A method for improving the recovery of silver from zinc concentrates

Abstract

The present application relates to a method for improving the silver recovery rate in zinc concentrate, belonging to the field of zinc hydrometallurgy. The technical scheme of the present application comprises: controlling the roasting temperature to be 960-1050 DEG C; collecting the electric field dust and the secondary-rotation dust separately to obtain high-silver roasting dust; mixing the boiler dust, the primary-rotation dust and the roasting sand to obtain low-silver roasting ore; separately leaching the high-silver roasting dust to obtain silver concentrate 1; subjecting the low-silver roasting ore to neutral and acid leaching to obtain leaching residue; subjecting the leaching residue to flotation+rotary kiln treatment or fuming furnace treatment to produce silver concentrate 2 or silver-containing zinc oxide dust. The present application increases the proportion of silver into the secondary-rotation dust and the electric field dust, separately treats the secondary-rotation dust and the electric field dust containing high silver to produce silver concentrate, and makes the silver in the zinc concentrate have an additional channel, so that even if the silver content of the zinc concentrate into the furnace is increased, the silver content of the leaching residue can be kept stable, and then the silver loss into the rotary kiln residue or into the fuming furnace water-quenched residue when the tailings of the rotary kiln treatment is reduced, the silver recovery rate in the zinc concentrate is improved, and the economic benefit of the enterprise is increased.

Description

Technical Field

[0001] This invention belongs to the field of zinc hydrometallurgy, and particularly relates to the recovery and utilization of associated silver in zinc concentrate. Background Technology

[0002] Silver is a valuable and rare metal widely used in industry, photography, jewelry, and currency manufacturing. Silver is often found as a byproduct of zinc concentrate production. After desulfurization through a roasting system, zinc concentrate is used to produce roasted zinc ore. This roasted zinc ore is then leached to extract zinc, resulting in leaching residue. The silver in the zinc concentrate is ultimately concentrated in this leaching residue, which is one source of commercial silver ingots. In existing technologies, the Ag content in zinc concentrate is typically controlled to be 60-140 g / t, corresponding to an Ag content of 150-350 g / t in the resulting leaching residue. If the Ag content in the zinc concentrate is greater than 140 g / t, the Ag content in the leaching residue will be greater than 350 g / t.

[0003] Currently, zinc volatilization of leaching residue is mainly carried out using rotary kilns or fuming furnaces. When using rotary kilns for zinc volatilization, the proportion of silver volatilized into the secondary zinc oxide dust is extremely low, with most of the silver being lost in the rotary kiln slag. Therefore, before using rotary kilns for volatilization, the leaching residue needs to be first floated to extract most of the silver into silver concentrate, and the flotation tailings are then volatilized in the rotary kiln to recover zinc. When using fuming furnaces for zinc volatilization, most of the silver volatilizes into the secondary zinc oxide dust, and a small portion of the silver is lost in the fuming furnace water-quenched slag.

[0004] When flotating silver from leaching residue, approximately 80% of the silver enters the silver concentrate, and the tailings produced after flotation contain 40-50 g / t of Ag. However, the silver flotation process has limitations on the silver content of the leaching residue. When the Ag content in the leaching residue is greater than 350 g / t, the proportion of silver entering the silver concentrate during flotation no longer increases, and the silver content in the tailings increases. When the Ag content in the leaching residue is greater than or equal to 400 g / t, the silver content in the tailings rises to greater than or equal to 80 g / t, ultimately increasing the amount of silver lost to the rotary kiln slag.

[0005] The fuming furnace also has restrictions on the silver content of the leaching residue, requiring the silver content to be ≤300g / t. When the silver content of the leaching residue is 100-300g / t, the water-quenched residue produced by the fuming furnace contains ≤30g / t of Ag; when the silver content of the leaching residue is >300g / t, the silver content of the water-quenched residue increases; when the silver content of the leaching residue is 400g / t, the silver content of the water-quenched residue increases to 70-100g / t, thus increasing the amount of silver lost into the water-quenched residue.

