CN114686682B - Comprehensive smelting method of molybdenite - Google Patents

Comprehensive smelting method of molybdenite Download PDF

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CN114686682B
CN114686682B CN202011630432.2A CN202011630432A CN114686682B CN 114686682 B CN114686682 B CN 114686682B CN 202011630432 A CN202011630432 A CN 202011630432A CN 114686682 B CN114686682 B CN 114686682B
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leaching
molybdenum
molybdenite
rhenium
copper
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CN114686682A (en
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李永立
赵中伟
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Zhengzhou University
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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
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B15/00—Obtaining copper
    • C22B15/0063—Hydrometallurgy
    • C22B15/0065—Leaching or slurrying
    • C22B15/0067—Leaching or slurrying with acids or salts thereof
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B15/00—Obtaining copper
    • C22B15/0063—Hydrometallurgy
    • C22B15/0065—Leaching or slurrying
    • C22B15/0067—Leaching or slurrying with acids or salts thereof
    • C22B15/0073—Leaching or slurrying with acids or salts thereof containing nitrogen
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B15/00—Obtaining copper
    • C22B15/0063—Hydrometallurgy
    • C22B15/0084—Treating solutions
    • C22B15/0089—Treating solutions by chemical methods
    • C—CHEMISTRY; METALLURGY
    • C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B15/00—Obtaining copper
    • C22B15/0063—Hydrometallurgy
    • C22B15/0084—Treating solutions
    • C22B15/0089—Treating solutions by chemical methods
    • C22B15/0093—Treating solutions by chemical methods by gases, e.g. hydrogen or hydrogen sulfide
    • 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
    • C22B3/065—Nitric 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
    • C22B3/42—Treatment or purification of solutions, e.g. obtained by leaching by ion-exchange extraction
    • 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
    • C22B30/00—Obtaining antimony, arsenic or bismuth
    • C22B30/06—Obtaining bismuth
    • 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
    • C22B34/00—Obtaining refractory metals
    • C22B34/30—Obtaining chromium, molybdenum or tungsten
    • C22B34/34—Obtaining molybdenum
    • 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
    • C22B34/00—Obtaining refractory metals
    • C22B34/30—Obtaining chromium, molybdenum or tungsten
    • C22B34/36—Obtaining tungsten
    • 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
    • C22B61/00—Obtaining metals not elsewhere provided for in this subclass
    • 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

Abstract

The invention relates to a comprehensive smelting method of molybdenite, belongs to the technical field of metal smelting, and solves the problems that ammonia water or sodium hydroxide is used in the recovery process of molybdenum in the prior art; the technology using high-temperature roasting has high energy consumption and difficult temperature control; nitric acid is used for leaching under high temperature and high pressure, so that the consumption of the nitric acid is high, and the requirement on corrosion resistance of equipment is high; the recovery of valuable metals in molybdenite is difficult and the process is complex. According to the comprehensive smelting method of molybdenite, primary leaching and secondary leaching are sequentially carried out, after the primary leaching, filter residues obtained by filtering after the primary leaching and an auxiliary leaching agent system are subjected to oxygen pressure boiling under the oxygen pressure and water conditions, and secondary leaching is carried out; the leaching aid is one or a combination of phosphoric acid and calcium phosphate. Realizes the efficient recycling of various metal elements in molybdenite.

Description

Comprehensive smelting method of molybdenite
Technical Field
The invention relates to the technical field of metal smelting, in particular to a comprehensive smelting method of molybdenite.
Background
Molybdenite is a typical sulphide ore in which copper, bismuth, lead, rhenium and other valuable metals are associated. More than 99% of molybdenum in nature exists in the form of molybdenite, which is the most important mineral raw material for extracting molybdenum. Due to the influence of the lanthanide shrinkage effect, the atomic radius and chemical properties of tungsten and molybdenum are very similar, so that tungsten and molybdenum tend to be symbiotic and the later smelting and separation difficulty is very high. Meanwhile, besides tungsten impurities in molybdenum ore, metal impurities such as iron and manganese are removed. In the prior art, the molybdenite smelting process mostly adopts an oxidizing roasting-ammonia leaching process, a large amount of sulfur dioxide smoke is generated in the oxidizing roasting process, so that sulfur elements in the molybdenite cannot be effectively recycled, a large amount of air is polluted, and sublimation loss of valuable metal rhenium in the molybdenite is difficult to recycle in the oxidizing roasting process. Meanwhile, the temperature control difficulty is high in the roasting process, and the temperature is too high in the roasting process, so that molybdenum can be sublimated and lost; if the temperature in the roasting process is too low, incomplete molybdenum oxidation can be caused, and leaching difficulty is increased.
