Method for comprehensively utilizing oxygen-sulfur mixed lead-zinc ore resources
By employing lead sulfide flotation, zinc sulfide flotation, pyrite flotation, and oxidized ore gravity separation processes, the problem of difficult recovery of oxidized ore in oxygen-sulfur mixed lead-zinc ores has been solved, achieving low-cost and high-efficiency recovery of oxidized ore and quartz resources, and improving metal recovery rate and product added value.
Patent Information
- Authority / Receiving Office
- CN · China
- Patent Type
- Applications(China)
- Current Assignee / Owner
- KUNMING METALLURGY INST
- Filing Date
- 2026-02-26
- Publication Date
- 2026-05-26
AI Technical Summary
In the beneficiation process of oxygen-sulfur mixed lead-zinc ore, ore oxidation causes clay minerals to float, reducing selectivity, resulting in high reagent costs, inconvenience in using production recycled water, difficulty in recovering oxidized ore, and serious waste of resources.
The process involves lead sulfide flotation, zinc sulfide flotation, pyrite flotation, oxidized ore gravity separation, and quartz leaching. It includes crushing and grinding, reagent addition flotation, shaking table gravity separation, and heated alkaline leaching. Sulfide minerals are first floated, followed by oxidized minerals gravity separation, and quartz is recovered through shaking table gravity separation and quartz leaching.
Reduce reagent costs, improve metal recovery rate, reduce land occupation, achieve effective recovery of oxide ores, increase product added value, and realize comprehensive utilization of resources.
Smart Images
Figure CN122076597A_ABST
Abstract
Description
Technical Field
[0001] This invention relates to the field of metallurgical engineering technology, specifically a method for the comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources. Background Technology
[0002] The beneficiation process for mixed oxygen-sulfur lead-zinc ores typically relies on the floatability of the minerals. First, sulfide lead-zinc minerals are floated, followed by oxide lead-zinc minerals. This method offers the advantage of high lead and zinc metal recovery rates and is technically feasible. However, actual production faces numerous challenges. There are three main reasons: First, the flotation process is difficult to control. Due to ore oxidation, the ore contains a certain amount of clay minerals such as kaolinite, causing these clay minerals to float during oxide ore flotation, reducing selectivity. Second, the reuse of production wastewater is inconvenient. The reagent systems for sulfide and oxide ore flotation differ significantly, making it impossible to reuse them together. Production wastewater must be reused separately or treated before reuse. Third, reagent costs are high. Oxidation ore flotation requires large amounts of modifiers, sulfidizing agents, and collectors, making reagent costs far higher than for sulfide ore flotation. For these reasons, if the oxide ore content is low, the production cost exceeds the ore value, leading mines to typically only recover sulfide ore and not oxide ore, resulting in a significant waste of resources. Summary of the Invention
[0003] The purpose of this invention is to provide a method for the comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources.
[0004] The objective of this invention is achieved as follows: the method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources includes lead sulfide flotation, zinc sulfide flotation, pyrite flotation, oxidized ore gravity separation, and quartz leaching steps, specifically including: A. Lead sulfide flotation: 1) The oxygen-sulfur mixed lead-zinc ore to be processed is crushed and ground to -200 mesh, accounting for 60-70%, to obtain material a; 2) Add flotation reagents zinc sulfide inhibitor and lead sulfide collector to material a, and obtain lead sulfide concentrate b and lead tailings c through one roughing, three scavenging and three cleaning processes; B. Zinc sulfide flotation: Pyrite inhibitor, zinc sulfide activator and zinc sulfide collector are added to lead tailings c. After one roughing, three scavenging and three cleaning processes, zinc sulfide concentrate d and zinc tailings e are obtained. C. Pyrite flotation: Add iron sulfide activator and iron sulfide collector to zinc tailings e, and obtain sulfur concentrate g and sulfur tailings h through one roughing, two scavenging and two cleaning processes; D. Oxidized ore gravity separation: The sulfur tailings h are classified into three levels: +200 mesh, -200+400 mesh, and -400 mesh. Then, the sulfur tailings of the three levels are subjected to shaking table gravity separation. Each level of shaking table gravity separation produces three products: heavy minerals are concentrate, medium minerals are medium minerals, and light minerals are tailings. The concentrates from the three levels of gravity separation are combined into oxidized ore concentrate, and the remaining medium minerals and tailings are combined into gravity separation tailings i. E. Quartz leaching: 1) After dehydration, the gravity separation tailings i are placed in a reactor with water and sodium hydroxide for heated leaching. The leaching residue and leaching solution are obtained by solid-liquid separation. 2) Add sodium sulfate electrolyte to the leachate, add sulfuric acid solution dropwise, control the pH value at the end of the reaction to 7, and after the reaction is completed, age and dry to obtain hydrated silica, i.e. white carbon black product.
