Antimony-gold concentrate flotation antimony removal process
By preparing arsenic-directed collectors, antimony-selective collectors, and lead chelating dispersants, and by stepwise regulating the slurry chemical environment and interfacial reactions, the problems of insufficient gold enrichment and antimony removal performance in existing antimony-gold concentrate flotation were solved, and the efficient preparation of high-purity gold concentrate was achieved.
Patent Information
- Authority / Receiving Office
- CN · China
- Patent Type
- Applications(China)
- Current Assignee / Owner
- GUANGDONG SHENGXIANG NEW MATERIAL TECH CO LTD
- Filing Date
- 2025-12-29
- Publication Date
- 2026-05-05
AI Technical Summary
In existing antimony-gold concentrate flotation technologies, it is difficult to further improve the enrichment and antimony removal performance of gold, and it is difficult to achieve high-selectivity interface differentiation of arsenic, antimony, and lead in stages. This results in limited improvement of gold grade, and impurities are prone to co-flotation and reflux, making it difficult to obtain high-purity gold concentrate.
By preparing arsenic-directed collectors, antimony-selective collectors, and lead-chelating dispersants, and by progressively controlling the slurry chemical environment and interfacial reactions, selective separation of arsenic, antimony, and lead is achieved. Specific steps include grinding, flotation, and pH adjustment, using specific collectors and dispersants for multi-stage flotation to form a stable interfacial separation layer.
This improved the floatability of gold, enhanced the interfacial cleanliness and responsiveness of gold, and achieved highly selective separation of arsenic, antimony, and lead, resulting in high-purity gold concentrate.
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Figure CN121972305A_ABST
Abstract
Description
Technical Field
[0001] This invention relates to the field of ore flotation technology, specifically to a process for removing antimony from antimony-gold concentrate flotation. Background Technology
[0002] Existing flotation technologies for antimony-containing gold concentrates mainly revolve around gold enrichment and selective antimony removal. Regarding gold enrichment, with the increasing proportion of fine-grained and encapsulated gold, traditional collecting systems are susceptible to interfacial coating and pulp chemical environment effects under polysulfide coexistence conditions, leading to a decrease in gold floatability. To improve gold interfacial exposure and adsorption selectivity, research has gradually developed methods such as multi-collector formulation, interfacial activation, and multi-stage enrichment to improve the flotation response of gold. Regarding antimony removal, since antimony minerals and gold have similar interfacial properties and are prone to co-flotation, related technologies often focus on controlling pulp pH, ionic composition, reagent structure, and process segmentation to differentiate the adsorption and bubble attachment behavior of antimony from that of gold, thereby constructing a controllable antimony removal window. Overall, gold enrichment and antimony removal technologies are developing towards more refined interfacial control and more synergistic process control.
[0003] In existing flotation processes for gold concentrates containing arsenic, antimony, and lead, insufficient interfacial control means often result in arsenic-containing minerals covering gold particles in the early stages, limiting the gold's interfacial exposure. Subsequently, the antimony phase exhibits strong competitive adsorption in the pulp, leading to instability in the liquid environment surrounding the gold and a decrease in interfacial responsiveness. Fine impurities are prone to secondary adsorption in later stages, resulting in insufficient gold surface cleanliness and making it difficult to continuously enhance overall floatability throughout the process, thus limiting further improvement in gold grade.
[0004] Meanwhile, existing processes struggle to achieve phased, highly selective interface differentiation of arsenic, antimony, and lead. Arsenic-containing minerals exhibit poor separation under near-neutral conditions and are prone to re-adhesion during the process. The interface properties of the antimony phase and gold are similar, resulting in a narrow separation window. The floatability of lead minerals fluctuates easily with changes in the slurry's chemical environment and is affected by the obscuring effect of other fine particles, making it difficult to form stable impurity removal layers. Multiple impurities may re-aggregate or co-float in the later stages of the process, limiting the acquisition of high-purity gold concentrate.
[0005] To address this technical deficiency, a solution is proposed. Summary of the Invention
[0006] The purpose of this invention is to provide a flotation process for antimony removal from antimony-gold concentrate, which addresses the technical problem that the gold enrichment and antimony removal performance of antimony-gold concentrate flotation in the prior art needs further improvement.
[0007] The objective of this invention can be achieved through the following technical solutions: A flotation process for removing antimony from antimony-gold concentrate includes the following steps: S1. Add gold concentrate and deionized water to the grinding equipment and wet grind until the 74μm particle size accounts for 70-80wt%. After grinding, adjust the pH of the slurry to 6.5-7.0 to obtain the slurry to be purified. S2. Add the slurry to be purified to the flotation cell, add the arsenic-directed collector, stir for 3-5 minutes, add methyl isobutyl methanol, float for 6-8 minutes, collect the foam as arsenic concentrate, and obtain the arsenic-removed slurry. S3. Add the arsenic-removed slurry to the mixing tank, adjust the pH to 7.0-7.5, add the antimony selective collector, stir for 3-5 minutes, then add methyl isobutyl methanol, float for 8-10 minutes, collect the foam as antimony concentrate, and obtain the arsenic-removed and antimony-removed slurry. S4. Add the arsenic- and antimony-removed slurry to the mixing tank, adjust the pH to 8.0-8.5, add lead chelating dispersant, stir for 5-8 minutes, then add methyl isobutyl methanol, float for 4-6 minutes, collect the foam as lead concentrate, and then process to obtain purified gold concentrate.
[0008] The reaction principle for preparing purified gold concentrate is as follows: After grinding, the surface of the minerals exposes active sites. Under near-neutral conditions, the arsenic-oriented collector preferentially forms a directional coordination structure with the arsenic-containing minerals, making their surfaces hydrophobic and achieving relatively independent interfacial separation. Subsequently, in a weakly alkaline environment, the antimony-selective collector, with its multi-site coordination ability of phosphoramide-dithio groups, forms a stable complex layer with the surface of antimony-containing minerals, making these minerals exhibit higher affinity for the air interface. Further increasing the pH of the system and introducing a lead chelating dispersant with multi-coordination groups alters the interfacial behavior of lead-containing minerals through surface chelation and dispersion effects, thereby creating differences in floatability with other minerals. Overall, this method, by stepwise controlling the slurry chemical environment and interfacial reactions, enables the sequential selective separation of associated components such as arsenic, antimony, and lead, providing an interfacial chemical basis for the subsequent enrichment of gold minerals.
[0009] Furthermore, in step S1, the ratio of gold concentrate to deionized water is 30-35 kg: 100 L, wherein the gold content is 0.004-0.06 wt%, the antimony content is 15-20 wt%, the arsenic content is 2-3 wt%, and the lead content is 1-2 wt%. Furthermore, in step S2, the ratio of the slurry to be purified, the arsenic-directed collector, and methyl isobutyl methanol is 15-25g:90L:2g. Furthermore, in S3, the ratio of arsenic-removed slurry, antimony selective collector, and methyl isobutyl methanol is 30-40g:80L:2g. Furthermore, in step S4, the ratio of the amount of arsenic- and antimony-removed slurry, lead chelating dispersant, and methyl isobutyl methanol is 20g:70L:2g. The post-treatment includes: filtering the slurry and collecting the filter cake, transferring the filter cake to a drying oven at 80°C and vacuum drying it to constant weight to obtain purified gold concentrate.
[0010] Furthermore, the arsenic-directed collector is prepared by the following method: A1. Add diethylphosphonite and anhydrous ethanol to a reaction vessel and stir. After mixing evenly, purge with nitrogen for protection and add acrylonitrile. Then heat the reaction vessel to 45-50℃ and stir for 3-4 hours. Post-treatment yields diethyl(2-cyanoethyl)phosphonite. A2. Hydroxylamine hydrochloride and deionized water are added to the reaction vessel and stirred until dissolved. The pH of the reaction system is adjusted to 6.5-7.5 using a 30-35 wt% sodium hydroxide aqueous solution, and the temperature of the reaction vessel is maintained at 20-30℃. Diethyl(2-cyanoethyl) phosphonate dispersion is added, and the mixture is kept at this temperature for 1-2 hours. Then, the temperature of the reaction vessel is raised to 45-50℃, and the mixture is kept at this temperature and stirred for 4-6 hours. The arsenic-directed collector is obtained by post-treatment.
