CN111926196A - Method for recovering zinc from smelting waste residues - Google Patents
Method for recovering zinc from smelting waste residues Download PDFInfo
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- CN111926196A CN111926196A CN202010815645.6A CN202010815645A CN111926196A CN 111926196 A CN111926196 A CN 111926196A CN 202010815645 A CN202010815645 A CN 202010815645A CN 111926196 A CN111926196 A CN 111926196A
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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
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/24—Obtaining zinc otherwise than by distilling with leaching with alkaline solutions, e.g. ammonia
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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
- C22B19/26—Refining solutions containing zinc values, e.g. obtained by leaching zinc ores
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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
- C22B19/34—Obtaining zinc oxide
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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
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- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
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- Life Sciences & Earth Sciences (AREA)
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Abstract
The invention discloses a method for recovering zinc from smelting waste residue, which comprises the steps of uniformly mixing the smelting waste residue with calcium hypochlorite, and roasting at high temperature to obtain calcine; uniformly mixing one complexing agent of citric acid or citrate with one complexing agent of tartaric acid or tartrate to obtain a mixed mixture, adding water, uniformly stirring, adding an ammonia water solution and ammonium sulfate, and uniformly mixing to form a coordination leaching agent; placing the obtained calcine in a coordination leaching agent for ultrasonic reaction to obtain a zinc leaching solution; adding zinc powder into the obtained leachate, reacting for a period of time, filtering, and calcining the obtained filtrate at high temperature to obtain the zinc oxide. The invention ensures the leaching rate of zinc in the smelting waste residue and prevents impurity ions from leaching, so that the obtained zinc leaching solution can directly obtain a high-purity zinc oxide product after high-temperature calcination, and the application of zinc recovery from the smelting waste residue in industry is facilitated.
Description
Technical Field
The invention relates to the technical field of metallurgy, in particular to a method for recovering zinc from smelting waste residues.
Background
Besides zinc mineral resources, metallurgical slag dust such as zinc-containing slag dust and the like are also extremely important utilizable resources in China, and mainly comprise zinc-containing gas mud and gas ash generated in the blast furnace smelting process, zinc-containing smoke dust generated in the electric arc furnace smelting process, zinc-containing smoke dust generated after lead smelting slag volatilization, zinc smelting slag obtained by a wet method and the like; the zinc hydrometallurgy accounts for more than 80 percent of the total zinc hydrometallurgy in the world, and the process generates a large amount of zinc-containing leaching residues, wherein the zinc-containing materials are secondary zinc oxide containing various impurities such as iron (up to 14 percent), calcium (up to 19 percent), chlorine (up to 12 percent), fluorine (up to 2 percent), and the like, zinc mainly exists in the forms of zinc oxide, zinc-iron spinel and zinc silicate, iron mainly exists in the forms of ferroferric oxide and zinc-iron spinel, and calcium mainly exists in the forms of calcium carbonate and silicate containing iron, zinc and calcium. Therefore, as with zinc mineral resources, effective utilization of zinc-containing waste residues, environmental stress reduction, and resource utilization rate improvement are problems that need to be solved at present.
In the prior art, an alkaline leaching method is an effective method for recovering zinc from smelting waste residues, mainly uses ammonia or ammonium salt water solution as a leaching agent to extract zinc or zinc compounds, and the publication number CN104975180A discloses a method and a device for leaching blast furnace gas ash by an ultrasonic-microwave combined ammonia method, so that the leaching rate of zinc is more than 87.4 percent, and the recovery and utilization of zinc in the zinc-containing waste residues are greatly promoted. However, in the prior art, the alkaline leaching method for preparing the zinc leaching solution leaches zinc ions, and simultaneously leaches other metal impurity ions in the smelting waste residues, such as copper, lead, iron, manganese and other ions, so that the zinc leaching solution obtained by ammonia leaching usually needs further purification and impurity removal operation to remove the impurity ions to obtain a high-purity zinc product, which inevitably increases the step of recovering zinc from the smelting waste residues, and influences the application of the recovered zinc in the smelting waste residues in the industry.
Disclosure of Invention
In order to solve the technical problems, the invention provides a method for recovering zinc from smelting waste residues, which can ensure the leaching rate of zinc in the smelting waste residues and prevent impurity ions from leaching, so that the obtained zinc leachate can be directly calcined at high temperature to obtain a high-purity zinc oxide product, and the zinc can be conveniently recovered from the smelting waste residues in the industry.