The pretreatment of molybdenite is generally carried out under alkaline conditions, for example, chinese patent CN201810597580 leaches with sodium hydroxide and soluble phosphates under alkaline conditions, mainly by the following chemical reactions (me=fe or Mn):
MeWO 4 +2NaOH＝Me(OH) 2 ↓+Na 2 WO 4
3MeWO 4 +2Na 3 (PO 4 ) 2 ＝Me 3 (PO 4 ) 2 ↓+3Na 2 WO 4
the alkaline pretreatment can only remove tungsten, and impurities such as ferro-manganese and the like generate insoluble hydroxide or phosphate which still remains in the slag, and the removal of ferro-manganese can be realized by acid leaching. The invention adopts a mixed solution of sulfuric acid and phosphoric acid, iron and manganese are leached into the solution together with tungsten in the pretreatment process, and the following chemical reaction (Me=Fe or Mn) mainly occurs:
12MeWO 4 +12H 2 SO 4 +H 3 PO 4 ＝12MeSO 4 +H 3 PW 12 O 40 +12H 2 O。
therefore, the difficulty of molybdenite smelting is that the symbiosis exists between molybdenum and tungsten, so that the molybdenum and the tungsten are separated and recovered, and meanwhile, the separation and impurity removal of iron and manganese elements are realized, and the influence of iron and manganese on the purification of the molybdenum and the tungsten is eliminated.
Disclosure of Invention
In view of the above analysis, the present invention aims to provide a comprehensive smelting method of molybdenite, which can solve at least one of the following technical problems: (1) Ammonia water or sodium hydroxide is used in the recovery process of molybdenum; (2) Nitric acid is used for leaching under high temperature and high pressure, so that the consumption of the nitric acid is high, and the requirement on corrosion resistance of equipment is high; (3) it is very difficult to effectively separate tungsten and molybdenum from molybdenite; (4) In the effective separation process of molybdenum and tungsten, iron and manganese impurity elements are difficult to separate, so that the purification of the molybdenum and the tungsten is influenced; (5) Other valuable metals in molybdenite are difficult to recover and the process is complex.
The invention provides a comprehensive smelting method of molybdenite, which sequentially carries out primary leaching and secondary leaching, after the primary leaching, under the conditions of oxygen pressure and water, filter residues obtained by filtering after the primary leaching and an auxiliary leaching agent system are subjected to oxygen pressure boiling, and secondary leaching is carried out; the leaching aid is one or a combination of phosphoric acid and calcium phosphate.
Further, the leaching agent for the primary leaching is a mixed solution of nitric acid and phosphoric acid.
Further, the nitric acid concentration is 20g/L to 50g/L.
Further, the concentration of the phosphoric acid is 10g/L to 30g/L.
Further, the temperature of the primary leaching is 80-100 ℃.
Further, the liquid-solid ratio of the primary leaching is 3L/kg-5L/kg.
Further, the time of one leaching is 2 to 5 hours.
Further, filtering after one-time leaching to obtain a filtrate, wherein the filtrate is a solution containing tungsten, iron and manganese elements.
Further, the filtrate is subjected to resin adsorption to realize the separation of tungsten.
Further, the oxygen pressure of the oxygen autoclaving is 0.8Mpa to 1.5Mpa.
Further, in the molybdenite and leaching aid system, the consumption of the leaching aid is 0.5-1.5 times of the mass of the molybdenite.
Further, the oxygen autoclaving time is 2 to 5 hours.
Further, the temperature of the oxygen autoclaving is 180-210 DEG C
Further, leaching liquid containing molybdenum, rhenium and copper and leaching slag enriched with bismuth are obtained after secondary leaching, and bismuth separation is realized.
Further, the leaching rate of molybdenum is more than 99%, the leaching rate of rhenium is more than 95%, and the leaching rate of copper is more than 98%.
Further, after obtaining a leaching solution containing molybdenum, rhenium and copper, the method comprises the following steps:
extracting molybdenum from leaching solution containing molybdenum, rhenium and copper by using an extractant to obtain raffinate and extract containing molybdenum, and separating molybdenum.
Further, the extractant is a cationic extractant.
Further, the cation extractant is one of P204 and P507.
Further, the concentration of the extractant is 10-50% by mass percent.
Further, the extraction of molybdenum is compared with O/a=4:1 to 2:1.
Further, the molybdenum extraction mode is countercurrent extraction.
