is magnesium oxide, the reaction temperature is controlled at 60°C, the pH is 8.0, and the reaction time is 5 hours.
  • the recovery rate of scandium in valuable metals is 93%; the leaching rate of nickel is 96%, and the recovery rate of the whole process is 93%; the leaching rate of cobalt is 97%, and the recovery rate is 94%; the leaching rate of manganese is 95%, and the recovery rate is 92%.
  • the process conditions of ball milling the time is 1.5h, the particle size is 270 mesh; the dense process conditions: the underflow concentration (by mass percentage) is 40%, and the content of suspended solids in the overflow is 50ppm; the added mass of sulfuric acid is the same as that of laterite The mass ratio of nickel ore is 1.2:1; the pre-dipping temperature is 90°C and the time is 4h.
  • step (2) adding the slurry obtained in step (1) into ore blending and performing pressure leaching to obtain the first leaching solution and the first leaching slag, carrying out the neutralization reaction of the first leaching slag, and performing solid-liquid separation to obtain the second leaching slag and the second leachate;
  • the ore blending is a limonite layer (including 0.9% nickel content, 35% iron content, and 4% magnesium content), and the mass ratio of the added quality of the blending ore to the laterite nickel ore of step (1) is 0.7:1; leaching The pressure is 2.2Mpa, the temperature is 200°C, and the leaching time is 2 hours;
  • the process conditions of the neutralization reaction are as follows: the neutralization reaction is carried out by using calcium carbonate slurry with a concentration of 25%, the pH of the neutralization reaction is 1.8, the neutralization reaction temperature is 90°C, and the neutralization reaction time is 1.5h.
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back extraction, and precipitation to obtain scandium hydroxide;
  • the process conditions of pickling are as follows: the second leaching residue is pickled with 98% sulfuric acid in a mass ratio of 1:4 to obtain a pickling solution, the pH of the pickling is 0.8, the pickling time is 3h, and the pickling temperature is 90°C , pickling times 5 times;
  • extracting agent and diluent are added to the pickling solution for extraction, and the obtained raffinate is incorporated into the second leachate; wherein, the extracting agent is P204 and P507, the diluent is sulfonated kerosene, and the extracting agent is 20% of the diluent, the ratio of water phase:oil phase is 20:1, the extraction time is 20min, and the standing time after extraction is 20min.
  • the detergent used is hydrochloric acid
  • the concentration is 6mol/L in terms of [H + ] concentration
  • the water phase: oil phase ratio of the washing process is 1:20
  • the washing temperature is 80 ° C
  • the washing time is 10 min
  • the washing 7 times the washing 7 times;
  • the back extraction agent used is liquid caustic soda, the concentration is calculated as 4mol/L in terms of the concentration of [OH - ], the water phase of back extraction: the oil phase ratio is 1:20, the back extraction time is 20min, the back extraction The post-resting time is 20min;
  • Precipitation process conditions the precipitating agent used is hydrochloric acid, the pH is controlled to be 9, and the reaction time is 3h.
  • the second leaching solution obtained in step (2) is subjected to alkalization, impurity removal and complex precipitation to obtain nickel cobalt manganese hydroxide, and the precipitation liquid produced by complex precipitation enters the ammonia recovery system for ammonia recovery, and the evaporation temperature is controlled to be 80 ° C , the recovered ammonia water is recycled;
  • the process conditions of alkalization and impurity removal are as follows: the alkalization impurity removal agent used is sodium hydroxide, the impurity removal temperature is 70 ° C, the impurity removal pH is 2.8, and the impurity removal time is 3 hours; the process conditions of complex precipitation are: : The used complexing agent is ammonia water, the added amount is 10% of the second leaching solution, the used precipitating agent is magnesium oxide, the reaction temperature is controlled at 70°C, the pH is 8.5, and the reaction time is 7 hours.
  • the recovery rate of scandium in valuable metals is 91%; the leaching rate of nickel is 96%, and the recovery rate of the whole process is 93%; the leaching rate of cobalt is 95%, and the recovery rate is 92%; the leaching rate of manganese is 95%, and the recovery rate is 92%.
  • the process conditions of ball milling the time is 1.7h, the particle size is 230 mesh; the dense process conditions: the underflow concentration is 45%, and the content of suspended solids in the overflow is 500ppm; the mass ratio of sulfuric acid added to laterite nickel ore is 1.4:1; prepreg temperature is 95°C, time is 5h.
  • step (2) adding the slurry obtained in step (1) into ore blending and performing pressure leaching to obtain the first leaching solution and the first leaching slag, carrying out the neutralization reaction of the first leaching slag, and performing solid-liquid separation to obtain the second leaching slag and the second leachate;
  • the ore blending is magnesia laterite nickel ore (including 2.5% nickel content, 7% iron content, 25% magnesium content), and the mass ratio of the added mass of the ore blending to the laterite nickel ore of step (1) is 0.5:1;
  • Oxygen pressure leaching was adopted, the temperature was 200°C, the leaching time was 1.7h, the total pressure was 1.8Mpa, and the oxygen partial pressure was 15% of the total pressure;
  • the technological conditions of the neutralization reaction are as follows: calcium carbonate slurry with a concentration of 18% is used for the neutralization reaction, the pH of the neutralization reaction is 2, the neutralization reaction temperature is 65°C, and the neutralization reaction time is 1.8h.
