CN105525093A - Method for simplifying zinc hydrometallurgy process - Google Patents
Method for simplifying zinc hydrometallurgy process Download PDFInfo
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- CN105525093A CN105525093A CN201510985762.6A CN201510985762A CN105525093A CN 105525093 A CN105525093 A CN 105525093A CN 201510985762 A CN201510985762 A CN 201510985762A CN 105525093 A CN105525093 A CN 105525093A
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/22—Obtaining zinc otherwise than by distilling with leaching with acids
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/20—Obtaining zinc otherwise than by distilling
- C22B19/26—Refining solutions containing zinc values, e.g. obtained by leaching zinc ores
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/04—Extraction of metal compounds from ores or concentrates by wet processes by leaching
- C22B3/06—Extraction of metal compounds from ores or concentrates by wet processes by leaching in inorganic acid solutions, e.g. with acids generated in situ; in inorganic salt solutions other than ammonium salt solutions
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B3/00—Extraction of metal compounds from ores or concentrates by wet processes
- C22B3/20—Treatment or purification of solutions, e.g. obtained by leaching
- C22B3/44—Treatment or purification of solutions, e.g. obtained by leaching by chemical processes
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- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
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- Manufacturing & Machinery (AREA)
- Metallurgy (AREA)
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Abstract
The invention discloses a method for simplifying a zinc hydrometallurgy process. The method includes the steps that firstly, neutral leaching is conducted on zinc roasted ore so as to obtain neutral leaching liquid and neutral leaching residues; secondly, the neutral leaching residues obtained in the first step are added into acid-containing waste liquid, hot acid leaching is conducted, and acid leaching liquid and acid leaching residues are obtained; thirdly, the acid leaching residues obtained in the second step and ZnS concentrate are added into acid-containing waste liquid, reduction leaching is conducted, and liquid after reduction and reduction residues are obtained; fourthly, zinc roasted ore serving as a neutralizer is added into the liquid after reduction obtained in the third step, neutralizing residues and neutralizing liquid are obtained, and the neutralizing residues are returned to the second step to be subjected to hot acid leaching; fifthly, the reduction residues obtained in the third step are added into the acid leaching liquid obtained in the second step, oxidizing leaching is conducted, oxidizing leaching liquid and lead-silver-sulfide residues are obtained, and the oxidizing leaching liquid is returned to the third step to be subjected to reduction leaching; and sixthly, the neutralizing liquid obtained in the fourth step is added into the zinc roasted ore, iron precipitation with goethite is conducted, and liquid after iron precipitation and goethite residues are obtained.
Description
Technical field
The present invention relates to colored metallurgical industrial field, particularly relate to a kind of method simplifying Zinc hydrometallurgy process.
Background technology
Leaching Technology of Zinc Hydrometallurgy has roasted ore conventional extract technology, roasted ore hot acid extract technology, zinc sulfide concentrates oxygen leaching technique, the ratio of zinc smelting dreg processing cost in zinc abstraction cost that the conventional extract technology of roasted ore produces is raising gradually, and existing above-mentioned technique all exists long flow path, the rate of recovery is low, Slag treatment cost is high, there is the problems such as environmental protection hidden danger.
Therefore, prior art has yet to be improved and developed.
Summary of the invention
In view of above-mentioned the deficiencies in the prior art, the object of the present invention is to provide a kind of method simplifying Zinc hydrometallurgy process, be intended to solve that existing technical process is long, the rate of recovery is low, Slag treatment cost is high, there is the problems such as environmental protection hidden danger.
Technical scheme of the present invention is as follows:
Simplify a method for Zinc hydrometallurgy process, wherein, comprising:
Step one, zinc roasted ore deposit and carry out meutral innersion and obtain middle immersion liquid and middle leaching slag;
Step 2, step one is obtained in soak slag and join and carry out hot acid leaching containing in acid waste liquid, obtain pickling liquor and acid leaching residue;
Step 3, acid leaching residue step 2 obtained and ZnS concentrate join and carry out reducing leaching containing in acid waste liquid, obtain reducing rear liquid and reducing slag;
Add zinc roasting ore deposit after step 4, the reduction that step 3 obtained in liquid and make neutralizing agent, in obtaining and slag and neutralizer, in and slag return step 2 hot acid and leach;
Step 5, reducing slag step 3 obtained join in the pickling liquor in step 2 and carry out Oxidation Leaching, and obtain Oxidation Leaching liquid and lead silver sulphur slag, Oxidation Leaching liquid turns back to step 3 reducing leaching;
Step 6, neutralizer step 4 obtained add and carry out pyrrhosiderite in zinc roasting ore deposit and sink iron, obtain liquid and pyrrhosiderite slag after heavy iron.
