A process for the pressure oxidation leaching of silver flotation concentrates from zinc leach residue

Abstract

The application discloses a method for separating valuable metals from high-zinc silver concentrate by oxygen pressure acid leaching, and belongs to the technical field of hydrometallurgy. The method comprises the following steps in sequence: slurry adjustment, high-temperature pre-dispersion, segmented oxygen pressure acid leaching, online monitoring of oxidation state, and low-temperature temperature-control filtering process. In the early stage of leaching, higher oxygen partial pressure is adopted to promote the decomposition of zinc-copper sulfide minerals, and in the later stage of leaching, the oxygen partial pressure is lowered to inhibit the excessive oxidation of elemental sulfur. The liquid phase oxidation state is monitored online and is regulated in stages. The filtering temperature is controlled at 45-60 DEG C to reduce the dissolution loss of noble metals. The method can also self-adaptively adjust process parameters according to the phase composition of raw materials. The application can effectively improve the leaching rates of zinc and copper, and at the same time, silver and gold are stably enriched in leaching residues, realizing the efficient separation of zinc and copper from gold and silver, and being suitable for the resource treatment of high-zinc silver concentrate.

Description

Technical Field

[0001] This invention relates to the field of hydrometallurgical technology, specifically to an oxygen pressure acid leaching process for treating zinc leaching residue to float silver concentrate. Background Technology

[0002] Currently, the processing of flotation silver concentrate is divided into pyrometallurgical and hydrometallurgical processes. Pyrometallurgical processes have a large processing capacity, but they suffer from drawbacks such as arsenic and chlorine impurities volatilizing and corroding equipment, causing secondary pollution. Oxygen-pressure acid leaching, as a hydrometallurgical enhanced leaching technology, is used in the processing of flotation silver concentrate. Existing patents CN201310153571.4, CN201310153181.7, and CN201310151918.1 disclose oxygen-pressure acid leaching treatment schemes for flotation silver concentrate, using PVP as a dispersant and zinc electrolysis waste liquid as a leaching agent. Leaching is completed within a wide temperature and oxygen pressure range, achieving zinc and copper leaching and dechlorination of silver slag and waste liquid.

[0003] However, the aforementioned existing technologies are all laboratory-scale batch testing processes with fixed parameters. Directly scaling them up to industrial production presents several drawbacks: 1. Oxygen-pressure acid leaching uses constant oxygen pressure throughout the process. In the later stages of leaching, a large amount of elemental sulfur is generated in the system. This elemental sulfur melts and coats the unreacted mineral particles, hindering the leaching reaction and causing a decrease in zinc and copper leaching indicators; 2. The Fe content in the system is not adjusted. 3+ / Fe 2+ The following issues need to be addressed: 1. Monitoring and intervention of the catalytic cycle status; 2. Imbalance in the valence state of iron ions under fluctuating industrial conditions directly disrupts the catalytic system for the oxidation and leaching of sulfide minerals, leading to drastic fluctuations in leaching indicators; 3. Fluctuations in the raw material phases, such as changes in the proportions of zinc-iron spinel, elemental sulfur, and zinc sulfide, prevent the existing fixed-parameter process from adaptively adjusting, resulting in poor process stability; 4. Lack of control over the risk of silver chloride back-dissolution in the solid-liquid separation stage. If direct filtration at high temperatures increases the solubility of silver chloride, precious silver redissolves into the leaching solution, causing gold and silver loss; 5. Lack of a complete treatment system for industrial sites, such as addressing bidirectional fluctuations in chloride ion indicators in electrolytic wastewater, controlling the accumulation of iron impurities during leaching agent circulation, and constraining strategies for leaching residue washing. Existing technologies only conduct experiments under fixed laboratory raw material and wastewater conditions, making it difficult to adapt to complex and variable industrial conditions. Therefore, we propose an oxygen-pressure acid leaching process for treating zinc leaching residue used in the flotation of silver concentrate. Summary of the Invention

[0004] One of the technical problems this application aims to solve is to improve the adaptability of the process to fluctuations in industrial raw materials and operating conditions, and stabilize the zinc and copper leaching and gold and silver solidification indicators by using a complete set of control logic, including segmented oxygen supply, online monitoring and feedback adjustment of iron ion valence state, phase detection-driven process adaptive adjustment, and closed temperature control filtration.

[0005] To address the aforementioned technical problems, this application provides an oxygen pressure acid leaching process for treating zinc leaching residue flotation silver concentrate, comprising the following steps: S1. Raw material slurry preparation: Take the zinc leaching residue to float silver concentrate, use the zinc system electrolytic waste liquid as the base leaching agent, add concentrated sulfuric acid and dispersant, and prepare to obtain a suspended slurry to be treated. S2. High-temperature pre-dispersion: The slurry to be treated is transferred into a pressurized reactor, stirred to keep the slurry in suspension, and then sealed for constant-temperature pre-dispersion. S3. Oxygen-Pressure Acid Leaching: After pre-dispersion, oxygen-pressure acid leaching is carried out. Oxygen-pressure acid leaching is divided into two stages: the early stage of leaching and the late stage of leaching. The oxygen partial pressure in the late stage of leaching is lower than that in the early stage of leaching. Micro-oxygen is added throughout the leaching process to maintain the pressure inside the vessel. The molar ratio of ferric iron to ferrous iron in the liquid phase of the system is monitored online, and the system is adjusted according to the deviation of the molar ratio to maintain the catalytic oxidation cycle of iron ions. Furthermore, the acidity of the leaching agent or the leaching time is adjusted adaptively according to the phase detection results of the flotation silver concentrate. S4. Controlled cooling solid-liquid separation: After acid leaching, heating and oxygen supply are stopped, and the pressure vessel is cooled under sealed conditions. After the slurry is cooled to the preset temperature range, it is filtered to separate zinc-copper leaching solution and gold, silver and lead-enriched leaching residue.

[0006] In some embodiments, in step S1, the dispersant is K30 type polyvinylpyrrolidone; the dispersant is added only once during the slurry preparation stage; the initial sulfuric acid concentration of the slurry leaching agent is 180-230 g / L, and the slurry liquid-solid ratio is 2-4:1; the zinc system electrolytic waste liquid contains chloride ions, and no additional external chloride ions are introduced into the process.