  • step (3) the second leaching residue obtained in step (2) is subjected to pickling, extraction, washing, back extraction, and precipitation to obtain scandium hydroxide;
  • the process conditions of pickling are as follows: the second leaching residue and the concentration of 98% sulfuric acid are pickled at a mass ratio of 1:3 to obtain a pickling solution, the pH of the pickling is 0.2, the pickling time is 4h, and the pickling temperature is 70 °C °C, pickling times 1;
  • the technological conditions of extraction adding extractant and diluent to the pickling solution for extraction, and the obtained raffinate is incorporated into the second leachate; wherein, the extractant is TBP, the diluent is sulfonated kerosene, and the extractant is a diluent 5%, the water phase:oil phase ratio was 25:1, the extraction time was 25min, and the standing time after extraction was 25min.
  • the detergent used is sulfuric acid
  • the concentration is 3mol/L in terms of [H + ] concentration
  • the water phase oil phase ratio of the washing process is 1:25
  • the washing temperature is 40 ° C
  • the washing time is 25 min, and the washing 3 times;
  • the back-extraction agent used is liquid caustic soda ammonia, and the concentration is calculated as 7 mol/L in terms of the concentration of [OH - ].
  • the standing time after extraction is 25min;
  • the precipitating agent used is hydrochloric acid
  • the pH is controlled to be 10
  • the reaction time is 3h.
  • the second leaching solution obtained in step (2) is subjected to alkalization, impurity removal and complex precipitation to obtain nickel cobalt manganese hydroxide.
  • the precipitation liquid produced by complex precipitation enters the ammonia recovery system for ammonia recovery, and the evaporation temperature is controlled to be 90 ° C. , the recovered ammonia water is recycled;
  • the process conditions of alkalization and impurity removal are: the alkalization impurity removal agent used is calcium carbonate, the impurity removal temperature is 50 ° C, the impurity removal pH is 3.7, and the impurity removal time is 3 hours; the process conditions of the complex precipitation reaction are: : The used complexing agent is ammonia water, the added amount is 10% of the second leaching solution, the used precipitating agent is magnesium oxide, the reaction temperature is controlled at 50°C, the pH is 8.5, and the reaction time is 6 hours.
  • the recovery rate of scandium in valuable metals is 94%; the leaching rate of nickel is 98%, and the recovery rate of the whole process is 95%; the leaching rate of cobalt is 98%, and the recovery rate is 95%; the leaching rate of manganese is 97%, and the recovery rate is 94%.
  • step (2) adding the slurry obtained in step (1) to the serpentine-type ore and performing pressure leaching to obtain leaching solution and leaching slag; the temperature is 170°C, and the leaching time is 4h;
  • step (3) The leaching solution obtained in step (2) is subjected to alkalization, impurity removal and complex precipitation to obtain nickel cobalt manganese hydroxide.
  • the leaching rate of nickel is 85%, and the recovery rate of the whole process is 80%; the leaching rate of cobalt is 87%, and the recovery rate is 85%; the leaching rate of manganese is 80%, and the recovery rate is 86%.

