A flotation method for comprehensive recovery of silver, zinc and lead from zinc smelting slag

Abstract

The present invention discloses a flotation method for comprehensively recovering silver, zinc, and lead from zinc smelting slag. This method addresses the problem of unsatisfactory recycling of valuable elements such as silver, zinc, and lead contained in zinc smelting slag. By regulating the slurry potential and pH, the method determines the principle process of flotation recovery of silver and zinc under acidic conditions and flotation recovery of lead under a weakly alkaline environment. In the lead flotation step, the dissolved oxygen in the slurry is driven out by a protective gas, the lead minerals are then sulfided with a sulfiding agent, and finally the lead is captured with a collector. This method has the characteristics of a stable process flow and low cost, solves the problem of low comprehensive recovery and utilization rate of silver, zinc, and lead in zinc smelting slag, and simultaneously realizes partial disposal of the zinc smelting slag.

Description

Technical Field

[0001] The present invention relates to the technical field of comprehensive utilization of nonferrous metal solid waste, and in particular to a flotation process for recovering valuable metals from zinc smelting slag. Background Art

[0002] Zinc is a vital nonferrous metal raw material. Approximately 85% of global zinc production is produced through hydrometallurgical zinc smelting, in which zinc sulfide concentrate is roasted and then subjected to hot and high-acid leaching to produce acid leaching slag. Every ton of electrolytic zinc produced generates over 0.3 tons of acid leaching slag, which contains a certain amount of metal elements such as lead, zinc, copper, silver, and cadmium. Currently, zinc smelting slag is primarily stockpiled, which poses a significant environmental hazard and carries high disposal costs. However, the metal components within it possess extremely high overall recovery value. Methods for recovering metal elements from zinc smelting slag primarily include hydrometallurgical leaching, pyrometallurgical processes, and combined flotation and metallurgical processes, with pyrometallurgical processes being the most commonly used comprehensive utilization method. For example, patent CN107326190B uses reduction roasting to change the state of valuable metals in smelting slag containing lead, zinc and silver, obtaining high-grade PbO and ZnO smoke and Ag-enriched roasting slag. The smoke enters the lead and zinc smelting system; the roasting slag is refined and flotated to obtain silver concentrate and tailings, ultimately realizing the resource utilization of valuable metals. However, the use of pyrometallurgical furnace technology requires the company to have supporting facilities for lead smelting. In areas without lead smelting, this method is difficult to apply. In addition, traditional pyrometallurgical processes (such as rotary kilns, fuming furnace volatilization treatment, etc.) are energy-intensive and polluting. With the increasing environmental protection requirements, the application of such methods is also subject to certain restrictions.

[0003] Due to the relatively low lead and silver content in acid leaching waste, direct recovery of lead and silver using traditional wet leaching is difficult. Flotation processes, however, offer the advantages of a simple process flow, low investment, and low production costs, making them an effective approach for the comprehensive recovery and disposal of acid leaching waste. For example, patent CN109261347B utilizes ultrasonic activation of smelting waste followed by mixed flotation using a novel collector to produce a mixed flotation concentrate. Consequently, smelting companies often use flotation to recover precious metals from acid leaching waste, particularly silver from zinc leaching waste. However, since lead in smelting slag is primarily present in the form of difficult-to-float silicates, sulfates, and oxides, acid leaching waste contains a large amount of impurity ions, making the flotation recovery of lead a long-standing technical challenge within the industry. Therefore, achieving the comprehensive recovery of silver, zinc, and lead from zinc smelting slag at room temperature using a low-cost flotation process is of great practical significance. Summary of the Invention

[0004] The purpose of the present invention is to provide a flotation process for comprehensively and efficiently recovering silver, zinc and lead from zinc smelting slag with low cost and simple process, and to enhance the lead flotation effect by introducing protective gas, adjusting agent and other slurry adjustment in the lead flotation link.