[0006] Rotary kiln slag or fuming furnace water-quenched slag is generally used to produce slag cement or as building material. The silver entering the kiln slag or water-quenched slag is ultimately lost, resulting in resource waste and limiting the amount of high-silver zinc concentrate that zinc smelting enterprises can purchase, thus affecting the economic benefits of zinc smelting enterprises. Summary of the Invention

[0007] To overcome the problems existing in the background technology, the present invention provides a method for improving the silver recovery rate in zinc concentrate. When the silver content in zinc concentrate increases, the silver content in flotation tailings can be controlled at Ag≤50g / t or the Ag content in fuming furnace water quenching slag can be controlled at Ag≤30g / t, thereby improving the silver recovery rate in zinc concentrate.

[0008] To achieve the above objectives, the present invention is implemented through the following technical solution: The method for improving silver recovery rate in zinc concentrate includes the following steps: (1) Collect the high-silver roasting dust from the zinc roasting system separately, and mix the low-silver roasting dust and roasting sand to obtain low-silver roasted ore; (2) The high silver roasting dust was leached and the liquid-solid separation was performed to obtain silver concentrate 1; (3) Leach the low-silver roasted ore, and after liquid-solid separation, obtain the leaching residue; (4) The leaching residue is floated to produce silver concentrate 2 and tailings respectively. The tailings are fed into a rotary kiln to volatilize zinc, producing zinc oxide dust and kiln slag respectively; or the leaching residue is fed into a fuming furnace for blowing, producing zinc oxide dust and water-quenched slag respectively.

[0009] Preferably, in step (1), the zinc concentrate used in the zinc roasting system is roasted after being batched.

[0010] As a preferred option, in step (1), the zinc roasting system has a zinc concentrate roasting temperature of 960-1050℃.

[0011] Preferably, the high-silver baking dust in step (1) is two-stage vortex dust and / or electric field dust; the low-silver baking dust is boiler dust and single-stage vortex dust.

[0012] Preferably, the low-silver roasted ore leaching in step (3) is neutral leaching + weak acid leaching or neutral leaching + weak acid leaching + subthermal acid leaching.

[0013] As a preferred option, the process also includes batching and roasting steps; the zinc concentrate is batched and roasted before entering step (1).

[0014] Preferably, in step (1), the roasted sand is ground with boiler dust and then mixed with a vortex dust.

[0015] As a preferred option, in step (2), the leaching conditions are: leaching temperature 70-90℃, leaching time 1-4h, and leaching endpoint acidity 0.005-100g / L.

[0016] Preferably, in step (2), the filtrate after liquid-solid separation enters the neutral leaching in step (3).

[0017] Preferably, in step (3), the neutral leachate is purified, electrolyzed, and cast to produce zinc ingots.

[0018] Preferably, in step (3), the acidic leachate is introduced into the neutral leachate.

[0019] Preferably, in steps (2) and (3), zinc electrolytic waste liquid or dilute sulfuric acid is used as the leaching agent.

[0020] The beneficial effects of this invention are: This invention addresses the main current zinc concentrate processing techniques: first, the zinc concentrate is roasted, then the roasted zinc ore is leached. The leaching residue is then either smelted in a fuming furnace to volatilize zinc, or the leaching residue is first floated for silver, and the flotation tailings are then volatilized in a rotary kiln. The invention resolves the problem in these two processes where increased silver content in the zinc concentrate leads to higher silver content in the leaching residue, resulting in silver loss into the flotation tailings or water-quenched slag. By increasing the zinc roasting temperature to obtain high-silver roasting dust, and then treating this high-silver roasting dust in an open-circuit manner, the silver content in the leaching residue is effectively reduced. If the resulting leaching residue is processed using flotation and rotary kiln silver volatilization, the silver content in the flotation tailings can be reduced. If a fuming furnace smelting process is used, the silver content in the fuming furnace water-quenched slag can be reduced, thus reducing silver loss, increasing the silver recovery rate in the zinc concentrate, reducing resource waste, and allowing zinc smelting companies to increase the proportion of high-silver zinc concentrate they purchase, thereby increasing the company's economic benefits. Attached Figure Description

[0021] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0022] To make the objectives, technical solutions, and beneficial effects of this invention clearer, the technical solutions of this invention will be described in detail below. Obviously, the described embodiments are only a part of the embodiments of this invention, and not all of them. Other embodiments obtained by those skilled in the art based on the embodiments of this invention without creative effort are all within the scope of protection of this invention.