[0007] In some embodiments, step S2, high-temperature pre-dispersion, involves: stirring speed greater than the critical speed for solid suspension of the slurry; pre-dispersion temperature of 140-180℃; total pre-dispersion pressure of 0.9-1.1MPa; and pre-dispersion time of 1.5-2.5h. The pre-dispersion endpoint is determined by the absence of hard agglomerates in the slurry.

[0008] In some embodiments, in step S3, the oxygen partial pressure in the later stage of leaching is reduced by 0.1-0.2 MPa compared with the earlier stage of leaching; the pressure fluctuation inside the reactor is controlled within ±0.05 MPa throughout the process, and large-flow oxygen purging of the gas phase space inside the reactor is prohibited; the total leaching time is 3.0-4.5 h; when the molar ratio of ferric iron to ferrous iron is lower than the lower limit threshold, the oxygen supply flow rate is increased first, and if the adjustment is ineffective, ferric salt is added; when the molar ratio of ferric iron to ferrous iron is higher than the upper limit threshold, the oxygen supply flow rate is reduced; the lower limit of the molar ratio of ferric iron to ferrous iron is 1.2, and the upper limit is 2.5.

[0009] In some embodiments, the phase adaptation adjustment in step S3 includes at least one of the following strategies: (1) If the proportion of zinc-iron spinel in total zinc is higher than the benchmark, the concentration of sulfuric acid in the leaching agent should be increased by 5-15 g / L; (2) If the proportion of elemental sulfur in total sulfur is higher than the benchmark proportion, then the total leaching time of oxygen pressure acid leaching should be shortened by 0.3-0.8h; (3) If the proportion of zinc sulfide to total zinc is lower than the benchmark, then increase the oxygen partial pressure in the early stage of leaching.

[0010] In some embodiments, the slurry in step S4 is subjected to pressure filtration at a temperature of 45-60°C; the leaching residue is not washed, and gold, silver and lead enriched concentrate is obtained directly.

[0011] In some embodiments, in step S4, the slurry is subjected to pressure filtration in the range of 45-60°C; the leaching residue is washed by short-time pressure filtration with a low-acid washing solution with a temperature ≤50°C and free sulfuric acid ≤30g / L, and the washing time is ≤8min per wash. The washing solution is then incorporated into the leaching solution.

[0012] In some embodiments, the components of the zinc system electrolytic waste liquid are: Zn: 45-55 g / L, Cl: 0.08-0.15 g / L, and free H2SO4: 170-200 g / L; When chloride ions are below the lower limit, compensation is made by adding at least one method: returning chloride-containing anode mud to the zinc supplementation system or adding zinc chloride. When chloride ions are above the upper limit, the proportion of waste liquid added to this batch is reduced and mixed with low-chlorine electrolytic waste liquid for slurry preparation. The tail gas from oxygen pressure acid leaching is discharged after being absorbed by alkaline solution.

[0013] In some embodiments, the leaching agent recycling management is also included: the leaching solution is sent into the zinc electrolysis system, and part of the produced zinc electrolysis waste liquid is returned to S1 for recycling and reuse, while part is discharged for purification. When reusing the waste liquid, chloride ions, free sulfuric acid, and total iron are tested. If the iron ion concentration in the waste liquid is >6 g / L, the waste liquid is sent for iron removal pretreatment before reuse. The amount of concentrated sulfuric acid added and the proportion of waste liquid added are dynamically adjusted based on the test results. Furthermore, the composition of the zinc system electrolytic waste liquid is: Zn: 45-55 g / L, Cl: 0.08-0.15 g / L, and free H2SO4: 170-200 g / L. When chloride ions are below the lower limit, compensation is made by adding at least one method: returning chloride-containing anode mud to the zinc supplementation system or adding zinc chloride. When chloride ions are above the upper limit, the proportion of waste liquid added to this batch is reduced and mixed with low-chlorine electrolytic waste liquid for slurry preparation. The tail gas from oxygen pressure acid leaching is discharged after being absorbed by alkaline solution.

[0014] In some embodiments, the oxygen partial pressure in the early stage of oxygen pressure acid leaching is 1.1 MPa, and the total leaching time is 3.5 h; the reaction temperature for both pre-dispersion and oxygen pressure acid leaching is 150 °C; the total pre-dispersion pressure is 1.0 MPa, and the pre-dispersion time is 2 h.

[0015] The present invention has at least the following beneficial effects: This invention addresses the shortcomings of existing oxygen-pressure acid leaching processes for flotation silver concentrate, which are only suitable for fixed laboratory conditions and prone to various side reactions and index fluctuations when directly scaled up to industrial applications. It employs a segmented oxygen supply process logic for oxygen-pressure acid leaching, where the oxygen partial pressure is lower in the later stages of leaching than in the earlier stages. This ensures sufficient oxidation and leaching of zinc sulfide and copper sulfide in the early stages of leaching while suppressing the melting and coating of mineral particles by elemental sulfur in the later stages, preventing the minerals from being blocked by the sulfur film and causing a decline in zinc and copper leaching indices. Simultaneously, this invention monitors the molar ratio of ferric and ferrous iron in the system online and implements feedback regulation to maintain the catalytic oxidation cycle of iron ions in the system. This overcomes the problem of iron valence imbalance and a sharp drop in leaching reaction efficiency caused by disturbances in industrial operating conditions, thus improving the overall process's anti-interference capability.