[0005] The purpose of the present invention is achieved by the following methods:

[0006] A flotation method for comprehensively recovering silver, zinc and lead from zinc smelting slag comprises the following steps:

[0007] (1) Grinding: Grinding zinc smelting slag slurry;

[0008] (2) Silver and zinc flotation: Adding the conditioning agent I, collector I and frother to the ore pulp, and then flotation to obtain silver and zinc concentrates;

[0009] (3) Lead flotation: For the tailings after silver and zinc flotation in step (2), pre-fill with protective gas and stir to mix the slurry, then add the adjusting agent II under the protection of the protective atmosphere, and then add the sulfiding agent to the slurry for sulfidation. After sulfidation, add the collector II and the frother, and then flotation is carried out to obtain the lead concentrate.

[0010] Further,

[0011] The zinc smelting slag slurry described in step (1) is prepared by adding water to zinc smelting slag, and the mass concentration is 20% to 50%.

[0012] Furthermore,

[0013] The zinc smelting slag described in step (1) is derived from a filter press sample of high-acid leaching slag from a zinc smelter, and water-soluble zinc in the zinc smelting slag is recovered by pulping before filter press.

[0014] In step (1), the grinding fineness is -0.074 mm and the content is more than 50%.

[0015] Further,

[0016] In step (2), the adjusting agent I is at least one of water glass and sodium hexametaphosphate, copper sulfate and sodium sulfide; the amount of at least one of water glass, sodium hexametaphosphate and copper sulfate is at least 300 g / t of dry zinc smelting slag, preferably in the range of 300 to 600 g / t of dry zinc smelting slag; the amount of sodium sulfide is at least 500 g / t of dry zinc smelting slag, preferably in the range of 800 to 1600 g / t of dry zinc smelting slag.

[0017] The collector I is at least one of butyl ammonium black powder, No. 25 black powder, No. 31 black powder, and butyl xanthate, and the dosage is at least 300g / t dry zinc smelting slag; the preferred dosage range is 500-800g / t dry zinc smelting slag.

[0018] Further,

[0019] In step (3), protective gas is pre-filled to adjust the slurry to drive out the dissolved oxygen in the slurry (the dissolved oxygen concentration in the slurry is reduced to within 2.0 mg / L, preferably controlled at 0.8-1.5 mg / L); and the charging of protective gas is stopped after the addition of the adjusting agent II and before the addition of the sulfiding agent.

[0020] Furthermore,

[0021] In step (3), the adjusting agent II is at least one of sodium citrate and disodium EDTA and sodium carbonate; the amount of at least one of sodium citrate and disodium EDTA is at least 300 g / t of dry zinc smelting slag; the amount of sodium carbonate is at least 1000 g / t of dry zinc smelting slag. The preferred amount range of sodium carbonate is 5000-10000 kg / t of dry zinc smelting slag; the preferred amount range of at least one of sodium citrate and sodium EDTA is 500-1300 g / t of dry zinc smelting slag.

[0022] The present invention finds that the reason why the sulfidation method is not ideal for flotation of lead in zinc smelting slag is the presence of oxygen in the system. Therefore, for the first time, a protective gas is used to adjust the slurry system before sulfidation to drive out the dissolved oxygen in the slurry, which will greatly improve the effect of sulfidation flotation of lead.

[0023] In addition, the present invention adopts a large dose of alkaline adjuster for the first time, especially the addition of sodium carbonate, which can not only effectively precipitate other metal ions but also continuously produce a large amount of carbon dioxide. After the introduction of protective gas is stopped, it continues to provide support for the oxygen-free environment, thereby ensuring the subsequent vulcanization effect.

[0024] In step (3), the sulfiding agent is at least one of sodium sulfide and sodium hydrosulfide, and the amount used is at least 1000g / t of zinc smelter high acid leaching slag, and the preferred amount of the sulfiding agent is in the range of 5000 to 20000kg / t; the collector II includes at least one of amyl xanthate, octyl xanthate, and mercaptobenzothiazole, and the amount used is at least 300g / t of zinc smelter dry slag, and the preferred amount used is 600 to 800g / t of zinc smelter dry slag.

[0025] Further,

[0026] The foaming agent used in steps (2) and (3) is at least one of 2# oil, pine oil, MIBC, and butyl ether alcohol, and the amount used is at least 10 g / t of dry zinc smelting slag; the preferred amount range is 10-20 g / t of dry zinc smelting slag.