[0011] The reaction principle for preparing arsenic-directed collectors is as follows: First, under suitable conditions, diethylphosphonite undergoes addition to the double bond of acrylonitrile, introducing a cyano group at the β-position of the phosphonite molecule to obtain diethyl(2-cyanoethyl)phosphonite with further reactivity. Subsequently, under near-neutral conditions, hydroxylamine undergoes nucleophilic attack on the cyano group and oxime transformation, introducing the original cyano group into an oxime group structure, forming a functionalized product with a phosphonic acid group and an oxime group dual coordination structure. This type of structure can exhibit high selectivity for arsenic-containing minerals in the slurry system. Its formation process is essentially the result of the continuous completion of the addition reaction after phosphonite activation and the oxime reaction of the cyano group, thus obtaining an arsenic-directed collector with specific selective recognition capabilities.
[0012] Further, in step A1, the ratio of diethylphosphonite, anhydrous ethanol and acrylonitrile is 25-30 mL: 200 mL: 30 mL. The post-treatment includes: after the reaction is completed, the solvent is removed by vacuum evaporation, the product is collected and placed in a drying oven at 50-60℃ for vacuum drying for 2-3 h to obtain diethyl(2-cyanoethyl)phosphonite. Further, in step A2, the ratio of hydroxylamine hydrochloride, deionized water, and diethyl(2-cyanoethyl)phosphonate dispersion is 12-16 g: 200 mL: 100 mL. The diethyl(2-cyanoethyl)phosphonate dispersion is obtained by mixing diethyl(2-cyanoethyl)phosphonate, anhydrous ethanol, and deionized water in a ratio of 20-25 mL: 80 mL: 20 mL. The post-treatment includes: adjusting the pH of the reaction system to 5-6 with 10-15 wt% hydrochloric acid aqueous solution after the reaction, allowing it to stand and cool, filtering and collecting the filter cake, rinsing the filter cake with deionized water until neutral, and placing it in a drying oven at 50-60℃ for vacuum drying for 4-6 h to obtain the arsenic-directed collector.
[0013] Furthermore, the antimony selective collector is prepared by the following method: B1. Add the polydimethylamine, anhydrous ethanol and deionized water to the reaction vessel and stir until dissolved. Cool the reaction vessel to 0-10℃ under ice bath conditions. Then add the calculated amount of carbon disulfide and use 30-35wt% sodium hydroxide aqueous solution to maintain the pH of the reaction system at 10-11. Keep the mixture warm and stir for 3-4 hours. Then heat the reaction vessel to 20-25℃ and keep it warm and stir for 0.5-1 hours. Post-processing yields sodium dithiocarbamate.
[0014] B2. Add sodium dithiocarbamate, deionized water and anhydrous ethanol to the reaction vessel and stir until dissolved. Then add anhydrous dichloromethane and purge with nitrogen for protection. Cool the reaction vessel to 0-5°C and add phosphorus oxychloride. After the addition is complete, stir for 1-2 hours. Then heat the reaction vessel to 25-30°C and keep it at that temperature for 3-4 hours. Post-processing yields the antimony selective collector.
[0015] The reaction principle for preparing antimony selective collectors is as follows: The amino group in the polyamine molecule first undergoes a nucleophilic attack on the electronegative carbon atom of carbon disulfide, generating a stable dithiocarbamate structure, thereby introducing a sulfur-coordinating functional group into the molecule. Subsequently, the dithiocarbamate undergoes a further phosphorylation reaction under the action of phosphorus oxychloride, with the nitrogen atom acting as the nucleophilic center to attack the phosphorus atom, causing the dithiocarbamate group to connect with the phosphoryl group, forming a phosphoramide structure with multi-site coordination characteristics. This reaction process is essentially a continuous transformation between the formation of dithiocarbamate and the phosphorylation reaction. Through the activation and reconstruction of nitrogen-sulfur-containing groups, the functional modification of the amine skeleton is achieved, resulting in an antimony selective collector with specific coordination characteristics.
[0016] Further, in step B1, the ratio of poly-reduced amine, anhydrous ethanol and deionized water is 18-20 g: 100 mL: 100 mL, wherein the amount of carbon disulfide added is 0.80-0.85 times the molar amount of amino in the reaction system. The post-treatment includes: after the reaction is completed, the mixture is allowed to stand and separate into layers and the aqueous phase is collected. After the solvent is removed by vacuum evaporation, the product is collected and placed in a drying oven at 50-60℃ for vacuum drying for 4-6 h to obtain sodium poly-dithiocarbamate. Further, in step B2, the ratio of sodium dithiocarbamate, deionized water, anhydrous ethanol, anhydrous dichloromethane, and phosphorus oxychloride is 15-20 g: 50 mL: 30 mL: 120 mL: 10-15 mL. The post-treatment includes: adding 0.25 times the volume of deionized water to the reaction solution and letting it stand for 10-20 min; adjusting the pH of the reaction system to 6-7 using a 10-15 wt% hydrochloric acid aqueous solution; collecting the organic phase after standing and separating the layers; washing with deionized water until neutral; removing the solvent under reduced pressure; collecting the product and placing it in a drying oven at 50-60℃ for vacuum drying for 4-6 h to obtain the antimony selective collector.
[0017] Furthermore, the preparation method of the multi-component reduced amine is as follows: dodecane, diethylenetriamine and anhydrous ethanol are added to a reaction vessel and stirred until they are mixed evenly. The reaction vessel is then placed in an ice bath at 0-5°C and stirred for 0.5-1.5 hours. Sodium borohydride is added in five batches while continuously stirring. The mixture is then stirred at 0-5°C for 1-2 hours. The reaction vessel is then heated to 25-30°C and stirred for 2-4 hours. The multi-component reduced amine is obtained after post-treatment.
[0018] The reaction principle for preparing polybasic reduced amines is as follows: Dodecaneal and diethylenetriamine first undergo condensation to form an unstable imine or hemiacetal intermediate. Subsequently, under the action of sodium borohydride, the carbonyl-related structure in the intermediate is selectively reduced to obtain a stable polyamine. This process is essentially a typical reductive amination reaction, achieving alkylation and diversification of polyamine structures through continuous transformation of condensation and reduction.
[0019] Furthermore, in the preparation of the polycyclic reduced amine, the ratio of dodecaneal, diethylenetriamine, anhydrous ethanol, and sodium borohydride is 20-24 mL:20 mL:100 mL:2.4-3.0 g. The post-treatment includes: adding 0.2-0.3 times the volume of the reaction liquid of deionized water and letting it stand for 10-20 min. After standing, the solvent is removed by vacuum evaporation, and the product is collected and placed in a drying oven at 50-60℃ for vacuum drying for 2-3 h to obtain the polycyclic reduced amine.
[0020] Furthermore, the preparation method of the lead chelating dispersant is as follows: deionized water is added to a reaction vessel and nitrogen gas is introduced under stirring for 20-30 minutes. Then, acrylic acid, 2-hydroxyethyl acrylate and 2-mercaptoethanol are added in sequence and stirred until uniform. The reaction vessel is heated to 60-70℃ and 2-4wt% ammonium persulfate aqueous solution is added. The mixture is kept warm and stirred for 3-4 hours. The lead chelating dispersant is then obtained after post-treatment.