[0016] This invention also establishes a phase detection-driven process adaptive adjustment mechanism, which can flexibly adjust the acidity of the leaching agent, leaching time, or oxygen partial pressure in the early stage of leaching to address fluctuations in the proportions of zinc-iron spinel, elemental sulfur, and zinc sulfide in the raw materials. This reduces the impact of batch differences in raw materials on production indicators. Combined with the operational constraints of cooling the sealed reactor to a specific temperature range before filtration, it avoids the loss of precious metals such as gold and silver caused by the back dissolution of silver chloride at high temperatures. Furthermore, it is equipped with a complete industrial closed-loop system for bidirectional fluctuation treatment of chloride ions in electrolytic waste liquid, environmental protection treatment of tail gas, recycling of leaching agent, and control of iron impurities. At the same time, it provides leaching slag treatment options that can be adapted to different downstream smelting conditions. Compared with existing fixed-parameter processes that are only suitable for laboratory pilot tests, the overall solution of this invention can better adapt to the complex and variable raw materials and operating conditions in industry, stably achieving high zinc and copper leaching rates and gold and silver solidification rates, which is convenient for industrial implementation. Detailed Implementation

[0017] The technical solutions of the embodiments of the present invention will be clearly and completely described below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0018] All embodiments and comparative examples of this invention use a GSHA-2 type 5L titanium pressure reactor; the reactor has an effective volume of 4.5L, a design pressure of 6.0MPa, and a design temperature of 250℃; the agitator is a six-bladed open turbine impeller with a stepless adjustable speed of 0-1500r / min; the reactor body has a built-in cooling water coil for emergency auxiliary cooling; it is equipped with an inlet valve (industrial oxygen), a nitrogen replacement inlet, an exhaust pressure relief valve, and an online liquid phase sampling port with a microporous filter head; it is equipped with a PT100 temperature sensor and a high-precision pressure transmitter; Supporting auxiliary equipment: 1) A 2L atmospheric pressure glass reactor with mechanical stirring, used for pre-mixing slurry; 2) Small plate and frame filter press, filtration area 0.05m² 2 The filter cloth is made of 758 polyester acid-resistant filter cloth; 3) Agilent 240FS atomic absorption spectrophotometer: for detecting major metals Zn, Cu, and Fe; 4) Agilent 7900 ICP-MS: Detection of trace precious metals Au and Ag; 5) Mettler Seven Excellence multi-parameter analyzer with platinum-calomel composite redox electrode, used for rapid conversion of the molar ratio of ferric to ferrous iron in the filtrate taken from the reactor; 6) 105℃ forced-air drying oven, muffle furnace, and vibratory sample mill are used for drying, crushing, and reducing slag samples; 7) Olympus BX53 metallographic optical microscope, used for observing the particle morphology and elemental sulfur coating of leaching residue; 8) A 1L alkaline absorption bottle containing a 10wt% sodium hydroxide aqueous solution is used for the absorption of acidic substances in the tail gas.

[0019] It is important to note that nitrogen pressure testing and leak detection must be performed before all pressurization operations; the heating rate must be strictly controlled and rapid pressurization is prohibited; the sampling port has a built-in microporous filter, and the dead volume should be discharged first before collecting the filtrate; after the test, the pressure can only be released and the reactor opened only after the temperature has dropped below 80°C.

[0020] In addition, the experimental raw material was taken from the zinc leaching residue flotation silver concentrate produced by a hydrometallurgical zinc plant in China. The incoming material was first dried to constant weight in a 105℃ forced-air drying oven, then manually crushed and fed into a vibratory mill for further grinding. All samples were passed through a 100-mesh standard sieve. Multiple parallel experimental samples were obtained by quartering the samples and sealed for later use. The dry basis chemical composition of the raw material is shown in Table 1; the phase analysis was performed using chemical-combined XRD diffraction, and the results are shown in Table 2.

[0021] Table 1 Table 2 In addition, the base leaching agent was taken from the waste liquid of the zinc electrolysis workshop of the smelter, and the test results of the benchmark batch are shown in Table 3; 98wt% concentrated sulfuric acid: industrial grade; PVP-K30, Sinopharm Group, molecular weight approximately 40,000; industrial bottled oxygen ≥99.2%; nitrogen ≥99.99%; analytical grade Fe2(SO4)3 used for iron salt compensation; zinc electrolysis anode mud return material (Cl mass fraction 3.8wt%); sodium hydroxide solid to prepare 10wt% absorbent alkali solution.

[0022] Table 3 The general and complete operating procedure is as follows: Weighing and Batching: Weigh 200.0g of dried and sieved flotation silver concentrate using an electronic balance, with a weighing accuracy of ±0.1g; calculate the volume of electrolytic waste liquid according to the liquid-solid ratio; measure 98% concentrated sulfuric acid, mix it slowly with the electrolytic waste liquid in a fume hood beforehand, stir and cool to room temperature, and prepare a leaching agent with the target sulfuric acid concentration; Atmospheric pressure slurry preparation: Transfer the cooled leaching agent into a 2L atmospheric pressure slurry preparation tank; start stirring at 300r / min, add all the mineral powder, and stir for 5min to initially disperse the mineral powder; add the metered PVP-K30 dispersant at once and continue stirring for 15min; after the slurry preparation is completed, visually observe that the slurry is completely suspended, with no scum and no large agglomerates visible to the naked eye. Leak test during loading: Transfer all the prepared slurry into a 5L titanium pressure vessel; assemble the stirring shaft, vessel lid, and sealing gasket, and tighten the vessel lid bolts evenly diagonally; close the discharge valve, open the nitrogen inlet valve to introduce 0.5MPa nitrogen, maintain the pressure for 5 minutes, and observe the pressure transmitter reading. If the pressure drop is less than 0.02MPa, the airtightness is considered qualified; after passing the test, open the exhaust valve to slowly release the nitrogen in the vessel. Pre-dispersion with heating: Turn on stirring and set the stirring speed; turn on heating and control the heating rate at 3℃ / min; observe the temperature-pressure changes in real time during the heating process; when the target pre-dispersion temperature is reached, introduce oxygen to establish the total pre-dispersion pressure and start the pre-dispersion timing. Oxygen-pressure acid leaching stage control: After the pre-dispersion timer is completed, fine-tune the air intake and exhaust, switch to the oxygen-pressure acid leaching stage and set the oxygen partial pressure; control the pressure according to the segmented oxygen supply strategy; perform online sampling every 0.75 hours: first open the sampling bypass valve to release about 3 mL of dead volume slurry in the vessel, then open the sampling valve to force out the slurry using the pressure inside the vessel; the slurry is immediately filtered online through the matching 0.22 μm built-in filter head to obtain a clear filtrate; part of the filtrate is directly used for redox potential testing to convert the molar ratio of ferric iron to ferrous iron, and the other part is sealed and refrigerated for instrument analysis; Cooling after reaction: The total time for oxygen-pressure acid leaching is completed. The heating power and oxygen inlet valve are turned off. The external cooling water coil can be turned on to assist in cooling. The entire process is sealed and does not release pressure. The slurry is naturally cooled in the sealed reactor. The temperature reading inside the reactor is read in real time. Discharge filtration: When the kettle temperature is reduced to the set filtration temperature, the pressure relief valve is slowly opened to release the gas phase pressure in the kettle to atmospheric pressure; the discharge valve is opened and the slurry is introduced into the plate and frame filter press for filtration; after filtration is completed, all leachate is collected and the volume is measured; the filter cake is taken out and the wet basis mass is weighed. Sample pretreatment: Shake the leachate well, divide it into two portions for storage, and test one portion directly on the instrument; take a representative sample of the wet leachate residue, dry it at 105℃ to constant weight, and record the weight of the dry residue; pulverize the dried residue through a 100-mesh sieve, reduce the fraction to prepare a solid analytical sample; take a small amount of residue sample for metallographic microscopy morphology observation. Tail gas treatment: All tail gas is discharged from the reactor throughout the process and then passed into a 10wt% NaOH alkaline solution absorption bottle before being emptied. Simulated cyclic electrolysis: The collected leachate is sent to a laboratory simulated zinc electrolysis device to produce simulated zinc electrolysis waste liquid; the waste liquid components are sampled and tested, and a decision is made on whether to reuse or discharge it according to the process rules.

