Process for jointly recycling zinc and indium by means of open-circuit impurity removal of electrolyte in zinc smelting process
Technical Field
The invention relates to the technical field of chemical smelting treatment, in particular to a process for jointly recycling zinc and indium by removing impurities from electrolyte in an open circuit in a zinc smelting process.
Background
Metal indium is one of important strategic mineral resources, which is an indispensable raw material for the development of strategic emerging industries, and recovery of indium in zinc smelting processes is an important raw indium source. At present, part of smeltery adopts oxygen pressure leaching technology to treat high-iron zinc concentrate, after floatation and hot filtration of leached slag, tailings and sulfur slag are subjected to oxygen-enriched side-blown smelting technology to recover lead and zinc, and smoke dust is produced at the same time, and part of smeltery returns the lead and zinc to a main system. Because the smoke contains indium and fluorine, chlorine, arsenic and antimony, on one hand, the added value of the smoke is reduced, and on the other hand, harmful impurities such as fluorine, chlorine, arsenic and antimony are circulated in the system, so that the subsequent process is not facilitated. Therefore, how to increase the added value of the smoke dust and simultaneously the harmful impurities such as open-circuit fluorine, chlorine, arsenic, antimony and the like become the problems to be solved urgently by zinc smelters. In addition, in the zinc smelting wet method system, the waste electrolyte contains excessive acid and contains a large amount of magnesium ions, if the waste electrolyte is returned to the main system, the magnesium ions can cause adverse effects on subsequent electrodeposition, and the problems of unbalanced acid and excessive magnesium content of the main system are also solved by a zinc smelting plant.
Patent CN 108411128A discloses a method for efficiently enriching indium from low-grade indium-containing smelting slag, which comprises the steps of carrying out concentrated leaching on zinc-smelting underflow slag by adopting high-temperature high acid, and then neutralizing indium to recover indium. The method is mainly aimed at zinc smelting underflow slag, and does not relate to open-circuit impurity removal of waste electrolyte magnesium and comprehensive recovery of zinc in zinc smelting. The CN 104141046A patent discloses a method for recovering indium and zinc from indium-containing zinc oxide soot, which adopts the processes of high-temperature volatilization enrichment, high-indium soot leaching, neutral leaching residue leaching and acid leaching liquid extraction and indium extraction to recover indium, and does not relate to open-circuit impurity removal of waste electrolyte magnesium in zinc smelting and comprehensive recovery of zinc. CN 109628745A patent discloses a method for removing magnesium ions from waste electrolytic solution of zinc hydrometallurgy, which adopts a low-temperature crystallization-centrifugal separation method to remove magnesium from waste electrolytic solution of zinc hydrometallurgy, but does not involve comprehensive recovery of indium and zinc. The CN 108624755A patent discloses a method for opening a circuit of impurities Mg and Cl in a zinc hydrometallurgy system, which utilizes zinc oxide materials (calcine, zinc oxide smoke dust and the like) in the flow of a zinc smelting enterprise to treat zinc sulfate solution with high fluorine, chlorine and magnesium, and simultaneously, the impurities Cl and Mg in the open circuit are not related to the comprehensive recovery of indium and zinc when zinc is recovered.
In view of the above, a plurality of problems existing in the zinc smelting process are necessary to be considered together, so that four problems of indium zinc magnesium recovery, excessive system acid, excessive solution magnesium content and open circuit of fluorine, chlorine, arsenic and antimony are comprehensively solved, and the economic benefit of zinc smelting enterprises is improved.
Disclosure of Invention
The invention aims to solve the technical problem of providing a process which has simpler treatment flow and can simultaneously solve four problems of comprehensive recovery of indium and zinc in smoke dust, excessive acid in a zinc smelting process by oxygen pressure leaching, excessive magnesium content in solution and open circuit of fluorine, chlorine, arsenic and antimony, and the like.
