Comprehensive utilization methods of sulfide tailings

Abstract

This invention discloses a method for the comprehensive utilization of sulfide tailings, relating to the field of sulfide tailings treatment technology. This method uses lead-zinc tailings as raw material and innovatively employs a combined "two-stage acid leaching-gravity separation-flotation" process. First, two stages of acid leaching selectively leach calcite associated with barite in the tailings, reducing the influence of muddy calcite while activating pyrite. Subsequently, the acid leaching residue undergoes gravity separation to further separate undissolved calcite, quartz, and other silica-containing gangue minerals. The gravity concentrate is then subjected to flotation to achieve the separation of barite and pyrite, ultimately obtaining high-purity barite and pyrite products respectively. This invention features a reasonable process route and simple operation, effectively solving the separation problem of fine-grained, high-viscosity lead-zinc tailings, achieving full-component purification and utilization of lead-zinc tailings, significantly enhancing the economic value of tailings resources, and reducing the environmental pressure caused by tailings stockpiling, thus demonstrating good economic, environmental, and social benefits.

Description

Technical Field

[0001] This invention relates to the field of sulfide tailings treatment technology, and in particular to a comprehensive utilization method for sulfide tailings, which is especially suitable for the purification, separation and full utilization of barite and pyrite in fine-grained, high-viscosity lead-zinc tailings. Background Technology

[0002] Although my country possesses abundant barite resources, its endowment characteristics have significant shortcomings. High-quality, single-type barite deposits account for a low proportion; most deposits are associated or co-occurring minerals closely linked to salt minerals such as fluorite, calcite, dolomite, and quartz, or polymetallic sulfide minerals such as lead-zinc, copper-iron (primarily pyrite). Furthermore, a large amount of barite resources are lost to tailings during lead-zinc ore beneficiation processes. Lead-zinc mining and beneficiation generate substantial tailings, which not only occupy significant land resources but also pose serious environmental pollution risks due to heavy metal ion leakage and dust emissions. Simultaneously, lead-zinc tailings often contain associated minerals with high economic value, such as barite and pyrite. Effective recycling and utilization of these minerals could not only improve resource utilization but also reduce the environmental pressure of tailings storage.

[0003] However, existing lead-zinc tailings generally exhibit fine particle size (a high proportion of particles ≤0.074mm) and high overall viscosity, posing numerous challenges to mineral separation. On the one hand, fine-grained tailings particles have a large specific surface area and high surface energy, making them prone to agglomeration, which greatly hinders physical sieving and separation from gangue minerals such as calcite. On the other hand, the highly viscous surface of these fine-grained tailings exhibits strong polarity, non-selectively adsorbing onto the surfaces of barite and pyrite, further impeding effective contact between reagents and target minerals, consuming large amounts of flotation reagents, and resulting in low separation efficiency and product purity.

[0004] Currently, the main methods for recovering barite and pyrite from lead-zinc tailings include single flotation and acid leaching, but these methods all have significant drawbacks: First, the single flotation process has significant shortcomings in recovering barite from lead-zinc ore tailings. During the primary flotation of lead-zinc ore, pyrite is inhibited by depressants and remains in the tailings. The depressants adsorbed on its surface hinder the action of subsequent flotation reagents. At the same time, calcite that has not been removed from the tailings competes with barite for collectors, causing mechanical entrainment. In addition, the agglomeration effect of fine tailings ultimately leads to low barite recovery rate and poor product purity in the single flotation process, making it difficult to achieve industrial application.

[0005] Secondly, there are few research reports on the acid leaching separation of barite and calcite. Calcite and barite, which are commonly found in lead-zinc mine tailings, have similar flotation characteristics, making it difficult for conventional flotation processes to achieve efficient separation of the two.

[0006] Third, the synergistic effect of acid leaching pretreatment has not been fully utilized. Removing calcite through acid leaching before flotation not only eliminates calcite interference at the source, but more importantly, the hydrogen ions in the acid leaching system can decompose the inhibitor film adsorbed on the pyrite surface, activating the inhibited pyrite and restoring its hydrophobicity. This creates favorable conditions for subsequent flotation removal of pyrite. However, existing technologies often overlook this crucial synergistic effect, leading to incomplete pyrite removal. While traditional single acid leaching can remove some calcite, incomplete leaching leaves residual calcite that still affects subsequent flotation and can easily lead to over-dissolution of the target mineral, reducing recovery rates.

