CN117965896A - Recycling treatment process of lead smelting anode slime later-stage slag - Google Patents
Recycling treatment process of lead smelting anode slime later-stage slag Download PDFInfo
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- CN117965896A CN117965896A CN202410383474.2A CN202410383474A CN117965896A CN 117965896 A CN117965896 A CN 117965896A CN 202410383474 A CN202410383474 A CN 202410383474A CN 117965896 A CN117965896 A CN 117965896A
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- 238000000034 method Methods 0.000 title claims abstract description 73
- 239000002893 slag Substances 0.000 title claims abstract description 64
- 238000003723 Smelting Methods 0.000 title claims abstract description 28
- 238000004064 recycling Methods 0.000 title claims description 8
- 238000002386 leaching Methods 0.000 claims abstract description 58
- 238000011084 recovery Methods 0.000 claims abstract description 50
- 238000005868 electrolysis reaction Methods 0.000 claims abstract description 21
- HEMHJVSKTPXQMS-UHFFFAOYSA-M Sodium hydroxide Chemical compound [OH-].[Na+] HEMHJVSKTPXQMS-UHFFFAOYSA-M 0.000 claims description 30
- 239000007788 liquid Substances 0.000 claims description 26
- 229910000510 noble metal Inorganic materials 0.000 claims description 22
- 239000003792 electrolyte Substances 0.000 claims description 15
- 239000002994 raw material Substances 0.000 claims description 15
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 claims description 14
- XLYOFNOQVPJJNP-UHFFFAOYSA-N water Substances O XLYOFNOQVPJJNP-UHFFFAOYSA-N 0.000 claims description 12
- 239000003638 chemical reducing agent Substances 0.000 claims description 11
- 239000010970 precious metal Substances 0.000 claims description 11
- 239000012535 impurity Substances 0.000 claims description 10
- 238000002156 mixing Methods 0.000 claims description 10
- 239000004568 cement Substances 0.000 claims description 8
- 239000000571 coke Substances 0.000 claims description 8
- 238000000926 separation method Methods 0.000 claims description 8
- 239000004566 building material Substances 0.000 claims description 7
- 229910000029 sodium carbonate Inorganic materials 0.000 claims description 7
- 239000007787 solid Substances 0.000 claims description 7
- 238000001035 drying Methods 0.000 claims description 6
- 238000005406 washing Methods 0.000 claims description 6
- 238000012216 screening Methods 0.000 claims description 5
- 238000001816 cooling Methods 0.000 claims description 4
- 229910021645 metal ion Inorganic materials 0.000 claims description 4
- 239000000203 mixture Substances 0.000 claims description 4
- 238000001556 precipitation Methods 0.000 claims description 4
- 238000002791 soaking Methods 0.000 claims description 3
- 239000002699 waste material Substances 0.000 abstract description 9
- 239000000463 material Substances 0.000 abstract description 3
- 238000006243 chemical reaction Methods 0.000 abstract 1
- 230000010354 integration Effects 0.000 abstract 1
- 238000005457 optimization Methods 0.000 abstract 1
- KDLHZDBZIXYQEI-UHFFFAOYSA-N Palladium Chemical compound [Pd] KDLHZDBZIXYQEI-UHFFFAOYSA-N 0.000 description 24
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 description 15
- 229910052797 bismuth Inorganic materials 0.000 description 15
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 description 15
- 229910052709 silver Inorganic materials 0.000 description 15
- 239000004332 silver Substances 0.000 description 15
- 229910052763 palladium Inorganic materials 0.000 description 11
- 238000006722 reduction reaction Methods 0.000 description 10
- 238000004519 manufacturing process Methods 0.000 description 9
- 239000002253 acid Substances 0.000 description 8
- 235000011121 sodium hydroxide Nutrition 0.000 description 8
- 238000009853 pyrometallurgy Methods 0.000 description 7
- 239000003795 chemical substances by application Substances 0.000 description 6
- 239000002245 particle Substances 0.000 description 5
- 230000000694 effects Effects 0.000 description 4
- 230000007613 environmental effect Effects 0.000 description 4
- 229910052751 metal Inorganic materials 0.000 description 4