Further, the extraction level of the extracted molybdenum is 4-7.
Further, after the separation of molybdenum is achieved, it comprises:
and (3) adsorbing rhenium in the raffinate by using ion exchange resin to obtain an exchange raffinate, so as to realize separation of rhenium.
Further, the ion exchange resin is designated 201.
Further, after separation of rhenium is achieved, it comprises:
adding soluble sulfide into the exchange raffinate to precipitate and enrich copper ions, so as to realize copper separation.
Further, the extracting molybdenum from the leaching solution containing molybdenum, rhenium and copper specifically comprises:
extracting molybdenum from the leaching solution by using a cationic extractant to obtain raffinate and a molybdenum-loaded cationic extractant;
and back-extracting molybdenum from the molybdenum-loaded cationic extractant by using a back-extracting agent to obtain molybdenum-containing extract.
Further, the stripping agent is hydrogen peroxide.
Further, the mass fraction of the hydrogen peroxide is 10% -20%.
Further, the extraction of the strip molybdenum is compared with O/a=7:1 to 5:1.
Further, the extraction mode of the back-extracted molybdenum is countercurrent extraction.
Further, the extraction level of the back-extracted molybdenum is 2-4.
Further, the raffinate is recycled to obtain the leaching aid after being added with the neutralizer, and the leaching aid is recycled for oxygen autoclaving and leaching of molybdenite.
Further, the main components of the molybdenite are 45-55% of molybdenum, 1-1.5% of copper, 0.04-0.05% of rhenium, 2.3-2.7% of bismuth, 0.1-0.5% of tungsten and 3-8% of ferromanganese in total by mass percent.
Further, the recovery rate of molybdenum is 99% or more, the recovery rate of rhenium and tungsten is 85% or more, and the recovery rate of copper is 90% or more.
Further, the comprehensive smelting method of molybdenite provided by the invention comprises the following steps:
leaching molybdenite by using a leaching agent for primary leaching, and filtering;
under the conditions of oxygen pressure and water, carrying out oxygen pressure boiling on filter residues obtained after primary leaching and a leaching aid system, and carrying out secondary leaching;
after secondary leaching, obtaining leaching liquid containing molybdenum, rhenium and copper and leaching slag enriched with bismuth, and realizing bismuth separation;
extracting molybdenum from leaching solution containing molybdenum, rhenium and copper by using a cationic extractant to obtain raffinate and a molybdenum-loaded cationic extractant;
back-extracting molybdenum from the molybdenum-loaded cationic extractant by using a back-extracting agent to obtain molybdenum-containing extract;
adsorbing rhenium in the raffinate by using ion exchange resin to obtain an exchange raffinate, and separating rhenium;
adding soluble sulfide into the exchange raffinate to precipitate and enrich copper ions, so as to realize copper separation. Compared with the prior art, the invention has at least one of the following beneficial effects:
(1) Compared with the existing technology that alkali is required to be used for separating tungsten in the treatment of molybdenite, the method provided by the invention has the advantages that the mixed acid of nitric acid and phosphoric acid is used for leaching the molybdenite once, the tungsten, iron and manganese are leached, the tungsten is separated through ion exchange resin, the effective separation of the tungsten and the molybdenum is realized, the iron and manganese impurity elements are removed, and the influence of iron and manganese on the separation and recovery of the molybdenum and the tungsten is effectively avoided.
(2) Compared with the existing leaching technology using inorganic strong acid, the leaching technology of the invention does not consume a large amount of inorganic strong acid in the leaching technology of the molybdenite by oxygen pressure boiling, and the leaching technology of the invention realizes the mild leaching and high leaching rate leaching of the molybdenite (the leaching rate of molybdenum is up to 99%, the leaching rate of rhenium is more than 95% and the leaching rate of copper is more than 98%) by adding phosphoric acid or calcium phosphate as an auxiliary leaching agent to perform oxygen pressure boiling on the molybdenite.
(3) In the leaching treatment process of molybdenite, molybdenum is converted into molybdenum acyl cations, rhenium is converted into rhenate ions, copper is converted into copper sulfate and enters a solution, and then molybdenum, rhenium and copper in the leaching solution are sequentially enriched and recovered; bismuth is fully enriched in slag for recycling or directly selling, so that the recovery of molybdenum, rhenium, bismuth and copper valuable metals in molybdenite is realized.
(4) The method has the advantages that the neutral phosphine extractant is used for extracting molybdenum, the hydrogen peroxide is used for back extraction of molybdenum, the molybdenum is efficiently recovered by simple extraction by fully utilizing different existence modes (phosphomolybdic heteropolyacid, molybdenum acyl cation and peroxymolybdic acid anion) of molybdenum under different environments, and the use of strong alkali and the discharge of ammonia nitrogen wastewater are avoided.
