CN108950195B - Method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater - Google Patents
Method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater Download PDFInfo
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- CN108950195B CN108950195B CN201811003061.8A CN201811003061A CN108950195B CN 108950195 B CN108950195 B CN 108950195B CN 201811003061 A CN201811003061 A CN 201811003061A CN 108950195 B CN108950195 B CN 108950195B
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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
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/14—Agglomerating; Briquetting; Binding; Granulating
- C22B1/24—Binding; Briquetting ; Granulating
- C22B1/242—Binding; Briquetting ; Granulating with binders
- C22B1/243—Binding; Briquetting ; Granulating with binders inorganic
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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
- C22B1/00—Preliminary treatment of ores or scrap
- C22B1/02—Roasting processes
- C22B1/08—Chloridising roasting
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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
- C22B11/00—Obtaining noble metals
- C22B11/04—Obtaining noble metals by wet processes
- C22B11/042—Recovery of noble metals from waste materials
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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
- C22B15/00—Obtaining copper
- C22B15/0002—Preliminary treatment
- C22B15/0004—Preliminary treatment without modification of the copper constituent
- C22B15/0008—Preliminary treatment without modification of the copper constituent by wet processes
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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
- C22B15/00—Obtaining copper
- C22B15/0063—Hydrometallurgy
- C22B15/0065—Leaching or slurrying
- C22B15/0067—Leaching or slurrying with acids or salts thereof
- C22B15/0069—Leaching or slurrying with acids or salts thereof containing halogen
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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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
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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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/30—Obtaining zinc or zinc oxide from metallic residues or scraps
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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
- C22B23/02—Obtaining nickel or cobalt by dry processes
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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
- C22B23/0423—Halogenated acids or salts thereof
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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
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/005—Separation by a physical processing technique only, e.g. by mechanical breaking
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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
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/007—Wet processes by acid leaching
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- 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 method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater, which comprises the steps of sequentially carrying out fine grinding, screening, stirring and granulating to obtain pellets, predrying, high-temperature roasting to obtain high-temperature flue gas (13) containing valuable metals and roasting slag, stirring to obtain reduction roasting slag, preserving heat to obtain heat preservation materials and heat preservation flue gas, crushing, fine grinding to obtain ground heat preservation materials, carrying out magnetic separation to obtain magnetite concentrate and silico-aluminum-calcium residues, leaching and collecting washing liquid containing valuable metal elements, heating and concentrating to obtain washing liquid for removing hydrogen chloride gas, recycling iron-containing products, copper-containing products, cobalt-nickel products and filtrate in sections, not only recycling the valuable metal elements such as iron, zinc, copper, cobalt, nickel and the like in pressure oxidation waste slag, but also reducing the treatment cost of waste water difficult to treat, and having small potential heavy metal pollution risk of the chlorine-containing waste slag, the method has the advantages of low acid consumption, low cost, high recovery rate, short process flow, low equipment investment, low maintenance cost and the like, and is suitable for comprehensive recovery and application of the solid waste residues in the nonferrous metallurgy.
Description
Technical Field
The invention relates to a method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater, which is suitable for comprehensive recovery and application of non-ferrous metallurgy solid waste slag.
Background
In general, zinc concentrate mainly contains zinc sulfide, and the pressure oxidation leaching process of zinc sulfide concentrate is actually a gas-liquid-solid three-phase reaction process, and the liquid-solid nucleation reaction of solid product formation includes external diffusion of liquid reactants or products through a liquid boundary layer, internal diffusion of liquid reactants or products through a solid product layer, and interfacial chemical reaction. In fact, the reaction process involves Fe in solution3+Diffuse to the unreacted nuclear interface through a solid sulfur film, and then Fe3+Reacts with zinc sulfide in unreacted core to generate elemental sulfur and Zn2+And then from the Zn formed2+Out-diffusion through a solid sulfur film. The basic reaction equation for pressure oxidation leaching of zinc sulfide:
2ZnS+2H2SO4+O2→ZnSO4+H2O+S0
the reaction process is very slow in the absence of substances which transfer oxygen molecules, and the oxygen molecules transfer when the iron element in the zinc sulfide concentrate enters the solution and is oxidized into ferric ions by oxygen, greatly accelerating the reaction process.