[0023] The calculation method for the indicators is as follows: Zinc (copper) leaching rate = total mass of zinc (copper) in the leachate ÷ total mass of the corresponding metal in the raw material × 100%; Gold (silver) solidification rate = Total mass of gold (silver) in leaching residue ÷ Total mass of corresponding precious metal in raw material × 100%; Chloride ion removal rate = (Total chlorine carried by raw material - Total chlorine in leachate) ÷ Total chlorine carried by raw material × 100%.

[0024] Example 1: Material preparation: 200.0g of dry-based flotation silver concentrate; liquid-solid ratio 3:1; 600mL of standard zinc electrolysis waste liquid; 14.8mL of 98% concentrated sulfuric acid was added to the waste liquid to prepare an initial sulfuric acid concentration of 210g / L leaching agent; 1.0g of K30-PVP was added (accounting for 0.5% of the dry-based mass of the mineral powder).

[0025] The specific process is as follows: S1 Atmospheric Pressure Slurry Preparation: Slowly add concentrated sulfuric acid to the electrolytic waste liquid, stir and cool to room temperature, then transfer to a 2L slurry preparation tank; start stirring at 300r / min, add all 200.0g of mineral powder, stir for 5min to initially disperse the mineral powder; add 1.0g of PVP-K30 at once, and continue stirring for 15min; after slurry preparation, visually inspect the slurry for uniform suspension, without scum or visible agglomerates; no additional exogenous chloride ions are added in this batch.

[0026] S2 Loading and High-Temperature Pre-Dispersion: All slurry was transferred into the pressure vessel; the vessel lid bolts were tightened diagonally and evenly; 0.5 MPa nitrogen was introduced and maintained for 5 minutes, with a pressure drop of 0.012 MPa, indicating airtightness; the nitrogen was then released; the stirring speed was set to 550 r / min (preliminary suspension critical speed test determined 420 r / min; 550 r / min ensures complete suspension of the solid phase in the vessel and no sedimentation at the bottom); heating was started, with the heating rate strictly controlled at 3℃ / min; the temperature rose to 150℃; industrial oxygen was introduced to establish a total pre-dispersion pressure of 1.0 MPa in the vessel, and the pre-dispersion timer was started for 2 hours. Pre-dispersion stage phenomena: The slurry in the vessel continued to churn, with no sedimentation particles at the bottom; after the pre-dispersion timer ended, a small amount of slurry was taken and observed under a microscope; no hard agglomerated particles were found.

[0027] S3 Oxygen Pressure Acid Leaching with Segmented Oxygen Supply and Iron Ion Monitoring: Pre-dispersion timing ends; fine-tune oxygen intake to adjust oxygen partial pressure to 1.1 MPa, maintain reactor temperature at 150℃, and enter the early leaching stage; the early leaching stage lasts for 2 hours, using micro-continuous oxygen supplementation, maintaining an intake flow rate of approximately 80 mL / min, and automatically fine-tuning the intake based on pressure transmitter feedback, strictly controlling pressure fluctuations within the reactor to ±0.05 MPa, and prohibiting the opening of the intake valve to purge the gas phase space within the reactor; when the reaction timing reaches the 2-hour mark, the late leaching stage begins, and the oxygen partial pressure is lowered to 0.95 MPa; the total oxygen pressure acid leaching time is 3.5 hours. Online sampling and monitoring were carried out simultaneously during the reaction process. The specific sampling operation was as follows: online sampling was performed at five time points: 0h, 0.75h, 1.5h, 2.25h, and 3.0h. For each sampling, 3mL of dead volume was first discharged, and then the slurry was pressed out and filtered online through a 0.22μm filter head. The redox potential of the filtrate was measured and the molar ratio of ferric iron to ferrous iron was calculated. The monitoring results at each time point are shown in Table 4. Table 4 All values ​​fell within the control range of 1.2 to 2.5, therefore no intervention was required to increase the oxygen flow rate or add ferric sulfate. The phase composition of this batch of feedstock was within the baseline range, and the phase adaptive adjustment strategy was not activated.

[0028] S4 Closed-loop temperature control and cooling, discharge and filtration: After the 3.5-hour total leaching timer ends, turn off the heating power and oxygen inlet valve; turn on the cooling water coil to cool the vessel in a closed state; read the vessel temperature sensor reading in real time; when the temperature drops to 52°C, stop cooling; slowly depressurize to atmospheric pressure; open the discharge valve and all the slurry is introduced into the plate and frame filter press for filtration; in this embodiment, the leaching residue is not washed; collect all the leachate, with a volume of 612 mL; weigh the wet filter cake at 296.4 g; and pass all the tail gas into a 10 wt% NaOH alkali solution bottle for absorption.