[0007] Fourth, the traditional process does not add a gravity separation step after acid leaching. Acid leaching alone cannot completely dissolve the calcite intergrowths embedded in barite. The remaining undissolved calcite, quartz and other silica gangue minerals will directly enter the flotation stage, interfering with the flotation separation of barite and pyrite. As a result, the two cannot be completely separated, and it is difficult to obtain high-purity barite products or recover high-purity pyrite products.

[0008] For a long time, barite purification has primarily relied on high-quality primary ore as raw material, employing processes such as single acid leaching and conventional flotation. Research on barite recovery technologies from low- and medium-grade associated ores and tailings has lagged behind. Especially for complex secondary resources like lead-zinc mine tailings containing pyrite and calcite, existing processes cannot effectively separate barite from pyrite and calcite, nor can they overcome the bottleneck of separating fine-grained, high-viscosity tailings. This results in the waste of large quantities of valuable barite resources due to tailings stockpiling, causing resource idleness and exacerbating the environmental pressure from tailings accumulation. Therefore, developing a co-sorting process that efficiently solves the separation problem of barite from pyrite and calcite, and is adapted to the characteristics of lead-zinc mine tailings, is of significant practical and industrial value for improving the recovery rate of barite resources in tailings and alleviating the pressure of high-quality barite resource scarcity. Summary of the Invention

[0009] This invention provides a method for the comprehensive utilization of sulfide ore tailings, aiming to overcome the shortcomings of existing technologies where the fine particle size and high viscosity of sulfide ore tailings make it difficult to separate from gangue minerals such as calcite, resulting in low recovery rates of barite and pyrite. This method employs a combined "two-stage acid leaching-gravity separation-flotation" process. First, two stages of acid leaching decalcify and activate pyrite. Then, after acid leaching, gravity separation further separates undissolved calcite, quartz, and other silica-containing gangue minerals, completely removing calcite from the tailings. Subsequently, the gravity concentrate is subjected to flotation to achieve complete separation of barite and pyrite, simultaneously obtaining high-purity barite and pyrite products. This also reduces the environmental pressure of tailings storage, achieving a win-win situation for resource recycling and environmental benefits in the mineral processing industry.

[0010] To achieve this technical objective, the present invention adopts the following solution: The comprehensive utilization method of sulfide tailings includes the following steps: First-stage acid leaching treatment: The lead-zinc tailings are mixed with the first acid solution in a certain proportion, stirred and reacted, and after the reaction is completed, solid-liquid separation is carried out to obtain first-stage acid leaching residue and first-stage acid leaching solution; Two-stage acid leaching treatment: The first-stage acid leaching residue and the second acid solution are mixed in proportion, stirred and reacted, and after the reaction is completed, solid-liquid separation is performed to obtain the second-stage acid leaching residue and the second-stage acid leaching solution. Acid leaching followed by gravity separation: The acid leaching residue from the two stages is mixed with water to prepare a gravity separation slurry, which is then subjected to gravity separation treatment. The feed concentration, shaking table stroke, number of strokes, and water volume are controlled. By utilizing the density difference between barite, pyrite, and undissolved calcite and quartz, the undissolved silica gangue minerals are further separated to obtain gravity concentrate and gravity tailings. Flotation pretreatment: The gravity concentrate is washed with water until the pH is 5.5~7.0, and then deionized water is added to prepare a slurry with a mass fraction of 20%~40%. A pH adjuster is added to the slurry to adjust the pH to 6.0~8.0 to obtain the pretreated slurry. Pyrite flotation: Collector and frother are added to the pretreated slurry, and flotation is carried out to obtain pyrite concentrate and flotation tailings; Barite purification: The flotation tailings are filtered and dried to obtain barite products. Further, the lead-zinc tailings contain 30%–50% barite, 3%–6% pyrite, and 10%–20% calcite, with particles ≤0.074mm accounting for ≥60% of the total.