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Description

一种从红土镍矿中综合提取有价金属的方法 技术领域
本发明涉及属于湿法冶金领域，具体涉及一种从红土镍矿中综合提取有价金属的方法。
背景技术
镍资源主要有硫化镍矿与氧化镍矿，其中硫化镍矿占比40％，氧化镍矿占比60％。随着社会的发展，高品位易处理的硫化镍矿日益消耗与枯竭，人们对镍资源的需求逐渐转向储量较多的氧化型红土镍矿。红土镍矿矿层分布自上而下分别为褐铁矿层、过渡层和残积层，主要包含镍、钴、锰、钪等有价金属，其中褐铁矿层主要为镍含量0.9～1.5％，铁含量在35～45％之间，镁含量小于5％的高铁型红土镍矿；过渡层主要为含镍量在1.3～1.7％之间，铁含量在20～30％之间，镁含量在5～10％的过渡层红土镍矿；残积层主要为镍含量1.7％～2.5％，铁含量在5～20％，镁含量在8～30％的镁质红土镍矿。镁质红土镍矿常采用火法冶炼用于不锈钢的生产，过渡层红土镍矿常用于不锈钢生产的配矿原料，而褐铁矿层主要采用湿法工艺或者是火法-湿法联合工艺制备镍中间品。
红土镍矿的湿法冶炼工艺主要有：加压氨浸、加压酸浸、常压浸出和堆浸等。其中，加压氨浸的代表工艺为Caron工艺，即红土镍矿经干燥、研磨等预处理后进行还原，还原后进行氨性浸出。主要存在高能耗、低镍钴回收率、只能处理铁含量大于35％的红土镍矿等缺点。加压酸浸工艺主要流程为矿样准备、加压酸浸、CCD洗涤(连续逆流洗涤)、中和与除杂、沉镍钴等步骤。投资成 本高、操作工况苛刻、设备易腐蚀、高压釜的维修成本高、调试时间长、不适用于镁含量高的矿物。常压浸出主要是在常压条件下，引入海水球磨与还原浸出相结合进行酸性浸出，浸出过程酸耗高且引入大量的铁铝铬等杂质。而堆浸工艺也存在酸耗高、周期长、金属溶液含量低、高铁铝杂质等问题。
发明内容
针对上述已有技术存在的不足，本发明提供一种投资成本低、酸耗低、有价金属浸出率高的从红土镍矿中综合提取有价金属的方法。
本发明是通过以下技术方案实现的。
一种从红土镍矿中综合提取有价金属的方法，其特征在于，所述方法包括：
(1)将红土镍矿经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；其中，所述配矿的加入质量与步骤(1)所述的红土镍矿(干矿)的质量比为0.4～1.5:1，浸出温度为185～210℃，浸出时间1.5～2h，压力为1.6～2.5Mpa；
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰。
进一步地，所述步骤(1)球磨：时间为0.5h～2h，粒径为200目～300目；浓密：底流浓度(按质量百分比计)20％～45％，溢流中的悬浮物含量为0～1000ppm；硫酸(浓硫酸)的加入质量与红土镍矿的质量比为1:2～1.5:1；预浸温度为70～100℃，时间2～6h。
进一步地，所述步骤(2)配矿为褐铁矿层、过渡层红土镍矿或者镁质红土镍矿。
进一步地，当进行预浸的红土镍矿和配矿均为褐铁矿层时，压力为1.6～2.5Mpa，浸出时间为1.5～2h；当进行预浸的红土镍矿和配矿中的任意一种为过渡层红土镍矿或者镁质红土镍矿时，浸出时间为1.5～2h，总压为1.6MPa～2.5MPa，氧分压为总压的0～35％。
进一步地，所述褐铁矿层包括(按质量百分比计)：镍含量0.9～1.5％，铁含量35～45％，镁含量小于5％；过渡层红土镍矿包括(按质量百分比计)：镍含量1.3～1.7％，铁含量20～30％，镁含量5～10％；镁质红土镍矿包括(按质量百分比计)：镍含量1.7～2.5％，铁含量5～20％，镁含量8～30％。
进一步地，所述步骤(2)第一浸出渣采用浓度为15～30％碳酸钙浆料进行中和反应，所述中和反应的工艺条件为：中和反应的pH为1～2，中和反应温度60～100℃，中和反应时间0.5～2h。
进一步地，所述步骤(3)将经步骤(2)得到的第二浸出渣与硫酸按固液比(质量比)1:1～1:5进行酸洗，得到酸洗液，所述酸洗的工艺条件为：酸洗pH为0～1，酸洗时间0.5～4h，酸洗温度60～100℃，酸洗次数1～5次。
进一步地，所述步骤(3)萃取：将萃取剂、稀释剂加入酸洗液中进行萃取，得到负载有机和萃余液，萃余液并入第二浸出液中；其中，所述萃取剂为P204(二(2-乙基己基)磷酸酯)、P507(2-乙基己基磷酸单-2-乙基己酯)、TBP(磷酸三丁酯)中的一种或者几种混合，所述稀释剂为磺化煤油，所述萃取剂为稀释剂的(按体积百分比计)1％～30％，水相(酸洗液)：油相(萃取剂)相比为10:1～30:1，萃取时间为3～30min，萃取后的静置时间为5～30min。