[0027] By the above method, a silver-zinc concentrate with a silver grade of 700-1000 g / t, a zinc grade of 6-8%, a silver recovery rate of 69-74%, and a zinc recovery rate of 35-40% can be obtained, as well as a lead concentrate with a lead grade of 19-23% and a lead recovery rate of 65-70%.

[0028] The dried zinc smelter slag described in this article is obtained by drying filter press samples of high-acid leaching slag from zinc smelters to a constant weight. In actual production, the moist filter press samples of the smelter slag are used directly, rather than the dried slag. The dried slag is used to facilitate the description of reagent dosages.

[0029] The advantages of the present invention are:

[0030] 1. Considering the characteristics of zinc smelting slag, such as strong acidity, fine particle size, and rich valuable metals that are difficult to recover, a flotation process was proposed to control the pulp potential and pH. The process flow is determined to flotate silver and zinc under acidic conditions and lead under a weak alkaline environment. The operation is continuous and highly operable.

[0031] 2. In the lead flotation process, the dissolved oxygen in the ore pulp is driven away by protective gas, the lead minerals are sulfided with a sulfiding agent, and finally the lead is collected with a collector, which reduces the cost of reagents.

[0032] 3. Use large doses of alkaline adjusters to effectively precipitate metal ions such as iron, zinc, and calcium, continuously produce large amounts of carbon dioxide, and cooperate with protective gases to provide protection for subsequent vulcanization effects.

[0033] 4. This method has the characteristics of stable process flow and low cost, solves the problem of low comprehensive recovery and utilization rate of silver, zinc and lead in zinc smelting slag, and realizes partial disposal of zinc smelting slag.

[0034] The following descriptions in conjunction with the accompanying drawings and embodiments are intended to further illustrate the present invention but are not intended to limit the present invention. BRIEF DESCRIPTION OF THE DRAWINGS

[0035] Figure 1 This is a test flow chart of Example 1 of the present invention; DETAILED DESCRIPTION

[0036] The raw ore used in the following examples and comparative examples is a filter press sample of high-acid leaching slag from a zinc smelter in Yunnan. Before filter pressing, the plant had recovered water-soluble zinc from the smelting slag by pulp washing. In the examples, the silver grade of the smelting slag is about 256g / t, the zinc grade is about 4.38%, and the lead grade is about 8.50%. The sodium sulfide used is industrial sodium sulfide with a purity of about 60%. The sodium carbonate used is industrial sodium carbonate with a purity of ≥98.8%. The amyl xanthate used is industrial grade with a purity of ≥99%. The water glass used is industrial grade with a purity of ≥99%. The pine oil used is industrial grade with a purity of ≥99%. The nitrogen used has a purity of ≥99%.

[0037] Example 1

[0038] The test flow chart of this embodiment is shown in Figure 1 , the detailed steps are as follows:

[0039] (1) After the raw ore is slurried at a concentration of 30%, it is ball milled for 2 minutes. The grinding fineness is -0.074mm and the content is more than 50%.

[0040] 300 g / t of copper sulfate, 300 g / t of sodium hexametaphosphate and 1000 g / t of sodium sulfide were added and reacted for 3 minutes. Then, 800 g / t of butyl ammonium chloride and 15 g / t of pine oil were added and reacted for 3 minutes. Then, silver and zinc were roughly separated for 6 minutes to obtain silver-zinc concentrate.

[0041] (2) Nitrogen is pressurized to 0.2m 3 / h flow rate of slurry pretreatment for 4 minutes, the slurry dissolved oxygen concentration is 2.0mg / L. Then, 1000g / t sodium citrate and 8000g / t sodium carbonate are added, reacting for 3 minutes, adjusting the slurry pH to 6.4 and the potential to 320mV. At this point, nitrogen injection is terminated, and 10kg / t sodium sulfide is added to the slurry. The reaction is continued for 3 minutes, adjusting the slurry pH to 7.5 and the potential to 40mV. Finally, 800g / t amyl xanthate and 15g / t pine oil are added, reacting for 3 minutes.

[0042] (3) After the above steps are completed, lead roughing is carried out for 8 minutes to obtain lead concentrate.

[0043] The test results of Example 1 are shown in Table 1.