The method of described simplification Zinc hydrometallurgy process, wherein, when controlling meutral innersion, initial pH is 2.5, and terminal pH is 4.0 ~ 4.2.
The method of described simplification Zinc hydrometallurgy process, wherein, the residual acid concentration controlling hot acid leaching is 70 ~ 80g/L.
The method of described simplification Zinc hydrometallurgy process, wherein, controls the Fe of the rear liquid of reduction
3+< 2g/L.
The method of described simplification Zinc hydrometallurgy process, wherein, the residual acid concentration that controlled oxidization leaches is 60 ~ 70g/L.
The method of described simplification Zinc hydrometallurgy process, wherein, in control and time, initial pH is 1.0, and terminal pH is 2.4 ~ 2.6.
The method of described simplification Zinc hydrometallurgy process, wherein, control pyrrhosiderite sink iron time, initial pH is 3.5 ~ 4.0, and terminal pH is 2.5.
Beneficial effect: high temperature high acidic oils section and reducing leaching section are merged into a procedure by the present invention, shorten technical process and reaction times, utilize the Fe that flow process self produces under the condition of additionally not adding other oxygenants
3+as oxygenant, to unreacted in reducing leaching slag, ZnS concentrate and sulfur-bearing metallic compound carry out Oxidation Leaching, raising valuable metal leaching yield completely.The invention solves conventional wet lay and leach flow process slag rate greatly, the problem that pyrogenic attack cost is high, solves the problem of Strengthen education long flow path simultaneously.
Accompanying drawing explanation
Fig. 1 is a kind of schema simplifying the method preferred embodiment of Zinc hydrometallurgy process of the present invention.
Embodiment
The invention provides a kind of method simplifying Zinc hydrometallurgy process, for making object of the present invention, technical scheme and effect clearly, clearly, the present invention is described in more detail below.Should be appreciated that specific embodiment described herein only in order to explain the present invention, be not intended to limit the present invention.
A kind of method simplifying Zinc hydrometallurgy process provided by the present invention, it comprises:
Step one (meutral innersion), zinc roasted ore deposit and carry out meutral innersion and obtain middle immersion liquid and middle leaching slag;
Step 2 (hot acid leaching), step one is obtained in soak slag and join and carry out hot acid leaching containing in acid waste liquid, obtain pickling liquor and acid leaching residue;
Step 3 (reducing leaching), acid leaching residue step 2 obtained and ZnS concentrate join and carry out reducing leaching containing in acid waste liquid, obtain reducing rear liquid and reducing slag;
Add zinc roasting ore deposit after step 4 (neutralization), the reduction that step 3 obtained in liquid and make neutralizing agent, in obtaining and slag and neutralizer, in and slag return step 2 hot acid and leach;
Step 5 (Oxidation Leaching), reducing slag step 3 obtained join in the pickling liquor in step 2 and carry out Oxidation Leaching, and obtain Oxidation Leaching liquid and lead silver sulphur slag, Oxidation Leaching liquid turns back to step 3 reducing leaching;
Step 6 (pyrrhosiderite sink iron), neutralizer step 4 obtained add in zinc roasting ore deposit and carry out pyrrhosiderite and sink iron, obtain liquid and pyrrhosiderite slag after heavy iron.
The present invention adopts under the low sour condition of high temperature, processes zinc roasting ore deposit, and the pickling liquor utilizing hot acid to leach to obtain is (containing Fe
3+) make oxygenant, shorten the reaction times of whole flow process, reduce leached mud rate and slag containing zinc, recovery rate of valuable metals of the present invention is high, cost recovery is low, and treating processes environmental protection, environmental pollution is little.
The principal reaction principle related in the present invention has:
(1)
(2)
(3)
(4)
(5)
(6)
(7)
(8)
In reducing leaching of the present invention (step 3) process, temperature chooses 85 ~ 90 DEG C.It is 1.2 that ZnS concentrate adds coefficient.
Oxidation Leaching section (step 5) temperature chooses 85 ~ 90 DEG C, and adds pickling liquor, utilizes the Fe in pickling liquor
3+come ZnS concentrate superfluous in redox slag and sulfur-bearing metallic compound as oxygenant, to reduce plumbous silver-colored sulphur slag containing zinc, reduce metal loss amount.Find in test that acidity is to Oxidation Leaching DeGrain, front liquid Fe
3+concentration has certain influence to oxidation effectiveness, ensures Fe in test
3+amount is the superfluous concentrate theoretical amount at least 2 times of reduction.