进一步地，所述步骤(3)洗涤：向经萃取分相得到的负载有机中加入洗涤剂，采用的洗涤剂为硫酸或者盐酸，所述洗涤剂的浓度以[H +]浓度计为1～8mol/L，洗涤过程的水相(盐酸或者硫酸)：油相(萃取剂)相比为1:1～1:30，洗涤温度10～90℃，洗涤时间5～30min，洗涤次数1～10次；反萃：向经洗涤分相后得到的负载有机中加入反萃剂，采用的反萃剂为液碱，所述液碱浓度以[OH -]的浓度计为1～10mol/L，反萃水相(反萃剂)：油相(洗涤分相后的负载有机)相比为1:1～1:30，反萃时间3～30min，反萃后静置时间为5～30min；沉淀：向经反萃分相得到的水相中加入沉淀剂，采用的沉淀剂为硫酸或者盐酸，控制pH为8～11，反应时间0.5～4h。
进一步地，所述步骤(3)碱化除杂的工艺条件为：采用的碱化除杂剂为碳酸钙或者氢氧化钠，除杂温度为40～80℃，除杂pH为2.8～5.5，除杂时间为2～4h；络合沉淀反应的工艺条件为：向经碱化除杂得到的除杂后液中加入络合剂和沉淀剂，采用的络合剂为氨水，氨水的加入量为第二浸出液的10％(按体积百分比)，采用的沉淀剂为氧化镁，控制反应温度40～80℃，pH为7.0～9.0，反应时间为2h～8h。
进一步地，所述步骤(3)络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为80～150℃，回收后的氨水循环使用。
本发明的有益技术效果，本发明提供一种红土镍矿综合提取有价金属的方法，投资成本低、工况操作良好、酸耗低、维修成本低、镍钴锰浸出率高、原料适用范围广。有价金属中钪的回收率>90％；镍的浸出率≥95％，全流程回收率≥92％；钴的浸出率≥95％，回收率≥92％；锰的浸出率≥95％，回收率≥92％。
附图说明
图1为本发明的工艺流程图。
具体实施方式
下面结合附图和具体实施方式对本发明进行详细说明。
如图1所示，一种从红土镍矿中综合提取有价金属的方法，包括：
(1)将红土镍矿(红土镍矿为褐铁矿层、过渡层红土镍矿、镁质红土镍矿中的一种或多种)经球磨、浓密后得到的浓度(按质量百分比计)20％～45％的底流(红土镍矿矿浆)，优选20％～35％，优选35％～40％，加入硫酸进行预浸得到料浆；其中，球磨时间为0.5h～2h，优选0.5～1h，优选1～2h，粒径要求为200目～300目；浓密得到的溢流中的悬浮物含量为0～1000ppm；硫酸(浓硫酸)的加入质量与红土镍矿的质量比为1:2～1.5:1，优选1:3～1.5:1，最优选1.2：1；预浸温度为70℃～100℃，优选85℃～100℃，时间2～6h。
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液(排放或者回用)、第一浸出渣，将第一浸出渣采用浓度为15～30％碳酸钙浆料进行中和反应后经固液分离得到第二浸出渣和第二浸出液；其中，配矿为褐铁矿层、过渡层红土镍矿或者镁质红土镍矿；褐铁矿层包括(按质量百分比计)：镍含量0.9～1.5％，铁含量35～45％，镁含量小于5％；过渡层红土镍矿包括(按质量百分比计)：镍含量1.3～1.7％，铁含量20～30％，镁含量5～10％；镁质红土镍矿包括(按质量百分比计)：镍含量1.7～2.5％，铁含量5～20％，镁含量8～30％；配矿的加入质量与步骤(1)所述的红土镍矿(干矿)的质量比为0.4～1.5:1，优选0.8～1.4:1，浸出温度为185℃～210℃，优选温度为190℃～210℃，浸出时间1.5～2h，压力为1.6～2.5Mpa，优选1.8～2.2Mpa；当进行预浸的红土镍矿和配矿均为褐铁矿层时，压力为1.6～2.5Mpa，浸出时间为1.5～2h； 当进行预浸的红土镍矿和配矿中的任意一种为过渡层红土镍矿或者镁质红土镍矿时，浸出时间为1.5～2h，总压为1.6MPa～2.5MPa，氧分压为总压的0～35％。中和反应的工艺条件为：中和反应的pH为1～2，中和反应温度60℃～100℃，中和反应时间0.5～2h。
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰；络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为80℃～150℃，回收后的氨水循环使用。
其中，酸洗：是将经步骤(2)得到的第二浸出渣与98％硫酸按固液比(质量比)1:1～1:5(优选1:3～1:5)进行酸洗，得到酸洗液，酸洗的工艺条件为：酸洗pH为0～1，酸洗时间0.5～4h，酸洗温度60℃～100℃，酸洗次数1～5次。
萃取：是将萃取剂、稀释剂加入酸洗液中进行萃取，分相得到负载有机和萃余液，得到的萃余液并入第二浸出液中回收钴镍锰；其中，萃取剂为P204(二(2-乙基己基)磷酸酯)、P507(2-乙基己基磷酸单-2-乙基己酯)、TBP(磷酸三丁酯)中的一种或者几种混合，稀释剂为磺化煤油，萃取剂为稀释剂的(按体积百分比计)1％～30％，优选5％～20％，水相(酸洗液)：油相(萃取剂)相比为10:1～30:1，优选15:1～25:1，萃取时间为3～30min，萃取后的静置时间为5～30min。
洗涤：向经萃取分相得到的负载有机中加入洗涤剂，采用的洗涤剂为硫酸或者盐酸，洗涤剂的浓度以[H +]浓度计为1～8mol/L，洗涤过程的水相(盐酸或者硫酸)：油相(萃取剂)相比为1:1～1:30，优选1:15～1:20，洗涤温度10℃～90℃，洗涤时间5～30min，洗涤次数1～10次；