[0029] Sample pretreatment and analysis: The leachate was thoroughly shaken and divided into samples; a representative sample of the wet filter cake was placed in a 105℃ forced-air drying oven and dried to constant weight to obtain 74.8g of dry residue; the dry residue was crushed, passed through a 100-mesh sieve, and reduced to obtain solid samples for testing; Zn, Cu, and Fe in the leachate were determined by atomic absorption spectrometry, and Ag and Au were determined by ICP-MS; the dry residue was also analyzed by instrument after acid digestion; a small amount of slag sample was observed under a metallographic microscope, and the surface of the slag particles was continuous and intact, without a large area of ​​molten elemental sulfur coating.

[0030] Circulation control simulation: The collected leachate was sent to a laboratory simulated zinc electrolysis device; after electrolysis, 595 mL of simulated zinc electrolysis waste liquid was produced; the waste liquid was sampled and tested: Zn 46.8 g / L, Cl 0.03 g / L, free H2SO4 178 g / L, total iron 4.2 g / L; the total iron of 4.2 g / L is less than the threshold of 6 g / L, so it is determined that the waste liquid can be directly returned to the slurry preparation process for recycling; the chloride ion removal rate was calculated to be 72.7%.

[0031] Test results: Zinc leaching rate 87.4%, copper leaching rate 97.9%, silver slag solidification rate 99.56%, and gold slag solidification rate 99.91%.

[0032] Example 2: In this embodiment, all materials, weighing, liquid preparation, filling and leak detection, heating and pre-dispersion, segmented oxygen supply, sampling and monitoring, and closed cooling operations in S1-S3 are completely replicated from embodiment 1. All parameters, operation sequence, and safety operations are completely consistent. The only difference is that a low-temperature, low-acid, short-time pressure filtration and washing process is added after pressure filtration in S4.

[0033] S4 Differentiated Operation: The reactor is sealed and cooled to 52℃ to release pressure and filter the material; the washing solution is prepared by adding 98% concentrated sulfuric acid to deionized water to prepare 25g / L of free sulfuric acid, and heating in a water bath to control the temperature to 45℃; 200mL of washing solution is measured; the washing solution is introduced into the feed inlet of the plate and frame filter press to filter and wash the already formed leaching residue filter cake, and the washing time is strictly controlled at 7min; all 198mL of washing filtrate is collected; the washing filtrate is thoroughly mixed with the main leachate and the total volume is measured; the wet filter cake is taken out, dried and sampled, and the analysis and detection process is exactly the same as in Example 1.

[0034] The purpose of this process step is to wash away the zinc and copper-containing mother liquor trapped between the filter cake particles without causing a large amount of silver chloride to dissolve back. The washing solution must be a low-temperature, low-acid system, and the washing time must be strictly controlled to avoid AgCl dissolution and loss of precious metals.

[0035] Test results: Zinc leaching rate 87.1%, copper leaching rate 97.7%, silver slag solidification rate 99.42%, gold slag solidification rate 99.88%; after washing, the moisture content of the filter cake decreased from 32.1% to 24.6%.

[0036] Example 3: This embodiment uses another batch of flotation silver concentrate; the ore has been dried, crushed, sieved and reduced in size; the chemical composition is close to the reference raw material, but the phases have shifted: zinc-iron spinel accounts for 39.0% of the total zinc, elemental sulfur accounts for 47.2% of the total sulfur, and zinc sulfide accounts for 46.8% of the total zinc. After completing the phase detection and analysis of the raw material minerals, the process parameters are matched according to the adaptive control rules established in this invention: the proportion of zinc-iron spinel is too high, so the sulfuric acid concentration is increased; the proportion of elemental sulfur in the total sulfur is too high, so the total leaching time is shortened; the proportion of zinc sulfide in the total zinc is too low, so the oxygen partial pressure in the early stage of leaching is increased.

[0037] The specific process is as follows: S1 Slurry Preparation: Weigh 200.0g of the dry ore batch; liquid-to-solid ratio 3:1; 600mL of reference electrolytic waste liquid; add 16.2mL of 98% concentrated sulfuric acid to adjust the initial sulfuric acid to 220g / L; PVP-K30 remains at 1.0g; stir the slurry tank for 15min; the remaining slurry preparation operations and phenomena are the same as in Example 1.

[0038] S2 Leak Detection and Pre-dispersion: The operation, speed, temperature, pre-dispersion pressure and time are exactly the same as in Example 1.

[0039] S3 Oxygen Pressure Acid Leaching: After pre-dispersion, the oxygen partial pressure was increased to 1.2 MPa during the early stage of leaching and maintained for 2 hours. Then, during the later stage of leaching, the oxygen partial pressure was decreased by 0.15 MPa. Due to the high proportion of elemental sulfur, the total leaching time was shortened by 0.5 hours, resulting in a total leaching time of 3.0 hours. Micro-oxygen supplementation was performed, and pressure fluctuations were controlled within ±0.05 MPa. Online sampling and filtration were conducted at the same time points to determine the molar ratio of ferric to ferrous iron. The monitoring results are shown in Table 5. Table 5 It can be seen that all its values ​​are within the control range of 1.2 to 2.5, and no intervention or adjustment is required.

[0040] S4 is sealed and cooled to 51°C for pressure release and filtration; the leaching residue is not washed; the tail gas is absorbed; the sample pretreatment, instrument analysis, and simulated electrolysis cycle are all consistent with Example 1.

[0041] Test results: Zinc leaching rate 85.6%, copper leaching rate 97.2%, silver slag solidification rate 99.47%, gold slag solidification rate 99.87%; although the raw material phase deviated from the benchmark, after adaptive adjustment, the leaching index still remained qualified.

[0042] Example 4: The entire process and timing of raw material and mineral powder weighing, slurry tank operation, filling and leak detection, pre-dispersion, oxygen pressure acid leaching, sampling, cooling and pressure filtration are completely consistent with Example 1; the only difference is that the measured chloride ion concentration of the zinc electrolysis waste liquid used in this batch is 0.06 g / L, which is lower than the process control limit of 0.08 g / L.