[0005] The specific steps are as follows: 1. Lead sulfide flotation. The oxygen-sulfur mixed lead-zinc ore is crushed and ground to a fineness of -200 mesh (60-70%). A preferential flotation process is used to first float the lead sulfide minerals, adding 500-1000 g / t of zinc sulfate and 25% zinc sulfate as flotation reagents. # The black powder is 60~120g / t. After one roughing, three scavenging and three cleaning in a closed-circuit cycle, lead sulfide concentrate and lead tailings are obtained. The lead content of lead sulfide concentrate can reach more than 60%. 2. Zinc sulfide flotation. Zinc sulfide minerals are further floated in the lead tailings using a preferential flotation process. Pyrite inhibitor lime (3000-6000 g / t), zinc sulfide activator copper sulfate (100-300 g / t), and zinc sulfide collector butyl xanthate (100-200 g / t) are added. After a closed-circuit cycle of one roughing, three scavenging, and three cleaning stages, zinc sulfide concentrate and zinc tailings are obtained. The zinc content of the zinc sulfide concentrate can reach over 50%. 3. Pyrite Flotation. The zinc tailings are further treated using a preferential flotation process to flotate pyrite. First, 600-1200 g / t of sulfuric acid is added to adjust the pulp pH to approximately 7. Then, 50-100 g / t of copper sulfate (an activator for iron sulfide), 100-200 g / t of copper sulfate (an activator for iron sulfide), and 50-100 g / t of butyl xanthate (an iron sulfide collector) are added. After a closed-circuit cycle of one roughing, two scavenging, and two cleaning stages, pyrite concentrate and sulfur tailings are obtained. The sulfur content of the concentrate can reach over 45%. 4. Oxidized Ore Gravity Separation. A shaking table gravity separation process is used to separate oxidized ore from sulfur beneficiation tailings. Before gravity separation, the sulfur beneficiation tailings need to be classified into three grades: +200 mesh, -200+400 mesh, and -400 mesh. Then, the tailings of each grade are subjected to shaking table gravity separation. Each grade produces three products: heavy minerals as concentrate, medium minerals as middlings, and light minerals as tailings. The concentrates from the three grades are combined to form an oxidized ore concentrate, and the remaining medium minerals and tailings are combined to form gravity separation tailings. The lead and zinc grade of the oxidized ore concentrate can reach over 30%. 5. Quartz Leaching. After dehydration, the gravity separation tailings are placed in a reactor with water and sodium hydroxide for heated leaching at 220℃ for 120 minutes. The alkali-to-residue mass ratio is 2:1, and the liquid-to-solid ratio is 5:1. After cooling to room temperature, the solid and liquid are separated to obtain leaching residue and leachate. 100 ml of the leachate is placed in a beaker and then in a constant-temperature water bath. A certain amount of sodium sulfate electrolyte is added, and a prepared sulfuric acid solution is added dropwise while stirring at a controlled rate. The pH at the reaction endpoint is controlled to be 7, and the reaction time is 1-2 hours. After the reaction, the mixture is aged at a constant temperature for 2 hours. After aging, the solid and liquid are separated. The solid is repeatedly washed with water and finally dried in a 100℃ oven to obtain hydrated silica, i.e., white carbon black.