- DGAQECJNVWCQMB-PUAWFVPOSA-M Ilexoside XXIX Chemical compound C[C@@H]1CC[C@@]2(CC[C@@]3(C(=CC[C@H]4[C@]3(CC[C@@H]5[C@@]4(CC[C@@H](C5(C)C)OS(=O)(=O)[O-])C)C)[C@@H]2[C@]1(C)O)C)C(=O)O[C@H]6[C@@H]([C@H]([C@@H]([C@H](O6)CO)O)O)O.[Na+] DGAQECJNVWCQMB-PUAWFVPOSA-M 0.000 description 3
- VYPSYNLAJGMNEJ-UHFFFAOYSA-N Silicium dioxide Chemical compound O=[Si]=O VYPSYNLAJGMNEJ-UHFFFAOYSA-N 0.000 description 3
- 238000005363 electrowinning Methods 0.000 description 3
- 238000003912 environmental pollution Methods 0.000 description 3
- 150000002500 ions Chemical class 0.000 description 3
- 239000002184 metal Substances 0.000 description 3
- 150000002739 metals Chemical class 0.000 description 3
- 238000002360 preparation method Methods 0.000 description 3
- 238000000746 purification Methods 0.000 description 3
- 238000007670 refining Methods 0.000 description 3
- 229910052708 sodium Inorganic materials 0.000 description 3
- 239000011734 sodium Substances 0.000 description 3
- 238000012360 testing method Methods 0.000 description 3
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 description 2
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 description 2
- 238000010521 absorption reaction Methods 0.000 description 2
- 239000003513 alkali Substances 0.000 description 2
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- 238000000605 extraction Methods 0.000 description 2
- 230000002349 favourable effect Effects 0.000 description 2
- 229910000464 lead oxide Inorganic materials 0.000 description 2
- PIJPYDMVFNTHIP-UHFFFAOYSA-L lead sulfate Chemical compound [PbH4+2].[O-]S([O-])(=O)=O PIJPYDMVFNTHIP-UHFFFAOYSA-L 0.000 description 2
- WABPQHHGFIMREM-UHFFFAOYSA-N lead(0) Chemical compound [Pb] WABPQHHGFIMREM-UHFFFAOYSA-N 0.000 description 2
- YEXPOXQUZXUXJW-UHFFFAOYSA-N oxolead Chemical compound [Pb]=O YEXPOXQUZXUXJW-UHFFFAOYSA-N 0.000 description 2
- 238000012545 processing Methods 0.000 description 2
- 239000000126 substance Substances 0.000 description 2
- 235000008733 Citrus aurantifolia Nutrition 0.000 description 1
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 description 1
- 235000011941 Tilia x europaea Nutrition 0.000 description 1
- HCHKCACWOHOZIP-UHFFFAOYSA-N Zinc Chemical compound [Zn] HCHKCACWOHOZIP-UHFFFAOYSA-N 0.000 description 1
- 229910052782 aluminium Inorganic materials 0.000 description 1
- XAGFODPZIPBFFR-UHFFFAOYSA-N aluminium Chemical compound [Al] XAGFODPZIPBFFR-UHFFFAOYSA-N 0.000 description 1
- 239000002585 base Substances 0.000 description 1
- 230000002457 bidirectional effect Effects 0.000 description 1
- 239000006227 byproduct Substances 0.000 description 1
- 238000004364 calculation method Methods 0.000 description 1
- 239000003518 caustics Substances 0.000 description 1
- 239000003153 chemical reaction reagent Substances 0.000 description 1
- 239000002817 coal dust Substances 0.000 description 1
- 238000011109 contamination Methods 0.000 description 1
- 229910052802 copper Inorganic materials 0.000 description 1
- 239000010949 copper Substances 0.000 description 1
- 238000007405 data analysis Methods 0.000 description 1
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- 230000003009 desulfurizing effect Effects 0.000 description 1
- 238000011161 development Methods 0.000 description 1
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Classifications
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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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- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention discloses a recovery treatment process of lead smelting anode slime later-stage slag, which realizes the optimized integration of the whole process from pretreatment, leaching and electrolysis to waste slag utilization. All the procedures are buckled, the characteristics and reaction mechanism of materials are fully considered, and the overall recovery efficiency is improved to the greatest extent. Our integrated optimization scheme has higher technical advances and economic feasibility than conventional routes employing a single process or simple series connection.