[0023] The method for improving silver recovery in zinc concentrate includes the following steps: (1) Zinc concentrate batching and roasting High-silver-zinc concentrate and low-silver-zinc concentrate are mixed and then roasted and desulfurized in a fluidized bed furnace.

[0024] During the roasting process, roasted sand and roasted dust are produced, each accounting for approximately 50%. The roasted dust includes boiler dust, primary vortex dust, secondary vortex dust, and electric field dust. The inventors discovered that the silver content in roasted dust is higher than that in roasted sand, especially in secondary vortex dust and electric field dust, where the silver content is higher and the dust volume is smaller. The inventors also found that increasing the roasting temperature further increases the silver content in secondary vortex dust and electric field dust, while correspondingly decreasing the silver content in roasted sand. However, roasting temperatures above 1050℃ increase the amount of zinc ferrite generated during roasting, leading to a decrease in zinc leaching rate. Therefore, the roasting temperature of zinc concentrate was increased from the conventional 890-950℃ to 960-1050℃ to increase the proportion of silver entering secondary vortex dust and electric field dust in the zinc concentrate and control the amount of zinc ferrite generated, thereby increasing the silver content in the roasted dust without affecting the zinc leaching rate.

[0025] This invention can use zinc concentrate with high silver content as feedstock, and even if the silver content of the zinc concentrate is increased after feedstock preparation, the silver content in the leaching residue can still be stably controlled.

[0026] (2) Separation of high-silver roasted dust and low-silver roasted ore Among the zinc concentrate roasting products, roasted sand has the lowest silver content; among the roasting dust, the proportions of boiler dust, primary vortex dust, secondary vortex dust and electric field dust decrease from large to small, and the silver content also increases from low to high.

[0027] High-silver-content electric field dust and secondary vortex dust are collected separately to obtain high-silver roasting dust; boiler dust and roasted sand are ball-milled and then mixed with primary vortex dust to obtain low-silver roasted ore.

[0028] (3) High silver dust leach out separately High-silver roasted dust was leached separately using dilute sulfuric acid or electrolytic waste liquid as the leaching agent. The leaching conditions were: leaching temperature 70-90℃, leaching time 1-4 hours, and final acidity of 0.005-100 g / L. After liquid-solid separation, silver concentrate was obtained. Because the leaching pulp contained Cl... - In high-silver roasting dust, a small amount of Ag2SO4 will precipitate into the slag in the form of AgCl. Therefore, all the silver in high-silver roasting dust is enriched in silver concentrate, and the filtrate enters the neutral leaching of low-silver roasted ore.

[0029] Leaching temperature, leaching time, and final acidity at the leaching endpoint determine the yield and silver grade of silver concentrate. When all three are taken at their lower limits, the yield of silver concentrate is the highest, but the silver grade of silver concentrate is the lowest; when all three are taken at their upper limits, the yield of silver concentrate is the lowest, but the silver grade of silver concentrate is the highest.

[0030] (4) Leaching of low-silver roasted ore Using conventional methods, dilute sulfuric acid or electrolytic waste liquid is used as the leaching agent. Low-silver roasted ore undergoes neutral leaching followed by weak acid leaching or neutral leaching followed by weak acid leaching followed by subthermal acid leaching. After liquid-solid separation, leaching residue is obtained. All the silver in the low-silver roasted ore is concentrated in the leaching residue. The principle is also based on Cl... -Ag2SO4 was precipitated; the acidic leaching solution was then introduced into the neutral leaching solution, and the supernatant produced by the neutral leaching solution was purified, electrolyzed, and cast to produce zinc ingots.