[0011] Furthermore, the ratio of the lead-zinc tailings to the first acid solution is 1g:(2~5)mL; and / or, the ratio of the first-stage acid leaching residue to the second acid solution is 1g:(2~5)mL; and / or, the stirring temperature of the first-stage acid leaching treatment and the second-stage acid leaching treatment is 20℃~50℃, the stirring rate is 200~400r / min, and the reaction time is 1~1.5h.

[0012] Furthermore, the first acid solution is dilute hydrochloric acid with a mass concentration of 6% to 9%; and / or, the second acid solution is dilute hydrochloric acid with a mass concentration of 9.5% to 12.5%.

[0013] Furthermore, the solid-liquid separation is performed using any one of vacuum filtration, pressure filtration, or centrifugal separation.

[0014] Furthermore, the mass fraction of the gravity separation slurry is 25% to 35%, and the gravity separation operation is carried out on a shaking table with a stroke of 10 mm to 18 mm, a stroke rate of 220 r / min to 320 r / min, and a water flow rate of 8 L / min to 18 L / min.

[0015] Furthermore, the pH adjuster is any one or a mixture of sodium carbonate, sodium bicarbonate, or sodium hydroxide.

[0016] Furthermore, the gravity concentrate is composed of barite and pyrite, and the gravity tailings mainly contain undissolved calcite and quartz.

[0017] Furthermore, the collector is any one or a mixture of butyl xanthate, pentyl xanthate, or isobutyl xanthate; and / or, the amount of the collector added is 80-120 g / t slurry; and / or, the frother is any one or a mixture of No. 2 oil, pine oil, or methyl isobutyl methanol; and / or, the amount of the frother added is 10-50 g / t slurry.

[0018] Furthermore, the flotation process includes roughing, 1-3 cleaning stages, and 1-2 scavenging stages. The roughing concentration is 20%-30% and the time is 3-8 minutes; the cleaning concentration is 20%-25% and the cleaning time is 2-5 minutes; the scavenging concentration is 20%-25% and the time is 2-5 minutes. Multi-stage flotation can improve the purity and recovery rate of pyrite concentrate.

[0019] Furthermore, the drying temperature is 90~120℃, and the drying time is 2~6 hours. Controlling the drying temperature and time ensures thorough removal of moisture while preventing damage to the barite crystal structure.

[0020] Compared with the prior art, the beneficial effects of the present invention are as follows: 1) This invention uses lead-zinc mine tailings as a secondary resource, not only recovering the associated barite in a targeted manner, but also simultaneously recovering pyrite resources through a synergistic design of "two-stage acid leaching-gravity separation-flotation" to obtain high-purity pyrite products, achieving full component recovery of valuable minerals in the tailings; at the same time, the gravity separation tailings generated after acid leaching can be reused as building aggregate, realizing tailings reduction and resource cascade utilization, which is in line with the resource recycling development concept of the mineral processing industry; 2) This invention innovatively designs a "two-stage acid leaching-gravity separation-flotation" synergistic separation system. First, the two-stage gradient acid leaching (8.5% hydrochloric acid pre-leaching → 10.5% hydrochloric acid deep leaching) achieves stepwise and precise dissolution of fine-grained calcite, while thoroughly activating the suppressed pyrite, laying the foundation for subsequent flotation separation. Second, a gravity separation stage is added after acid leaching to precisely remove undissolved calcite, quartz, and other silica-containing gangue minerals, preventing them from interfering with subsequent flotation separation and improving flotation accuracy from the source. Third, the flotation process is specifically designed to achieve complete separation of barite and pyrite, solving the core pain point of incomplete separation in traditional processes. Fourth, the final shaking table gravity separation deeply purifies barite, ensuring product purity. 3) This invention performs flotation pretreatment after acid leaching, removes residual acid and soluble salts by washing with water, adjusts the pH of the pulp, ensures the effective action of flotation reagents on the target mineral, significantly improves flotation separation efficiency, and achieves complete separation of barite and pyrite. 4) The unit operations such as gravity separation, acid leaching, and flotation used in this invention are all mature technologies in the field of mineral processing. The equipment selection is conventional, the process parameters are easy to control, and the raw material is low-cost tailings. The overall processing cost is low and it is suitable for large-scale industrial promotion. 5) This invention reduces the amount of tailings stockpiled and the area occupied by tailings through resource utilization and volume reduction, and reduces the environmental risk of pollutant migration and diffusion in tailings; the acid leaching waste liquid can be discharged in compliance with standards after neutralization treatment, and the flotation reagents are used in small quantities and are easily degradable, which meets the industry requirements of green mineral processing. 6) The process route of this invention is reasonable and easy to operate. It can realize the purification and utilization of valuable components of lead-zinc tailings, and obtain high-purity barite and pyrite products respectively, which significantly improves the economic value of tailings resources. Attached Figure Description