[0044] Table 1 Experimental results of Example 1

[0045]

[0046] Comparative Example 1

[0047] The difference between Comparative Example 1 and Example 1 is that nitrogen is not used in Comparative Example 1, and the other experimental conditions are completely the same.

[0048] The test results of Comparative Example 1 are shown in Table 2.

[0049] Table 2 Test results of comparative example 1

[0050]

[0051] From the results in Table 2, it can be seen that when nitrogen is not used, the silver and lead grades decrease compared with Example 1, especially the lead recovery rate decreases significantly.

[0052] Comparative Example 2

[0053] The difference between Comparative Example 2 and Example 1 is that, in Comparative Example 2, the nitrogen gas was terminated before adding water glass, sodium citrate and sodium carbonate, and the other experimental conditions were completely the same.

[0054] The test results of Comparative Example 1 are shown in Table 3.

[0055] Table 3 Comparative Example 2 Test Results

[0056]

[0057] It can be seen from the results in Table 3 that when the nitrogen gas is stopped prematurely, the lead grade and lead recovery rate of the lead concentrate are significantly reduced compared with Example 1.

[0058] Comparative Example 3

[0059] The difference between Comparative Example 3 and Example 1 is that the amount of sodium carbonate used in Comparative Example 3 is 6000 g / t, the pH is 6.0, and the other experimental conditions are completely the same.

[0060] The test results of Comparative Example 3 are shown in Table 4.

[0061] Table 4 Comparative Example 3 Test Results

[0062]

[0063] As can be seen from the results in Table 4, when the amount of sodium carbonate is reduced, the lead grade and lead recovery rate of the lead concentrate are significantly reduced compared with Example 1.

[0064] Example 2

[0065] The detailed steps of this embodiment are as follows:

[0066] (1) After the raw ore is slurried at a concentration of 30%, it is ball milled for 2 minutes. The grinding fineness is -0.074mm and the content is more than 50%.

[0067] 300 g / t of copper sulfate, 300 g / t of sodium hexametaphosphate and 1000 g / t of sodium sulfide were added and reacted for 3 minutes. Then, 800 g / t of butyl ammonium chloride and 15 g / t of pine oil were added and reacted for 3 minutes. Then, silver and zinc were roughly separated for 6 minutes to obtain silver-zinc concentrate.

[0068] (2) Nitrogen is pressurized to 0.2m 3 / h flow rate of slurry was added for 4 minutes to pre-treat the slurry, achieving a dissolved oxygen concentration of 2.0 mg / L. Then, 1000g / t sodium citrate and 9000g / t sodium carbonate were added and reacted for 3 minutes, adjusting the slurry pH to 6.5 and the potential to 310mV. Nitrogen injection was terminated at this point, and 11kg / t sodium sulfide was added to the slurry, reacting for 4 minutes until the slurry pH was adjusted to 7.7 and the potential to 33mV. Finally, 800g / t amyl xanthate and 15g / t pine oil were added and reacted for 3 minutes.

[0069] (3) After the above steps are completed, lead roughing is carried out for 8 minutes to obtain lead concentrate.

[0070] The test results are shown in Table 5.

[0071] Table 5 Experimental results of Example 2

[0072]

[0073] Comparative Example 4

[0074] The difference between Comparative Example 2 and Example 2 is that, in Comparative Example 2, the pH and the pulp potential are modulated to the same level as in Example 2 by adjusting the dosage of the reagent without using nitrogen.

[0075] (1) After the raw ore is pulped at a concentration of 30%, it is ground with iron balls for 2 minutes. The grinding fineness is -0.074mm and the content is more than 50%.

[0076] 300 g / t of copper sulfate, 300 g / t of sodium hexametaphosphate and 1000 g / t of sodium sulfide were added and reacted for 3 minutes. Then, 800 g / t of butyl ammonium chloride and 15 g / t of pine oil were added and reacted for 3 minutes. Then, silver and zinc were roughly separated for 6 minutes to obtain silver-zinc concentrate.

[0077] (2) First, add 1000g / t sodium citrate and 9500g / t sodium carbonate, react for 3 minutes, and adjust the slurry pH to 6.5 and the potential to 310mV. Then, add 16kg / t sodium sulfide to the slurry and react for 4 minutes. Then, adjust the slurry pH to 7.7 and the potential to 33mV. Finally, add 800g / t amyl xanthate and 20g / t pine oil and react for 3 minutes.