Specifically, as shown in Figure 1, in step one, first control the whole acidacidity (namely last acid concentration) of the rear liquid of mixing for 30-80g/L(is as 50g/L by after the ratio mixing of electrolysis waste solution (also can claim containing acid waste liquid) and zinc hydrometallurgy mixed solution 2:1 in mass ratio), Fe
2+< 0.1g/L, then storage tank is placed in, liquid after the mixing obtained and zinc are roasted together with ore deposit (the roasting ore deposit namely in figure) and slowly join reactive tank (having " crystal seed "), regulate pH values of pulp 2.5 and pass into oxygen reaction 40min, after adding the mixing of specified quantity in the 40min time, liquid and zinc roast ore deposit, whole meutral innersion reaction times 2h, temperature is 65-70 DEG C (as 68 DEG C), terminal pH control 4.0-4.2(is as 4.0), reaction terminates rear ore pulp and carries out solid-liquor separation, obtain meutral innersion liquid (middle immersion liquid) and underflow (middle leaching slag), meutral innersion liquid send cleaning section process, underflow carries out hot acid leaching.
In step 2, the underflow that step one is obtained in liquid-solid ratio 10-12:1(as 11:1) ratio join in electrolysis waste solution (namely new containing acid waste liquid), control beginning acidacidity (acid concentration when namely starting) 160-180g/L(as 170g/L), temperature of reaction is 85-90 DEG C (as 88 DEG C), reaction times 3h, whole acidacidity (namely last acid concentration, also residual acid concentration can be claimed): 70-80g/L(is as 75g/L), reaction terminates rear ore pulp and carries out solid-liquor separation, obtain hot acid leach liquor (pickling liquor) and underflow (acid leaching residue), hot acid leach liquor is left to Oxidation Leaching section in advance and uses, underflow carries out reducing leaching.
In step 3, underflow step 2 obtained joins in electrolysis waste solution (new contains acid waste liquid) in the ratio of liquid-solid ratio 10:1, add required ZnS concentrate by Theoretical Calculation simultaneously, control ZnS concentrate add-on coefficient of excess is that namely 1.2(counts in mass ratio, add-on is theoretical amount 1.2 times), temperature of reaction 85-90 DEG C (85 DEG C), reaction times, 3-4h(was as 3h), whole acidacidity: 50-60g/L(55g/L), reaction terminates rear ore pulp and carries out solid-liquor separation, obtain reducing leaching liquid (after reduction liquid) and underflow (reducing slag, also reducing leaching slag can be claimed), reducing leaching liquid is left to neutralizing zone in advance and uses, underflow carries out Oxidation Leaching.
The generation of alum is avoided, liquid Fe after control reduction when liquid carries out neutralization test after considering reduction
3+< 2g/L.
In step 4, reducing leaching liquid step 3 obtained and zinc roast together with ore deposit and slowly join in reactive tank (having " crystal seed "), control pH1.0 reacts 5min, then under pH2.0 condition, 10min is reacted, terminal pH2.4-2.6(is as 2.5), then continue to react (temperature 85-90 DEG C, as 87 DEG C), reaction total time 1h, reaction terminates rear ore pulp and carries out solid-liquor separation, obtain neutralizer and underflow (in and slag), neutralizer is left to the heavy iron section of hydrolysis in advance and uses, and underflow turns back to hot acid extracting stage and carries out hot acid leaching.
In step 5, the underflow that step 3 is obtained in liquid-solid ratio 10-8:1(as 9:1) ratio add in the hot acid leach liquor of step 2 output, control temperature of reaction 85-90 DEG C (as 86 DEG C), reaction times, 3-4h(was as 3h), whole acidacidity: 60-70g/L(65g/L), reaction terminates rear ore pulp and carries out solid-liquor separation, obtain Oxidation Leaching liquid and underflow, Oxidation Leaching liquid turns back to reducing leaching section and uses, and underflow filters and obtains plumbous silver-colored sulphur slag, sells as outside byproduct.
In step 6, neutralizer step 4 obtained and zinc roast together with ore deposit and slowly join in reactive tank (having " crystal seed "), and control pH3.5-4.0(is as 3.6), pass into oxygen reaction 60min simultaneously, period adds neutralizer and required zinc roasting ore deposit, continues to pass into liquid Fe after Oxygen control
2+< 2.5g/L gets final product stopped reaction, terminal pH2.4-2.6(is as 2.5), temperature of reaction 85-90 DEG C (as 87 DEG C), reaction terminates rear ore pulp and carries out solid-liquor separation, liquid (being namely hydrolyzed rear liquid) and pyrrhosiderite slag (i.e. hydrolysis residue) after obtaining heavy iron, after heavy iron, liquid turns back to meutral innersion section and cooks mixed solution use (i.e. step one), sells outside pyrrhosiderite slag.