In the invention, the technical schemes can be mutually combined to realize more preferable combination schemes. Additional features and advantages of the invention will be set forth in the description which follows, and in part will be obvious from the description, or may be learned by practice of the invention. The objects and other advantages of the invention may be realized and obtained by means of the instrumentalities and combinations particularly pointed out in the specification.
Drawings
FIG. 1 is a flow chart of a comprehensive smelting process of molybdenite.
Detailed Description
The invention provides a comprehensive smelting method of molybdenite, which is shown in figure 1, wherein a process flow chart of the method is shown in the figure 1, primary leaching is carried out on the molybdenite by utilizing a first leaching agent, filtrate and filter residues are obtained through filtration and separation, and under the conditions of oxygen pressure and water, the filter residues and an auxiliary leaching agent system are subjected to oxygen pressure boiling to realize secondary leaching of metals in the molybdenite; the leaching aid is one or a combination of phosphoric acid and calcium phosphate.
Specifically, the leaching agent for primary leaching is a mixed solution of nitric acid and phosphoric acid.
Specifically, the concentration of nitric acid is 20g/L to 50g/L.
Specifically, the concentration of phosphoric acid is 10g/L to 30g/L.
Specifically, the temperature of one leaching is 80-100 ℃.
Specifically, the liquid-solid ratio of the primary leaching is 3L/kg-5L/kg.
Specifically, the time of one leaching is 2 to 5 hours.
Specifically, the filtrate is obtained by filtering after one-time leaching, the filtrate is a solution containing tungsten, iron and manganese elements, and the separation of tungsten can be realized by carrying out resin adsorption on the filtrate.
Due to the influence of the lanthanide shrinkage effect, the atomic radius and chemical properties of tungsten and molybdenum are very similar, so that tungsten and molybdenum tend to be symbiotic and the later smelting and separation difficulty is very high. Meanwhile, besides tungsten impurities in molybdenum ore, metal impurities such as iron and manganese are removed, and if the impurities cannot be well removed, the subsequent purification of molybdenum and tungsten is greatly influenced. Through researches, tungsten, iron and manganese can be leached out through primary leaching of mixed acid of nitric acid and phosphoric acid, and elements such as molybdenum, rhenium, bismuth and copper in ores cannot be leached out. Filtering the mixture after primary leaching to obtain filtrate and filter residue. The filtrate is a solution containing tungsten, iron and manganese elements, the tungsten is recovered through ion exchange resin, so that iron and manganese impurity elements are effectively separated from undigested molybdenum, rhenium, bismuth and copper, and the tungsten is recovered through an ion exchange method, so that the iron and manganese elements are effectively removed.
In the primary leaching, elements such as molybdenum, rhenium, bismuth, copper and the like in the ore cannot be leached out, remain in filter residues, and can be separated only by effectively treating the filter residues. For the filter residue after primary leaching, the treatment method combining oxygen pressure boiling and leaching is adopted, phosphoric acid and/or calcium phosphate serve as an auxiliary leaching agent to play roles in assisting leaching and ore digestion in a system, under the condition of oxygen, oxygen oxidizes molybdenum disulfide in molybdenite, sulfur in the molybdenite is oxidized into IV-valent sulfur and is further oxidized into VI-valent sulfuric acid, and under the effect of sulfuric acid generated by oxidizing sulfur in the filter residue, molybdenum in the filter residue is converted into soluble molybdenum acyl cations so as to be completely dissolved in a liquid phase to form a solution.
Through experimental study, the filter residue cannot be digested by using the leaching aid phosphoric acid and/or calcium phosphate alone; oxygen is independently introduced into the water environment for oxypoaching, and the ore cannot be digested. Therefore, under the condition of oxygen, phosphoric acid and/or calcium phosphate are used as auxiliary leaching agents to realize digestion and leaching of filter residues, and realize leaching of molybdenum, rhenium and copper. While bismuth is not leached, is totally enriched in residual leaching residues and can be used for recovery or directly sold.