Description
Technical Field
The invention relates to the technical field of metallurgy, in particular to a recovery treatment process for lead smelting anode slime later-stage slag.
Background
Lead anode slime is a byproduct in the lead electrolysis process and mainly comes from the production process of lead refining and secondary lead. During the electrowinning of lead, lead metal is deposited on the cathode, while impurity metals and certain oxides of lead form anode sludge on the anode. The anode mud contains a large amount of valuable metals such as lead, silver, bismuth and the like, and has higher recovery value.
However, the treatment and recovery of lead anode slime has been a technical challenge. Because the anode slime contains complex components and various metal elements and oxides, the treatment process needs to comprehensively consider the recovery rate, the cost and the environmental factors. In the prior art, the method for treating the lead anode slime mainly comprises pyrometallurgy, wet smelting, bioleaching and the like. The pyrometallurgy can realize the recovery of lead, but has high energy consumption and serious environmental pollution; in the wet smelting process, the selection of the leaching agent and the control of the leaching condition are difficult, so that the treatment efficiency is low; although the bioleaching method is environment-friendly, the treatment period is long, and the method is not suitable for mass production.
In addition, the prior art has the problem of low resource recovery rate. The recovery rate of valuable metals, especially noble metals such as silver, bismuth, etc., is not high due to the selection of the treatment method and the improper control of the treatment conditions. This not only wastes valuable resources, but can also create a potential hazard to the environment.
In addition, environmental pollution problems are also a major problem in the prior art. Waste gas and waste residue generated in the pyrometallurgy process cause serious pollution to the environment; the waste liquid produced in the wet smelting process contains a large amount of harmful substances, and the improper treatment can cause serious influence on the environment. These environmental pollution problems not only damage the ecological environment but also may cause harm to human health.
Finally, the difficulty in tailings disposal is also an important issue in the prior art. In the prior art, the tailings are treated by a single method, and most of the tailings are buried or piled up, so that not only are land resources occupied, but also potential safety hazards exist. The existence of the problems severely restricts the efficient and environment-friendly recycling and utilization of the lead anode slime. Therefore, developing a processing method with high efficiency, environmental protection and comprehensive utilization of resources has important significance for solving the problems.
Disclosure of Invention
In order to solve or partially solve the problems existing in the related art, the invention provides a recovery treatment process of lead smelting anode slime later-stage slag.
The method comprises the following steps:
(1) Crushing and screening the lead anode slime later-stage slag to a grain size of 2-10mm;
(2) Reducing roasting the crushed and screened slag at 700-1000 ℃ for 2-5 hours, wherein a reducing agent accounting for 1-10% of the mass of the slag is added in the reducing roasting process;
(3) Cooling the slag after reduction roasting to 500-800 ℃, adding sodium hydroxide accounting for 5-20% of the mass of the slag, and fully mixing;
(4) Soaking the mixture obtained in the step (3) in water for 1-5 hours under the conditions that the liquid-solid ratio is 3-10 and the temperature is 40-90 ℃;
(5) Carrying out solid-liquid separation on the leaching solution and leaching slag, wherein the leaching solution enters an electrolysis process, electrolysis is carried out under the condition that the cathode current density is 100-500A/m, and anode mud containing noble metals is obtained through anode enrichment in the electrolysis process;
(6) Purifying the electrolyte tail liquid, adding sodium carbonate accounting for 1-5% of the mass of the tail liquid into the tail liquid, and recycling the electrolyte after removing impurity metal ions by precipitation;
(7) Washing and drying the leaching slag obtained in the step (5), and mixing the leaching slag with cement raw materials or preparing building material aggregates.