[0021] Figure 1 A flowchart of the comprehensive utilization method of sulfide tailings provided by the present invention. Detailed Implementation

[0022] To fully understand the purpose, features and effects of the present invention, the present invention will be described in detail through the following specific embodiments, but the present invention is not limited thereto.

[0023] As described in the background section, existing single flotation technologies fail to remove calcite beforehand, and fine particle agglomeration and high viscosity affect the separation effect; traditional single acid leaching methods do not completely remove calcite, and residual gangue interferes with subsequent separation. It is impossible to simultaneously achieve efficient purification and separation of barite and pyrite in fine-grained, high-viscosity lead-zinc tailings, as well as the utilization of all components. To solve the above technical problems, this invention provides a method for comprehensive utilization of tailings.

[0024] The equipment used in this invention are all conventional industrial equipment in the field of mineral processing, and the reagents used are all industrial grade with a purity ≥98%. The detection methods are as follows: barite purity is determined by EDTA complexometric titration, pyrite purity is determined by potassium dichromate redox titration, mineral recovery rate = (concentrate mass × concentrate purity) ÷ (raw ore mass × raw ore mineral content) × 100%, calcite removal rate = ((raw ore mass × raw ore calcite content) - (leaching residue mass × leaching residue calcite content)) ÷ (raw ore mass × raw ore calcite content) × 100%, and tailings viscosity is determined by rotational viscometer (25℃, 60 r / min).

[0025] The raw materials used in the examples and comparative examples were flotation tailings from a lead-zinc mine. The main chemical components (mass fraction) of the tailings were tested and found to be: BaSO4 31.2%, CaCO3 15.5%, FeS2 4.8%, SiO2 23.3%, MgCO3 (dolomite) 7.2%, and other impurities 3.0%. The tailings particles with a size ≤0.074mm accounted for 72%, and the viscosity was 38mPa・s, which is consistent with the characteristics of fine-grained, high-viscosity lead-zinc tailings.

[0026] Example 1

[0027] The specific steps for the comprehensive utilization of sulfide ore tailings are as follows: S1. First-stage acid leaching treatment: Take 100.0 kg of the above-mentioned lead-zinc ore flotation tailings and put them into a 1000L stainless steel stirred reaction tank. Add 300.0 L of dilute hydrochloric acid (first acid solution) with a mass concentration of 8.5%, control the solid-liquid ratio at 1:3 (g:mL), the stirring speed at 300 r / min, and leach at a constant temperature for 1.0 h. After the reaction is completed, use a XAY200 / 1250 plate and frame filter press for solid-liquid separation to obtain a first-stage acid leaching residue and a first-stage acid leaching solution (calcium chloride solution). The mass of the first-stage acid leaching residue is 78.6 kg, and the CaCO3 content is reduced to 8.2%.

[0028] S2, Second-stage acid leaching treatment: The first-stage acid leaching residue obtained in S1 was fed into the same reaction tank, and 493.0 L of 10.5% dilute hydrochloric acid (second acid solution) was added. The solid-liquid ratio was controlled at 1:5 (g:mL), the acid leaching temperature at 25℃, and the stirring speed at 300 r / min. The leaching was carried out at a constant temperature for 1.0 h. After the reaction, a plate and frame filter press was used again for solid-liquid separation to obtain the second-stage acid leaching residue and the second-stage acid leaching solution (calcium chloride solution). The second-stage acid leaching solution and the first-stage acid leaching solution were combined and collected for subsequent calcium chloride recovery. The mass of the second-stage acid leaching residue was 52.3 kg, with the CaCO3 content reduced to 0.8%, the tailings viscosity reduced to 8 mPa·s, and the calcite removal rate 96.1%.