[0021] The reaction principle for preparing lead chelating dispersants is as follows: Acrylic acid, 2-hydroxyethyl acrylate, and a thiol-containing chain transfer agent participate in polymerization under free radical conditions, introducing functional groups with metal coordination capabilities, such as carboxyl, hydroxyl, and thiol groups, into the same polymer chain in a copolymer form. This system obtains polymers with multi-coordination structures through parallel free radical polymerization and chain transfer, thereby forming a lead chelating dispersant with chelating capabilities.
[0022] Furthermore, the ratio of deionized water, acrylic acid, 2-hydroxyethyl acrylate, 2-mercaptoethanol, and 2-4wt% ammonium persulfate aqueous solution is 400mL:30mL:20-30mL:8mL:20-30mL. The post-treatment includes: cooling to room temperature after the reaction, adjusting the pH of the reaction system to 7.0-7.5 with 30-35wt% sodium hydroxide aqueous solution, filtering to remove the gel-like substance, and then distilling under reduced pressure until no liquid is collected to obtain the lead chelating dispersant.
[0023] The present invention has the following beneficial effects: The arsenic-directed collector prepared in this invention initially participates in interface regulation. Its dual-coordination groups preferentially exfoliate arsenic-containing minerals, reducing their coverage and adsorption on gold particles and increasing the exposure of gold at the interface level. Subsequently, the introduced antimony-selective collector further reduces the competitive adhesion of the antimony phase within the interface through multi-site complexation, maintaining a stable and dispersed liquid environment around the gold particles. On this basis, the lead chelating dispersant removes residual fine impurities through weakly coordinated chain segments, resulting in higher interface cleanliness on the gold surface and improving its responsiveness at the air / water interface. The interfacial stratification effect formed by the three materials at different stages collectively reduces the interference of non-target minerals on gold, gradually enhancing the floatability of gold during system evolution, thereby achieving a multiple-level increase in gold grade.
[0024] The arsenic-directed collector prepared in this invention exhibits significant directional complexation characteristics of arsenic-containing minerals in a near-neutral system through the structural synergy of phosphonic acid and oxime groups, establishing the first interfacial layer for selective arsenic removal. As the pH increases, the phosphoramide-dithio structure of the antimony-selective collector has higher affinity, enabling it to rapidly establish a second interfacial selectivity after arsenic removal, promoting the concentrated separation of antimony-containing minerals. During this process, although the lead chelating dispersant does not dominate the target separation, its weakly coordinated fragments can weaken the secondary adsorption of arsenic, antimony, and other fine particles, maintaining the stability of the interfacial gradient. The interfacial barrier continuously constructed by the three materials through structural differences makes it difficult for arsenic and antimony to re-adhere in the slurry system, thereby achieving highly selective and deep impurity removal.
[0025] After arsenic and antimony are separated in sequence, the slurry interface has a high degree of openness. Under these conditions, the lead chelating dispersant, through its multi-coordination mechanism of carboxyl, hydroxyl and mercapto groups, transforms the lead-bearing minerals from a potentially floatable phase to a stable hydrophilic equilibrium state, forming a selective separation layer at the third interface. At the same time, the residual hydrophilic fragments of the arsenic directional collector and the antimony selective collector continue to inhibit the aggregation between heterogeneous fine particles, preventing the lead separation process from being affected by shielding or covering. The residual coordination effect of the three agents jointly enhances the stable structure of the slurry interface, making it difficult for various impurities in the system to be re-enriched or co-floated, achieving synergistic control of lead and other associated impurities, and ultimately obtaining high-purity gold concentrate. Attached Figure Description
[0026] To more clearly illustrate the technical solutions in the embodiments of the present invention or the prior art, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.
[0027] Figure 1 The infrared absorption spectrum of the arsenic-directed collector prepared in Example 3 of this invention; Figure 2 The infrared absorption spectrum of the sodium dithiocarbamate salt prepared in Example 6 of this invention; Figure 3 The infrared absorption spectrum of the lead chelating dispersant prepared in Example 3 of this invention is shown. Detailed Implementation
[0028] The technical solution of the present invention will be clearly and completely described below with reference to the embodiments. 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.
[0029] Example 1 This embodiment provides a method for preparing an arsenic-directed collector, including the following steps: Step (1): Preparation of diethyl(2-cyanoethyl)phosphonate Weigh 25.0 mL of diethylphosphonite and 200.0 mL of anhydrous ethanol and add them to the reaction vessel. Stir until the mixture is homogeneous, then purge with nitrogen for protection and add 30.0 mL of acrylonitrile. Heat the reaction vessel to 45°C and stir for 3 hours. After the reaction is complete, remove the solvent by vacuum evaporation, collect the product and place it in a drying oven at 50°C for vacuum drying for 2 hours to obtain diethyl(2-cyanoethyl)phosphonite.
[0030] Step 2: Preparation of arsenic-directed collector Weigh out 20.0 mL of diethyl(2-cyanoethyl) phosphonate, 80.0 mL of anhydrous ethanol and 20.0 mL of deionized water and mix them to obtain a diethyl(2-cyanoethyl) phosphonate dispersion; Weigh 12.0 g of hydroxylamine hydrochloride and 200.0 mL of deionized water and add them to the reaction vessel. Stir until dissolved. Adjust the pH of the reaction system to 6.5 using 30 wt% sodium hydroxide aqueous solution and maintain the temperature of the reaction vessel at 20 °C. Add 100.0 mL of diethyl(2-cyanoethyl) phosphonate dispersion and keep warm for 1 h. Then raise the temperature of the reaction vessel to 45 °C and keep warm and stir for 4 h. After the reaction is completed, adjust the pH of the reaction system to 5 using 10 wt% hydrochloric acid aqueous solution. After standing and cooling, filter and collect the filter cake. Rinse the filter cake with deionized water until neutral and place it in a drying oven at 50 °C for vacuum drying for 4 h to obtain the arsenic-directed collector.
[0031] Example 2 This embodiment provides a method for preparing an arsenic-directed collector, including the following steps: Step (1): Preparation of diethyl(2-cyanoethyl)phosphonate Weigh out 30.0 mL of diethylphosphonite and 200.0 mL of anhydrous ethanol and add them to the reaction vessel. Stir until the mixture is homogeneous, then purge with nitrogen for protection and add 30.0 mL of acrylonitrile. Heat the reaction vessel to 50°C and stir for 4 hours. After the reaction is complete, remove the solvent by vacuum evaporation, collect the product and place it in a drying oven at 60°C for vacuum drying for 3 hours to obtain diethyl(2-cyanoethyl)phosphonite.
[0032] Step 2: Preparation of arsenic-directed collector Weigh out 25.0 mL of diethyl(2-cyanoethyl) phosphonate, 80.0 mL of anhydrous ethanol and 20.0 mL of deionized water and mix them to obtain a diethyl(2-cyanoethyl) phosphonate dispersion; Weigh 16.0 g of hydroxylamine hydrochloride and 200.0 mL of deionized water and add them to the reaction vessel. Stir until dissolved. Adjust the pH of the reaction system to 7.5 using 35 wt% sodium hydroxide aqueous solution and maintain the temperature of the reaction vessel at 30 °C. Add 100.0 mL of diethyl(2-cyanoethyl) phosphonate dispersion and keep warm for 2 h. Then raise the temperature of the reaction vessel to 50 °C and keep warm and stir for 6 h. After the reaction is completed, adjust the pH of the reaction system to 6 using 15 wt% hydrochloric acid aqueous solution. After standing and cooling, filter and collect the filter cake. Rinse the filter cake with deionized water until neutral and place it in a drying oven at 60 °C for vacuum drying for 6 h to obtain the arsenic-directed collector.