[0006] This invention involves first flotating sulfide lead-zinc ore, followed by gravity separation of oxidized lead-zinc ore. It's important to note that pyrite must be flotated (or removed) before gravity separation of the oxidized lead-zinc ore because pyrite has a high specific gravity (4.9-5.2), which would affect the separation of oxidized lead-zinc ore during gravity separation. The gravity separation tailings are then processed using heated alkaline leaching and sulfuric acid precipitation to recover quartz minerals. Compared to traditional flotation methods, this method has three advantages: firstly, lower reagent costs, as gravity separation does not require flotation reagents; secondly, the water used in gravity separation requires no treatment and can be directly reused; and thirdly, it can recover ore regardless of its oxide content. There are two disadvantages: firstly, lower recovery rate compared to flotation; and secondly, larger footprint, as the shaking table requires more space than the flotation machine. In summary, this method, while keeping costs under control, can effectively improve metal recovery rates and increase product added value, achieving comprehensive utilization of mineral resources. Attached Figure Description
[0007] Figure 1 This is a schematic diagram of the process flow of the present invention. Detailed Implementation
[0008] The present invention will be further described below with reference to embodiments, but this is not intended to limit the present invention in any way. Any modifications or substitutions made based on the teachings of the present invention shall fall within the protection scope of the present invention.
[0009] The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources described in this invention includes steps such as lead sulfide flotation, zinc sulfide flotation, pyrite flotation, oxidized ore gravity separation, and quartz leaching, specifically including: A. Lead sulfide flotation: 1) The oxygen-sulfur mixed lead-zinc ore to be processed is crushed and ground to -200 mesh, accounting for 60-70%, to obtain material a; 2) Add flotation reagents zinc sulfide inhibitor and lead sulfide collector to material a, and obtain lead sulfide concentrate b and lead tailings c through one roughing, three scavenging and three cleaning processes; B. Zinc sulfide flotation: Pyrite inhibitor, zinc sulfide activator and zinc sulfide collector are added to lead tailings c. After one roughing, three scavenging and three cleaning processes, zinc sulfide concentrate d and zinc tailings e are obtained. C. Pyrite flotation: Add iron sulfide activator and iron sulfide collector to zinc tailings e, and obtain sulfur concentrate g and sulfur tailings h through one roughing, two scavenging and two cleaning processes; D. Oxidized ore gravity separation: The sulfur tailings h are classified into three levels: +200 mesh, -200+400 mesh, and -400 mesh. Then, the sulfur tailings of the three levels are subjected to shaking table gravity separation. Each level of shaking table gravity separation produces three products: heavy minerals are concentrate, medium minerals are medium minerals, and light minerals are tailings. The concentrates from the three levels of gravity separation are combined into oxidized ore concentrate, and the remaining medium minerals and tailings are combined into gravity separation tailings i. E. Quartz leaching: 1) After dehydration, the gravity separation tailings i are placed in a reactor with water and sodium hydroxide for heated leaching. The leaching residue and leaching solution are obtained by solid-liquid separation. 2) Add sodium sulfate electrolyte to the leachate, add sulfuric acid solution dropwise, control the pH value at the end of the reaction to 7, and after the reaction is completed, age and dry to obtain hydrated silica, i.e. white carbon black product.
[0010] The amount of zinc sulfate added in step 2) is 500~1000g / t.
[0011] The amount of 25# black powder added in step 2) is 60~120g / t.
[0012] The pyrite inhibitor mentioned in step B is lime, and the amount added is 3000~6000g / t.
[0013] The zinc sulfide activator mentioned in step B is copper sulfate, with an addition amount of 100~300g / t; the zinc sulfide collector is butyl xanthate, with an addition amount of 100~200g / t.
[0014] The activator for iron sulfide mentioned in step C is sulfuric acid, and the amount added is 50~300g / t.
[0015] The iron sulfide collector mentioned in step C is butyl xanthate, and the addition amount is 50~100g / t.