反萃：向经洗涤分相后得到的负载有机中加入反萃剂，采用的反萃剂为液碱，液碱浓度以[OH -]的浓度计为1～10mol/L，优选5～8mol/L，反萃水相(反萃剂)：油相(洗涤分相后的负载有机)相比为1:1～1:30，优选1:15～1:20，反萃时间3～30min，反萃后静置时间为5～30min；沉淀：向经反萃分相得到的水相加入沉淀剂，采用的沉淀剂为硫酸或者盐酸，控制pH为8～11，反应时间0.5～4h。
碱化除杂的工艺条件为：采用的碱化除杂剂为碳酸钙或者氢氧化钠，除杂温度为40℃～80℃，除杂pH为2.8～5.5，优选3.2～5.0，除杂时间为2～4h；络合沉淀反应的工艺条件为：向经碱化除杂得到的除杂后液中加入络合剂和沉淀剂，采用的络合剂为氨水，氨水的加入量为第二浸出液的10％(按体积百分比)，采用的沉淀剂为氧化镁，控制反应温度40℃～80℃，pH为7.0～9.0，反应时间为2h～8h。
实施例1
(1)将红土镍矿(为褐铁矿层)经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；其中，球磨：时间为0.5h，粒径为200目；浓密：底流浓度45％，溢流中的悬浮物含量为900ppm；硫酸的加入质量与红土镍矿的质量比为1:2；预浸温度为90℃，时间6h。
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；其中，配矿为褐铁矿层(包括镍含量1.5％，铁含量45％，镁含量4％)，配矿的加入质量与步骤(1)的红土镍矿的质量比为1.4:1，浸出压力为2.5Mpa，温度为210℃，浸出时间为1.5小时；中和反应：采用浓度为15％碳酸钙浆料进 行中和反应，中和反应的pH为1，中和反应温度80℃，中和反应时间2h。
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；
其中，酸洗：第二浸出渣与硫酸按质量比1:1进行酸洗，得到酸洗液，酸洗pH为1，酸洗时间0.5h，酸洗温度80℃，酸洗次数2次；
萃取：将萃取剂、稀释剂加入酸洗液中进行萃取，得到的萃余液并入第二浸出液中；其中，萃取剂为P204，稀释剂为磺化煤油，萃取剂为稀释剂的1％，水相：油相相比为10:1，萃取时间为3min，萃取后的静置时间为30min。
洗涤：向经萃取分相得到的负载有机中加入洗涤剂，采用的洗涤剂为硫酸，洗涤剂的浓度以[H +]浓度计为8mol/L，洗涤过程的水相：油相相比为1:1，洗涤温度10℃，洗涤时间30min，洗涤次数5次；
反萃的工艺条件：向经洗涤分相得到的负载有机中加入反萃剂，采用的反萃剂为液碱，液碱浓度以[OH -]浓度计为2mol/L，反萃水相：油相相比为1:1，反萃时间3min，反萃后静置时间为30min；
沉淀：向经反萃分相得到的水相中加入沉淀剂，采用的沉淀剂为硫酸，控制pH为11，反应时间0.5h。
将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰，络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为150℃，回收后的氨水循环使用；其中，碱化除杂：采用的碱化除杂剂为碳酸钙，除杂温度为40℃，除杂pH为3.2，除杂时间为2小时；络合沉淀：除杂后液中加入络合剂和沉淀剂，采用的络合剂为氨水，氨水的加入量为第二浸出液的10％。采用的沉淀剂为氧化镁，控制反应温度40℃，pH为7.0，反应时间为 8小时。
有价金属中钪的回收率91％；镍的浸出率95％，全流程回收率92％；钴的浸出率95％，回收率92％；锰的浸出率95％，回收率92％。
实施例2
(1)将红土镍矿(为过渡层红土镍矿)经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
其中，球磨：时间为2h，粒径为300目；浓密的工艺条件：底流浓度35％，溢流中的悬浮物含量为10ppm；硫酸的加入质量与红土镍矿的质量比为1.5:1；预浸温度为100℃，时间6h。
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；
其中，配矿为过渡层红土镍矿(包括镍含量1.5％，铁含量25％，镁含量8％)，配矿的加入质量与步骤(1)的红土镍矿的质量比为1.2:1；采用氧压浸出，温度为200℃，浸出时间为2h，总压为2.0Mpa，氧分压为总压的35％；
中和反应的工艺条件为：采用浓度为30％碳酸钙浆料进行中和反应，中和反应的pH为2，中和反应温度100℃，中和反应时间1h。
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；
其中，酸洗的工艺条件为：第二浸出渣与硫酸按质量比1:5进行酸洗，得到酸洗液，酸洗pH为0.5，酸洗时间2h，酸洗温度60℃，酸洗次数4次；