[0033] Example 3 This embodiment provides a method for preparing an arsenic-directed collector, including the following steps: Step (1): Preparation of diethyl(2-cyanoethyl)phosphonate Weigh 28.0 mL of diethylphosphonite and 200.0 mL of anhydrous ethanol and add them to the reaction vessel. Stir until the mixture is homogeneous, then purge with nitrogen for protection and add 30.0 mL of acrylonitrile. Heat the reaction vessel to 50°C and stir for 4 hours. After the reaction is complete, remove the solvent by vacuum evaporation, collect the product and place it in a drying oven at 55°C for vacuum drying for 3 hours to obtain diethyl(2-cyanoethyl)phosphonite.
[0034] Step 2: Preparation of arsenic-directed collector Weigh out 24.0 mL of diethyl(2-cyanoethyl) phosphonate, 80.0 mL of anhydrous ethanol and 20.0 mL of deionized water and mix them to obtain a diethyl(2-cyanoethyl) phosphonate dispersion; Weigh 14.0 g of hydroxylamine hydrochloride and 200.0 mL of deionized water and add them to the reaction vessel. Stir until dissolved. Adjust the pH of the reaction system to 7.0 using 35 wt% sodium hydroxide aqueous solution and maintain the temperature of the reaction vessel at 25 ℃. Add 100.0 mL of diethyl(2-cyanoethyl) phosphonate dispersion and keep warm for 2 h. Then raise the temperature of the reaction vessel to 50 ℃ and keep warm and stir for 5 h. After the reaction is completed, adjust the pH of the reaction system to 5 using 12 wt% hydrochloric acid aqueous solution. After standing and cooling, filter and collect the filter cake. Rinse the filter cake with deionized water until neutral and place it in a drying oven at 55 ℃ for vacuum drying for 5 h to obtain the arsenic-directed collector.
[0035] like Figure 1 As shown, the spectrum of the arsenic-directed collector at 1650 cm⁻¹... -1 A strong absorption peak appears at 930 cm⁻¹, which is the characteristic stretching vibration of the oxime group C=NOH, indicating the presence of the characteristic structure after oximeization in the sample; in addition, a strong absorption peak appears at 930 cm⁻¹. -1 A NO vibrational absorption peak can be observed nearby, around 1650 cm⁻¹. -1The characteristic peaks together constitute the criterion for oxime groups, and the simultaneous appearance of both verifies that the cyano group has been completely converted into an oxime group structure. At 1240cm -1 The P=O stretching vibration observed at this point is a typical signal of phosphonate ester structure, with stable peak shape and intensity, indicating that the phosphonic acid group remains intact during the reaction; the fingerprint region at 1050 cm⁻¹... -1 The presence of POC stretching absorption peaks further confirms the existence and absence of breakage of the phosphonate ester framework. The appearance of these two absorption peaks is completely consistent with the phosphonate ester portion in the designed structure. The sample exhibits a wavelength of 3200-3000 cm⁻¹ in the high wavenumber region. -1 The broad absorption band is attributed to the OH stretching vibration, a signal generated by the oxime hydroxyl group and trace amounts of adsorbed water. A slight perturbation in the spectral background is also present, consistent with common infrared characteristics of solid samples. At 2940 cm⁻¹... -1 A weak CH stretching absorption peak appears at the position, which is consistent with the phosphonate side chain and the methylene structure after addition; In summary, the infrared spectrum of the sample simultaneously exhibits oxime characteristic peaks (1650, 930 cm⁻¹). -1 ) and phosphonate characteristic peaks (1240, 1050 cm⁻¹) -1 All key absorption signals correspond to the structural composition of the target arsenic-directed collector. Spectroscopic evidence shows that the prepared material has a clear combination of dual-coordination characteristic groups, verifying that the sample structure is consistent with the expected product.
[0036] Example 4 This embodiment provides a method for preparing an antimony selective collector, including the following steps: Step I: Preparation of multi-component reduced amines Weigh out 20.0 mL of dodecane, 20.0 mL of diethylenetriamine, and 100.0 mL of anhydrous ethanol and add them to the reaction vessel. Stir until the mixture is homogeneous. Place the reaction vessel in an ice bath at 0°C and stir for 0.5 h. Add 2.4 g of sodium borohydride in five batches while stirring continuously. Continue stirring at 0°C for 1 h. Then, heat the reaction vessel to 25°C and stir for 2 h. Add 0.2 times the volume of deionized water and let stand for 10 min. After standing, remove the solvent by vacuum evaporation. Collect the product and place it in a drying oven at 50°C for vacuum drying for 2 h to obtain the polycyclic reduced amine.
[0037] Step II: Preparation of sodium dithiocarbamate Weigh out 18.0 g of polydithiocarbamate, 100.0 mL of anhydrous ethanol and 100.0 mL of deionized water and add them to the reaction vessel. Stir until dissolved and cool the reaction vessel to 0°C under ice bath conditions. Then add carbon disulfide with a molar amount of 0.80 times that of amino in the reaction system and maintain the pH of the reaction system at 10 using 30 wt% sodium hydroxide aqueous solution. Keep the reaction vessel heated and stirred for 3 h. Then heat the reaction vessel to 20°C and keep it heated and stirred for 0.5 h. After the reaction is completed, allow it to stand and separate into layers and collect the aqueous phase. After removing the solvent under reduced pressure, collect the product and place it in a drying oven at 50°C for vacuum drying for 4 h to obtain sodium dithiocarbamate.
[0038] Step III: Preparation of Antimony Selective Collector Weigh out 15.0 g of sodium dithiocarbamate, 50.0 mL of deionized water, and 30.0 mL of anhydrous ethanol and add them to a reaction vessel. Stir until dissolved, then add 120.0 mL of anhydrous dichloromethane and purge with nitrogen for protection. Cool the reaction vessel to 0°C and add 10.0 mL of phosphorus oxychloride. After the addition is complete, stir for 1 h. Then heat the reaction vessel to 25°C and keep it at that temperature for 3 h. Add 0.25 times the volume of deionized water and let it stand for 10 min. Adjust the pH of the reaction system to 6 using 10 wt% hydrochloric acid aqueous solution. After standing and separating the layers, collect the organic phase and wash it with deionized water until neutral. Remove the solvent under reduced pressure and collect the product. Place it in a drying oven at 50°C and vacuum dry for 4 h to obtain the antimony selective collector.
[0039] Example 5 This embodiment provides a method for preparing an antimony selective collector, including the following steps: Step I: Preparation of multi-component reduced amines Weigh out 24.0 mL of dodecane, 20.0 mL of diethylenetriamine, and 100.0 mL of anhydrous ethanol and add them to the reaction vessel. Stir until the mixture is homogeneous. Place the reaction vessel in an ice bath at 5°C and stir for 1.5 h. Add 3.0 g of sodium borohydride in five batches while stirring continuously. Continue stirring at 5°C for 2 h. Then, heat the reaction vessel to 30°C and stir for 4 h. Add 0.3 times the volume of deionized water and let it stand for 20 min. After standing, remove the solvent by vacuum evaporation. Collect the product and place it in a drying oven at 60°C for vacuum drying for 3 h to obtain the polycyclic reduced amine.
[0040] Step II: Preparation of sodium dithiocarbamate Weigh out 20.0 g of polydithiocarbamate, 100.0 mL of anhydrous ethanol and 100.0 mL of deionized water and add them to the reaction vessel. Stir until dissolved, then cool the reaction vessel to 10 °C under ice bath conditions. Then add carbon disulfide with a molar amount of 0.85 times that of the amino group in the reaction system and maintain the pH of the reaction system at 11 using 35 wt% sodium hydroxide aqueous solution. Keep the reaction vessel heated and stirred for 4 h. Then heat the reaction vessel to 25 °C and keep it heated and stirred for 1 h. After the reaction is completed, allow it to stand and separate into layers and collect the aqueous phase. After removing the solvent under reduced pressure, collect the product and place it in a drying oven at 60 °C for vacuum drying for 6 h to obtain sodium dithiocarbamate.