Further, the particle diameter in the step (1) is 3-8mm.
Further, the reduction roasting temperature in the step (2) is 800-900 ℃ and the time is 2-3 hours.
Further, the reducing agent in the step (2) is coke, and the adding amount is 3-8% of the mass of slag.
Further, the adding amount of the sodium hydroxide in the step (3) is 8-15% of the mass of the slag.
Further, the water leaching temperature in the step (4) is 60-80 ℃, the liquid-solid ratio is 5:1-8:1, and the leaching time is 1.5-3 hours.
Further, the cathode current density in the step (5) is 200-400A/m.
Further, the adding amount of the sodium carbonate in the step (6) is 2-4% of the mass of the tail liquid.
Further, the mixing amount of the leaching slag obtained in the step (7) to the cement raw material is 5-20%, or the prepared building material aggregate is light aggregate.
Further, the anode slime rich in noble metals obtained in the step (5) is used as a raw material for noble metal smelting after washing and drying.
It is to be understood that both the foregoing general description and the following detailed description are exemplary and explanatory only and are not restrictive of the invention as claimed.
The beneficial technical effects of the invention are as follows:
1. Particle size control: the process of the invention particularly emphasizes the control of the particle size of the raw materials, and the anode slime later-stage slag is crushed and sieved to the particle size range of 2-10 mm. This step not only helps to improve the efficiency and uniformity of the roasting and leaching process, but also provides convenience for subsequent industrial production.
2. And (3) reduction roasting: our process performs the reduction roasting at 800-900 ℃ and adds a certain amount of coke or coal dust as reducing agent. This step can effectively reduce lead oxide to metallic lead, improves lead recovery and creates favorable conditions for subsequent leaching processes.
3. Alkaline-assisted leaching: adding proper amount of caustic soda into the roasted material, reacting with impurities such as lead sulfate in the slag, and converting into soluble sodium plumbate. This step not only creates conditions for subsequent lead leaching, but also helps to reduce insoluble materials in subsequent processing, thereby reducing processing difficulty and cost.
4. Leaching with water: the leaching is carried out in 60-80 ℃ water solution, so that soluble substances such as sodium plumbate and the like are effectively dissolved into the water phase, and indissolvable impurities such as silicon dioxide and the like remain in the slag phase. The separation effect is beneficial to the subsequent electrolysis process, reduces the impurity content in the electrolysis process, and improves the quality and recovery rate of electrolytic lead.
5. The electrolysis process comprises the following steps: during electrolysis, lead ions migrate to the cathode and deposit, forming electrolytic lead. Meanwhile, impurities in the electrolyte, such as silver, bismuth and other noble metals, are converted into anode slime. This step not only increases the recovery of lead, but also allows for efficient recovery of precious metals, adding additional economic benefits.
6. Treatment of tail liquid and tail slag: the process can be used for recycling the electrolyte tail liquid after purifying treatment, so that the addition amount of new reagents is reduced, and the cost is reduced. Meanwhile, the leached slag is mixed with cement raw materials or made into light aggregate after washing and drying, and is used for building material production, thereby realizing harmless and resource utilization of tailings.
Detailed Description
Alternative embodiments of the present application will be described in more detail below. While alternative embodiments of the present application have been described, it should be understood that the present application may be embodied in various forms and should not be limited to the embodiments set forth herein. Rather, these embodiments are provided so that this disclosure will be thorough and complete, and will fully convey the scope of the disclosure to those skilled in the art.