The invention provides a method for recovering zinc from smelting waste residues, which comprises the following steps:
(1) uniformly mixing the smelting waste residue and calcium hypochlorite, and roasting at high temperature to obtain calcine;
(2) uniformly mixing one complexing agent of citric acid or citrate and one complexing agent of tartaric acid or tartrate according to the molar ratio of (1-30) to obtain a mixed mixture, adding water, uniformly stirring, adding an ammonia water solution and ammonium sulfate, and uniformly mixing to form a coordination leaching agent;
(3) placing the calcine obtained in the step (1) in the coordination leaching agent prepared in the step (2) for ultrasonic reaction to obtain a zinc leaching solution;
(4) adding zinc powder into the obtained leachate, reacting for a period of time, filtering, crystallizing, drying and calcining at high temperature to obtain the zinc oxide.
Preferably, in the step (1), the addition amount of the calcium hypochlorite is 5-10% of the mass of the smelting waste residue, the high-temperature roasting is microwave roasting, the microwave heating rate is 10-20 ℃/min, the roasting temperature is 300-.
Preferably, in the step (2), the concentration of citric acid or citrate in the coordination leaching agent is 0.05-1.5mol/L, the concentration of tartaric acid or tartrate is 0.05-1.5mol/L, the molar ratio of the ammonia water solution to the ammonium sulfate is (1-10): (1-10), the total ammonia concentration in the coordination leaching agent is 3-8mol/L, and the pH value is 7-10.
Preferably, the solid-to-liquid ratio of the calcine and the coordination leaching agent in the step (3) is 1.5-10g/ml, the ultrasonic reaction temperature is normal temperature, the ultrasonic power is 100-1000W, and the leaching time is 10-30 min.
Preferably, the material of the zinc powder in the step (4) is 1-3g/L, the reaction temperature is 80-90 ℃, and the reaction time is 1-2 h.
Preferably, in the step (4), the calcination is inert atmosphere microwave calcination, the calcination temperature is 100-300 ℃, and the calcination time is 10-30 min.
Preferably, ammonia distillation is performed before the calcination in the step (4), and ammonia in the ammonia distillation is recovered.
Compared with the prior art, the invention has the following beneficial effects:
(1) under an ammoniacal leaching system, ferrous iron and divalent manganese can complex with ammonia to form coordination ions to enter a leaching solution, and ferric iron and trivalent manganese cannot complex with ammonia to form coordination ions, so that before leaching, an oxidation pretreatment mode is adopted, calcium hypochlorite is selected as an oxidant, high-temperature roasting is carried out, ferrous iron and divalent manganese in waste residues can be oxidized into a form of ferric iron and trivalent manganese, and re-leaching of the waste residues in a subsequent ammonia process basic process is avoided; further, hypochlorous acid and carbon dioxide in the air generate calcium carbonate in the roasting process, and the calcium carbonate further reacts with zinc silicate and zinc ferrite which cannot be leached and dissolved in the ammonia leaching process, so that the zinc silicate and the zinc ferrite which are difficult to leach are converted into zinc oxide; meanwhile, calcium hypochlorite is adopted for pretreatment before ammonia leaching, the reaction between the hypochlorous acid and ammonia is avoided, and the consumption of ammonia water in the ammonia leaching process is reduced. Through one-step reaction, the dissolution of impurity ions in the waste residue is avoided on the premise of realizing the high-efficiency dissolution of zinc in the waste residue.
(2) The waste residue is preheated under the microwave condition, so that the surface layer of the waste residue has an impact destruction effect, the coating on the surface of the solid is favorably cracked and scattered, the zinc reaction interface in the waste residue is exposed, and the subsequent zinc leaching reaction is favorably realized.
(3) And (3) carrying out ammonia leaching reaction on the pretreated waste residue in an alkaline environment, so that impurities, calcium ions and the like introduced in the step (1) can be generated into calcium hydroxide, and the calcium hydroxide is prevented from being immersed into the leaching solution in the process. Due to Zn2+Is stronger Lewis acid and is easy to react with biological ligand, and citrate and tartrate are stronger Lewis base, and carboxyl of the Lewis base can provide electron pair, so that Zn2+The zinc ion complex forms a complex with citrate (tartrate) to be dissolved, other impurities are not dissolved, the mutual contact between complex ions and the generation of the complex are facilitated by leaching under the ultrasonic condition, the coordination efficiency is improved, the leaching rate of zinc ions is accelerated, the leaching temperature is reduced, and the energy consumption is saved.
(4) And adding zinc powder into the leachate, and performing a displacement reaction with copper, lead, cadmium and other ions leached from the leachate, so as to further remove impurity ions in the leachate and improve the purity of the calcined product zinc oxide.
Detailed Description
Reference will now be made in detail to various exemplary embodiments of the invention, the detailed description should not be construed as limiting the invention but as a more detailed description of certain aspects, features and embodiments of the invention.