萃取的工艺条件：将萃取剂、稀释剂加入酸洗液中进行萃取，得到的萃余 液并入第二浸出液中；其中，萃取剂为P507，稀释剂为磺化煤油，萃取剂为稀释剂的30％，水相：油相相比为30:1，萃取时间为15min，萃取后的静置时间为5min。
洗涤的工艺条件：采用的洗涤剂为盐酸，洗涤剂的浓度以[H +]浓度计为1mol/L，洗涤过程的水相：油相相比为1:30，洗涤温度90℃，洗涤时间20min，洗涤次数1次；
反萃的工艺条件：采用的反萃剂为液碱，浓度以[OH -]的浓度计为10mol/L，反萃水相：油相相比为1:30，反萃时间30min，反萃后静置时间为5min；
沉淀的工艺条件：采用的沉淀剂为盐酸，控制pH为8，反应时间4h。
将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰，络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为100℃，回收后的氨水循环使用；
其中，碱化除杂的工艺条件为：采用的碱化除杂剂为氢氧化钠，除杂温度为80℃，除杂pH为5.5，除杂时间为4小时；络合沉淀的工艺条件为：采用的络合剂为氨水，氨水的加入量为第二浸出液的10％，采用的沉淀剂为氧化镁，控制反应温度80℃，pH为9.0，反应时间为2小时。
有价金属中钪的回收率92％；镍的浸出率96％，全流程回收率93％；钴的浸出率97％，回收率94％；锰的浸出率96％，回收率93％。
实施例3
(1)将红土镍矿(为镁质红土镍矿)经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
其中，球磨的工艺条件：时间为1h，粒径为250目；浓密的工艺条件：底 流浓度25％，溢流中的悬浮物含量为100ppm；硫酸的加入质量与红土镍矿的质量比为1.3:1；预浸温度为85℃，时间3h。
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；
其中，配矿为镁质红土镍矿(包括镍含量1.7％，铁含量20％，镁含量15％)，配矿的加入质量与步骤(1)的红土镍矿的质量比为0.8:1；采用氧压浸出，温度为190℃，浸出时间为1.8h，总压为1.8Mpa，氧分压为总压的30％；
中和反应的工艺条件为：采用浓度为20％碳酸钙浆料进行中和反应，中和反应的pH为1.5，中和反应温度80℃，中和反应时间0.5h。
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；
其中，酸洗的工艺条件为：第二浸出渣与浓度为98％硫酸按质量比为1:3进行酸洗，得到酸洗液，酸洗pH为0.2，酸洗时间2.5h，酸洗温度85℃，酸洗次数3次；
萃取的工艺条件：将萃取剂、稀释剂加入酸洗液中进行萃取，得到的萃余液并入第二浸出液中；其中，萃取剂为TBP，稀释剂为磺化煤油，萃取剂为稀释剂的25％，水相：油相相比为15:1，萃取时间为30min，萃取后的静置时间为15min。
洗涤的工艺条件：采用的洗涤剂为硫酸，浓度以[H +]浓度计为4mol/L，洗涤过程的水相：油相相比为1:15，洗涤温度50℃，洗涤时间5min，洗涤次数8次；
反萃的工艺条件：采用的反萃剂为液碱，液碱[OH -]的浓度计为5mol/L，反 萃水相：油相相比为1:15，反萃时间15min，反萃后静置时间为15min；
沉淀的工艺条件：采用的沉淀剂为硫酸，控制pH为10，反应时间2h。
将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰，络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为130℃，回收后的氨水循环使用；
其中，碱化除杂的工艺条件为：采用的碱化除杂剂为碳酸钙，除杂温度为60℃，除杂pH为4，除杂时间为3小时；络合沉淀反应的工艺条件为：采用的络合剂为氨水，加入量为第二浸出液的10％，采用的沉淀剂为氧化镁，控制反应温度60℃，pH为8.0，反应时间为5小时。
有价金属中钪的回收率93％；镍的浸出率96％，全流程回收率93％；钴的浸出率97％，回收率94％；锰的浸出率95％，回收率92％。
实施例4
(1)将红土镍矿(为褐铁矿层)经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
其中，球磨的工艺条件：时间为1.5h，粒径为270目；浓密的工艺条件：底流浓度(按质量百分比计)40％，溢流中的悬浮物含量为50ppm；硫酸的加入质量与红土镍矿的质量比为1.2:1；预浸温度为90℃，时间4h。
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；
其中，配矿为褐铁矿层(包括镍含量0.9％，铁含量35％，镁含量4％)，配矿的加入质量与步骤(1)的红土镍矿的质量比为0.7:1；浸出压力为2.2Mpa， 温度为200℃，浸出时间为2小时；