ZnS+2Fe2(SO4)3→ZnSO4+FeSO4+S0
H2SO4+2FeSO4+0.5O2→Fe2(SO4)3+H2O
The oxidation waste slag generated in the zinc smelting industry refers to oxidation slag obtained by filtering and washing pressure leached ore pulp and feeding the slag for slag flotation, and the main mineral composition of the oxidation waste slag comprises elemental sulfur, grass yellow iron vanadium, sodium iron vanadium, yellow potassium iron vanadium, lead iron vanadium, a small amount of quartz, residual zinc blende and valuable metals containing more iron, zinc, copper, cobalt, nickel and the like. At present, the oxidation waste residue is not further utilized in the industry, but is neutralized and then stored in a tailing pond, and the problems of low utilization rate of valuable metal resources in zinc concentrate pressure oxidation slag, heavy metal pollution in the storage process, incapability of treating and disposing chlorine-containing wastewater and the like exist.
The current processing methods of zinc sulfide concentrate include roasting method, pressure oxidation method and biological oxidation method. The roasting method is as disclosed in Chinese patent CN201010611051.X, and the method is characterized in that the high-iron zinc sulfide concentrate is subjected to roasting-fine grinding-wet magnetic separation to realize iron-zinc separation, so that low-iron zinc concentrate and high-iron zinc concentrate are obtained. The pressure oxidation method is a method for simultaneously recovering valuable metals by performing pressure oxygen leaching on zinc sulfide concentrate under high temperature and high pressure, as disclosed in Chinese patent CN201410112953.7, aiming at zinc sulfide concentrate rich in lead silver and rare metals and zinc sulfide concentrate free of lead silver and rare metals, zinc sulfide is firstly oxidized at high temperature and high pressure, leaching liquid and leaching residues are obtained by filtering and separating, lead silver-containing oxidation residues and iron residues are obtained, the leaching liquid is neutralized, the neutralized residues of the rare metals are recovered, and then the supernatant is subjected to zinc recovery; the process for recovering nickel, cobalt and iron by high-temperature chlorination and volatilization of nickel oxide ore disclosed in Chinese patent CN201410504145.5 adopts one or more of calcium chloride, sodium chloride and magnesium chloride as chlorinating agent; chinese patent CN201210017910.1 discloses a complex chloridizing metallurgical method for fully utilizing complex poor tin middlings, which adopts ferrous chloride as a reduction chlorinating agent and calcium chloride as an oxidation chlorinating agent; the comprehensive recovery method of pyrite cinder disclosed in Chinese patent CN200510021005.3 adopts ferric chloride and magnesium chloride as chlorinating agents; the Chinese patent CN201510401560.2 discloses a method and a device for extracting metal indium in LCD by chlorination and volatilization by taking PVC as a chlorinating agent, wherein the PVC is used as the chlorinating agent; chinese patent CN201210017933.2 discloses a method for producing iron pellets and enriching valuable metals by using high-impurity sulfuric acid residues, which adopts ferrous chloride as a chlorinating agent; the Chinese patent CN200410081722.0 discloses a chlorination treatment method of Se-containing substances, which adopts sodium chloride as a chlorinating agent; chinese patent CN201410757942.4 discloses a method for synchronously reducing and recovering gold and iron by chlorination roasting of gold concentrate cyanidation tailings, which adopts calcium chloride and sodium chloride as chlorinating agents and applies a biological oxidation method. However, the literature reports that "the oxidized waste residue generated after the zinc concentrate is subjected to pressure oxidation is subjected to resource comprehensive utilization" and chlorine-containing waste water is used as a chlorinating agent are not retrieved.
Therefore, the method for extracting the valuable metals in the zinc concentrate oxidizing slag by using the chlorine-containing wastewater is particularly urgent.
Disclosure of Invention
The invention aims to overcome the defects of the prior art, provides a method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater, can obviously improve the comprehensive utilization rate of the pressure oxidizing slag, solves the problems of valuable metal recycling and heavy metal pollution, and makes the chlorine-containing wastewater be recycled.