[0041] Step III: Preparation of Antimony Selective Collector Weigh out 20.0 g of sodium dithiocarbamate, 50.0 mL of deionized water, and 30.0 mL of anhydrous ethanol and add them to a reaction vessel. Stir until dissolved, then add 120.0 mL of anhydrous dichloromethane and purge with nitrogen for protection. Cool the reaction vessel to 5°C and add 15.0 mL of phosphorus oxychloride. After the addition is complete, stir for 2 h. Then heat the reaction vessel to 30°C and keep it at that temperature for 4 h. Add 0.25 times the volume of deionized water and let it stand for 20 min. Adjust the pH of the reaction system to 7 using 15 wt% hydrochloric acid aqueous solution. After standing and separating the layers, collect the organic phase and wash it with deionized water until neutral. Remove the solvent under reduced pressure and collect the product. Place it in a drying oven at 60°C and vacuum dry for 6 h to obtain the antimony selective collector.
[0042] Example 6 This embodiment provides a method for preparing an antimony selective collector, including the following steps: Step I: Preparation of multi-component reduced amines Weigh out 21.0 mL of dodecane, 20.0 mL of diethylenetriamine, and 100.0 mL of anhydrous ethanol and add them to the reaction vessel. Stir until the mixture is homogeneous. Place the reaction vessel in an ice bath at 3°C and stir for 1 hour. Add 2.7 g of sodium borohydride in five batches while stirring continuously. Continue stirring at 3°C for 2 hours. Then, heat the reaction vessel to 30°C and stir for 3 hours. Add 0.3 times the volume of deionized water and let it stand for 20 minutes. After standing, remove the solvent by vacuum evaporation. Collect the product and place it in a drying oven at 55°C for vacuum drying for 3 hours to obtain the polycyclic reduced amine.
[0043] Step II: Preparation of sodium dithiocarbamate Weigh out 20.0 g of polydithiocarbamate, 100.0 mL of anhydrous ethanol and 100.0 mL of deionized water and add them to the reaction vessel. Stir until dissolved and cool the reaction vessel to 5 °C under ice bath conditions. Then add carbon disulfide with a molar amount of 0.80 times that of amino in the reaction system and use 32 wt% sodium hydroxide aqueous solution to maintain the pH of the reaction system at 11. Keep the reaction vessel heated and stirred for 4 h. Then heat the reaction vessel to 25 °C and keep it heated and stirred for 1 h. After the reaction is completed, let it stand and separate into layers and collect the aqueous phase. After removing the solvent under reduced pressure, collect the product and place it in a drying oven at 55 °C for vacuum drying for 5 h to obtain sodium dithiocarbamate.
[0044] Step III: Preparation of Antimony Selective Collector Weigh out 18.0 g of sodium dithiocarbamate, 50.0 mL of deionized water, and 30.0 mL of anhydrous ethanol and add them to a reaction vessel. Stir until dissolved, then add 120.0 mL of anhydrous dichloromethane and purge with nitrogen for protection. Cool the reaction vessel to 3°C and add 12.0 mL of phosphorus oxychloride. After the addition is complete, stir for 2 h. Then heat the reaction vessel to 30°C and keep it at that temperature for 4 h. Add 0.25 times the volume of deionized water and let it stand for 15 min. Adjust the pH of the reaction system to 7 using 12 wt% hydrochloric acid aqueous solution. After standing and separating the layers, collect the organic phase and wash it with deionized water until neutral. Remove the solvent under reduced pressure and collect the product. Place it in a drying oven at 55°C and vacuum dry for 5 h to obtain the antimony selective collector.
[0045] like Figure 2 As shown, at 1500cm -1 A strong absorption peak is visible nearby, and its peak shape is consistent with the CN stretching characteristics of dithiocarbamate compounds. This signal indicates that the sample contains a stable NC(=S)-S group, verifying that the dithiocarbamate structure remains intact after the reaction; simultaneously, at 1050 cm⁻¹... -1 The presence of a moderate-intensity absorption peak is attributed to the C=S / CS related stretching vibration, and is related to the 960 cm⁻¹ peak. -1 The absorption peaks that appeared nearby together constituted the characteristic fingerprint region of the dithiocarbamic acid group, further confirming that this functional group does indeed exist in the target product; At 1250cm -1 A significant characteristic absorption peak was observed, belonging to the P=O stretching vibration, which is a core marker of phosphoramide structures. This peak was absent in the precursor, and its appearance in the final product indicates that the phosphorylation structure was successfully constructed. Furthermore, the fingerprint region is 960 cm⁻¹. -1 The shift of the absorption peak relative to the precursor can be attributed to the coupling absorption of the PN vibration and the neighboring CS unit. This change is direct spectroscopic evidence of the bonding between the nitrogen atom and the phosphoryl group, which is consistent with the phosphoramide structure present in the product. Absorption in the high wavenumber region also reflects the retention of the product skeleton; at 2930 cm⁻¹ -1 With 2850cm -1 The presence of distinct CH stretching absorption bands indicates that the alkylated polyamine backbone still exists; simultaneously, at approximately 3230 cm⁻¹... -1 The broad absorption band can be attributed to the residual NH vibration signal. The presence of such absorption peaks indicates that incompletely functionalized nitrogen sites are still retained in the molecule, which is consistent with the polynitrogen and polysulfide coordination framework of the target collector. In summary, CN (1500 cm⁻¹) in the infrared spectrum -1 C=S / CS (1050, 960cm) -1 The characteristic peaks confirmed the presence of the dithiocarbamic acid group; P=O (1250cm) -1 ) and PN-related characteristics (960cm) -1 The successful construction of the phosphoramide structure was confirmed by joint analysis; the CH and NH absorption bands in the high wavenumber region proved that the alkylated polyamine skeleton was preserved; all characteristic peaks appeared simultaneously and corresponded completely to the functional units of the target structure, and the spectroscopic results showed that the final product had been successfully prepared as an antimony selective collector.
[0046] Example 7 This embodiment provides a method for preparing a lead chelating dispersant, including the following steps: Weigh 400.0 mL of deionized water and add it to the reaction vessel. After purging with nitrogen for 20 min while stirring, add 30.0 mL of acrylic acid, 20.0 mL of 2-hydroxyethyl acrylate and 8.0 mL of 2-mercaptoethanol in sequence and stir until well mixed. Heat the reaction vessel to 60℃ and add 20.0 mL of 2wt% ammonium persulfate aqueous solution. Keep warm and stir for 3 h. After the reaction is completed, cool to room temperature and adjust the pH of the reaction system to 7.0 using 30wt% sodium hydroxide aqueous solution. After filtering to remove the gel-like substance, distill under reduced pressure until no liquid is collected to obtain the lead chelating dispersant.
[0047] Example 8 This embodiment provides a method for preparing a lead chelating dispersant, including the following steps: Weigh out 400.0 mL of deionized water and add it to the reaction vessel. After purging with nitrogen for 30 min while stirring, add 30.0 mL of acrylic acid, 30.0 mL of 2-hydroxyethyl acrylate and 8.0 mL of 2-mercaptoethanol in sequence and stir until well mixed. Heat the reaction vessel to 70℃ and add 30.0 mL of 4wt% ammonium persulfate aqueous solution. Keep the temperature and stir for 4 h. After the reaction is completed, cool to room temperature and adjust the pH of the reaction system to 7.5 using 35wt% sodium hydroxide aqueous solution. After filtering to remove the gel-like substance, distill under reduced pressure until no liquid is collected to obtain the lead chelating dispersant.