The terminology used in the present application is for the purpose of describing particular embodiments only and is not intended to be limiting of the present application. As used in this application and the appended claims, the singular forms "a," "an," and "the" are intended to include the plural forms as well, unless the context clearly indicates otherwise. It should also be understood that the term "and/or" as used herein refers to and encompasses any or all possible combinations of one or more of the associated listed items.
The invention provides a recovery treatment process for lead smelting anode slime later-stage slag.
The method comprises the following steps:
(1) Crushing and screening the lead anode slime later-stage slag to a grain size of 2-10mm;
The traditional treatment process usually adopts a direct leaching or roasting mode to treat anode slime, and the extraction rate and the separation effect of lead are unsatisfactory. Coke is introduced as a reducing agent in the reduction roasting pretreatment, and lead oxide is converted into elemental lead by utilizing a reduction reaction, so that favorable conditions are created for subsequent leaching and electrolysis. Meanwhile, the slag is loose and porous in the roasting process, so that the leaching agent is in full contact with the slag, and the leaching rate of lead is remarkably improved;
(2) Reducing roasting the crushed and screened slag at 700-1000 ℃ for 2-5 hours, wherein a reducing agent accounting for 1-10% of the mass of the slag is added in the reducing roasting process;
(3) Cooling the slag after reduction roasting to 500-800 ℃, adding sodium hydroxide accounting for 5-20% of the mass of the slag, and fully mixing;
Prior to the leaching process we innovatively introduced a caustic pretreatment step. The caustic soda reacts with impurities such as lead sulfate and the like remained in the slag to generate soluble sodium plumbate, thereby greatly improving the leaching rate of lead in the subsequent water leaching process. The pretreatment measures make up for the defects of the traditional leaching process when complex anode slime is treated, so that lead is extracted more efficiently and completely;
(4) Soaking the mixture obtained in the step (3) in water for 1-5 hours under the conditions that the liquid-solid ratio is 3-10 and the temperature is 40-90 ℃;
(5) Carrying out solid-liquid separation on the leaching solution and leaching slag, wherein the leaching solution enters an electrolysis process, electrolysis is carried out under the condition that the cathode current density is 100-500A/m, and anode mud containing noble metals is obtained through anode enrichment in the electrolysis process;
We have adopted a two-way electrowinning mode for recovery of lead by cathodic electrodeposition and for enrichment of noble metals by the anode. In the electrolytic process, lead ions are reduced and deposited at a cathode, so that the electrowinning recovery of lead is realized, and noble metal ions such as silver, bismuth and the like in the electrolyte are enriched at an anode to form anode mud, and the anode mud is collected to become a high-grade raw material for noble metal smelting. The bidirectional extraction mode furthest realizes the separation and recovery of lead and noble metal, and the comprehensive recovery effect is far superior to that of the traditional single leaching or electrolysis process;
(6) Purifying the electrolyte tail liquid, adding sodium carbonate accounting for 1-5% of the mass of the tail liquid into the tail liquid, and recycling the electrolyte after removing impurity metal ions by precipitation;
Conventional electrolytic processes typically suffer from electrolyte contamination and handling difficulties. The method adopts the measure of precipitation and purification of sodium carbonate for the electrolyte tail liquid, and effectively removes impurity metal ions such as zinc, copper and the like remained in the electrolyte liquid. The purified electrolyte can be recycled, so that not only is the emission of pollutants reduced, but also the water resource and the electrolyte consumption are saved, and the concept of green environmental protection and sustainable development is reflected;
(7) Washing and drying the leaching slag obtained in the step (5), and mixing the leaching slag with cement raw materials or preparing building material aggregates.
For the leached waste residue, the traditional process often adopts a simple piling or landfill disposal mode, so that not only the land resource is occupied, but also secondary pollution is possibly caused. The leached waste residue is mixed with cement raw material or made into building aggregate, so that the components such as silicon oxide, aluminum and the like remained in the slag are fully utilized, and the high-value utilization of the waste residue is realized. The recycling mode eliminates the environmental risk of waste residues, creates economic benefits and embodies the concept of circular economy.