[0016] The reaction time in step 2) is 1-2 hours.
[0017] In step 2), the aging time is 1-3 hours.
[0018] In step 2), the drying process is carried out in an oven at a temperature of 95~105℃.
[0019] The present invention will be further described below with reference to specific embodiments. Example 1
[0020] Take 500g of oxygen-sulfur mixed lead-zinc ore (lead content 5.52%, zinc content 12.38%, sulfur content 12.84%, iron grade 11.31%, silica grade 40.36%, lead mainly exists in the form of galena, zinc mainly exists in the form of sphalerite and smithsonite, iron mainly exists in the form of pyrite, gangue minerals are mainly silicate minerals such as quartz, zinc oxidation rate is 24%), grind to -0.074mm (70%), mix the ore with water in a 1.5L flotation cell, liquid-solid ratio approximately 2:1; first mix and float galena, add 1000g / t of zinc sulfate before flotation, stir for 3 minutes, then add 25g / t of zinc sulfate. # Add 80 g / t of black powder, stir for 1 minute, and then begin flotation. The flotation froth is the lead rough concentrate. Place the lead rough concentrate in a 0.5L flotation cell and perform three blank cleaning cycles to obtain a lead concentrate with a lead content of 64.82%. Next, float sphalerite in the lead tailings. Before flotation, add 4500 g / t of lime to suppress pyrite, control the pH at around 11, stir for 3 minutes, add 180 g / t of copper sulfate to activate the sphalerite, stir for 3 minutes, and then add 120 g / t of butyl xanthate and 2... # Add 30g / t of oil, stir for 1 minute, and then begin flotation. The flotation foam is the zinc rough concentrate. Place the zinc rough concentrate in a 0.5L flotation cell for three refining processes (500g / t of lime added for refinement 1, 300g / t of lime added for refinement 2, and no lime added for refinement 3). This yields a zinc concentrate containing 52.56% zinc, with a zinc recovery rate of 72.14%. The zinc roughing tailings undergo three scavenging processes, with a small amount of copper sulfate and butyl xanthate added during scavenging. The scavenged tailings are then the zinc tailings. The zinc tailings are then used for pyrite flotation. Before flotation, the pulp is pre-treated by adding 850g / t of sulfuric acid to adjust the pulp pH to approximately 7. 80g / t of copper sulfate is added to activate the pyrite, and the mixture is stirred for 3 minutes. Finally, 80g / t of butyl xanthate and 2... #30g / t of oil is stirred for 1 minute before flotation begins. The flotation foam is the sulfur rough concentrate. The sulfur rough concentrate is placed in a 0.5L flotation cell for two blank cleaning cycles to obtain a sulfur concentrate with a sulfur content of 46.33% and a sulfur recovery rate of 85%. The sulfur roughing tailings are scavenged twice, with a small amount of copper sulfate and butyl xanthate added during scavenging. The scavenged tailings are the sulfur tailings. For sulfur tailings, a shaking table gravity separation process is used to separate smithsonite (due to the difference in specific gravity between smithsonite (4.4) and gangue (2.7)). Before gravity separation, the sulfur tailings need to be classified into three levels: +200 mesh, -200+400 mesh, and -400 mesh. Then, the sulfur tailings of the three levels are subjected to shaking table gravity separation of smithsonite. Each level of shaking table gravity separation produces three products: heavy minerals as concentrate, medium minerals as middlings, and light minerals as tailings. The concentrates from the three levels of gravity separation are combined into zinc oxide concentrate, and the remaining medium minerals and tailings are combined into gravity separation tailings. The zinc grade of the zinc oxide concentrate can reach 31.28%, and the operating recovery rate is 60%. The zinc recovery rate of the original ore is 86%, which is 14 percentage points higher than that of the zinc concentrate. After dehydration, sulfur tailings were leached in a reactor with water and sodium hydroxide at 220°C for 120 minutes. The alkali-to-residue ratio was 2:1, and the liquid-to-solid ratio was 5:1. After cooling to room temperature and solid-liquid separation, leaching residue and leachate were obtained. 100 ml of the leachate was placed in a beaker and then in a constant-temperature water bath at 80°C. 3% sodium sulfate electrolyte was added, and 10% sulfuric acid solution was added dropwise with a certain stirring rate, controlling the pH at the final reaction point to be 7. The reaction time was 1-2 hours. After the reaction, the mixture was aged at a constant temperature for 2 hours. After aging, the solid and liquid were separated. The solid was repeatedly washed with water and finally dried in an oven at 100°C to obtain hydrated silica, i.e., white carbon black. The specific surface area of white carbon black is 204.71 m². 2 / g, this product meets the highest category A requirements in the national standard GB / T10722.