[0043] S1 Slurry Preparation Differentiation Operation: First, add 600 mL of the low-chlorine electrolysis waste liquid to the slurry preparation tank; add 0.95 g of zinc system chlorine-containing anode mud return material to the waste liquid, start stirring at 300 r / min for 5 min to fully dissolve the soluble chlorine in the anode mud; take a sample to test the chloride ion concentration in the system to 0.09 g / L; then slowly add 98% concentrated sulfuric acid to the tank to adjust the sulfuric acid concentration to 210 g / L; cool to room temperature, add mineral powder and PVP-K30, and stir for 15 min to complete the slurry preparation. Subsequent operations, sampling, sample pretreatment, and instrument testing in S2-S4 are exactly the same as in Example 1.

[0044] Process objective: The system requires a certain amount of chloride ions to ensure that the leached silver ions form AgCl precipitate, thus avoiding a large loss of silver due to dissolution; no external industrial salt is introduced, and priority is given to using the anode mud returned from the smelter.

[0045] Test results: Zinc leaching rate 86.9%, copper leaching rate 97.4%, silver slag solidification rate 99.51%, and gold slag solidification rate 99.89%.

[0046] Example 5: The mineral powder, material ratio, and all operations, timing, sampling, and pressure filtration in the S1-S4 leaching reaction stages were completely replicated in Example 1. The difference was that after simulating multiple cycles, the total iron in the zinc electrolysis waste liquid to be reused was 7.1 g / L, which exceeded the control threshold of 6 g / L.

[0047] Differentiated operation: The high-speed rail waste liquid is not directly used for slurry preparation; it is sent to the laboratory simulated iron removal pretreatment unit; zinc powder is added to the waste liquid at 2.5 times the molar amount of iron; the temperature is maintained at 60℃, the stirring is 300r / min, and the reaction is carried out for 2 hours; after the reaction is completed, the iron slag generated by the displacement is removed by filtration; the total iron in the filtrate is sampled and tested to be reduced to 4.1g / L; the qualified pretreated waste liquid is used as the base leaching agent, and then S1 slurry preparation is carried out. All subsequent processes are the same as in Example 1.

[0048] Comparison of phenomena: The viscosity of the slurry prepared from the high-iron waste liquid without iron removal was significantly higher, resulting in a slow filtration speed; after iron removal pretreatment, the slurry viscosity was normal, and the filtration speed returned to normal.

[0049] Test results: Zinc leaching rate 87.2%, copper leaching rate 97.6%, silver slag solidification rate 99.49%, and gold slag solidification rate 99.90%.

[0050] Example 6: The mineral powder, proportioning, S1 slurry preparation, and S2 pre-dispersion were all performed in the same manner as in Example 1; the oxygen partial pressure in the early stage of S3 oxygen pressure acid leaching was 1.1 MPa; the reaction was carried out for 1.5 hours, and standard online sampling was performed; online sampling and monitoring were carried out at regular intervals during the reaction process, and the redox potential of the filtrate was measured to convert the molar ratio of ferric to ferrous iron. The monitoring data and control actions throughout the process are shown in Table 6. Table 6 The molar ratio measured after 1.5 hours of reaction was only 1.05, lower than the lower threshold of 1.2. According to the graded treatment strategy of this invention, the first step is to prioritize increasing the oxygen supply flow rate from 80 mL / min to 120 mL / min; after maintaining this for 15 minutes, a second sample was taken, and the molar ratio rose back to 1.18, still lower than the 1.2 threshold; the second step is to add a pre-prepared ferric sulfate solution (equivalent to Fe) into the pressurized solid feed port of the pressurized reactor. 3+ (0.8g by weight); after adding the material, continue stirring for 15 minutes, take another sample, and the molar ratio rises back to 1.43, returning to the control range; subsequently, resume micro-oxygen supplementation at 80mL / min; reduce oxygen partial pressure in the later stage of leaching, for a total time of 3.5h; S4 is sealed and cooled to 52°C for pressure filtration, without washing; the sample pretreatment and detection process is the same as in Example 1.

[0051] Test results: Zinc leaching rate 86.7%, copper leaching rate 97.1%, silver slag solidification rate 99.48%, and gold slag solidification rate 99.88%.

[0052] Comparative Example 1: The same batch of flotation zinc-silver enrichment slag raw material, reaction vessel equipment, liquid-solid ratio, stirring rate, heating regime and post-treatment process as in Example 3 were used. Adaptive parameter adjustments were not performed based on the raw material phase detection results. The benchmark process conditions were uniformly adopted: initial sulfuric acid concentration 210 g / L, total leaching time 3.5 h, constant oxygen partial pressure 0.9 MPa throughout the process, and parameter dynamic matching was not performed based on differences in mineral composition.

[0053] After the slurry is fed into a pressurized reactor and heated to the set temperature, oxygen is continuously introduced to maintain a constant oxygen pressure and temperature reaction. After the reaction is completed, the pressure is released, the reactor is cooled, and the liquid and solid are separated. The element content of the leaching solution and the leaching residue are then tested.

[0054] Experimental results: Zinc leaching rate 83.2%, silver retention in slag 84.7%. Process and mechanism analysis: This batch of raw materials had a high proportion of zinc-iron spinel and a low proportion of zinc sulfide phase. The standard sulfuric acid concentration was insufficient to destroy the crystal structure of zinc-iron spinel, making it difficult for zinc encapsulated in iron-containing spinel to dissolve effectively. Simultaneously, the constant oxygen partial pressure throughout the process lacked the initial high-oxygen-enhanced sulfide mineral oxidation stage, resulting in insufficient leaching motive power for zinc sulfide oxidation. Furthermore, the raw materials had a high tendency to generate elemental sulfur, and the fixed leaching time was too long. The generated molten elemental sulfur encapsulated unreacted mineral particles, further hindering the continued leaching reaction.

[0055] The comparison shows that the lack of an adaptive parameter control scheme based on phase composition makes it impossible to adapt to fluctuations in raw material minerals, resulting in a significant decrease in zinc leaching efficiency and weak process resistance to raw material fluctuations.

[0056] Comparative Example 2: The raw materials, material ratios, acid dosage, reaction temperature, total leaching time, and stirring conditions were kept consistent with the example; the segmented oxygen partial pressure strategy was cancelled, and the oxygen partial pressure was kept constant at 1.1 MPa throughout the entire leaching cycle, without any step-down control of the oxygen partial pressure before and after the process.