In one embodiment of the present application, the particle size in step (1) is 3-8mm.
In one embodiment of the present application, the reduction roasting temperature in step (2) is 800-900 ℃ and the time is 2-3 hours.
In one embodiment of the present application, the reducing agent in step (2) is coke, and the addition amount is 3-8% of the slag mass.
In one embodiment of the present application, the sodium hydroxide in the step (3) is added in an amount of 8-15% of the slag mass.
In one embodiment of the present application, the water immersion temperature in step (4) is 60-80 ℃, the liquid-solid ratio is 5:1-8:1, and the leaching time is 1.5-3 hours.
In one embodiment of the present application, the cathode current density in step (5) is 200-400A/m.
In one embodiment of the present application, the sodium carbonate in the step (6) is added in an amount of 2-4% of the tail liquid by mass.
In one embodiment of the present application, the leaching residue of step (7) is blended into the cement raw material in an amount of 5-20%, or the produced building material aggregate is a lightweight aggregate.
In one embodiment of the present application, the precious metal-enriched anode slime obtained in step (5) is washed and dried to be used as a precious metal smelting raw material.
For the sake of clarity, the following test examples are described in detail.
Test example 1
The purpose of the experiment is as follows:
The difference in lead recovery from the present process (including the steps of reduction roasting, alkali-assisted leaching, electrolysis, etc.) is compared to the conventional acid leaching process.
Experimental materials:
lead anode slime later-stage slag sample
Coke (as reducing agent)
Sodium hydroxide (as an alkaline auxiliary leaching agent)
Acid leaching agent (such as sulfuric acid or hydrochloric acid)
Electrolysis apparatus
Analytical instrument (e.g. atomic absorption spectrometer)
The experimental steps are as follows:
1. sample preparation:
And taking the lead anode slime later slag samples of the same batch, and dividing the lead anode slime later slag samples into two groups, wherein the quality of each group of samples is the same.
2. The traditional acid leaching process is as follows:
mixing a group of samples with an acid leaching agent according to a certain proportion, and carrying out acid leaching treatment under certain temperature and time conditions.
After acid leaching treatment, solid-liquid separation is carried out, and leaching liquid is collected.
3. The method of the invention comprises the following steps:
crushing and screening the other group of samples, then reducing and roasting at 800-900 ℃, adding 3-8% of coke as a reducing agent, and roasting for 2-3 hours.
Cooling the roasted sample to 600-700 ℃, adding 8-15% sodium hydroxide, and fully mixing.
Leaching the mixture in water at 60-80deg.C and liquid-solid ratio of 5:1-8:1 for 1.5-3 hr.
And (3) carrying out solid-liquid separation after leaching, and collecting leaching liquid.
4. Lead recovery rate determination:
and (5) measuring the lead content in the two groups of leaching solutions respectively by using an atomic absorption spectrometer.
And (3) calculating the lead recovery rate of the two groups of samples, wherein the recovery rate calculation formula is as follows:
5. data analysis:
Lead content in the original lead anode slime post slag sample: 980 g/t
Lead content in leachate treated by the traditional acid leaching process: 632.5 g/t
Lead content in the leachate treated by the method: 885.6 g/t
By comparing the lead recovery of the two processes, we can see that the recovery of the process (about 90.40%) is significantly higher than that of the conventional acid leaching process (about 64.59%). This result shows that the present method has significant advantages in terms of improved lead recovery due to the effectiveness of pretreatment steps such as reduction roasting and alkali-assisted leaching.