[0029] S3. Gravity Separation After Acid Leaching: The second-stage acid leaching residue obtained in S2 was added to a stirring tank, and water was added to prepare a gravity separation slurry with a mass fraction of 30%. This slurry was then fed into an LY-1.2 type shaking table for gravity separation. The gravity separation parameters were controlled as follows: feed concentration 30%, shaking table stroke 14 mm, stroke rate 280 r / min, and water flow rate 12 L / min. After gravity separation, 36.8 kg of gravity concentrate and 15.5 kg of gravity tailings were obtained. Analysis showed that the gravity tailings mainly contained 72.3% SiO2 and 0.6% CaCO3, while the gravity concentrate was mainly a mixture of BaSO4 and FeS2.

[0030] S4. Flotation pretreatment: The gravity concentrate obtained in S3 is added to a washing tank and washed with clean water until the pH of the washing liquid reaches 6.2 (monitored in real time by a PHS-3C pH meter) to remove residual acid and soluble salts from the surface. Then, the washed concentrate is mixed with water to prepare a slurry with a mass fraction of 30%, and industrial-grade sodium carbonate is added as a pH adjuster to adjust the pH of the slurry to 7.0. The mixture is then stirred evenly and set aside.

[0031] S5. Pyrite Flotation: The slurry pretreated by S4 is fed into a KYF-2.8 type aerated mechanically stirred flotation cell, and stirring is started (stirring speed 260 r / min); first, 100 g / t of butyl xanthate (collector) (based on the weight of the raw ore, i.e., 10.0 g) is added, and stirred for 4 min to allow the reagent to be fully adsorbed; then, 30 g / t of pine oil (frother) (based on the weight of the raw ore, i.e., 3.0 g) is added, and stirred for 2 min; then, flotation is carried out: one roughing stage (flotation time 5 min, roughing froth is skimmed off). Two cleaning processes were performed (first cleaning: the froth concentration in the roughing process was adjusted to 20%, no additional reagents were added, and the flotation time was 3 minutes; second cleaning: the froth concentration in the first cleaning process was adjusted to 18%, no additional reagents were added, and the froth product was pyrite concentrate), and one scavenging process was performed (the product in the roughing cell was adjusted to 20%, butyl xanthate 20 g / t and pine oil 5 g / t were added, the flotation time was 3 minutes, and the froth was returned to the roughing operation); finally, 4.2 kg of pyrite concentrate and 22.1 kg of slurry (barite crude product) were obtained.

[0032] S6. Barite purification: The slurry obtained in S5 is sent to a plate and frame filter press for filtration to obtain a barite filter cake; the filter cake is placed in a 101-3 type electric constant temperature oven and dried at 110℃ for 3.0h to obtain a high-purity barite product.

[0033] X-ray fluorescence spectroscopy was used to detect the viscosity of the obtained barite product, pyrite concentrate, final tailings, and calcite removal rate. The results were as follows: the purity of the barite product was 97.5%, the actual recovery amount was 21.5475 kg, and the total recovery rate of barite was 94.0%; the purity of the pyrite concentrate was 93.2%, the actual recovery amount was 3.9144 kg, and the total recovery rate of pyrite was 81.6%; the final tailings discharge was reduced by 68.2% compared with the original ore, and the tailings viscosity was reduced to 5 mPa·s, achieving complete removal of calcite, reduction of tailings volume, and efficient resource recovery.

[0034] Example 2

[0035] The method for comprehensive utilization of sulfide tailings, and the specific mineral processing steps are as follows: S1. First-stage acid leaching treatment: Take 200.0 kg of the above-mentioned lead-zinc ore flotation tailings and feed it into a 2000L stainless steel stirred reaction tank. Add 400.0 L of 7.0% hydrochloric acid (first acid solution), control the solid-liquid ratio at 1:2 (g:mL), the acid leaching temperature at 40℃, and the stirring speed at 250 r / min, and leach at a constant temperature for 1.2 h. After the reaction, use a centrifuge (speed 3000 r / min) to separate the solid and liquid, obtaining a first-stage acid leaching residue and a first-stage acid leaching solution. The mass of the first-stage acid leaching residue was 155.3 kg, the CaCO3 content was reduced to 7.8%, and the tailings viscosity was reduced to 16 mPa·s.