As the oxygen autoclaving continues, molybdenum sulfide in the filter residue is continuously oxidized and combined with phosphate radical under the combined action of oxygen and the leaching aid, and is converted into phosphomolybdic heteropolyacid and sulfuric acid, and the acidity is gradually increased. Under a strong acid environment, the reaction balance of the mutual conversion of the phosphomolybdic heteropolyacid and the molybdenum acyl cations continuously moves to the direction of generating the molybdenum acyl cations, and molybdenum in filter residues is dissolved and converted into a state of the molybdenum acyl cations with good solubility. The phosphoric acid or calcium phosphate assists molybdenum to be converted into phosphomolybdic heteropolyacid in the digestion and leaching process of the filter residue, and is released in the process of converting the phosphomolybdic heteropolyacid into molybdenum acyl cations, thereby playing a role in assisting leaching in the digestion and leaching process of the filter residue.
Specifically, the oxygen pressure is 0.8Mpa to 1.5Mpa.
Oxygen is a key factor playing an oxidation role in the oxygen autoclaving process, and because the oxygen autoclaving leaching process is a heterogeneous reaction of gas-solid-liquid, the general gas participation reaction mainly depends on the heterogeneous reaction of the interface of gas-solid and gas-liquid, and the reaction rate of the interface reaction of gas participation is slow, so that the oxygen autoclaving leaching efficiency is seriously affected. The oxygen is provided with enough pressure, so that the oxygen is better dissolved in the liquid, and a gas-solid-liquid heterogeneous reaction is generated at a heterogeneous reaction interface of the liquid and the solid, thereby greatly improving the efficiency of oxygen autoclaving leaching.
Specifically, in the filter residue and leaching aid system, the consumption of the leaching aid is 0.5-1.5 times of the mass of the filter residue.
Specifically, the autoclaving time is 2-5 hours.
Specifically, the temperature of the oxygen leaching and autoclaving is 180-210 ℃.
The longer the oxygen autoclaving time, the more thorough the reaction, but at the same time, too long autoclaving results in higher energy consumption, so that autoclaving times of 2 to 5 hours are chosen.
The higher the leaching oxygen autoclaving temperature is, the faster the autoclaving speed is, but at the same time, the higher energy consumption is caused, so that the production efficiency and the energy consumption of autoclaving are comprehensively considered, and the leaching oxygen autoclaving temperature is determined to be 180-210 ℃ according to the efficiency-cost ratio.
Specifically, the leaching rate of molybdenum is more than 99%, the leaching rate of rhenium and tungsten is more than 95%, and the leaching rate of copper is more than 98%.
Specifically, the filter residue and the leaching aid system are subjected to oxygen pressure boiling to obtain leaching liquid containing molybdenum, rhenium and copper.
Specifically, after obtaining a leaching solution containing molybdenum, rhenium and copper, the leaching solution comprises:
extracting molybdenum from the leaching solution by using an extractant to obtain raffinate and extract containing molybdenum, and separating molybdenum;
adsorbing rhenium in the raffinate by using ion exchange resin to obtain an exchange raffinate, and separating rhenium;
adding soluble sulfide into the exchange raffinate to precipitate and enrich copper ions, so as to realize copper separation.
In one possible embodiment, the ion exchange resin is identified by the numeral 201.
The method for recovering copper is a precipitation method, namely, negative bivalent sulfur and copper are added into the exchange residual liquid to generate copper sulfide precipitate, so that copper enrichment recovery is realized. Wherein the negative divalent sulfur is a soluble sulfide; in one possible embodiment, the soluble sulfide includes sodium sulfide, and like sulfide salts; in another possible embodiment, the soluble sulfide may also include hydrogen sulfide gas.
Specifically, the extracting molybdenum from the leaching solution by using the extractant specifically comprises:
extracting molybdenum from the leaching solution by using a cationic extractant to obtain raffinate and a molybdenum-loaded cationic extractant;
and back-extracting molybdenum from the molybdenum-loaded cationic extractant by using a back-extracting agent to obtain molybdenum-containing extract.
Specifically, the cation extractant is one or a combination of P204 and P507, and the back extractant is hydrogen peroxide.
Specifically, the concentration of the extractant is 10-50% by mass percent.
Specifically, the extraction ratio O/a=4:1 to 2:1 of the extracted molybdenum.
The ratio of the extraction is an important factor of the extraction, and when the ratio of the O/A ratio of the extraction is less than 2:1, the organic extraction is insufficient, and part of molybdenum cannot be transferred into the organic phase, so that from the viewpoint of the recovery rate of molybdenum, the larger the ratio of the O/A, the more thoroughly the molybdenum is transferred into the organic phase, and the less residues remain in the aqueous phase. However, too much organic phase would increase the cost of solvents and processes, and when O/A is greater than 4:1, the cost ratio of extraction is severely reduced, so that the extraction of molybdenum is selected to be compared with O/A=4:1-2:1.