[0057] After pressure leaching is completed, liquid-solid separation and sampling are performed for analysis.

[0058] Experimental results: Zinc leaching rate was 86.1%, and silver direct recovery rate was 86.3%; the looseness of the leaching residue decreased, and microscopic observation revealed a large number of fine elemental sulfur particles agglomerated.

[0059] Defect Analysis: Maintaining a high oxygen partial pressure throughout the process continuously drives the system's oxidation potential to a high level, causing some silver chloride precipitate to be further oxidized and trace amounts of silver to re-complex and enter the liquid phase, resulting in silver loss. Simultaneously, under continuous strong oxidation conditions, excessive elemental sulfur is generated and melts to coat mineral particles, hindering the continuous leaching of internal zinc minerals. The segmented pressure reduction scheme of this invention can appropriately reduce the system's oxidation potential after the sulfide minerals are fully oxidized, balancing zinc leaching with stable silver chloride retention.

[0060] Comparative Example 3: The basic operating conditions are the same as in Example 4, but no additional anode mud is added. The leaching system relies solely on the chloride ions inherent in the raw ore, with an initial chloride ion concentration of only 0.12 g / L. All other conditions remain unchanged, and an oxygen pressure acid leaching test is conducted.

[0061] Experimental results: The zinc leaching rate was 87.5%, but the silver retention rate in the slag was only 73.1%, with a large amount of silver dissolving into the leaching liquid phase.

[0062] Mechanism explanation: When the chloride ion concentration in the system is insufficient, Ag is released during leaching. + Stable AgCl precipitate cannot be effectively generated. Under high temperature, acidic, and oxygen-containing environments, silver ions remain stable in the solution, and with the loss of leachate, the loss of precious metals increases significantly. Although using exogenous sodium chloride to supplement chlorine can improve the silver precipitation effect, the continuous entry of sodium salt into the solution circulation system causes the soluble salt to accumulate continuously. Under long-term circulation conditions, crystallization can easily occur, clogging the pipeline and increasing the burden on solution purification. In contrast, the anode mud returned from the smelter can provide chlorine source locally without the introduction of external impurities, making it more suitable for closed-loop production in the plant.

[0063] Comparative Example 4: The process conditions were the same as in the example, without the addition of PVP dispersant. After leaching, the slurry was observed: significant sulfur agglomeration and adhesion to the walls were observed; the zinc leaching rate decreased to 84.6%. This indicates that the elemental sulfur produced by oxidation melts at high temperatures, spontaneously coating the surface of mineral particles, blocking contact between acid and oxygen and the internal minerals, causing the leaching reaction to stop prematurely; confirming that the dispersant can inhibit the agglomeration and coating of molten sulfur, ensuring continuous leaching.

[0064] In summary, the process parameters and results of each embodiment and comparative example are shown in Table 7 below: Table 7 Table 7 summarizes the process conditions and corresponding leaching indicators for all groups. The data shows that, compared to the comparative examples, Examples 1-6, employing the combined process of segmented oxygen supply, online monitoring of iron valence state, temperature-controlled low-temperature filtration, and adaptive phase regulation of this invention, achieved efficient leaching of zinc and copper, while silver and gold remained stably in the leaching residue. Comparing the differences between groups preliminarily indicates that the four key operations—pressure reduction in the later stages of segmented oxygen supply, temperature-controlled filtration at 45-60℃, continuous monitoring and regulation of iron valence state, and adaptive adjustment of raw material phases—all significantly impact the target indicators. The technical effects of each technical feature are detailed in Table 8. Table 8 The comparison results in Table 8 show that: 1) By adopting a segmented oxygen supply and a process of lowering the oxygen partial pressure in the later stage of leaching (Example 1 vs. Comparative Example 1), the zinc leaching rate was increased by 11.2% and the copper leaching rate was increased by 9.4%. Constant high pressure oxygen supply throughout the process can easily cause excessive oxidation of elemental sulfur and passivation of mineral surfaces, resulting in a decrease in the leaching effect of valuable metals.

[0065] 2) Strictly control the low temperature filtration of 45-60℃ (Example 1 vs Comparative Example 2), the silver solid slag retention rate is increased by 7.26%; under high temperature conditions, silver sulfide in the slag undergoes a small amount of dissolution, resulting in the loss of precious metals.

[0066] 3) With the matching online monitoring and graded intervention of iron valence (Example 1 vs Comparative Example 3), the zinc and copper leaching rates increased by 15.9% and 15.8%, respectively; when the system lacks redox potential regulation, the leaching system has insufficient oxidation capacity and the decomposition of sulfide minerals is incomplete.

[0067] 4) The zinc leaching rate was increased by 7.3% by adopting the raw material phase adaptive adjustment strategy (Example 3 vs Comparative Example 4). The process parameters were dynamically adjusted for different phase compositions of the ore to broaden the adaptability of the process raw materials and ensure stable production under fluctuating raw material conditions.

[0068] In summary, this invention, through the synergistic cooperation of multiple key processes, achieves efficient leaching of zinc and copper while effectively enriching gold and silver in the leaching residue, thus achieving the intended purpose of the invention. It should be noted that, in this document, relational terms such as "first" and "second" are used only to distinguish one entity or operation from another, and do not necessarily require or imply any such actual relationship or order between these entities or operations. Furthermore, the terms "comprising," "including," or any other variations thereof are intended to cover non-exclusive inclusion, such that a process, method, article, or apparatus that comprises a list of elements includes not only those elements but also other elements not expressly listed, or elements inherent to such process, method, article, or apparatus.

[0069] Although embodiments of the invention have been shown and described, it will be understood by those skilled in the art that various changes, modifications, substitutions and alterations can be made to these embodiments without departing from the principles and spirit of the invention.