[0021] Example 2
[0022] Take 500g of oxygen-sulfur mixed lead-zinc ore (lead content 3.11%, zinc content 15.40%, sulfur content 9.29%, iron grade 13.08%, silica grade 42.28%, lead mainly exists in the form of galena, zinc mainly exists in the form of sphalerite and smithsonite, iron mainly exists in the form of pyrite, gangue minerals are mainly silicate minerals such as quartz, and zinc oxidation rate is 15%), grind it to -0.074mm (65%), mix the ore with water in a 1.5L flotation cell, liquid-solid ratio approximately 2:1; first mix and float the galena, add 1000g / t of zinc sulfate before flotation, stir for 3 minutes, then add 25g / t of zinc sulfate. #Add 60 g / t of black powder, stir for 1 minute, and then begin flotation. The flotation froth is the lead rough concentrate. Place the lead rough concentrate in a 0.5L flotation cell and perform three blank cleaning cycles to obtain a lead concentrate with a lead content of 61.36%. Next, float sphalerite in the lead tailings. Before flotation, add 4000 g / t of lime to suppress pyrite, control the pH at around 11, stir for 3 minutes, add 150 g / t of copper sulfate to activate the sphalerite, stir for 3 minutes, and then add 150 g / t of butyl xanthate and 2... # Add 30g / t of oil, stir for 1 minute, and then begin flotation. The flotation foam is the zinc rough concentrate. Place the zinc rough concentrate in a 0.5L flotation cell for three refining processes (500g / t of lime added for refinement 1, 300g / t of lime added for refinement 2, and no lime added for refinement 3). This yields a zinc concentrate containing 53.12% zinc, with a zinc recovery rate of 81.07%. The zinc roughing tailings are then subjected to three scavenging processes, with a small amount of copper sulfate and butyl xanthate added during scavenging. The scavenged tailings are the zinc tailings. The zinc tailings are then used for pyrite flotation. Before flotation, the pulp is pre-treated by adding 750g / t of sulfur to adjust the pulp pH to approximately 7. 70g / t of copper sulfate is added to activate the pyrite, and the mixture is stirred for 3 minutes. 80g / t of butyl xanthate and 2... #30g / t of oil is stirred for 1 minute before flotation begins. The flotation froth is the sulfur rough concentrate. The sulfur rough concentrate is then placed in a 0.5L flotation cell for two blank cleaning cycles to obtain a sulfur concentrate containing 45.49% sulfur, with a sulfur recovery rate of 87%. The sulfur roughing tailings are then scavenged twice, with a small amount of copper sulfate and butyl xanthate added during scavenging. The scavenged tailings are the sulfur tailings. The sulfur tailings are then separated into smithsonite using a shaking table gravity separation process. Before gravity separation, the sulfur tailings are first classified into three levels: +200 mesh, -200+400 mesh, and -400 mesh. Then, the sulfur tailings of each level are subjected to shaking table gravity separation for smithsonite. Each level of shaking table gravity separation produces three products: heavy minerals as concentrate, medium minerals as middlings, and light minerals as tailings. The concentrates from the three levels of gravity separation are combined to form zinc oxide concentrate, and the remaining medium minerals and tailings are combined to form gravity separation tailings. The zinc grade of zinc oxide concentrate can reach 30.55%, with an operational recovery rate of 50%, compared to 88% for the original ore, representing a 7.5 percentage point increase in zinc recovery. After dehydration, the sulfur tailings were placed in a reactor with water and sodium hydroxide for heated leaching at 220℃ for 120 minutes. The alkali-to-residue mass ratio was 2:1, and the liquid-to-solid ratio was 5:1. After cooling to room temperature and solid-liquid separation, leaching residue and leachate were obtained. 100 ml of the leachate was placed in a beaker and then in a constant-temperature water bath at 80℃. 3% sodium sulfate electrolyte was added, and 10% sulfuric acid solution was added dropwise with a certain stirring rate, controlling the pH at the reaction endpoint to 7. The reaction time was 1-2 hours, followed by aging at a constant temperature for 2 hours. After aging, the solid and liquid were separated. The solid was repeatedly washed with water and finally dried in a 100℃ oven to obtain hydrated silica, i.e., white carbon black. The specific surface area of white carbon black is 195.29 m². 2 / g, this product meets the highest category A requirements in the national standard GB / T10722.