In order to solve the technical problems, the invention adopts the following technical scheme: a process for jointly recycling zinc and indium by using an electrolyte in an open-circuit impurity removal mode in a zinc smelting process is characterized by comprising the following steps of: the method comprises the following steps:
(1) And (3) leaching: directly adopting magnesium-containing waste electrolyte in a zinc-oxygen pressure leaching system to directly carry out primary neutral leaching on smoke dust, wherein the leaching temperature is 60-80 ℃, the leaching time is 1-2 h, and primary leaching liquid and primary leaching slag are obtained after solid-liquid separation; in the first leaching process, zinc oxide reacts with sulfuric acid to form Zn 2+ Is brought into solution. In the first neutral leaching process, a small amount of indium is added In 3+ The form goes into solution and mainly the following reactions occur: znO+H 2 SO 4 =ZnSO 4 +H 2 O,In 2 O 3 +3H 2 SO 4 =In 2 (SO4) 3 +3H2O。
(2) Neutralizing and precipitating zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the neutralizing and precipitating zinc temperature is 20-30 ℃, the neutralizing and precipitating zinc end point pH is 6.8-7.8, and the alkaline zinc carbonate and the precipitating zinc post-liquid are obtained after solid-liquid separation; the step is a neutralization zinc precipitation process, and mainly comprises the following reactions: 2Zn 2+ +CO 3 2- +2OH - =Zn 2 CO 3 (OH) 2 。
(3) Evaporating and crystallizing the zinc-precipitated liquid obtained in the step (2) to obtain magnesium sulfate crystals;
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 60-80 ℃, the leaching time is 4-5 h, and the solid ratio of leaching liquid is 4:1, obtaining a second-stage leaching solution and lead silver slag after solid-liquid separation, wherein the final pH value is 1.0-1.2; during the second-stage acid leaching, zinc and indium enter the solution in the same ionic form as each element during the first-stage leaching, and as most zinc enters the solution during the first-stage neutral leaching, indium enters the second-stage acid leaching along with the underflow, so that the leaching of indium is mainly performed during the second-stage acid leaching.
(5) Extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of a load phase to HCl of 1:1, so as to obtain indium-rich liquid and raffinate;
(6) Neutralizing and precipitating indium in the indium-rich liquid obtained in the step (5), wherein the indium precipitating process firstly adopts NaOH for preneutralization, and the pH value is adjusted to be 1; and then zinc oxide is utilized to carry out indium precipitation, the pH value of the indium precipitation is controlled to be 3.5-4.0, and the indium-rich slag is obtained. The indium deposition process mainly comprises the following reactions: 2In 3+ +6OH - =2In(OH) 3 。
The smoke dust in the step (1) is derived from oxygen pressure leaching slag generated in the process of oxygen pressure leaching zinc smelting, and tailings and sulfur slag obtained after flotation and hot filtration are generated in the process of lead and zinc recovery by adopting oxygen-enriched side-blown smelting and fuming processes; the components of the composition comprise: zn: 50.5-55%, pb:20.1 to 27.43%, in:0.3 to 0.5%, as 0.3 to 0.35%, sb: 0.25-0.28%, F: 0.03-0.06%, cl: 0.1-0.16%, fe: 4.1-6.2%, and 400-600 g/t of Ag.
The magnesium-containing waste electrolyte component in the step (1) comprises the following components: zn: 44.5-47.8 g/L, mg: 25-28 g/L, H 2 SO 4 :160~170g/L。
In the step (1), the pH of the neutral leaching end point is controlled to be 4.2-4.4, the leaching rate of zinc in smoke dust is more than 80%, the leaching rate of indium is less than 8%, and the solid ratio of leaching liquid is 4.5:1.
in the step (2), sodium carbonate is adopted as a precipitator, and the dosage of the sodium carbonate is controlled to be 2.3t/t.Zn, so that the zinc precipitation rate reaches more than 98 percent, and the magnesium entrainment rate is less than 6 percent.
In the step (2), the obtained basic zinc carbonate has a zinc content of more than 45%, a magnesium content of less than 1.5%, a fluorine content of less than 0.009%, a chlorine content of less than 0.005%, an arsenic content of less than 0.007% and an antimony content of less than 0.002%.
In the step (2) and the step (4), most fluorine and antimony enter lead silver slag, part chlorine enters basic zinc carbonate, and arsenic is in a zinc precipitation solution.