The task of the invention is completed by the following technical scheme:
the method for extracting the valuable metals from the zinc concentrate oxidizing slag by using the chlorine-containing wastewater sequentially comprises the following process steps and conditions:
(A) fine grinding-screening, namely crushing and fine grinding the zinc concentrate pressure oxidation slag (1) to screen out fine oxidation slag powder and an oxidation slag oversize material (2);
(B) stirring and granulating, namely adding 50-100 kg of binder (4) and 3-15 kg of reducing agent (3) into each ton of fine oxidation slag powder, uniformly mixing, adding 400-600L of chlorine-containing wastewater (5) into each ton of total mixed solid materials, stirring, mixing and granulating, wherein the molar mass ratio of the chlorine element added into each ton of oxidation slag (1) to the total molar mass of valuable metals of zinc, copper, cobalt and nickel is 1.10-1.50: 1, obtaining pellets;
(C) pre-drying, namely naturally airing and pre-drying the pellets, wherein the pre-drying temperature is 70-150 ℃, and the pre-drying time is 1.0-4.0 h, so as to obtain dried pellets and low-temperature flue gas (6);
(D1a) high-temperature roasting, namely roasting the dried pellets at 1050-1250 ℃ for 4-12 h to obtain high-temperature flue gas (13) containing valuable metals and roasting slag with the discharging temperature of 850-1150 ℃;
(D1b), stirring, adding the reducing material (14) into the roasting slag, and stirring to obtain reduced roasting slag;
(D1c) preserving heat, and preserving heat of the reduction roasting slag at the temperature of 550-750 ℃ for 1.0-6.0 h to obtain heat preservation materials and heat preservation flue gas;
(D1D) crushing and fine grinding, namely crushing and fine grinding the heat insulation material to obtain a ground heat insulation material;
(D1e) magnetic separation, wherein the levigated heat-insulating material is subjected to magnetic separation to obtain a product magnetite concentrate (15) and a silico-aluminum-calcium residue (16);
(D2a) leaching and collecting, and carrying out multi-stage spray washing on the low-temperature flue gas (6) pre-dried in the step (C) by using dilute hydrochloric acid leacheate (7) to obtain washing liquid containing valuable metal elements;
(D2b) heating and concentrating, namely heating and concentrating the washing liquid containing valuable metal elements, heating and concentrating at the temperature of 60-90 ℃, heating and concentrating for 5-30 min, and heating and recovering hydrogen chloride gas (8) by using heat collected by a heat exchange system to obtain washing liquid without the hydrogen chloride gas (8);
(D2c) recycling in sections, adding an alkaline reagent (9) into the washing liquid without the hydrogen chloride gas (8) for recycling in sections to obtain an iron-containing product (10), a copper-containing product (11), a cobalt-nickel product (12) and filtrate.
The percentages referred to in the specification are mass percentages, the temperature of the high-temperature flue gas is about 1000-1300 ℃, the temperature of the low-temperature flue gas is about 60-80 ℃, and the temperature of the heat-preservation flue gas is about 600-800 ℃.
Compared with the prior art, the invention has the following advantages and effects:
(1) valuable metals in the pressure oxidation waste residues are recycled, so that the utilization rate of zinc concentrate resources is improved, the occupied area of the waste residues is reduced, the service life of tailings storage capacity is prolonged, the potential heavy metal pollution risk of the oxidation waste residues is reduced, and better economic benefit can be obtained;
(2) the method adopts high temperature (1150 +/-100 ℃) for chlorination roasting, realizes the recovery of valuable metal elements such as iron, zinc, copper, cobalt and nickel in the slag, and has the advantages of small residual slag amount, low acid consumption, low cost, high recovery rate, environment-friendly production process and the like;
(3) the method realizes the environmental protection idea of treating wastes with wastes, successfully utilizes the chlorine-containing wastewater as a chlorinating agent of a chlorination volatilization method, adds the chlorine-containing wastewater into the oxidized waste residue for valuable metal recovery, can realize the recovery and utilization of valuable resources in the solid waste residue, reduces the treatment cost of the chlorine-containing wastewater difficult to treat, shortens the whole process flow, reduces the equipment investment and maintenance cost, and the like.
Drawings
FIG. 1 is a flow chart of a flotation process of the method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater.
In the drawings, the designations represent:
1. oxidizing slag 2, oxidizing slag oversize material 3, reducing agent 4, binder 5, chlorine-containing wastewater 6, low-temperature flue gas 7, dilute hydrochloric acid leacheate 8, hydrogen chloride gas 9, alkaline reagent 10, iron-containing product 11, copper-containing product 12, cobalt-nickel product 13, high-temperature flue gas 14, reducing material 15, magnetite concentrate 16, silicon-aluminum-calcium residue
The description is described in further detail below with reference to the accompanying drawings.