Claims
1. A method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources, characterized in that, The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources includes lead sulfide flotation, zinc sulfide flotation, pyrite flotation, oxidized ore gravity separation, and quartz leaching steps, specifically including: A. Lead sulfide flotation: 1) The oxygen-sulfur mixed lead-zinc ore to be processed is crushed and ground to -200 mesh, accounting for 60-70%, to obtain material a; 2) Add flotation reagents zinc sulfide inhibitor and lead sulfide collector to material a, and obtain lead sulfide concentrate b and lead tailings c through one roughing, three scavenging and three cleaning processes; B. Zinc sulfide flotation: Pyrite inhibitor, zinc sulfide activator and zinc sulfide collector are added to lead tailings c. After one roughing, three scavenging and three cleaning processes, zinc sulfide concentrate d and zinc tailings e are obtained. C. Pyrite flotation: Add iron sulfide activator and iron sulfide collector to zinc tailings e, and obtain sulfur concentrate g and sulfur tailings h through one roughing, two scavenging and two cleaning processes; D. Oxidized ore gravity separation: The sulfur tailings h are classified into three levels: +200 mesh, -200+400 mesh, and -400 mesh. Then, the sulfur tailings of the three levels are subjected to shaking table gravity separation. Each level of shaking table gravity separation produces three products: heavy minerals are concentrate, medium minerals are medium minerals, and light minerals are tailings. The concentrates from the three levels of gravity separation are combined into oxidized ore concentrate, and the remaining medium minerals and tailings are combined into gravity separation tailings i. E. Quartz leaching: 1) After dehydration, the gravity separation tailings i are placed in a reactor with water and sodium hydroxide for heated leaching. The leaching residue and leaching solution are obtained by solid-liquid separation. 2) Add sodium sulfate electrolyte to the leachate, add sulfuric acid solution dropwise, control the pH value at the end of the reaction to 7, and after the reaction is completed, age and dry to obtain hydrated silica, i.e. white carbon black product.
2. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The amount of zinc sulfate added in step 2) is 500~1000g / t.
3. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The 25 mentioned in step 2) # The amount of black medicine added is 60~120g / t.
4. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The pyrite inhibitor mentioned in step B is lime, and the amount added is 3000~6000g / t.
5. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The zinc sulfide activator mentioned in step B is copper sulfate, with an addition amount of 100~300g / t; the zinc sulfide collector is butyl xanthate, with an addition amount of 100~200g / t.
6. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The activator for iron sulfide mentioned in step C is sulfuric acid, and the amount added is 50~300g / t.
7. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The iron sulfide collector mentioned in step C is butyl xanthate, and the amount added is 50~100g / t.
8. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, The reaction time in step 2) is 1-2 hours.
9. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, In step 2), the aging time is 1-3 hours.
10. The method for comprehensive utilization of oxygen-sulfur mixed lead-zinc ore resources according to claim 1, characterized in that, In step 2), the drying process is carried out in an oven at a temperature of 95~105℃.