Specifically, the molybdenum extraction mode is countercurrent extraction.
Specifically, the extraction level of molybdenum is 4-7.
Specifically, the mass fraction of the hydrogen peroxide is 10% -20%.
Specifically, the extraction ratio O/a=7:1 to 5:1 of the stripping molybdenum.
Specifically, the extraction mode of the back-extracted molybdenum is countercurrent extraction.
Specifically, the extraction level of the back-extracted molybdenum is 2-4.
The leaching solution is subjected to chemical extraction by the cation extractant, hydrogen ions in the cation extractant and leached molybdenum acyl cations are subjected to cation exchange, and the molybdenum acyl cations are transferred to an organic phase.
In the chemical equilibrium process of the molybdenum acyl cations in the organic phase, a small amount of molybdenum acyl cations enter the water phase and are converted into molybdate ions, in the back extraction process, a small amount of molybdic acid radicals which enter the water phase and are converted into peroxomonosylate anions by the hydrogen peroxide serving as a back extractant promote the chemical equilibrium to move towards the direction of the molybdenum acyl cations converted into molybdate ions, and all the molybdenum acyl cations are converted into peroxomonosylate anions which are thoroughly separated from the organic phase, so that the back chemical extraction of molybdenum from the molybdenum-loaded cationic extractant is realized, and all the molybdenum ions are back extracted into the water phase of the back extractant.
Specifically, the raffinate is recycled after being added with the neutralizer to obtain the leaching aid, and the leaching aid can be reused for oxygen autoclaving and leaching of molybdenite and is recycled for oxygen autoclaving and leaching of molybdenite.
Specifically, the neutralizer is one or more of calcium oxide, calcium hydroxide, calcium carbonate and calcium phosphate.
The neutralizing agent neutralizes excessive hydrogen ions in the system and retains phosphoric acid or calcium phosphate in the system, so that the neutralizing agent can recover the leaching aid after neutralization and is directly used for the leaching oxygen autoclaving process in the step 1 and is used for digestion and leaching of ores.
The addition of the neutralizer can simultaneously precipitate the impurity iron enriched in the system, remove the impurity iron from the system and prevent Fe 3+ The process of extracting molybdenum acyl cations from the cations in the system is interfered.
Specifically, the main components of molybdenite comprise 45 to 55 percent of molybdenum, 1 to 1.5 percent of copper, 0.04 to 0.05 percent of rhenium, 2.3 to 2.7 percent of bismuth, 0.1 to 0.5 percent of tungsten and 3 to 8 percent of iron in percentage by mass.
Specifically, the recovery rate of molybdenum is 99% or more, the recovery rate of rhenium and tungsten is 85% or more, and the recovery rate of copper is 90% or more.
The following detailed description of the preferred embodiments of the invention illustrates the principles of the invention and is not intended to limit the scope of the invention.
Example 1
In one embodiment of the invention, a method for recovering metallic elements and copper from molybdenite is disclosed.
The main components of the molybdenite raw material are 45.1% of molybdenum, 1.3% of copper, 0.04% of rhenium, 2.59% of bismuth, 0.1% of tungsten content and 3% of ferromanganese in total by mass percent.
The mixed solution of 20g/L nitric acid and 10g/L phosphoric acid is used as a primary leaching agent, heated to 80 ℃, and reacted with molybdenite for 5 hours under the condition that the liquid-solid ratio is 1:1.
And after the primary leaching is finished, filtering to obtain filtrate and filter residue.
The obtained filtrate is a solution in which the elements tungsten, iron, manganese and the like are dissolved, and the 201 resin adsorption method is adopted to recover the valuable elements tungsten, wherein the adsorption rate of the tungsten is 95.5%.
Phosphoric acid is used as an auxiliary leaching agent, the consumption of the auxiliary leaching agent is 0.5 times of the mass of filter residues, the oxygen partial pressure is 1Mpa, and the leaching is carried out in an autoclave for 2 hours under the condition of the temperature of 190 ℃.
After oxygen autoclaving leaching, cooling to room temperature and filtering to obtain leaching slag containing 16.1% of bismuth and leaching liquid containing molybdenum, rhenium and copper.
Wherein, the leaching rate of molybdenum is up to 99%, the leaching rate of rhenium is 95%, and the leaching rate of copper is up to 98%.
For leaching solution containing molybdenum, rhenium and copper, firstly, extracting molybdenum by using 40% of cationic extractant P204 to obtain molybdenum-loaded P204, wherein the extraction ratio is O/A=2:1, and the extraction stages are five countercurrent stages.