Test example 2
The content of noble metals in the lead smelting anode slime post slag before treatment is respectively as follows: silver (Ag) 6256.78 g/t, bismuth (Bi) 150.29 g/t, palladium (Pd) 51.63 g/t
The noble metal content after the treatment of the invention is silver (Ag) 5632.1 g/t, bismuth (Bi) 127.8 g/t and palladium (Pd) 41.3 g/t;
Precious metal content after traditional pyrometallurgy: silver 4379.74 g/t, bismuth 90.17: 90.17 g/t, palladium 25.82: 25.82 g/t;
Precious metal content after traditional hydrometallurgy: silver 5005.42 g/t, bismuth 112.71 g/t, palladium 36.14 g/t.
The recovery rate of noble metal in the process is about 90% of silver, 85% of bismuth and 80% of palladium;
the recovery rate of noble metal in traditional pyrometallurgy is about 70% of silver, 60% of bismuth and 50% of palladium;
the recovery rate of noble metal in traditional wet smelting is about 80% of silver, 75% of bismuth and 70% of palladium;
the price of the noble metal is 3 yuan/g of silver, 0.5 yuan/g of bismuth and 200 yuan/g of palladium.
The new process of the invention recovers noble metal value:
the recovery silver value is 5632.1 X0.9X3= 15206.67 yuan/t;
the recovered bismuth value is 127.8x0.85x0.5= 54.32 yuan/t;
the recovery palladium value is 41.3 multiplied by 0.8 multiplied by 200=6608 yuan/t;
total recovered noble metal value: 15206.67+54.32+6608= 21868.99 yuan/t.
Precious metal value recovered by traditional pyrometallurgy:
The recovery silver value is 5632.1 X0.7X3= 11826.81 yuan/t;
The recovered bismuth value is 127.8x0.6x0.5= 38.34 yuan/t;
the recovery palladium value is 41.3 multiplied by 0.5 multiplied by 200=4130 yuan/t;
The total recovery precious metal value is 11826.81+38.34+4130= 15995.15 yuan/t.
Precious metal value recovered by traditional wet smelting:
The recovery silver value is 5632.1 X0.8X3= 13517.04 yuan/t;
The recovered bismuth value is 127.8x0.75x0.5= 47.93 yuan/t;
the recovery palladium value is 41.3 multiplied by 0.7 multiplied by 200=5782 yuan/t;
The total recovery precious metal value is 13517.04+47.93+5782= 19346.97 yuan/t.
The novel process has the production cost:
Crushing and screening cost is 12 yuan/t;
Reducing roasting cost is 325 yuan/t;
The alkaline leaching cost is 130 yuan/t;
The electrolysis cost is 1210 yuan/t;
the anode mud treatment cost is 6 yuan/t;
The purification cost of the electrolyte tail liquid is 22 yuan/t;
Equipment depreciation and labor cost are 325 yuan/t;
total cost is 12+325+130+1210+6+22+325=2030 elements/t.
The traditional pyrometallurgy production cost is as follows:
raw material preparation cost is 200 yuan/t;
smelting cost is 1800 yuan/t (1200 yuan/t for fuel and 600 yuan/t for electric power);
reducing agent (coke) cost 400 yuan/t;
the cost of the desulfurizing agent (lime) is 150 yuan/t;
the smoke dust collecting and treating cost is 300 yuan/t;
the slag treatment cost is 200 yuan/t;
the refining cost of the crude lead is 550 yuan/t;
Equipment depreciation and labor cost are 325 yuan/t;
Total cost of 200+1800+400+150+300+200+550+325=3925 units/t.
The traditional wet smelting production cost is as follows:
the preparation cost of the raw materials is 150 yuan/t;
leaching cost is 600 yuan/t (acid-base medicament 400 yuan/t, electric power 200 yuan/t);
the purification cost is 450 yuan/t (medicament 300 yuan/t, electric power 150 yuan/t);
The electrolysis cost is 1000 yuan/t (electric power 950 yuan/t, cathode 50 yuan/t);
the anode slime treatment cost is 200 yuan/t;
the waste liquid treatment cost is 500 yuan/t;
the refining cost of electrolytic lead is 500 yuan/t;
Equipment depreciation and labor cost are 325 yuan/t;
Total cost 150+600+450+1000+200+500+500+325=3725 yuan/t.