[0031] The neutral leaching conditions described in this step are: leaching temperature 70-90℃, leaching time 1.0-3.0h, and final pH value 4.8-5.4; the weak acid leaching conditions are: leaching temperature 70-90℃, leaching time 1.0-4h, and final pH value 2.0-3.5; the subthermal acid leaching conditions are: leaching temperature 70-90℃, leaching time 1.0-4h, and final acidity 20-70g / L.

[0032] (5) Subsequent treatment of leaching residue The leaching residue is treated by either flotation followed by zinc volatilization in a rotary kiln or by zinc volatilization in a fuming furnace. The rotary kiln method involves flotating silver from the leaching residue, producing silver concentrate 2 and tailings. The tailings are then fed into the rotary kiln for zinc volatilization, producing zinc oxide dust and kiln slag. The fuming furnace method involves feeding the leaching residue into the fuming furnace for zinc volatilization, producing zinc oxide dust and water-quenched slag. Most of the silver in the leaching residue enters the zinc oxide dust, which is then treated using conventional methods to recover the silver.

[0033] To illustrate the present invention more clearly, the following embodiments will be described in detail. Example 1

[0034] The zinc concentrate composition (%wt) used in this embodiment after batching is: Zn 46.72, H2O 10.81, Fe 11.06, Ag 210 g / t.

[0035] The zinc concentrate after batching was added to a fluidized bed furnace for roasting and desulfurization, with the roasting temperature controlled at 1050℃. Secondary vortex dust and electric field dust were collected separately to obtain high-silver roasted dust, with the following composition (%wt): Zn 42.36, Fe 12.53, Ag 960 g / t. Roasted sand and boiler dust were ball-milled and then mixed with primary vortex dust to obtain low-silver roasted ore, with the following composition (%wt): Zn 55.85, Fe 12.66, Ag 140 g / t.

[0036] The high-silver roasted dust was leached separately using zinc electrolysis waste liquid as the leaching agent. The leaching conditions were: leaching temperature 70℃, leaching time 1h, and leaching endpoint acidity 0.005g / L (i.e., pH value 4.0). After liquid-solid separation, silver concentrate 1 was obtained, with a yield of 34.22%. The composition of silver concentrate 1 (%wt) was: Zn 20.09, Fe 23.36, Ag 2810g / t. All the silver in the high-silver roasted dust was enriched in silver concentrate 1. Silver concentrate 1 was then used to recover silver in the crude lead smelting unit. The filtrate was then used for neutral leaching of low-silver roasted ore.

[0037] Using dilute sulfuric acid as the leaching agent, the low-silver roasted ore was subjected to neutral and acidic leaching, respectively. The acidic leaching was a weak acid leaching. The neutral leaching conditions were: 80℃, 2.0h, with a final pH of 5.0; the weak acid leaching conditions were: 80℃, 2h, with a final pH of 2.0. After liquid-solid separation, leaching residue was obtained, with the following composition (%wt): Zn 16.32, Fe 24.53, Ag 280g / t. All the silver in the low-silver roasted ore was enriched in the leaching residue. The acidic leachate was returned to the neutral leaching solution, which was then purified, electrolyzed, and cast to produce zinc ingots.

[0038] Silver is floated from the leaching residue to produce silver concentrate 2 and tailings, with the tailings containing 41 g / t of Ag. The silver in silver concentrate 1 and silver concentrate 2 is recovered, and the tailings are fed into a rotary kiln to volatilize zinc. The kiln slag contains 38 g / t of Ag, and the silver loss rate in the kiln slag is 5.31%.

[0039] Comparative Example 1

[0040] Using the zinc concentrate composition after batching in Example 1, the fluidized bed roasting technical conditions are the same as in Example 1. The vortex dust and electric field dust are not collected separately. The vortex dust and electric field dust are mixed with the roasted ore to obtain roasted ore. The roasted ore composition (%wt): Zn 54.45, Fe 12.31, Ag 240g / t.