The specific embodiment one of step 3: carry out reducing leaching to acid leaching residue and Oxidation Leaching liquid and obtain reducing slag and the rear liquid of reduction under 80 DEG C of conditions, wherein, in 4 hours reaction times, it is 1.5 that ZnS concentrate adds coefficient, liquid Fe after reduction
3+concentration 0.53g/L, acidity (H
+) be 81.6g/L.
The specific embodiment two of step 3: carry out reducing leaching to acid leaching residue and Oxidation Leaching liquid and obtain reducing slag and the rear liquid of reduction under 85 DEG C of conditions, wherein, in 3 hours reaction times, it is 1.3 that ZnS concentrate adds coefficient, liquid Fe after reduction
3+concentration 2.26g/L, acidity (H
+) be 74.25g/L.
The specific embodiment three of step 3: carry out reducing leaching to acid leaching residue and Oxidation Leaching liquid and obtain reducing slag and the rear liquid of reduction under 90 DEG C of conditions, wherein, in 2 hours reaction times, it is 1.2 that ZnS concentrate adds coefficient, liquid Fe after reduction
3+concentration 2.02g/L, acidity (H
+) be 65.25g/L.
Increase ZnS concentrate and add coefficient, liquid Fe after reduction
3+< 2g/L far away, but Problems existing adds excessive ZnS, and subsequent oxidation extracting stage is not as having abundant Fe
3+be oxidized too much ZnS, can there is plumbous silver-colored sulphur slag containing the high problem of zinc, therefore, it is 1.2 ~ 1.3 to be advisable that ZnS concentrate adds coefficient.
In the above embodiments, Fe3+ reduction rate and recovery time are all greatly improved relative to prior art.
The present invention adopts the metallurgical mode of Strengthen education, conventional leached mud rate is reduced to 15% from 50%, greatly reduces slag rate, thus reduces pyrogenic process Slag treatment cost.High temperature high acidic oils and reducing leaching two procedures are merged into a procedure by the present invention innovatively, when not adding other additives, utilize the Fe that flow process self produces
3+as oxygenant, adopt zinc ore concentrate reduction Fe
3+avoid the generation of alum and realization response is complete, to improve the leaching yield of roasting ore deposit, Zinc in Zinc Concentrates, copper, indium, reduce the acidity of acidleach, Oxidation Leaching, each operation of reducing leaching, to improve the work-ing life of equipment and to reduce maintenance cost.Solve the long flow path of Strengthen education, the series of problems such as alum, slag rate are comparatively large, indium leaching yield is low are occurred to high ferro, the inadaptability of high silicon raw material, leaching process.
In sum, the present invention is directed to Zinc hydrometallurgy process long flow path in existing technique, operation is loaded down with trivial details, and energy consumption is large, equipment is perishable, the problems such as valuable metal loss is large, adopt the mode of Strengthen education metallurgy, greatly reduce the wet-leaching quantity of slag, reduce pyrogenic process Slag treatment cost, shorten Zinc hydrometallurgy process flow process, improve service life of equipment, reduce valuable metal loss.The inventive method is simple to operate, and technique is easy to control, and the oxygenant utilizing flow process self to produce when not adding other additives can realize the maximized recovery of valuable metal.
Should be understood that, application of the present invention is not limited to above-mentioned citing, for those of ordinary skills, can be improved according to the above description or convert, and all these improve and convert the protection domain that all should belong to claims of the present invention.
Claims (7)
1. simplify a method for Zinc hydrometallurgy process, it is characterized in that, comprising:
Step one, zinc roasted ore deposit and carry out meutral innersion and obtain middle immersion liquid and middle leaching slag;
Step 2, step one is obtained in soak slag and join and carry out hot acid leaching containing in acid waste liquid, obtain pickling liquor and acid leaching residue;
Step 3, acid leaching residue step 2 obtained and ZnS concentrate join and carry out reducing leaching containing in acid waste liquid, obtain reducing rear liquid and reducing slag;
Add zinc roasting ore deposit after step 4, the reduction that step 3 obtained in liquid and make neutralizing agent, in obtaining and slag and neutralizer, in and slag return step 2 hot acid and leach;
Step 5, reducing slag step 3 obtained join in the pickling liquor in step 2 and carry out Oxidation Leaching, and obtain Oxidation Leaching liquid and lead silver sulphur slag, Oxidation Leaching liquid turns back to step 3 reducing leaching;
Step 6, neutralizer step 4 obtained add and carry out pyrrhosiderite in zinc roasting ore deposit and sink iron, obtain liquid and pyrrhosiderite slag after heavy iron.