中和反应的工艺条件为：采用浓度为25％碳酸钙浆料进行中和反应，中和反应的pH为1.8，中和反应温度90℃，中和反应时间1.5h。
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；
其中，酸洗的工艺条件为：第二浸出渣与浓度为98％硫酸按质量比1:4进行酸洗，得到酸洗液，酸洗pH为0.8，酸洗时间3h，酸洗温度90℃，酸洗次数5次；
萃取的工艺条件：将萃取剂、稀释剂加入酸洗液中进行萃取，得到的萃余液并入第二浸出液中；其中，萃取剂为P204和P507，稀释剂为磺化煤油，萃取剂为稀释剂的20％，水相：油相相比为20:1，萃取时间为20min，萃取后的静置时间为20min。
洗涤的工艺条件：采用的洗涤剂为盐酸，浓度以[H +]浓度计为6mol/L，洗涤过程的水相：油相相比为1:20，洗涤温度80℃，洗涤时间10min，洗涤次数7次；
反萃的工艺条件：采用的反萃剂为液碱，浓度以[OH -]的浓度计为4mol/L，反萃水相：油相相比为1:20，反萃时间20min，反萃后静置时间为20min；
沉淀的工艺条件：采用的沉淀剂为盐酸，控制pH为9，反应时间3h。
将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰，络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为80℃，回收后的氨水循环使用；
其中，碱化除杂的工艺条件为：采用的碱化除杂剂为氢氧化钠，除杂温度 为70℃，除杂pH为2.8，除杂时间为3小时；络合沉淀的工艺条件为：采用的络合剂为氨水，加入量为第二浸出液的10％，采用的沉淀剂为氧化镁，控制反应温度70℃，pH为8.5，反应时间为7小时。
有价金属中钪的回收率91％；镍的浸出率96％，全流程回收率93％；钴的浸出率95％，回收率92％；锰的浸出率95％，回收率92％。
实施例5
(1)将红土镍矿经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
其中，球磨的工艺条件：时间为1.7h，粒径为230目；浓密的工艺条件：底流浓度45％，溢流中的悬浮物含量为500ppm；硫酸的加入质量与红土镍矿的质量比为1.4:1；预浸温度为95℃，时间5h。
(2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；
其中，配矿为镁质红土镍矿(包括镍含量2.5％，铁含量7％，镁含量25％)，配矿的加入质量与步骤(1)的红土镍矿的质量比为0.5:1；采用氧压浸出，温度为200℃，浸出时间为1.7h，总压为1.8Mpa，氧分压为总压的15％；
中和反应的工艺条件为：采用浓度为18％碳酸钙浆料进行中和反应，中和反应的pH为2，中和反应温度65℃，中和反应时间1.8h。
(3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；
其中，酸洗的工艺条件为：第二浸出渣与浓度为98％硫酸按质量比为1:3进行酸洗，得到酸洗液，酸洗pH为0.2，酸洗时间4h，酸洗温度70℃，酸洗次 数1次；
萃取的工艺条件：将萃取剂、稀释剂加入酸洗液中进行萃取，得到的萃余液并入第二浸出液中；其中，萃取剂为TBP，稀释剂为磺化煤油，萃取剂为稀释剂的5％，水相：油相相比为25:1，萃取时间为25min，萃取后的静置时间为25min。
洗涤的工艺条件：采用的洗涤剂为硫酸，浓度以[H +]浓度计为3mol/L，洗涤过程的水相：油相相比为1:25，洗涤温度40℃，洗涤时间25min，洗涤次数3次；
反萃的工艺条件：采用的反萃剂为液碱氨，浓度以[OH -]的浓度计为7mol/L，反萃水相：油相相比为1:5，反萃时间25min，反萃后静置时间为25min；
沉淀的工艺条件：采用的沉淀剂为盐酸，控制pH为10，反应时间3h。
将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰，络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为90℃，回收后的氨水循环使用；
其中，碱化除杂的工艺条件为：采用的碱化除杂剂为碳酸钙，除杂温度为50℃，除杂pH为3.7，除杂时间为3小时；络合沉淀反应的工艺条件为：采用的络合剂为氨水，加入量为第二浸出液的10％，采用的沉淀剂为氧化镁，控制反应温度50℃，pH为8.5，反应时间为6小时。
有价金属中钪的回收率94％；镍的浸出率98％，全流程回收率95％；钴的浸出率98％，回收率95％；锰的浸出率97％，回收率94％。
对比例
(1)将红土镍矿(为褐铁矿层)经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
(2)将经步骤(1)得到的料浆加入蛇纹石型矿后进行加压浸出得到浸出液、浸出渣；温度为170℃，浸出时间为4h；