[0078] (3) After the above steps are completed, lead roughing is carried out for 8 minutes to obtain lead concentrate.

[0079] The test results of Comparative Example 4 are shown in Table 6.

[0080] Table 6 Test results of Comparative Example 4

[0081]

[0082] As shown in Table 6, in the absence of nitrogen, Comparative Example 4 used an additional 500 g / t of sodium carbonate and 5000 g / t of sodium sulfide, and adjusted the pH and slurry potential to the same levels as in Example 2. Compared with Example 2, the lead grade and lead recovery rate of the lead concentrate in Comparative Example 2 decreased significantly.

[0083] Comparative Example 5

[0084] The difference between Comparative Example 5 and Example 2 is that the amount of sodium citrate used in Comparative Example 5 is 300 g / t, and the other experimental conditions are completely the same.

[0085] The test results of Comparative Example 5 are shown in Table 7.

[0086] Table 7 Comparative Example 5 Test Results

[0087]

[0088] As shown in Table 7, when the amount of sodium citrate was reduced, the lead grade and lead recovery rate of the lead concentrate decreased slightly compared with those in Example 2.

Claims

1. A flotation method for comprehensively recovering silver, zinc and lead from zinc smelting slag, characterized in that: The following steps are involved: (1) Grinding: Grinding zinc smelting slag slurry; (2) Silver and zinc flotation: Adding adjusting agent I, collecting agent I and frother to the ore pulp, and then flotation to obtain silver and zinc concentrates; (3) Lead flotation: For the tailings after silver and zinc flotation in step (2), pre-fill with protective gas and stir to mix the slurry, then add the adjusting agent II under the protection of the protective atmosphere, then add the sulfiding agent to the slurry for sulfidation, add the collector II and the frother after sulfidation, and then flotation to obtain the lead concentrate; In step (3), protective gas is pre-filled to adjust the slurry to drive away the dissolved oxygen in the slurry; and the filling of protective gas is stopped after the addition of the adjusting agent II and before the addition of the sulfiding agent; The adjusting agent II is at least one of sodium citrate and disodium edetate and sodium carbonate; the amount of at least one of sodium citrate and disodium edetate is at least 300 g / t of dry zinc smelting slag; the amount of sodium carbonate is at least 1000 g / t of dry zinc smelting slag.

2. The method according to claim 1, wherein: The zinc smelting slag slurry described in step (1) is prepared by adding water to zinc smelting slag, and the mass concentration is 20% to 50%.

3. The method according to claim 1 or 2, characterized in that: The zinc smelting slag described in step (1) is derived from a filter press sample of high-acid leaching slag from a zinc smelter, and water-soluble zinc in the zinc smelting slag is recovered by pulp washing before filter press.

4. The method according to claim 1, wherein: In step (1), the grinding fineness is -0.074 mm and the content is more than 50%.

5. The method according to claim 1, wherein: In step (2), the adjusting agent I is at least one of water glass and sodium hexametaphosphate, copper sulfate and sodium sulfide, and the amount of at least one of water glass and sodium hexametaphosphate and copper sulfate is at least 300g / t of dry zinc smelting slag; the amount of sodium sulfide is at least 500g / t of dry zinc smelting slag; the collecting agent I is at least one of butyl ammonium black medicine, No. 25 black medicine, No. 31 black medicine and butyl xanthate, and the amount is at least 300g / t of dry zinc smelting slag.

6. The method according to claim 1, wherein: In step (3), the sulfiding agent is at least one of sodium sulfide and sodium hydrosulfide, and the amount used is at least 1000 g / t of dry zinc smelting slag; the collector II includes at least one of amyl xanthate, octyl xanthate, and mercaptobenzothiazole, and the amount used is at least 300 g / t of dry zinc smelting slag.

7. The method according to claim 1, wherein: The foaming agent used in steps (2) and (3) is at least one of 2# oil, pine oil, MIBC, and butyl ether alcohol, and the amount used is at least 10 g / t dry zinc smelting slag.

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