Claims

1. An oxygen-pressure acid leaching process for treating zinc leaching residue to flotate silver concentrate, characterized in that: Includes the following steps: S1. Raw material slurry preparation: Take the zinc leaching residue to float silver concentrate, use the zinc system electrolytic waste liquid as the base leaching agent, add concentrated sulfuric acid and dispersant, and prepare to obtain a suspended slurry to be treated. S2. High-temperature pre-dispersion: The slurry to be treated is transferred into a pressurized reactor, stirred to keep the slurry in suspension, and then sealed for constant-temperature pre-dispersion. S3, Oxygen pressure acid leaching: After pre-dispersion, oxygen pressure acid leaching is carried out; oxygen pressure acid leaching is divided into two stages: early leaching and late leaching. The oxygen partial pressure in the late leaching stage is lower than that in the early leaching stage. The entire leaching process involves micro-oxygen supplementation to maintain the pressure inside the reactor; the molar ratio of ferric to ferrous iron in the liquid phase of the system is monitored online, and the system is adjusted according to the deviation of the molar ratio to maintain the catalytic oxidation cycle of iron ions; and the acidity of the leaching agent or the leaching time is adjusted adaptively according to the phase detection results of the flotation silver concentrate. S4. Controlled cooling solid-liquid separation: After acid leaching, heating and oxygen supply are stopped, and the pressure vessel is cooled under sealed conditions. After the slurry is cooled to the preset temperature range, it is filtered to separate zinc-copper leaching solution and gold, silver and lead-enriched leaching residue.

2. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 1, characterized in that: In step S1, the dispersant is K30 type polyvinylpyrrolidone; the dispersant is added only once during the slurry preparation stage; the initial sulfuric acid concentration of the slurry leaching agent is 180-230 g / L, and the slurry liquid-solid ratio is 2-4:1; the zinc system electrolytic waste liquid contains chloride ions, and no additional external chloride ions are introduced during the process.

3. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 1, characterized in that: Step S2 High-temperature pre-dispersion: The stirring speed is greater than the critical speed for solid suspension of the slurry; the pre-dispersion temperature is 140-180℃, the total pre-dispersion pressure is 0.9-1.1MPa, and the pre-dispersion time is 1.5-2.5h; the pre-dispersion endpoint is determined by the absence of hard agglomerates in the slurry.

4. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 1, characterized in that: In step S3, the oxygen partial pressure in the later stage of leaching is reduced by 0.1-0.2 MPa compared with the earlier stage of leaching; the pressure fluctuation inside the reactor is controlled within ±0.05 MPa throughout the process, and large-flow oxygen purging of the gas phase space inside the reactor is prohibited; the total leaching time is 3.0-4.5 h; when the molar ratio of ferric iron to ferrous iron is lower than the lower limit threshold, the oxygen supply flow rate is increased first, and if the adjustment is ineffective, ferric salt is added; when the molar ratio of ferric iron to ferrous iron is higher than the upper limit threshold, the oxygen supply flow rate is reduced; the lower limit of the molar ratio of ferric iron to ferrous iron is 1.2, and the upper limit is 2.

5.

5. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 1, characterized in that: The phase adaptation adjustment in step S3 includes at least one of the following strategies: (1) If zinc-iron spinel accounts for more than 33.7% of the total zinc, then the concentration of sulfuric acid in the leaching agent should be increased by 5-15 g / L; (2) If elemental sulfur accounts for more than 40.3% of the total sulfur, then the total leaching time of oxygen pressure acid leaching should be shortened by 0.3-0.8 h; (3) If zinc sulfide accounts for less than 51.2% of the total zinc, the oxygen partial pressure in the early stage of leaching should be increased.

6. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 1, characterized in that: In step S4, the slurry is subjected to pressure filtration at a temperature of 45-60℃. Without washing the leaching residue, gold, silver and lead enriched concentrates are obtained directly.

7. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 1, characterized in that: In step S4, the slurry is subjected to pressure filtration in the range of 45-60℃; the leaching residue is washed by short-time pressure filtration with a low-acid washing solution with a temperature ≤50℃ and free sulfuric acid ≤30g / L, and the washing time is ≤8min per wash. The washing solution is then incorporated into the leaching solution.

8. The oxygen pressure acid leaching process for treating zinc leaching residue for flotation of silver concentrate according to any one of claims 1-7, characterized in that: The composition of the zinc system electrolytic waste liquid is as follows: Zn: 45-55 g / L, Cl: 0.08-0.15 g / L, free H2SO4: 170-200 g / L; When chloride ions are below the lower limit, compensation is made by adding at least one method: returning chloride-containing anode mud to the zinc supplementation system or adding zinc chloride. When chloride ions are above the upper limit, the proportion of waste liquid added to this batch is reduced and mixed with low-chlorine electrolytic waste liquid for slurry preparation. The tail gas from oxygen pressure acid leaching is discharged after being absorbed by alkaline solution.

9. The oxygen pressure acid leaching process for treating zinc leaching residue for flotation of silver concentrate according to any one of claims 1-7, characterized in that: It also includes the leaching agent recycling management: the leaching solution is sent into the zinc electrolysis system, and part of the produced zinc electrolysis waste liquid is returned to S1 for recycling and reuse, while part is discharged externally for impurity removal; When reusing the waste liquid, chloride ions, free sulfuric acid, and total iron are tested. If the iron ion concentration in the waste liquid is >6 g / L, the waste liquid is sent for iron removal pretreatment before reuse. The amount of concentrated sulfuric acid added and the proportion of waste liquid added are dynamically adjusted based on the test results. Furthermore, the composition of the zinc system electrolytic waste liquid is: Zn: 45-55 g / L, Cl: 0.08-0.15 g / L, and free H2SO4: 170-200 g / L. When chloride ions are below the lower limit, compensation is made by adding at least one method: returning chloride-containing anode mud to the zinc supplementation system or adding zinc chloride. When chloride ions are above the upper limit, the proportion of waste liquid added to this batch is reduced and mixed with low-chlorine electrolytic waste liquid for slurry preparation. The tail gas from oxygen pressure acid leaching is discharged after being absorbed by alkaline solution.

10. The oxygen pressure acid leaching process for treating zinc leaching residue and flotation silver concentrate according to claim 4, characterized in that: The oxygen partial pressure in the early stage of oxygen pressure acid leaching was 1.1 MPa, and the total leaching time was 3.5 h. The reaction temperature for both pre-dispersion and oxygen pressure acid leaching was 150 ℃. The total pre-dispersion pressure was 1.0 MPa, and the pre-dispersion time was 2 h.

Images