The invention has the advantages that: 1. the invention adopts the magnesium-containing waste electrolyte in the zinc-oxygen pressure leaching system to directly leach the soot, and uses hazardous waste to treat hazardous waste, thereby avoiding the problem of separate treatment of the waste electrolyte and the soot, obviously reducing the cost, simplifying the treatment process, and being environment-friendly and efficient;
2. the invention develops a smoke dust and waste electrolyte cooperative treatment technology in the zinc smelting process of oxygen pressure leaching, can simultaneously solve four problems of comprehensive recovery of indium and zinc in the smoke dust, excessive acid content in a system in the zinc smelting process of oxygen pressure leaching, excessive magnesium content in a solution and open circuit of fluorine, chlorine, arsenic and antimony, not only improves the added value of the smoke dust, but also solves the difficult problem in the industrial production of zinc oxygen pressure leaching.
Drawings
FIG. 1 is a process flow diagram of the present invention.
Description of the embodiments
The invention is further illustrated by the following specific examples in connection with fig. 1:
example 1: the method is characterized in that the ash and the waste electrolyte of a certain zinc smelting plant are treated according to the process of the invention, and the ash component and the waste electrolyte component are respectively shown in the following table I and table II:
list one
Element(s)
|
Zn,%
|
In,%
|
Fe,%
|
F,%
|
Cl,%
|
As,%
|
Sb,%
|
Content of
|
54.1
|
0.49
|
4.4
|
0.06
|
0.16
|
0.35
|
0.28 |
Watch II
Element(s)
|
Zn,%
|
Mg,%
|
H 2 SO 4 ,g/L
|
Content of
|
46.79
|
27.88
|
165 |
The method comprises the following specific steps:
(1) Directly adopting magnesium-containing waste electrolyte to carry out one-stage neutral leaching on smoke dust, wherein the leaching temperature is 80 ℃, the leaching time is 1h, and the solid ratio of leaching liquid is 4.5:1, controlling the pH of a leaching end point at 4.3, and obtaining a first-stage leaching solution and a first-stage leaching slag after solid-liquid separation, wherein the leaching rate of Zn is 82.1%, the leaching rate of In is 6.8%, the leaching rate of F is 2.3%, and the leaching rate of Cl is as follows: 49.12%, and the leaching rate of As is As follows: 79.5 percent, the leaching rate of Sb is as follows: 4.58%.
(2) And (3) neutralizing and precipitating the zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the dosage of the neutralizing agent is 2.3t/t.Zn, the neutralizing and precipitating temperature is 20 ℃, the end point pH of the neutralizing and precipitating zinc is 7.4, and the alkaline zinc carbonate and the post-precipitating zinc solution are obtained after solid-liquid separation, wherein the precipitation rate of zinc is 98.9%, the entrainment rate of magnesium is 5.2%, the zinc content of the alkaline zinc carbonate is 45.6%, the magnesium content is 1.16%, the chlorine content is 0.0012%, the arsenic content is 0.0068%, the antimony content is 0.0019% and the fluorine content is 0.0089%.
(3) And (3) evaporating and crystallizing the zinc-precipitated solution obtained in the step (2) to obtain magnesium sulfate crystals.
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 80 ℃, the leaching time is 5h, and the solid ratio of leaching liquid is 4:1, the final pH value is 1.13, and the second-stage leaching solution and lead-silver slag are obtained after solid-liquid separation, wherein the lead-silver slag contains 0.0037 percent of chlorine, 0.11 percent of arsenic, 0.42 percent of antimony and 0.028 percent of fluorine.
(5) And (3) extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of the load phase to the HCl of 1:1, so as to obtain indium-rich liquid and raffinate.
(6) And (3) neutralizing the indium-rich liquid obtained in the step (5) to precipitate indium, wherein the indium precipitation process firstly adopts NaOH to perform preneutralization, the pH value is regulated to be about 1.0, then zinc oxide is utilized to precipitate indium, and the pH value of the indium precipitation is controlled to be 3.5, so that the indium-rich slag is obtained.