It is to be understood that the terminology used herein is for the purpose of describing particular embodiments only and is not intended to be limiting of the invention. Further, for numerical ranges in this disclosure, it is understood that each intervening value, between the upper and lower limit of that range, is also specifically disclosed. Every smaller range between any stated value or intervening value in a stated range and any other stated or intervening value in a stated range is encompassed within the invention. The upper and lower limits of these smaller ranges may independently be included or excluded in the range.
Unless defined otherwise, all technical and scientific terms used herein have the same meaning as commonly understood by one of ordinary skill in the art to which this invention belongs. Although only preferred methods and materials are described herein, any methods and materials similar or equivalent to those described herein can be used in the practice or testing of the present invention. All documents mentioned in this specification are incorporated by reference herein for the purpose of disclosing and describing the methods and/or materials associated with the documents. In case of conflict with any incorporated document, the present specification will control.
It will be apparent to those skilled in the art that various modifications and variations can be made in the specific embodiments of the present disclosure without departing from the scope or spirit of the disclosure. Other embodiments will be apparent to those skilled in the art from consideration of the specification. The specification and examples are exemplary only.
As used herein, the terms "comprising," "including," "having," "containing," and the like are open-ended terms that mean including, but not limited to.
Example 1
(1) Mixing the smelting waste residue with calcium hypochlorite, wherein the addition amount of the calcium hypochlorite accounts for 10% of the mass of the smelting waste residue, placing the mixture in a microwave reactor, heating to 300 ℃ at the heating rate of 20 ℃/min, roasting for 20min, and taking out;
(2) uniformly mixing citric acid and tartaric acid according to a molar ratio of 1: 1 to obtain a mixed mixture, adding water until the concentration of the citric acid is 1mol/L, adding ammonia water and ammonium sulfate according to a molar ratio of 1: 1 to ensure that the total ammonia concentration in the solution is 5mol/L, and adjusting the pH value to 8 to obtain a coordination leaching agent;
(3) uniformly mixing the roasted product in the step (1) and the coordination leaching agent in the step (2) according to a solid-to-liquid ratio of 10g/ml, and leaching at normal temperature (25 ℃) for 20min under 1000W of ultrasonic power to obtain a leaching solution;
(4) adding zinc powder into the obtained leachate according to the using amount of 1g/L, stirring and reacting for 1h at 85 ℃, filtering, performing ammonia evaporation treatment on the obtained filtrate, crystallizing, drying, and performing inert atmosphere microwave calcination at the calcination temperature of 200 ℃ for 10min to obtain the zinc oxide.
The ammonia distillation step specifically comprises the following steps: and introducing the filtrate into an ammonia still, distilling ammonia at 100 ℃ for 1h, filtering, and performing microwave drying on the obtained filter cake at 100 ℃ for 20min to obtain the product.
Example 2
The difference from example 1 is that the amount of calcium hypochlorite added is 20% by mass of the slag.
Example 3
The difference from example 1 is that the amount of calcium hypochlorite added is 3% by mass of the slag.
Example 4
The same as example 1 except that calcium hypochlorite was replaced with an equal amount of calcium carbonate.
Example 5
The difference from example 1 is that the temperature increase rate in step (1) was 5 ℃/min.
Example 6
The difference from example 1 is that the temperature increase rate in step (1) was 30/min.
Example 7
The difference from example 1 is that citric acid and tartaric acid were not added in step (2).
Example 8
The difference from example 1 is that the reaction process in step (3) is not sonicated.
Example 9
The difference from example 1 is that the firing process in step (1) is not carried out in a microwave environment.
Example 10
The difference from example 1 is that step (1) is omitted.
Example 11
The difference from example 1 is that zinc powder was not added to the zinc leachate.
Effect verification
The leaching rates and the purity of the zinc oxide after calcination in the leachate treated in examples 1 to 10 were measured, and the results are shown in Table 1.
TABLE 1
| Leaching rate% | Purity% | |
| Example 1 | 95.2 | 99.9 |
| Example 2 | 96.1 | 95.4 |
| Example 3 | 93.7 | 98.7 |
| Example 4 | 90.4 | 94.8 |
| Example 5 | 92.8 | 95.3 |
| Example 6 | 90.4 | 96.7 |
| Example 7 | 85.7 | 92.4 |
| Example 8 | 91.4 | 98.6 |
| Example 9 | 92.7 | 97.4 |
| Example 10 | 86.4 | 85.7 |
| Example 11 | 96.3 | 98.4 |
It can be seen from table 1 that the added amount of calcium hypochlorite, microwave environment, temperature rise rate, etc. greatly affect the purity of the final product and the leaching rate of zinc, and the analysis may be that when the added amount of calcium hypochlorite is too large, the concentration of introduced calcium ions is too high, which makes the subsequent reaction difficult to remove, thereby affecting the product purity, and when the added amount of calcium hypochlorite is too small, the conversion of zinc ferrite and other hard-to-convert opposite zinc oxide phases in the waste residue and the oxidation process requirements of manganese, iron, etc. as impurity ions therein cannot be satisfied. When the temperature rising rate is too fast or too slow, it is difficult to balance the reaction process of calcium hypochlorite, carbon dioxide and impurity ions therein, so that one of them does not react completely, and thus the optimal leaching rate and product purity cannot be achieved. And the omission of the microwave environment causes that substances in impurities cannot fully participate in the reaction process, thereby affecting the leaching rate and the product purity. The addition of citric acid and tartaric acid and ultrasonic treatment also greatly reduce the dissolution of impurity ions, so that the purity of the product is higher and the dissolution rate of zinc is higher.