2. the method for simplification Zinc hydrometallurgy process according to claim 1, is characterized in that, when controlling meutral innersion, initial pH is 2.5, and terminal pH is 4.0 ~ 4.2.
3. the method for simplification Zinc hydrometallurgy process according to claim 1, is characterized in that, the residual acid concentration controlling hot acid leaching is 70 ~ 80g/L.
4. the method for simplification Zinc hydrometallurgy process according to claim 1, is characterized in that, controls the Fe of the rear liquid of reduction
3+< 2g/L.
5. the method for simplification Zinc hydrometallurgy process according to claim 1, is characterized in that, the residual acid concentration that controlled oxidization leaches is 60 ~ 70g/L.
6. the method for simplification Zinc hydrometallurgy process according to claim 1, is characterized in that, in control and time, initial pH is 1.0, and terminal pH is 2.4 ~ 2.6.
7. the method for simplification Zinc hydrometallurgy process according to claim 1, is characterized in that, control pyrrhosiderite sink iron time, initial pH is 3.5 ~ 4.0, and terminal pH is 2.5.
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| CN107728664A (en) * | 2017-05-25 | 2018-02-23 | 中南大学 | A kind of metallurgical process multiple reactor cascades pH value optimal control method |
| CN109971961A (en) * | 2019-04-09 | 2019-07-05 | 云南驰宏锌锗股份有限公司 | A method of handling germanic zinc leaching residue |
| CN110093506A (en) * | 2019-04-09 | 2019-08-06 | 云南驰宏锌锗股份有限公司 | Valuable metal high efficiency extraction and its minimizing processing method in germanic zinc leaching residue |
| CN110295290A (en) * | 2019-07-08 | 2019-10-01 | 云南驰宏资源综合利用有限公司 | A kind of method of acid leaching residue collocation one three sections of oxygen leaching zinc of kettle of lead skim |
| CN110372038A (en) * | 2019-08-15 | 2019-10-25 | 衢州华友资源再生科技有限公司 | A kind of method of raw material containing manganese preparation LITHIUM BATTERY manganese sulfate and its LITHIUM BATTERY manganese sulfate of preparation |
| CN114438340A (en) * | 2022-01-11 | 2022-05-06 | 云南云铜锌业股份有限公司 | Wet zinc smelting leaching process |
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| CN107728664A (en) * | 2017-05-25 | 2018-02-23 | 中南大学 | A kind of metallurgical process multiple reactor cascades pH value optimal control method |
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| CN110093506B (en) * | 2019-04-09 | 2021-03-26 | 云南驰宏锌锗股份有限公司 | Method for efficiently extracting valuable metals from germanium-containing zinc leaching residues and reducing valuable metals |
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| CN110093506A (en) * | 2019-04-09 | 2019-08-06 | 云南驰宏锌锗股份有限公司 | Valuable metal high efficiency extraction and its minimizing processing method in germanic zinc leaching residue |
| CN109971961B (en) * | 2019-04-09 | 2021-03-30 | 云南驰宏锌锗股份有限公司 | Method for treating germanium-containing zinc leaching residue |
| CN110295290A (en) * | 2019-07-08 | 2019-10-01 | 云南驰宏资源综合利用有限公司 | A kind of method of acid leaching residue collocation one three sections of oxygen leaching zinc of kettle of lead skim |
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| CN115011810A (en) * | 2022-01-05 | 2022-09-06 | 昆明理工大学 | Leaching process for improving copper recovery rate in zinc roasted ore |
| CN115011810B (en) * | 2022-01-05 | 2023-09-12 | 昆明理工大学 | Leaching process for improving copper recovery rate in zinc roasting ore |
| CN114438340A (en) * | 2022-01-11 | 2022-05-06 | 云南云铜锌业股份有限公司 | Wet zinc smelting leaching process |
| CN114438340B (en) * | 2022-01-11 | 2023-12-29 | 云南云铜锌业股份有限公司 | Zinc hydrometallurgy leaching process |
| CN114672663A (en) * | 2022-03-11 | 2022-06-28 | 云南云铜锌业股份有限公司 | Wet zinc smelting method |
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