[0036] S2. Second-stage acid leaching treatment: The first-stage acid leaching residue was fed into a reaction tank, and 776.5 L of 11.0% dilute hydrochloric acid (second acid solution) was added. The solid-liquid ratio was controlled at 1:4 (g:mL), the acid leaching temperature at 40℃, and the stirring speed at 250 r / min. The leaching was carried out at a constant temperature for 1.2 h. After the reaction, the residue was separated by centrifugation to obtain the second-stage acid leaching residue and the second-stage acid leaching solution. The two acid leaching solutions were combined for later use. The second-stage acid leaching residue had a mass of 103.6 kg, the CaCO3 content was reduced to 0.6%, the tailings viscosity was reduced to 7 mPa·s, and the calcite removal rate was 97.1%.

[0037] S3. Gravity separation after acid leaching: The acid leaching residue from the second stage is mixed with water to prepare a gravity slurry with a mass fraction of 35%, which is then fed into an LY-1.2 type shaking table for gravity separation. Control parameters: feed concentration 35%, shaking table stroke 16mm, stroke rate 300r / min, and water flow rate 15L / min. 72.5kg of gravity concentrate and 31.1kg of gravity tailings are obtained. The gravity tailings mainly contain 73.1% SiO2 and 0.5% CaCO3, and the valuable minerals have been fully enriched in the gravity concentrate.

[0038] S4. Flotation pretreatment: Wash the gravity concentrate with clean water until the washing liquid pH=6.8, prepare a slurry with a mass fraction of 35%, add sodium hydroxide as a pH adjuster to adjust the slurry pH to 7.5, and stir evenly.

[0039] S5. Pyrite Flotation: The pretreated slurry is fed into an XFD-3.0L single-cell flotation machine. 110 g / t of pentyl xanthate (collector) (based on the weight of the raw ore, i.e., 22.0 g) is added and stirred for 4 min. 40 g / t of No. 2 oil (frother) (based on the weight of the raw ore, i.e., 8.0 g) is added and stirred for 2 min. Flotation is carried out: one roughing (flotation time 6 min), three cleanings (each cleaning concentration is 20%, 18%, and 16% respectively, without additional reagents, for 3 min, 2 min, and 2 min respectively), and two scavengings (each time 15 g / t of pentyl xanthate and 5 g / t of No. 2 oil are added, for 3 min, and the froth is returned to the previous stage). 8.5 kg of pyrite concentrate and 43.8 kg of slurry (barite crude) are obtained.

[0040] S6. Barite purification: The slurry in the tank is filtered to obtain a barite filter cake, which is then dried at 105℃ for 4.0h to obtain a high-purity barite product.

[0041] The obtained barite product and pyrite concentrate were subjected to X-ray fluorescence spectroscopy. The results showed that the purity of the barite product was 97.8%, the actual recovery amount was 42.8364 kg, and the total recovery rate of barite was 95.0%; the purity of the pyrite concentrate was 93.8%, the actual recovery amount was 7.973 kg, and the total recovery rate of pyrite was 83.0%; the tailings discharge was reduced by 69.5%, and the tailings viscosity was reduced to 4.5 mPa·s.

[0042] Comparative Example 1 (omitting the two-stage acid leaching process, using only the one-stage acid leaching-gravity separation-flotation process) A simplified process of "one-stage acid leaching-gravity separation-flotation" is used to recover barite and pyrite from lead-zinc ore flotation tailings. The specific steps are as follows: S1. First stage acid leaching treatment: Take 100.0 kg of the above lead-zinc ore flotation tailings, add 300.0 L of dilute hydrochloric acid with a mass concentration of 8.5%, control the solid-liquid ratio at 1:3, the temperature at 25℃, and the stirring speed at 300 r / min, and leach for 1.0 h; plate and frame filter press to obtain 78.6 kg of acid leaching residue (consistent with the first stage acid leaching residue of Example 1, CaCO3 content 8.2%, viscosity 18 mPa·s), and the calcite removal rate is 56.8%.