Detailed Description
As shown in figure 1, the method for extracting valuable metals from zinc concentrate oxidizing slag by using chlorine-containing wastewater of the invention is sequentially carried out according to the following process steps and conditions:
(A) fine grinding-screening, namely crushing and fine grinding the zinc concentrate pressure oxidation slag (1) to screen out fine oxidation slag powder and an oxidation slag oversize material (2);
(B) stirring and granulating, namely adding 50-100 kg of binder (4) and 3-15 kg of reducing agent (3) into each ton of fine oxidation slag powder, uniformly mixing, adding 400-600L of chlorine-containing wastewater (5) into each ton of total mixed solid materials, stirring, mixing and granulating, wherein the molar mass ratio of the chlorine element added into each ton of oxidation slag (1) to the total molar mass of valuable metals of zinc, copper, cobalt and nickel is 1.10-1.50: 1, obtaining pellets;
(C) pre-drying, namely naturally airing and pre-drying the pellets, wherein the pre-drying temperature is 70-150 ℃, and the pre-drying time is 1.0-4.0 h, so as to obtain dried pellets and low-temperature flue gas (6);
(D1a) high-temperature roasting, namely roasting the dried pellets at 1050-1250 ℃ for 4-12 h to obtain high-temperature flue gas (13) containing valuable metals and roasting slag with the discharging temperature of 850-1150 ℃;
(D1b), stirring, adding the reducing material (14) into the roasting slag, and stirring to obtain reduced roasting slag;
(D1c) preserving heat, and preserving heat of the reduction roasting slag at the temperature of 550-750 ℃ for 1.0-6.0 h to obtain heat preservation materials and heat preservation flue gas;
(D1D) crushing and fine grinding, namely crushing and fine grinding the heat insulation material to obtain a ground heat insulation material;
(D1e) magnetic separation, wherein the levigated heat-insulating material is subjected to magnetic separation to obtain a product magnetite concentrate (15) and a silico-aluminum-calcium residue (16);
(D2a) leaching and collecting, and carrying out multi-stage spray washing on the low-temperature flue gas (6) pre-dried in the step (C) by using dilute hydrochloric acid leacheate (7) to obtain washing liquid containing valuable metal elements;
(D2b) heating and concentrating, namely heating and concentrating the washing liquid containing valuable metal elements, heating and concentrating at the temperature of 60-90 ℃, heating and concentrating for 5-30 min, and heating and recovering hydrogen chloride gas (8) by using heat collected by a heat exchange system to obtain washing liquid without the hydrogen chloride gas (8);
(D2c) recycling in sections, adding an alkaline reagent (9) into the washing liquid without the hydrogen chloride gas (8) for recycling in sections to obtain an iron-containing product (10), a copper-containing product (11), a cobalt-nickel product (12) and filtrate.
The process of the invention may further be:
and (D1a) returning the high-temperature flue gas (13) containing valuable metals roasted at high temperature in the step (C) to be pre-dried for recycling.
And (D1C) returning one part of the heat-preservation flue gas subjected to heat preservation in the step (C) to be pre-dried, and sending the other part of the heat-preservation flue gas to the step (D2b) for heating and concentrating, wherein the residual heat is recycled.
And (D2b) heating the concentrated hydrogen chloride gas (8) and returning to the step (D2a) for rinsing, collecting and recycling.
And (C) returning part of the filtrate recycled in the step (D2c) to the step (D2a) for leaching and collection after hydrogen chloride escapes at the temperature of 40-80 ℃, and returning the other part of the filtrate to the step (B) for stirring-granulating for recycling.
And (C) returning the oversize material (2) of the oxidation slag obtained by fine grinding-sieving in the step (A) to fine grinding for recycling.
The adhesive (4) for stirring and granulating in the step (B) is bentonite.
And (B) stirring-granulating the chlorine-containing wastewater (5) in the step (B) refers to any one or combination of electrodialysis produced water, hydrochloric acid wastewater and chlorine-containing brine, and the mass concentration of chloride ions is 5-300 g/L.
The reducing agent (3) for stirring and granulating in the step (B) is any one or combination of anthracite powder and blast furnace powder with the grain diameter of 0.01-1.00 mm accounting for 50-95%.
And (D2a) the dilute acid leacheate (7) collected by leaching in the step (D2a) is dilute hydrochloric acid, and the mass concentration of the hydrochloric acid is 1-15%.
And (D2a) washing and collecting, wherein the washing grade number is 3-8, and the retention time of each grade is 5-30 min.
The reducing material (14) stirred in the step (D1b) is any one or a combination of anthracite powder, activated carbon particles and activated carbon powder with the granularity of 0.01-2 mm accounting for 40-80%.
The following further describes embodiments of the present invention with reference to specific examples.