The above description is only for the purpose of illustrating the preferred embodiments of the present invention and is not to be construed as limiting the invention, and any modifications, equivalents and improvements made within the spirit and principle of the present invention are intended to be included therein.
Claims (7)
1. A method for recovering zinc from smelting waste slag is characterized by comprising the following steps:
(1) uniformly mixing the smelting waste residue and calcium hypochlorite, and roasting at high temperature to obtain calcine;
(2) uniformly mixing one complexing agent of citric acid or citrate and one complexing agent of tartaric acid or tartrate according to the molar ratio of (1-30) to obtain a mixed mixture, adding water, uniformly stirring, adding an ammonia water solution and ammonium sulfate, and uniformly mixing to form a coordination leaching agent;
(3) placing the calcine obtained in the step (1) in the coordination leaching agent prepared in the step (2) for ultrasonic reaction to obtain a zinc leaching solution;
(4) adding zinc powder into the obtained leachate, reacting for a period of time, filtering, crystallizing, drying and calcining at high temperature to obtain the zinc oxide.
2. The method for recovering zinc from smelting slag as claimed in claim 1, wherein in step (1), the amount of calcium hypochlorite added is 5-10% of the mass of the smelting slag, the high-temperature roasting is microwave roasting, the microwave heating rate is 10-20 ℃/min, the roasting temperature is 300-400 ℃, and the roasting time is 10-30 min.
3. The method for recovering zinc from smelting slag according to claim 1, wherein in the step (2), the concentration of citric acid or citrate in the coordination leaching agent is 0.05-1.5mol/L, the concentration of tartaric acid or tartrate in the coordination leaching agent is 0.05-1.5mol/L, the molar ratio of the ammonia water solution to the ammonium sulfate is (1-10) to (1-10), the total ammonia concentration in the coordination leaching agent is 3-8mol/L, and the pH value is 7-10.
4. The method for recovering zinc from smelting slag as claimed in claim 1, wherein the solid-to-liquid ratio of the calcine and the coordination leaching agent in the step (3) is 1.5-10g/ml, the ultrasonic reaction temperature is normal temperature, the ultrasonic power is 100-1000W, and the leaching time is 10-30 min.
5. The method for recovering zinc from smelting waste residue according to claim 1, wherein the amount of zinc powder used in step (4) is 1-3g/L, the reaction temperature is 80-90 ℃, and the reaction time is 1-2 h.
6. The method for recovering zinc from smelting slag according to claim 1, wherein in the step (4), the calcination is inert atmosphere microwave calcination, the calcination temperature is 100-300 ℃, and the calcination time is 10-30 min.
7. The method for recovering zinc from smelting slag according to claim 1, wherein ammonia distillation is performed before calcination in step (4) to recover ammonia therefrom.
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Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN113528844A (en) * | 2021-07-15 | 2021-10-22 | 江西理工大学 | Method for efficiently separating valuable metals from copper regeneration soot |
| CN115109929A (en) * | 2022-01-21 | 2022-09-27 | 昆明理工大学 | Method for directly preparing coarse germanium dioxide from tannin germanium slag |
| CN115109925A (en) * | 2022-06-20 | 2022-09-27 | 东北大学 | Method for treating transition metal oxide ore by citric acid system |
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Cited By (5)
| Publication number | Priority date | Publication date | Assignee | Title |
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| CN113528844A (en) * | 2021-07-15 | 2021-10-22 | 江西理工大学 | Method for efficiently separating valuable metals from copper regeneration soot |
| CN115109929A (en) * | 2022-01-21 | 2022-09-27 | 昆明理工大学 | Method for directly preparing coarse germanium dioxide from tannin germanium slag |
| CN115109929B (en) * | 2022-01-21 | 2023-11-10 | 昆明理工大学 | Method for directly preparing crude germanium dioxide from single Ning Zhe slag |
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| CN115109925B (en) * | 2022-06-20 | 2024-05-17 | 东北大学 | Method for treating transition metal oxide ore by using citric acid system |
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