[0048] Example 9 This embodiment provides a method for preparing a lead chelating dispersant, including the following steps: Weigh out 400.0 mL of deionized water and add it to the reaction vessel. After purging with nitrogen for 25 min while stirring, add 30.0 mL of acrylic acid, 25.0 mL of 2-hydroxyethyl acrylate and 8.0 mL of 2-mercaptoethanol in sequence and stir until well mixed. Heat the reaction vessel to 65℃ and add 25.0 mL of 3wt% ammonium persulfate aqueous solution. Keep the temperature and stir for 4 h. After the reaction is completed, cool to room temperature and adjust the pH of the reaction system to 7.5 using 32wt% sodium hydroxide aqueous solution. After filtering to remove the gel-like substance, distill under reduced pressure until no liquid is collected to obtain the lead chelating dispersant.
[0049] like Figure 3 As shown, at 1500cm -1 A strong absorption peak was detected at 1050 cm⁻¹, attributed to the CN stretching vibration of the dithiocarbamate group. The presence of this peak indicates the presence of a stable NC(=S)-S structure in the material, which is a core identifier for this type of ligand. Subsequently, a strong absorption peak was detected at 1050 cm⁻¹. -1 A moderate-intensity absorption peak appears at the position, corresponding to the C=S / CS related stretching vibration, and is associated with the 960 cm⁻¹ position. -1 The nearby absorption peaks together constitute the fingerprint region signal of the dithiocarbamic acid structure, and the simultaneous appearance of both verifies that the group is completely preserved in the final product. At 1250cm -1 A strong absorption peak can be observed at 960 cm⁻¹, which is the characteristic P=O stretching vibration of the phosphoramide structure and is the core spectral basis for determining whether the phosphorylation structure has been formed. This peak is not present in the precursor, and its appearance in the product indicates that the P=O structure has been successfully introduced. At the same time, the fingerprint region is at 960 cm⁻¹. -1 The absorption peaks of the material show changes in position and morphology compared to the precursor, exhibiting absorption characteristics of PN vibration coupled with adjacent sulfur groups. This further confirms the bonding relationship between nitrogen atoms and phosphoryl groups, thus proving that a multi-site coordinated phosphoramide unit has been constructed in the material. 2930cm -1 With 2850cm -1 The absorption peaks were attributed to the asymmetric and symmetric stretching vibrations of the alkyl chain, respectively, indicating that the alkylated polyamine skeleton remained intact during synthesis; furthermore, the peak at approximately 3230 cm⁻¹... -1 A broad absorption band was observed, which can be attributed to the residual NH stretching signal and weak hydrogen bonding, indicating that the molecule still contains incompletely substituted amino sites, consistent with the polynitrogen and polysulfide coordination characteristics of the product. In summary, the characteristic peaks of dithiocarbamic acid in the infrared spectrum (1500, 1050, 960 cm⁻¹) -1The presence of the sulfur-coordinating group was confirmed by the appearance of the characteristic absorption (1250 cm⁻¹) of the phosphoramide structure. -1 With 960cm -1 The phosphorylation reaction was confirmed to be complete; the CH and NH absorption bands in the high wavenumber region indicated that the alkylated polyamine skeleton was preserved; each absorption characteristic completely corresponded to the key functional groups of the target molecule structure, and the comprehensive spectral results proved that the prepared material was indeed an antimony selective collector.
[0050] Example 10 This embodiment provides a method for preparing a flotation process for removing antimony from antimony-gold concentrate, comprising the following steps: Step 1: Prepare the slurry to be purified Weigh out 30.0 kg of gold concentrate with a gold content of 0.004 wt%, antimony content of 15.125 wt%, arsenic content of 2.354 wt%, and lead content of 1.238 wt%, and add 100.0 L of deionized water to the grinding equipment. Wet grind until the 74 μm particle size accounts for 70 wt%. After grinding, adjust the pH of the slurry to 6.5 to obtain the slurry to be purified.
[0051] Step 2: Preparation of Arsenic-De-Arsenic Slurry Weigh out 90.0L of the slurry to be purified and add it to the flotation cell. Add 15g of the arsenic-directed collector prepared in Example 1, stir for 3 minutes, add 2.0g of methyl isobutyl methanol, float for 6 minutes, collect the foam as arsenic concentrate, and obtain the arsenic-removed slurry.
[0052] Step 3: Preparation of arsenic- and antimony-removed slurry Weigh out 80.0L of arsenic-removed slurry and add it to the stirring tank. Adjust the pH to 7.0, add 30.0g of the antimony selective collector prepared in Example 4, stir for 3 minutes, then add 2.0g of methyl isobutyl methanol, float for 8 minutes, collect the foam as antimony concentrate, and obtain the arsenic-removed and antimony-removed slurry.
[0053] Step 4: Preparation of purified gold concentrate Weigh out 70.0 L of arsenic- and antimony-removed slurry and add it to a stirring tank. Adjust the pH to 8.0, add 20.0 g of the lead chelating dispersant prepared in Example 7 and 2.0 g of methyl isobutyl methanol, stir for 5 min, float for 4 min, collect the foam as lead concentrate, filter the slurry and collect the filter cake, transfer the filter cake to a drying oven at 80 °C and vacuum dry to constant weight to obtain purified gold concentrate.
[0054] Example 11 This embodiment provides a method for preparing a flotation process for removing antimony from antimony-gold concentrate, comprising the following steps: Step 1: Prepare the slurry to be purified Weigh out 35.0 kg of gold concentrate with a gold content of 0.006 wt%, antimony content of 19.875 wt%, arsenic content of 2.687 wt%, and lead content of 1.945 wt%, and add 100.0 L of deionized water to the grinding equipment. Wet grind until the 74 μm particle size accounts for 80 wt%. After grinding, adjust the pH of the slurry to 7.0 to obtain the slurry to be purified.
[0055] Step 2: Preparation of Arsenic-De-Arsenic Slurry Weigh out 90.0L of the slurry to be purified and add it to the flotation cell. Add 25.0g of the arsenic-directed collector prepared in Example 2, stir for 5 minutes, add 2.0g of methyl isobutyl methanol, float for 8 minutes, collect the foam as arsenic concentrate, and obtain the arsenic-removed slurry.
[0056] Step 3: Preparation of arsenic- and antimony-removed slurry Weigh out 80.0L of arsenic-removed slurry and add it to the stirring tank. Adjust the pH to 7.5, add 40.0g of the antimony selective collector prepared in Example 5, stir for 5 minutes, then add 2.0g of methyl isobutyl methanol, float for 10 minutes, collect the foam as antimony concentrate, and obtain the arsenic-removed and antimony-removed slurry.
[0057] Step 4: Preparation of purified gold concentrate Weigh out 70.0 L of arsenic- and antimony-removed slurry and add it to a stirring tank. Adjust the pH to 8.5, add 20.0 g of the lead chelating dispersant prepared in Example 8 and 2.0 g of methyl isobutyl methanol, stir for 8 min, float for 6 min, collect the foam as lead concentrate, filter the slurry and collect the filter cake, transfer the filter cake to a drying oven at 80 °C and vacuum dry to constant weight to obtain purified gold concentrate.
[0058] Example 12 This embodiment provides a method for preparing a flotation process for removing antimony from antimony-gold concentrate, comprising the following steps: Step 1: Prepare the slurry to be purified Weigh out 32.0 kg of gold concentrate with a gold content of 0.005 wt%, antimony content of 18.842 wt%, arsenic content of 2.651 wt%, and lead content of 1.583 wt%, and add 100.0 L of deionized water to the grinding equipment. Wet grind until the 74 μm particle size accounts for 75 wt%. After grinding, adjust the pH of the slurry to 7.0 to obtain the slurry to be purified.
[0059] Step 2: Preparation of Arsenic-De-Arsenic Slurry Weigh out 90.0L of the slurry to be purified and add it to the flotation cell. Add 20g of the arsenic-directed collector prepared in Example 3, stir for 4 minutes, add 2.0g of methyl isobutyl methanol, float for 7 minutes, collect the foam as arsenic concentrate, and obtain the arsenic-removed slurry.