[0041] Using the leaching conditions of low-silver roasted ore in Example 1, the roasted ore was leached under neutral and weak acid conditions. After liquid-solid separation, leaching residue was obtained. The composition of the leaching residue (%wt) was: Zn 14.23, Fe 23.83, Ag 478 g / t. All the silver in the low-silver roasted ore was enriched in the leaching residue.

[0042] The leaching residue was floated with silver under the same technical conditions as in Example 1, producing silver concentrate 2 and tailings respectively. The tailings contained 82 g / t of Ag. Silver concentrate 2 was used to recover silver in the crude lead smelting unit, and the tailings were used to volatilize zinc in a rotary kiln. The kiln slag contained 80 g / t of Ag, and the silver loss rate in the kiln slag was 11.23%.

[0043] Comparative Example 1 shows that if the high silver content of the cyclone dust and electric field dust is not separated and leached separately, the silver content of the roasted ore increases, which in turn increases the silver content of the leaching residue, exceeding the requirement of Ag ≤ 350 g / t for flotation silver. This results in high silver content in the tailings, which in turn increases the silver content in the rotary kiln slag, and the silver loss rate entering the kiln slag increases by 5.92% compared to Example 1. Example 2

[0044] The zinc concentrate composition (%wt) used in this embodiment after batching is: Zn 49.69, H2O 10.01, Fe 7.96, Ag 160 g / t.

[0045] The zinc concentrate after batching was added to a fluidized bed furnace for roasting and desulfurization, with the roasting temperature controlled at 960℃. Secondary vortex dust and electric field dust were collected separately to obtain high-silver roasted dust, with the following composition (%wt): Zn 49.16, Fe 9.12, Ag 610 g / t. Roasted sand and boiler dust were ball-milled and then mixed with primary vortex dust to obtain low-silver roasted ore, with the following composition (%wt): Zn 59.89, Fe 9.30, Ag 124 g / t.

[0046] The high-silver roasted dust was leached separately using dilute sulfuric acid as the leaching agent. The leaching conditions were: leaching temperature 90℃, leaching time 4h, and final acidity 100g / L. After liquid-solid separation, silver concentrate 1 was obtained with a yield of 18.96%. The composition of silver concentrate 1 (%wt) was: Zn 9.26, Fe 11.33, Ag 3220g / t. All the silver in the high-silver roasted dust was enriched in silver concentrate 1. Silver concentrate 1 was then used to recover silver in the crude lead smelting unit. The filtrate was then used for neutral leaching of low-silver roasted ore.

[0047] The low-silver roasted ore was subjected to neutral and acidic leaching, with the acidic leaching consisting of weak acid leaching followed by subthermal acid leaching. Neutral leaching conditions: 70℃, 1.0h, final pH 4.8; weak acid leaching conditions: 70℃, 1.0h, final pH 2.0; subthermal acid leaching conditions: 70℃, 1.0h, final acidity 2g / L. After liquid-solid separation, leaching residue was obtained, with the following composition (%wt): Zn 11.76, Fe 21.50, Ag 280g / t. All silver in the low-silver roasted ore was enriched in the leaching residue. The acidic leachate was returned to the neutral leaching solution, which was then purified, electrolyzed, and cast to produce zinc ingots.

[0048] The leaching residue is added to the fuming furnace for blowing, producing secondary zinc oxide dust and water-quenched slag respectively. The water-quenched slag contains 25g / t of Ag, and the silver loss rate of the water-quenched slag is 4.26%. The zinc oxide dust is treated with conventional methods to recover zinc and silver, and the water-quenched slag is sold to cement plants.

[0049] Comparative Example 2

[0050] Using the zinc concentrate composition after batching in Example 2, the fluidized bed roasting temperature was controlled at 900°C. The secondary vortex dust and electric field dust were collected separately to obtain high-silver roasting dust, with the following composition (%wt): Zn 48.89, Fe 9.33, Ag 420 g / t. The roasted sand and boiler dust were ball-milled and then mixed with the primary vortex dust to obtain low-silver roasted ore, with the following composition (%wt): Zn 60.03, Fe 9.86, Ag 152 g / t.