The new process of the invention has ton profit:
The total recovery noble metal value is 21868.99 yuan/t;
The total production cost is 2030 yuan/t;
Ton profit, 21868.99-2030= 19838.99 yuan/t.
Ton surplus is smelted by traditional pyrogenic process:
the total recovery noble metal value is 15995.15 yuan/t;
total production cost is 3925 yuan/t;
ton profit 15995.15-3925= 12070.15 yuan/t.
Traditional hydrometallurgical ton surplus:
the total recovery noble metal value is 19346.97 yuan/t;
the total production cost is 3725 yuan/t;
ton profit 19346.97-3725= 15621.97 yuan/t.
Claims (10)
1. The recovery treatment process of the lead smelting anode slime later-stage slag is characterized by comprising the following steps of:
(1) Crushing and screening the lead anode slime later-stage slag to a grain size of 2-10mm;
(2) Reducing roasting the crushed and screened slag at 700-1000 ℃ for 2-5 hours, wherein a reducing agent accounting for 1-10% of the mass of the slag is added in the reducing roasting process;
(3) Cooling the slag after reduction roasting to 500-800 ℃, adding sodium hydroxide accounting for 5-20% of the mass of the slag, and fully mixing;
(4) Soaking the mixture obtained in the step (3) in water for 1-5 hours under the conditions that the liquid-solid ratio is 3-10 and the temperature is 40-90 ℃;
(5) Carrying out solid-liquid separation on the leaching solution and leaching slag, wherein the leaching solution enters an electrolysis process, electrolysis is carried out under the condition that the cathode current density is 100-500A/m, and anode mud containing noble metals is obtained through anode enrichment in the electrolysis process;
(6) Purifying the electrolyte tail liquid, adding sodium carbonate accounting for 1-5% of the mass of the tail liquid into the tail liquid, and recycling the electrolyte after removing impurity metal ions by precipitation;
(7) Washing and drying the leaching slag obtained in the step (5), and mixing the leaching slag with cement raw materials or preparing building material aggregates.
2. The recovery treatment process of lead smelting anode slime post-slag of claim 1, wherein the grain size in the step (1) is 3-8mm.
3. The recovery treatment process of the lead smelting anode slime post slag of claim 1, wherein the reduction roasting temperature of the step (2) is 800-900 ℃ for 2-3 hours.
4. The recovery treatment process of the post-slag of the lead smelting anode slime, which is characterized in that the reducing agent in the step (2) is coke, and the adding amount is 3-8% of the mass of slag.
5. The recovery treatment process of the lead smelting anode slime post-slag of claim 1, wherein the adding amount of sodium hydroxide in the step (3) is 8-15% of the mass of slag.
6. The recovery treatment process of the lead smelting anode slime post-slag according to claim 1, wherein the water leaching temperature in the step (4) is 60-80 ℃, the liquid-solid ratio is 5:1-8:1, and the leaching time is 1.5-3 hours.
7. The recovery treatment process of lead smelting anode slime post-slag of claim 1, wherein the cathode current density in the step (5) is 200-400A/m.
8. The recovery treatment process of the lead smelting anode slime post-slag of claim 1, wherein the adding amount of sodium carbonate in the step (6) is 2-4% of the mass of the tail liquid.
9. The recovery treatment process of the post-slag of the lead smelting anode slime according to claim 1, wherein the blending amount of the leached slag in the step (7) to cement raw materials is 5-20%, or the prepared building material aggregate is lightweight aggregate.
10. The recovery treatment process of the lead smelting anode slime post slag according to claim 1, wherein the precious metal-enriched anode slime obtained in the step (5) is used as a precious metal smelting raw material after washing and drying.
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