[0060] Step 3: Preparation of arsenic- and antimony-removed slurry Weigh out 80.0L of arsenic-removed slurry and add it to the stirring tank. Adjust the pH to 7.5, add 35.0g of the antimony selective collector prepared in Example 6, stir for 4 minutes, then add 2.0g of methyl isobutyl methanol, float for 9 minutes, collect the foam as antimony concentrate, and obtain arsenic-removed and antimony-removed slurry.
[0061] Step 4: Preparation of purified gold concentrate Weigh out 70.0 L of arsenic- and antimony-removed slurry and add it to a stirring tank. Adjust the pH to 8.5, add 20.0 g of the lead chelating dispersant prepared in Example 9 and 2.0 g of methyl isobutyl methanol, stir for 6 min, float for 8 min, collect the foam as lead concentrate, filter the slurry and collect the filter cake, transfer the filter cake to a drying oven at 80 °C and vacuum dry to constant weight to obtain purified gold concentrate.
[0062] Comparative Example 1 The difference between this comparative example and Example 12 is that step two is omitted.
[0063] Comparative Example 2 The difference between this comparative example and Example 12 is that step III is omitted in the preparation process of the antimony selective collector used in step three.
[0064] Comparative Example 3 The difference between this comparative example and Example 12 is that step four is omitted.
[0065] Performance testing: The gold content of the purified gold concentrates prepared in Examples 10-12 and Comparative Examples 1-3 was tested according to the standard GB / T 7739.1-2019 "Chemical Analysis Methods for Gold Concentrates - Part 1: Determination of Gold and Silver Contents", and the gold content enhancement factor was calculated. The arsenic content of the purified gold concentrates prepared in Examples 10-12 and Comparative Examples 1-3 was tested according to the standard GB / T 7739.3-2019 "Chemical Analysis Methods for Gold Concentrates - Part 3: Determination of Arsenic Content", and the reduction factor of arsenic content was calculated. The antimony content of the gold concentrates prepared in Examples 10-12 and Comparative Examples 1-3 was tested according to the standard GB / T 7739.10-2007 "Chemical Analysis Methods for Gold Concentrates - Part 10: Determination of Antimony Content", and the reduction factor of antimony content was calculated. The lead content of the gold concentrates prepared in Examples 10-12 and Comparative Examples 1-3 was tested according to the standard GB / T 7739.5-2021 "Chemical Analysis Methods for Gold Concentrates - Part 5: Determination of Lead Content" and the reduction factor of lead content was calculated. The specific data are shown in Table 1. Table 1 - Performance Test Data for Each Sample Project Group Example 10 Example 11 Example 12 Comparative Example 1 Comparative Example 2 Comparative Example 3 Gold content increased by multiple 12.483 13.124 14.536 7.842 10.317 12.004 Arsenic content reduction factor 23.417 24.687 25.926 9.473 14.628 18.205 Antimony content reduction factor 41.352 42.924 43.473 12.516 18.337 26.914 Lead content reduction factor 11.274 12.549 13.741 5.236 8.417 6.952 Data Analysis: Comparative analysis of the data in Table 1 reveals that the purified gold concentrate prepared by this invention exhibits a gold content increase of 14.536, an arsenic content decrease of 25.926, an antimony content decrease of 43.473, and a lead content decrease of 13.741. All these data are superior to the comparative example, indicating that: After step two was removed in Comparative Example 1, the slurry system lacked an interfacial pre-selection process before entering the multi-stage flotation sequence. The original impurity coating, interfacial adhesion, and microphase intercalation structure between minerals continued to exist. Since the interface did not undergo early stripping, the active sites on the surface of the mineral particles were not fully exposed, and the slurry as a whole exhibited a high interfacial impurity load. This prevented the subsequent interface control process from being carried out on the basis of an ideal interface. Under these conditions, the interfacial electrical distribution, short-range forces between particles, and surface hydration structure of the system were all in an unreorganized state, resulting in non-selective bubble-particle aggregation behavior during flotation. At the same time, the system could not establish the interfacial gradient structure on which the subsequent steps depended, causing the three-stage flotation to degenerate from a sequential separation mechanism to a passive stripping mode of a mixed system. Ultimately, this manifested as a decrease in interface adjustability, insufficient response of target particles, and a shrinking impurity migration window. In Comparative Example 2, step III was omitted, resulting in the collector molecules not undergoing phosphorylation conformational reconstruction. Consequently, the spatial arrangement, electron density distribution, and polar gradient of its coordination fragments did not achieve the structural stability established in the examples. The unreconstructed molecular framework could not form a multi-site synergistic adsorption mode that matched the antimony-containing mineral surface, causing the interface recognition process to degenerate from "configuration-driven" to "functional group coarse coordination." In the slurry system, the orientation, interfacial residence time, and adsorption layer density of such unreconstructed molecules all showed significant fluctuations, making it difficult to construct a continuous interfacial film on the particle surface. At the same time, the molecules lacked the interfacial repulsion and stratification capabilities unique to the reagents in the examples, failing to maintain the spatial separation between the antimony phase and associated impurities. This disrupted the controllable separation window and affected the subsequent system interface stability. The overall separation behavior exhibited selective decay due to the decrease in interface recognition. In Comparative Example 3, the elimination of step four resulted in the system losing the terminal interface redistribution stage in the flotation sequence. This prevented the fine impurities, heavy metal components, and micro-aggregates remaining after the first two flotation stages from undergoing the terminal interface decoupling process. Without this interface rearrangement effect, the particle electrical balance, surface hydration film thickness, and interface-bubble adhesion kinetics in the pulp system remained in an unconverged state, making it easy for impurity particles to undergo secondary aggregation at the interface, forming heterogeneous micro-aggregate structures. In addition, the interface selectivity window that should have been established in the terminal step was not formed, which compressed the overall interfacial energy gradient of the system. The flotation system entered the non-selective adhesion range, thereby weakening the spatial exclusion ability of associated heavy metals and fine impurities in the sequential flotation. The overall system showed a decline in interface control, particle dispersibility, and stratified flotation efficiency. Ultimately, the system behavior of the three comparative examples shows that the stepwise flotation process of the present invention relies on the continuous unfolding of the interface-level reconstruction mechanism. However, in Comparative Examples 1-3, the key interface evolution links were disturbed, resulting in the system being unable to establish the interface layering structure unique to the embodiments. In Comparative Example 1, since the slurry did not undergo primary interface reforming, the original impurity coating structure and surface hydration state were retained, making the subsequent interface control steps lack a stable base and weakening the separation window of the subsequent stages from the source of the system. In Comparative Example 2, the reagent molecules did not undergo the structure reconstruction stage, and their interface recognition ability and spatial arrangement did not have the selective configuration required by the embodiments, making it impossible for the second interface layering link to form an effective target phase-impurity phase interface segmentation. In Comparative Example 3, the terminal interface redistribution link was missing, and the system could not complete the final interface convergence, causing the residual impurity particles to be randomly aggregated on the interface, resulting in the overall interface energy gradient collapse. In summary, the common feature of the three comparative examples is that the continuity of the interface structure is interrupted, which manifests as: the interface substrate cannot be established, the interface selectivity cannot be enhanced, and the interface convergence cannot be completed. The above-mentioned interruption of interface evolution causes the multi-stage flotation to degenerate from a "sequential stratification mechanism" to a "non-selective stripping mechanism", ultimately resulting in an overall enrichment degree and separation efficiency lower than those of the examples.
[0066] The above description is merely an example and illustration of the structure of the present invention. Those skilled in the art can make various modifications or additions to the specific embodiments described, or use similar methods to replace them, as long as they do not deviate from the structure of the invention or exceed the scope defined in the claims, all of which should fall within the protection scope of the present invention.
[0067] In the description of this specification, references to terms such as "an embodiment," "example," "specific example," etc., indicate that a specific feature, structure, material, or characteristic described in connection with that embodiment or example is included in at least one embodiment or example of the invention. In this specification, illustrative expressions of the above terms do not necessarily refer to the same embodiment or example. Furthermore, the specific features, structures, materials, or characteristics described may be combined in any suitable manner in one or more embodiments or examples.