[0051] The leaching conditions for high-silver roasted dust were the same as in Example 2. After liquid-solid separation, silver concentrate 1 was obtained with a yield of 21.98%. The composition of silver concentrate (%wt) was: Zn 13.05, Fe 13.63, Ag 1910 g / t. All the silver in the high-silver roasted dust was enriched in silver concentrate 1. Silver concentrate 1 was then used to recover silver in the crude lead smelting unit.

[0052] Under the same conditions as in Example 2, the low-silver roasted ore was subjected to neutral and acidic leaching respectively. After liquid-solid separation, leaching residue was obtained. The composition of the leaching residue (%wt) was: Zn 11.44, Fe 20.77, Ag 360g / t. All the silver in the low-silver roasted ore was enriched in the leaching residue.

[0053] The leaching residue was added to a fuming furnace for blowing, producing secondary zinc oxide dust and water-quenched slag respectively. The water-quenched slag contained 67 g / t of Ag, and the silver loss rate of the water-quenched slag was 11.74%. Zinc and silver in the secondary zinc oxide dust were recovered using conventional methods, and the water-quenched slag was sold to cement plants.

[0054] Comparative Example 2 shows that if the conventional roasting temperature is controlled, the proportion of silver entering the vortex dust and electric field dust decreases, while the proportion entering the low-silver roasted ore increases. This results in the leaching residue containing 360 g / t of Ag, exceeding the requirement of ≤300 g / t for the fuming furnace. Consequently, the silver content in the water-quenched residue increases, and the silver loss rate in the water-quenched residue increases by 7.48% compared to Example 2. Therefore, increasing the roasting temperature is extremely important.

[0055] Finally, it should be noted that the above preferred embodiments are only used to illustrate the technical solutions of the present invention and are not intended to limit it. Although the present invention has been described in detail through the above preferred embodiments, those skilled in the art should understand that various changes can be made to it in form and detail without departing from the scope defined by the claims of the present invention.

Claims

1. A method for improving the silver recovery rate in zinc concentrate, characterized in that, Includes the following steps: (1) Collect the high-silver roasting dust from the zinc roasting system separately, and mix the low-silver roasting dust and roasting sand to obtain low-silver roasted ore; (2) The high silver roasting dust was leached and the liquid-solid separation was performed to obtain silver concentrate 1; (3) Leach the low-silver roasted ore, and after liquid-solid separation, obtain the leaching residue; (4) The leaching residue is floated to produce silver concentrate 2 and tailings respectively. The tailings are fed into a rotary kiln to volatilize zinc, producing zinc oxide dust and kiln slag respectively; or the leaching residue is fed into a fuming furnace for blowing, producing zinc oxide dust and water-quenched slag respectively. In step (1), the zinc roasting system has a roasting temperature of 960-1050℃; the high silver roasting dust mentioned in step (1) is two-stage dust and / or electric field dust, and the low silver roasting dust is boiler dust and single-stage dust.

2. The method according to claim 1, characterized in that, In step (1), the zinc concentrate used in the zinc roasting system is roasted after being batched.

3. The method according to claim 1, characterized in that, In step (1), the roasted sand is ground with boiler dust and then mixed with a vortex of dust to obtain low-silver roasted ore.

4. The method according to claim 1, characterized in that, In step (2), the leaching conditions are: leaching temperature 70-90℃, leaching time 1-4h, and leaching endpoint acidity 0.005-100g / L.

5. The method according to claim 1, characterized in that, The low-silver roasted ore leaching in step (3) is a combination of neutral leaching and weak acid leaching or neutral leaching, weak acid leaching, and subthermal acid leaching.

6. The method according to claim 1 or 4, characterized in that, In step (2), the filtrate after liquid-solid separation enters the neutral leaching in step (3).

7. The method according to claim 1 or 5, characterized in that, In step (3), the neutral leachate is purified, electrolyzed and cast to produce zinc ingots; the acidic leachate enters the neutral leachate.

8. The method according to claim 1, characterized in that, In steps (2) and (3), zinc electrolysis waste liquid or dilute sulfuric acid is used as the leaching agent.

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