In the embodiment, the basic zinc carbonate with the grade of 45.6 percent and the indium-rich slag with the grade of 63kg/t are finally obtained, and the recovery rates of zinc and indium are respectively as follows: 90.1%, 82.3%.
Example 2: the method is characterized in that the ash and the waste electrolyte of a certain zinc smelting plant are treated according to the process of the invention, and the ash component and the waste electrolyte component are respectively shown in the following table III and table IV:
watch III
Element(s)
|
Zn,%
|
In,%
|
Fe,%
|
F,%
|
Cl,%
|
As,%
|
Sb,%
|
Content of
|
54.9
|
0.47
|
4.7
|
0.05
|
0.15
|
0.33
|
0.27 |
Table four
Element(s)
|
Zn,%
|
Mg,%
|
H 2 SO 4 ,g/L
|
Content of
|
45.52
|
26.99
|
168 |
The method comprises the following specific steps:
(1) Directly adopting magnesium-containing waste electrolyte to carry out one-stage neutral leaching on smoke dust, wherein the leaching temperature is 60 ℃, the leaching time is 1.5h, and the solid-to-liquid ratio of the leaching liquid is 4.5:1, controlling the pH of a leaching end point at 4.25, and obtaining a first-stage leaching solution and a first-stage leaching slag after solid-liquid separation, wherein the leaching rate of Zn is 81.5%, the leaching rate of In is 6.1%, the leaching rate of F is 2.0%, and the leaching rate of Cl is as follows: 47.12%, and the leaching rate of As is As follows: 78.5% and the leaching rate of Sb is 4.6%.
(2) And (3) neutralizing and precipitating the zinc from the first-stage leaching solution obtained in the step (1), wherein the neutralizing agent is sodium carbonate, the dosage of the neutralizing agent is 2.3t/t.Zn, the neutralizing and precipitating temperature is 25 ℃, the end point pH of the neutralizing and precipitating zinc is 7.3, and the alkaline zinc carbonate and the post-precipitating zinc solution are obtained after solid-liquid separation, wherein the precipitation rate of zinc is 98.5%, the entrainment rate of magnesium is 5.0%, the zinc content of the alkaline zinc carbonate is 46.1%, the magnesium content is 1.10%, the chlorine content is 0.0011%, the arsenic content is 0.0065%, the antimony content is 0.0019% and the fluorine content is 0.0087%.
(3) And (3) evaporating and crystallizing the zinc-precipitated solution obtained in the step (2) to obtain magnesium sulfate crystals.
(4) Carrying out secondary acid leaching on the primary leaching slag obtained in the step (1), wherein the leaching temperature is 70 ℃, the leaching time is 4.5h, and the solid-to-liquid ratio of the leaching liquid is 4:1, the final pH value is 1.02, and a second-stage leaching solution and lead-silver slag are obtained after solid-liquid separation, wherein the lead-silver slag contains 0.0038% of chlorine, 0.12% of arsenic, 0.44% of antimony and 0.024% of fluorine.
(5) And (3) extracting and back-extracting the second-stage leaching solution obtained in the step (4), wherein the extracting agent is 10% of P204+90% of kerosene, and compared with the ratio of A/O=2:1, the back-extracting is carried out by adopting the volume ratio of the load phase to the HCl of 1:1, so as to obtain indium-rich liquid and raffinate.
(6) And (3) neutralizing the indium-rich liquid obtained in the step (5) to precipitate indium, wherein the indium precipitation process firstly adopts NaOH to perform preneutralization, the pH value is adjusted to be 1.1, then zinc oxide is used for precipitating indium, and the pH value of the precipitated indium is controlled to be 3.7, so that the indium-rich slag is obtained.
In the embodiment, the basic zinc carbonate with the grade of 45.3 percent and the indium-rich slag with the grade of 62.5kg/t are finally obtained, and the recovery rates of zinc and indium are respectively as follows: 89.98% and 81.12%.
The foregoing detailed description of the invention has been presented for purposes of illustration and description, but is not intended to limit the scope of the invention, i.e., the invention is not limited to the details shown and described.