[0043] S2, Direct Gravity Separation-Flotation: The two-stage acid leaching step is omitted. The above-mentioned first-stage acid leaching residue is directly subjected to gravity separation, flotation pretreatment, pyrite flotation and barite purification operations according to the parameters of Example 1. All operating parameters are completely consistent with those of Example 1.

[0044] X-ray fluorescence spectroscopy was performed on the barite product and pyrite concentrate obtained in Comparative Example 1. The results showed that the purity of the barite product was 85.3%, the actual recovery amount was 16.2513 kg, and the total recovery rate of barite was 76.9%; the purity of the pyrite concentrate was 78.6%, the actual recovery amount was 2.751 kg, and the total recovery rate of pyrite was 57.3%; the viscosity of the tailings was 12.0 mPa·s, and the residual CaCO3 content reached 7.5%. Due to the omission of the second-stage acid leaching, calcite was not completely removed. The residual calcite not only increased the viscosity of the tailings but also competed with barite and pyrite for reagents during the flotation process, resulting in a decrease in separation efficiency and a significantly lower product purity and recovery rate than in Example 1.

[0045] Comparative Example 2 (omitting the re-selection step after acid leaching, only two stages of acid leaching-flotation process) The "two-stage acid leaching-direct flotation" process is used to recover barite and pyrite from lead-zinc tailings, omitting the gravity separation step after acid leaching. The specific steps are as follows: Two-stage acid leaching treatment: Take 100.0 kg of the above lead-zinc tailings and complete the first-stage acid leaching and second-stage acid leaching operations according to the parameters of Example 1 to obtain 52.3 kg of second-stage acid leaching residue (CaCO3 content 0.8%, viscosity 8 mPa·s).

[0046] Direct flotation: The gravity separation step after acid leaching is omitted. The second-stage acid leaching residue is directly washed, pH adjusted, pyrite flotated and barite purified. All flotation and purification parameters are completely consistent with those in Example 1.

[0047] X-ray fluorescence spectroscopy was performed on the barite product and pyrite concentrate obtained from Comparative Example 2. The results showed that the purity of the barite product was 90.2%, the actual recovery amount was 18.0222 kg, and the total recovery rate of barite was 84.0%; the purity of the pyrite concentrate was 85.5%, the actual recovery amount was 3.249 kg, and the total recovery rate of pyrite was 67.7%; the SiO2 content in the tailings reached 28.3%. Because the gravity separation step after acid leaching was omitted, undissolved silica gangue minerals such as quartz directly entered the flotation system, resulting in a decrease in product purity. This fully demonstrates the key role of gravity separation after acid leaching in separating silica gangue and improving product purity.

[0048] Comparison Table of Effects between Examples and Comparative Examples

[0049] Comparative Analysis: Examples 1 and 2 employ the complete process of the present invention, namely "two-stage acid leaching - gravity separation after acid leaching - flotation - barite purification". Compared with Comparative Example 1 (omitting the two-stage acid leaching), the purity of barite is increased by 12.2-12.5 percentage points, and the recovery rate is increased by 17.1-18.1 percentage points; the purity of pyrite is increased by 14.6-15.2 percentage points, and the recovery rate is increased by 24.3-25.7 percentage points; the calcite removal rate is increased by 39.3-40.3 percentage points, and the tailings viscosity is reduced by 6.5-7.5 mPa·s. This fully demonstrates the key role of two-stage acid leaching in thoroughly removing calcite and reducing tailings viscosity. Compared to Comparative Example 2 (which omitted the re-separation after acid leaching), the purity of barite in Examples 1 and 2 increased by 7.3–7.6 percentage points, and the recovery rate increased by 10.0–11.0 percentage points; the purity of pyrite increased by 7.7–8.3 percentage points, and the recovery rate increased by 13.9–15.3 percentage points. This indicates that re-separation after acid leaching can effectively separate silica-containing gangue minerals, further improving product purity. In summary, this invention, through the synergistic cooperation of various process steps, successfully overcomes the shortcomings of existing technologies, such as the difficulty in separating lead-zinc tailings and the low recovery rate, achieving simultaneous recovery of high-purity barite and pyrite, and full-component purification and utilization of lead-zinc tailings.