(3)将经步骤(2)得到的浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰。
有价金属中；镍的浸出率85％，全流程回收率80％；钴的浸出率87％，回收率85％；锰的浸出率80％，回收率86％。
以上所述的仅是本发明的较佳实施例，并不局限发明。应当指出对于本领域的普通技术人员来说，在本发明所提供的技术启示下，还可以做出其它等同改进，均可以实现本发明的目的，都应视为本发明的保护范围。

Claims (10)

  1. 一种从红土镍矿中综合提取有价金属的方法，其特征在于，所述方法包括：
    (1)将红土镍矿经球磨、浓密后得到的底流，加入硫酸进行预浸得到料浆；
    (2)将经步骤(1)得到的料浆加入配矿后进行压力浸出得到第一浸出液、第一浸出渣，将第一浸出渣进行中和反应后经固液分离得到第二浸出渣和第二浸出液；其中，所述配矿的加入质量与步骤(1)所述的红土镍矿的质量比为0.4～1.5:1，浸出温度为185～210℃，浸出时间1.5～2h，压力为1.6～2.5Mpa；
    (3)将经步骤(2)得到的第二浸出渣经过酸洗、萃取、洗涤、反萃、沉淀后得到氢氧化钪；将经步骤(2)得到的第二浸出液经过碱化除杂、络合沉淀后得到氢氧化镍钴锰。
  2. 根据权利要求1所述的方法，其特征在于，所述步骤(1)球磨：时间为0.5h～2h，粒径为200目～300目；浓密：底流浓度20％～45％，溢流中的悬浮物含量为0～1000ppm；硫酸的加入质量与红土镍矿的质量比为1:2～1.5:1；预浸温度为70～100℃，时间2～6h。
  3. 根据权利要求1所述的方法，其特征在于，所述步骤(2)配矿为褐铁矿层、过渡层红土镍矿或者镁质红土镍矿。
  4. 根据权利要求3所述的方法，其特征在于，当进行预浸的红土镍矿和配矿均为褐铁矿层时，压力为1.6～2.5Mpa，浸出时间为1.5～2h；当进行预浸的红土镍矿和配矿中的任意一种为过渡层红土镍矿或者镁质红土镍矿时，浸出时间为1.5～2h，总压为1.6MPa～2.5MPa，氧分压为总压的0～35％。
  5. 根据权利要求3或4所述的方法，其特征在于，所述褐铁矿层包括：镍含量0.9～1.5％，铁含量35～45％，镁含量小于5％；过渡层红土镍矿包括：镍含量1.3～ 1.7％，铁含量20～30％，镁含量5～10％；镁质红土镍矿包括：镍含量1.7～2.5％，铁含量5～20％，镁含量8～30％。
  6. 根据权利要求1或2所述的方法，其特征在于，所述步骤(2)第一浸出渣采用浓度为15～30％碳酸钙浆料进行中和反应，所述中和反应的工艺条件为：中和反应的pH为1～2，中和反应温度60～100℃，中和反应时间0.5～2h。
  7. 根据权利要求1或2所述的方法，其特征在于，所述步骤(3)将经步骤(2)得到的第二浸出渣与硫酸按固液比1:1～1:5进行酸洗，得到酸洗液，所述酸洗的工艺条件为：酸洗pH为0～1，酸洗时间0.5～4h，酸洗温度60～100℃，酸洗次数1～5次。
  8. 根据权利要求7所述的方法，其特征在于，所述步骤(3)萃取：将萃取剂、稀释剂加入酸洗液中进行萃取，得到负载有机和萃余液，萃余液并入第二浸出液中；其中，所述萃取剂为P204、P507、TBP中的一种或者几种混合，所述稀释剂为磺化煤油，所述萃取剂为稀释剂的1％～30％，水相：油相相比为10:1～30:1，萃取时间为3～30min，萃取后的静置时间为5～30min。
  9. 根据权利要求8所述的方法，其特征在于，所述步骤(3)洗涤：向经萃取分相得到的负载有机中加入洗涤剂，采用的洗涤剂为硫酸或者盐酸，所述洗涤剂的浓度以[H +]浓度计为1～8mol/L，洗涤过程的水相：油相相比为1:1～1:30，洗涤温度10～90℃，洗涤时间5～30min，洗涤次数1～10次；反萃：向经洗涤分相后得到的负载有机中加入反萃剂，采用的反萃剂为液碱，所述液碱浓度以[OH -]的浓度计为1～10mol/L，反萃水相：油相相比为1:1～1:30，反萃时间3～30min，反萃后静置时间为5～30min；沉淀：向经反萃分相得到的水相中加入沉淀剂，采用的沉淀剂为硫酸或者盐酸，控制pH为8～11，反应时间0.5～4h。
  10. 根据权利要求1或2所述的方法，其特征在于，所述步骤(3)碱化除杂的工艺条件为：采用的碱化除杂剂为碳酸钙或者氢氧化钠，除杂温度为40～80℃，除杂pH为2.8～5.5，除杂时间为2～4h；络合沉淀反应的工艺条件为：向经碱化除杂得到的除杂后液中加入络合剂和沉淀剂，采用的络合剂为氨水，氨水的加入量为第二浸出液的10％，采用的沉淀剂为氧化镁，控制反应温度40～80℃，pH为7.0～9.0，反应时间为2h～8h；络合沉淀产生的沉淀后液进入氨回收系统进行氨水回收，控制蒸发温度为80～150℃，回收后的氨水循环使用。
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