The chemical composition (wt%) of the copper tailings used in the implementation is as follows: 0.10 percent of copper, 3.64 percent of sulfur, 0.11g/t of gold and 5.2g/t of silver. The sulfur concentrate is obtained by the technological process of 'two times of roughing, one time of scavenging, regrinding of rough concentrate, two times of concentration and sequential return of middlings', and the ultra-pure sulfur concentrate and the gold-bearing copper common sulfur concentrate are obtained by carrying out copper-sulfur separation on the sulfur concentrate under the weak acid condition.
Example 1
The zinc concentrate pressure oxidation waste residue comprises the following main components: 1.5 percent of zinc, 0.18 percent of copper, 2.1 percent of lead, 127g/t of nickel and 25g/t of cobalt. And crushing, ball-milling and fine-milling the oxidized waste residue to obtain 81% of fineness of-200 meshes.
Taking 100kg of oxidation waste residue, adding 3% of coal powder, then adding 25L of chlorine-containing wastewater, mixing, stirring, pelletizing, pre-drying, then feeding into a first-stage preheating furnace, roasting at 130 ℃, keeping the temperature for 2h, then feeding into a second-stage high-temperature roasting furnace, roasting at 1150 ℃, roasting for 8h, recovering 91% of copper, 92.5% of zinc, 86% of nickel, 90% of cobalt, 88% of silver and 88% of iron.
Example 2
The zinc concentrate pressure oxidation waste residue comprises the following main components: 1.05 percent of zinc, 0.29 percent of copper, 2.13 percent of lead, 115g/t of nickel and 36g/t of cobalt. And crushing, ball-milling and fine-milling the oxidized waste residue to obtain the oxidized waste residue with the fineness of-200 meshes accounting for 89% of the total ore sample.
Adding 3% of coal powder according to the mass ratio of the oxidized waste residues, uniformly mixing, adding 40L of chlorine-containing wastewater with the concentration of chloride ions of 61g/L, mixing, stirring, pelletizing, pre-drying and drying, then feeding the mixture into a first-stage preheating furnace, roasting at the temperature of 100 ℃ for 4h, and then feeding the mixture into a second-stage high-temperature roasting furnace at the roasting temperature of 1250 ℃ for 24h, wherein the copper recovery rate is 93%, the zinc recovery rate is 95.5%, the nickel recovery rate is 91%, the cobalt recovery rate is 93%, the silver recovery rate is 93% and the iron recovery rate is 91.5%.
Example 3
The zinc concentrate pressure oxidation waste residue comprises the following main components: 1.16% of zinc, 0.21% of copper, 1.75% of lead, 112g/t of nickel and 45g/t of cobalt. And crushing, ball milling and fine milling the oxidized waste residue to obtain a fineness of-200 meshes accounting for 75% of the total ore sample.
Adding 5% of coal powder according to the mass ratio of the oxidized waste residues, uniformly mixing, adding 20L of chlorine-containing wastewater with the concentration of chloride ions of 107g/L, mixing, stirring, pelletizing, pre-drying, feeding into a first-stage preheating furnace at the temperature of 180 ℃, roasting for 4 hours, feeding into a second-stage high-temperature roasting furnace at the roasting temperature of 1050 ℃, roasting for 12 hours, wherein the copper recovery rate is 85.4%, the zinc recovery rate is 90.7%, the nickel recovery rate is 84.9%, the cobalt recovery rate is 84%, the silver recovery rate is 82.9%, and the iron recovery rate is 85.5%.
As described above, the present invention can be preferably realized. The above embodiments are only preferred embodiments of the present invention, but the embodiments of the present invention are not limited by the above embodiments, and other changes, modifications, substitutions, combinations, and simplifications which do not depart from the spirit and principle of the present invention should be regarded as equivalent replacements within the protection scope of the present invention.