[0050] Finally, it should be noted that the above-listed embodiments are merely preferred embodiments of the present invention. Of course, those skilled in the art can make modifications and variations to the present invention. If such modifications and variations fall within the scope of the claims of the present invention and their equivalents, they should be considered as being within the protection scope of the present invention.

Claims

1. A method for comprehensive utilization of sulfide tailings, characterized in that, Includes the following steps: First-stage acid leaching treatment: The lead-zinc tailings are mixed with the first acid solution in a certain proportion, stirred and reacted, and after the reaction is completed, solid-liquid separation is carried out to obtain first-stage acid leaching residue and first-stage acid leaching solution; Two-stage acid leaching treatment: The first-stage acid leaching residue and the second acid solution are mixed in proportion, stirred and reacted, and after the reaction is completed, solid-liquid separation is performed to obtain the second-stage acid leaching residue and the second-stage acid leaching solution. Acid leaching followed by gravity separation: The acid leaching residue from the two stages is mixed with water to prepare a gravity separation slurry, which is then subjected to gravity separation treatment. The feed concentration, shaking table stroke, number of strokes, and water volume are controlled. By utilizing the density difference between barite, pyrite, and undissolved calcite and quartz, the undissolved silica gangue minerals are further separated to obtain gravity concentrate and gravity tailings. Flotation pretreatment: The gravity concentrate is washed with water until the pH is 5.5~7.0, and then deionized water is added to prepare a slurry with a mass fraction of 20%~40%. A pH adjuster is added to the slurry to adjust the pH to 6.0~8.0 to obtain the pretreated slurry. Pyrite flotation: Collector and frother are added to the pretreated slurry, and flotation is carried out to obtain pyrite concentrate and flotation tailings; Barite purification: The flotation tailings are filtered and dried to obtain barite products.

2. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The lead-zinc tailings contain 30% to 50% barite, 3% to 6% pyrite, and 10% to 20% calcite, with particles ≤0.074mm accounting for ≥60%.

3. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The ratio of the lead-zinc tailings to the first acid solution is 1g:(2~5)mL; and / or, The ratio of the first acid leaching residue to the second acid solution is 1g:(2~5)mL; and / or, The stirring temperature for the first-stage acid leaching treatment and the second-stage acid leaching treatment is 20℃~50℃, the stirring rate is 200~400r / min, and the reaction time is 1~1.5h.

4. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The first acid solution is dilute hydrochloric acid with a mass concentration of 6% to 9%; and / or, The second acid solution is dilute hydrochloric acid with a mass concentration of 9.5% to 12.5%.

5. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The mass fraction of the gravity separation slurry is 25% to 35%, and the gravity separation operation is carried out on a shaking table with a stroke of 10 mm to 18 mm, a stroke rate of 220 r / min to 320 r / min, and a water flow rate of 8 L / min to 18 L / min.

6. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The pH adjuster is any one or a mixture of sodium carbonate, sodium bicarbonate, or sodium hydroxide.

7. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The gravity concentrate consists of barite and pyrite, while the gravity tailings mainly contain undissolved calcite and quartz.

8. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The collector is any one or a mixture of butyl xanthate, pentyl xanthate, or isobutyl xanthate; and / or The collector is added at a rate of 80-120 g / t slurry; and / or, The foaming agent is any one or a mixture of No. 2 oil, pine oil, or methyl isobutyl methanol; and / or, The amount of foaming agent added is 10~50g / t slurry.

9. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The flotation process includes roughing, 1 to 3 cleaning stages, and 1 to 2 scavenging stages. The roughing concentration is 20% to 30% and the time is 3 to 8 minutes; the cleaning concentration is 20% to 25% and the cleaning time is 2 to 5 minutes; the scavenging concentration is 20% to 25% and the time is 2 to 5 minutes.

10. The method for comprehensive utilization of sulfide tailings according to claim 1, characterized in that, The drying temperature is 90~120℃, and the drying time is 2~6h.

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