Carrying out back extraction on the P204 loaded with molybdenum by using 12% hydrogen peroxide to obtain back extraction liquid and raffinate; the back extraction phase ratio is O/A=7:1, and the back extraction stage number is countercurrent three.
The total recovery rate of molybdenum in the extraction back-extraction process reaches 99 percent.
Evaporating and crystallizing the strip liquor to obtain molybdic acid, and calcining to obtain a molybdenum trioxide product.
The raffinate after molybdenum extraction adopts 201 resin to adsorb rhenium, and the recovery rate of rhenium can reach 85 percent.
And (3) enriching copper and recovering copper by adding sodium sulfide into the exchange raffinate after rhenium extraction, wherein the recovery rate of copper reaches 90%.
And adding calcium hydroxide into the residual liquid after back extraction, and filtering to precipitate when the pH value of the solution is regulated to about 1, wherein the filtered filtrate is used as an oxygen pressure cooking leaching liquid for recycling.
Example two
In one embodiment of the invention, a method for recovering metallic elements and copper from molybdenite is disclosed.
The main components of the molybdenite raw material are 50.2% of molybdenum, 1.2% of copper, 0.05% of rhenium, 2.42% of bismuth, 0.2% of tungsten and 5% of ferromanganese in total by mass percent.
The mixed solution of 50g/L nitric acid and 30g/L phosphoric acid is used as a primary leaching agent, heated to 90 ℃, and reacted with molybdenite for 2 hours under the condition that the liquid-solid ratio is 3:1.
And after the primary leaching is finished, filtering to obtain filtrate and filter residue.
The obtained filtrate is a solution in which the elements tungsten, iron, manganese and the like are dissolved, and the 201 resin adsorption method is adopted to recover the valuable elements tungsten, wherein the adsorption rate of the tungsten is more than 95.9 percent.
Calcium phosphate is used as an infusion aid, and the usage amount of the infusion aid is 1 time of the mass of filter residues. The mixture is leached by oxygen pressure in an autoclave for 5 hours under the condition of oxygen partial pressure of 0.8Mpa and temperature of 210 ℃.
After oxygen autoclaving leaching, cooling to room temperature and filtering to obtain leaching slag containing 16.1% of bismuth and leaching liquid containing molybdenum, rhenium and copper.
Wherein, the leaching rate of molybdenum is up to 99%, the leaching rate of rhenium is 95%, and the leaching rate of copper is up to 98%.
For leaching solution containing molybdenum, rhenium and copper, firstly, 50% of cationic extractant P204 is used for extracting molybdenum to obtain molybdenum-loaded P204, the extraction ratio is O/A=4:1, and the extraction stages are five countercurrent stages.
Back-extracting the P204 loaded with molybdenum by using 10% hydrogen peroxide to obtain back-extracted liquid and raffinate; the back extraction phase ratio is O/A=7:1, and the back extraction stage number is countercurrent three.
The total recovery rate of molybdenum in the extraction back-extraction process reaches 99.1 percent.
Evaporating and crystallizing the strip liquor to obtain molybdic acid, and calcining to obtain a molybdenum trioxide product.
The raffinate after molybdenum extraction adopts 201 resin to adsorb rhenium, and the recovery rate of rhenium can reach 86 percent.
And after rhenium extraction, the copper is enriched and recovered by introducing the exchange raffinate into the method of hydrogen sulfide copper precipitation, and the copper recovery rate reaches 91%.
Adding calcium carbonate into the residual liquid after back extraction, and filtering to precipitate when the pH value of the solution is regulated to about 1, wherein the filtered filtrate is used as oxygen pressure cooking leaching liquid for recycling.
Example III
The invention discloses a comprehensive smelting method of molybdenite.
The main components of the molybdenite raw material are 48.2% of molybdenum, 1.4% of copper, 0.045% of rhenium, 2.69% of bismuth, 0.3% of tungsten and 8% of ferromanganese in total.
The mixed solution of 40g/L nitric acid and 20g/L phosphoric acid is used as a primary leaching agent, heated to 100 ℃, and reacted with molybdenite for 3 hours under the condition that the liquid-solid ratio is 2:1.
And after the primary leaching is finished, filtering to obtain filtrate and filter residue.
The obtained filtrate is a solution in which the elements tungsten, iron, manganese and the like are dissolved, and the 201 resin adsorption method is adopted to recover the valuable elements tungsten, wherein the adsorption rate of the tungsten is 96.2 percent.