Claims (12)
1. The method for extracting the valuable metals from the zinc concentrate oxidizing slag by using the chlorine-containing wastewater is characterized by sequentially carrying out the following process steps and conditions:
(A) fine grinding-screening, namely crushing and fine grinding the zinc concentrate pressure oxidation slag (1) to screen out fine oxidation slag powder and an oxidation slag oversize material (2);
(B) stirring and granulating, namely adding 50-100 kg of binder (4) and 3-15 kg of reducing agent (3) into each ton of fine oxidation slag powder, uniformly mixing, adding 400-600L of chlorine-containing wastewater (5) into each ton of total mixed solid materials, stirring, mixing and granulating, wherein the molar mass ratio of the chlorine element added into each ton of oxidation slag (1) to the total molar mass of valuable metals of zinc, copper, cobalt and nickel is 1.10-1.50: 1, obtaining pellets;
(C) pre-drying, namely naturally airing and pre-drying the pellets, wherein the pre-drying temperature is 70-150 ℃, and the pre-drying time is 1.0-4.0 h, so as to obtain dried pellets and low-temperature flue gas (6);
(D1a) high-temperature roasting, namely roasting the dried pellets at 1050-1250 ℃ for 4-12 h to obtain high-temperature flue gas (13) containing valuable metals and roasting slag with the discharging temperature of 850-1150 ℃;
(D1b), stirring, adding the reducing material (14) into the roasting slag, and stirring to obtain reduced roasting slag;
(D1c) preserving heat, and preserving heat of the reduction roasting slag at the temperature of 550-750 ℃ for 1.0-6.0 h to obtain heat preservation materials and heat preservation flue gas;
(D1D) crushing and fine grinding, namely crushing and fine grinding the heat insulation material to obtain a ground heat insulation material;
(D1e) magnetic separation, wherein the levigated heat-insulating material is subjected to magnetic separation to obtain a product magnetite concentrate (15) and a silico-aluminum-calcium residue (16);
(D2a) leaching and collecting, and carrying out multi-stage spray washing on the low-temperature flue gas (6) pre-dried in the step (C) by using dilute hydrochloric acid leacheate (7) to obtain washing liquid containing valuable metal elements;
(D2b) heating and concentrating, namely heating and concentrating the washing liquid containing valuable metal elements, heating and concentrating at the temperature of 60-90 ℃, heating and concentrating for 5-30 min, and heating and recovering hydrogen chloride gas (8) by using heat collected by a heat exchange system to obtain washing liquid without the hydrogen chloride gas (8);
(D2c) recycling in sections, adding an alkaline reagent (9) into the washing liquid without the hydrogen chloride gas (8) for recycling in sections to obtain an iron-containing product (10), a copper-containing product (11), a cobalt-nickel product (12) and filtrate.
2. The method as set forth in claim 1, characterized in that the high-temperature flue gas (13) containing valuable metals calcined at high temperature in the step (D1a) is returned to the step (C) for pre-drying cycle.
3. The method as claimed in claim 1, wherein a part of the heat-insulating flue gas obtained in the step (D1C) is returned to the step (C) for pre-drying, and the other part is sent to the step (D2b) for heating and concentrating, and the residual heat is recycled.
4. The method as set forth in claim 1, characterized in that said step (D2b) of heating the concentrated hydrogen chloride gas (8) is returned to the step (D2a) of rinsing and collecting for recycling.
5. The method as set forth in claim 1, characterized in that a part of the filtrate recovered in the step (D2c) by stages is returned to the step (D2a) for leaching and collection after hydrogen chloride escapes from the filtrate at the temperature of 40-80 ℃, and the other part is returned to the step (B) for stirring-granulating for recycling.
6. The method as set forth in claim 1, characterized in that the fine ground-sieved material (2) of the oxidizing slag sieved in the step (a) is returned to the fine grinding for reuse.
7. The method according to claim 1, wherein the binder (4) for the stirring-granulation in step (B) is bentonite.
8. The method according to claim 1, wherein the chlorine-containing wastewater (5) obtained by the agitation-granulation in the step (B) is any one or a combination of electrodialysis produced water, hydrochloric acid wastewater, chlorine-containing brine, and has a mass concentration of chlorine ions of 5 to 300 g/L.
9. The method according to claim 1 or 7 or 8, wherein the stirring-granulating reducing agent (3) in step (B) is any one of anthracite coal powder and blast furnace coal powder or a combination thereof with a particle size of 0.01 to 1.00mm in proportion of 50 to 95%.
10. The method as claimed in claim 1, characterized in that the diluted acid eluent (7) collected in the step (D2a) is diluted hydrochloric acid, and the mass concentration of the hydrochloric acid is 1-15%.
11. The method as set forth in claim 1 or 10, characterized in that the rinsing collected in the step (D2a) is carried out in the number of stages of 3 to 8 and the retention time of each stage is 5 to 30 min.
12. The method according to claim 1, wherein the reducing material (14) stirred in step (D1b) is any one or a combination of anthracite powder, activated carbon particles, activated carbon powder with a particle size of 0.01-2 mm and 40-80%.
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