[0068] The preferred embodiments of the present invention disclosed above are merely illustrative of the invention. These preferred embodiments do not exhaustively describe all details, nor do they limit the invention to specific implementations. Clearly, many modifications and variations can be made based on the content of this specification. This specification selects and specifically describes these embodiments to better explain the principles and practical applications of the invention, thereby enabling those skilled in the art to better understand and utilize the invention. The invention is limited only by the claims and their full scope and equivalents.
Claims
1. A flotation process for removing antimony from antimony-gold concentrate, characterized in that, Includes the following steps: S1. Add gold concentrate and deionized water to the grinding equipment and wet grind until the 74μm particle size accounts for 70-80wt%. After grinding, adjust the pH of the slurry to 6.5-7.0 to obtain the slurry to be purified. S2. Add the slurry to be purified to the flotation cell, add the arsenic-directed collector, stir for 3-5 minutes, add methyl isobutyl methanol, float for 6-8 minutes, collect the foam as arsenic concentrate, and obtain the arsenic-removed slurry. S3. Add the arsenic-removed slurry to the mixing tank, adjust the pH to 7.0-7.5, add the antimony selective collector, stir for 3-5 minutes, then add methyl isobutyl methanol, float for 8-10 minutes, collect the foam as antimony concentrate, and obtain the arsenic-removed and antimony-removed slurry. S4. Add the arsenic- and antimony-removed slurry to the mixing tank, adjust the pH to 8.0-8.5, add lead chelating dispersant, stir for 5-8 minutes, then add methyl isobutyl methanol, float for 4-6 minutes, collect the foam as lead concentrate, and then process to obtain purified gold concentrate.
2. The antimony removal process for antimony-gold concentrate flotation according to claim 1, characterized in that, In step S1, the ratio of gold concentrate to deionized water is 30-35 kg: 100 L, wherein the gold content is 0.004-0.06 wt%, the antimony content is 15-20 wt%, the arsenic content is 2-3 wt%, and the lead content is 1-2 wt%. In step S2, the ratio of the slurry to be purified, the arsenic-directed collector, and methyl isobutyl methanol is 15-25 g: 90 L: 2 g. In step S3, the ratio of the arsenic-removed slurry, the antimony selective collector, and methyl isobutyl methanol is 30-40 g: 80 L: 2 g. In step S4, the ratio of the arsenic-removed and antimony-removed slurry, the lead chelating dispersant, and methyl isobutyl methanol is 20 g: 70 L: 2 g.
3. The antimony removal process for antimony-gold concentrate flotation according to claim 1, characterized in that, The arsenic-directed collector is prepared by the following method: A1. Add diethylphosphonite and anhydrous ethanol to a reaction vessel and stir. After mixing evenly, purge with nitrogen for protection and add acrylonitrile. Then heat the reaction vessel to 45-50℃ and stir for 3-4 hours. Post-treatment yields diethyl(2-cyanoethyl)phosphonite. A2. Hydroxylamine hydrochloride and deionized water are added to the reaction vessel and stirred until dissolved. The pH of the reaction system is adjusted to 6.5-7.5 using a 30-35 wt% sodium hydroxide aqueous solution, and the temperature of the reaction vessel is maintained at 20-30℃. Diethyl(2-cyanoethyl) phosphonate dispersion is added, and the mixture is kept at this temperature for 1-2 hours. Then, the temperature of the reaction vessel is raised to 45-50℃, and the mixture is kept at this temperature and stirred for 4-6 hours. The arsenic-directed collector is obtained by post-treatment.
4. The antimony removal process for antimony-gold concentrate flotation according to claim 3, characterized in that, In step A1, the ratio of diethylphosphonite, anhydrous ethanol, and acrylonitrile is 25-30 mL: 200 mL: 30 mL; in step A2, the ratio of hydroxylamine hydrochloride, deionized water, and diethyl(2-cyanoethyl)phosphonite dispersion is 12-16 g: 200 mL: 100 mL, wherein the diethyl(2-cyanoethyl)phosphonite dispersion is obtained by mixing diethyl(2-cyanoethyl)phosphonite, anhydrous ethanol, and deionized water in a ratio of 20-25 mL: 80 mL: 20 mL.
5. The antimony removal process for antimony-gold concentrate flotation according to claim 1, characterized in that, The antimony selective collector is prepared by the following method: B1. Add the polydimethylamine, anhydrous ethanol, and deionized water to a reaction vessel and stir until dissolved. Cool the reaction vessel to 0-10°C under ice bath conditions. Then add the calculated amount of carbon disulfide and maintain the pH of the reaction system at 10-11 using a 30-35 wt% sodium hydroxide aqueous solution. Keep the mixture warm and stir for 3-4 hours. Then heat the reaction vessel to 20-25°C and keep it warm and stir for 0.5-1 hour. Post-processing yields sodium dithiocarbamate. B2. Add sodium dithiocarbamate, deionized water and anhydrous ethanol to the reaction vessel and stir until dissolved. Then add anhydrous dichloromethane and purge with nitrogen for protection. Cool the reaction vessel to 0-5°C and add phosphorus oxychloride. After the addition is complete, stir for 1-2 hours. Then heat the reaction vessel to 25-30°C and keep it at that temperature for 3-4 hours. Post-processing yields the antimony selective collector.
6. The antimony removal process for antimony-gold concentrate flotation according to claim 5, characterized in that, In step B1, the ratio of poly-reduced amine, anhydrous ethanol, and deionized water is 18-20 g: 100 mL: 100 mL, wherein the amount of carbon disulfide added is 0.80-0.85 times the molar amount of amino in the reaction system; in step B2, the ratio of sodium poly-dithiocarbamate, deionized water, anhydrous ethanol, anhydrous dichloromethane, and phosphorus oxychloride is 15-20 g: 50 mL: 30 mL: 120 mL: 10-15 mL.
7. The antimony removal process for antimony-gold concentrate flotation according to claim 5, characterized in that, The preparation method of the multi-component reduced amine is as follows: dodecane, diethylenetriamine and anhydrous ethanol are added to a reaction vessel and stirred until they are mixed evenly. The reaction vessel is then placed in an ice bath at 0-5℃ and stirred for 0.5-1.5h. Sodium borohydride is added in five batches while continuously stirring. Stirring is continued at 0-5℃ for 1-2h. Then the reaction vessel is heated to 25-30℃ and stirred for 2-4h. The multi-component reduced amine is obtained after post-treatment.
8. The antimony removal process for antimony-gold concentrate flotation according to claim 7, characterized in that, In the preparation of polycyclic reducing amines, the ratio of dodecaneal, diethylenetriamine, anhydrous ethanol and sodium borohydride is 20-24 mL:20 mL:100 mL:2.4-3.0 g.
9. The antimony removal process for antimony-gold concentrate flotation according to claim 1, characterized in that, The lead chelating dispersant is prepared as follows: Deionized water is added to a reaction vessel and nitrogen gas is introduced under stirring for 20-30 minutes. Then, acrylic acid, 2-hydroxyethyl acrylate and 2-mercaptoethanol are added in sequence and stirred until uniform. The reaction vessel is heated to 60-70℃ and 2-4wt% ammonium persulfate aqueous solution is added. The mixture is kept warm and stirred for 3-4 hours. The lead chelating dispersant is then obtained after post-treatment.
10. The antimony removal process for antimony-gold concentrate flotation according to claim 9, characterized in that, In the preparation of lead chelating dispersant, the ratio of deionized water, acrylic acid, 2-hydroxyethyl acrylate, 2-mercaptoethanol and 2-4wt% ammonium persulfate aqueous solution is 400mL:30mL:20-30mL:8mL:20-30mL.