Phosphoric acid is used as an auxiliary leaching agent, the consumption of the auxiliary leaching agent is 1.5 times of the mass of filter residues, the oxygen partial pressure is 1.5Mpa, and the leaching is carried out in an autoclave for 3 hours under the condition of oxygen pressure and 190 ℃.
After oxygen pressure boiling leaching, cooling to room temperature and filtering to obtain leaching slag containing 17% of bismuth and leaching liquid containing molybdenum, rhenium and copper.
Wherein, the leaching rate of molybdenum is up to 99.1%, the leaching rate of rhenium is 95.5%, and the leaching rate of copper is up to 98.6%.
For leaching solution containing molybdenum, rhenium and copper, firstly, extracting molybdenum by using 30% of cationic extractant P507 to obtain molybdenum-loaded P507, wherein the extraction ratio is O/A=3:1, and the extraction stages are five countercurrent stages.
Back-extracting the molybdenum-loaded P507 with 10% hydrogen peroxide to obtain back-extracted liquid and raffinate; the back extraction phase ratio is O/A=7:1, and the back extraction stage number is countercurrent three.
The total recovery rate of molybdenum in the extraction back-extraction process reaches 99 percent.
Evaporating and crystallizing the strip liquor to obtain molybdic acid, and calcining to obtain a molybdenum trioxide product.
The raffinate after molybdenum extraction adopts 201 resin to adsorb rhenium, and the recovery rate of rhenium can reach 85.4 percent.
And (3) enriching copper and recovering copper by adding sodium sulfide into the exchange raffinate after rhenium extraction, wherein the recovery rate of copper reaches 90.2%.
Adding calcium oxide into the residual liquid after back extraction, and filtering to precipitate when the pH value of the solution is regulated to about 1, wherein the filtered filtrate is used as oxygen pressure cooking leaching liquid for recycling.
The present invention is not limited to the above-mentioned embodiments, and any changes or substitutions that can be easily understood by those skilled in the art within the technical scope of the present invention are intended to be included in the scope of the present invention.

Claims (7)

1. A comprehensive smelting method of molybdenite is characterized in that primary leaching and secondary leaching are sequentially carried out, after the primary leaching, filter residues obtained by filtering after the primary leaching and an auxiliary leaching agent system are subjected to oxygen pressure boiling under the oxygen pressure and water conditions, and secondary leaching is carried out; the leaching aid is one or a combination of phosphoric acid and calcium phosphate;
wherein, the oxygen pressure of the oxygen pressure cooking is 0.8 Mpa-1.5 Mpa, the oxygen pressure cooking time is 2 hours-5 hours, the temperature of the oxygen pressure cooking is 180 ℃ to 210 ℃, and the consumption of the leaching aid is 0.5 to 1.5 times of the mass of the molybdenite in the molybdenite and leaching aid system;
the leaching agent for primary leaching is a mixed solution of nitric acid and phosphoric acid;
filtering after primary leaching to obtain filtrate, wherein the filtrate is a solution containing tungsten, iron and manganese elements; and leaching liquid containing molybdenum, rhenium and copper and leaching slag enriched with bismuth are obtained after the secondary leaching, so that bismuth separation is realized.
2. The method for comprehensive smelting of molybdenite according to claim 1, wherein the filtrate is subjected to resin adsorption to realize tungsten separation.
3. The comprehensive smelting method of molybdenite according to claim 1, wherein,
extracting molybdenum from leaching solution containing molybdenum, rhenium and copper by using an extractant to obtain raffinate and extract containing molybdenum, and separating molybdenum.
4. A method for the integrated smelting of molybdenite according to claim 3, characterized in that after the separation of molybdenum is achieved, it comprises:
and (3) adsorbing rhenium in the raffinate by using ion exchange resin to obtain an exchange raffinate, so as to realize separation of rhenium.
5. The method for integrated smelting molybdenite according to claim 4, wherein after separating rhenium is achieved, the method comprises:
adding soluble sulfide into the exchange raffinate to precipitate and enrich copper ions, so as to realize copper separation.
6. The method of claim 3, wherein the extracting molybdenum from the leaching solution containing molybdenum, rhenium, and copper comprises:
extracting molybdenum from leaching solution containing molybdenum, rhenium and copper by using a cationic extractant to obtain raffinate and a molybdenum-loaded cationic extractant;
and back-extracting molybdenum from the molybdenum-loaded cationic extractant by using a back-extracting agent to obtain molybdenum-containing extract.
7. The method of claim 3, wherein the raffinate is recycled to the leaching aid after the neutralizing agent is added, and the leaching aid is recycled for oxygen autoclaving and leaching of the molybdenite.
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