CN114934178B - Method for recovering gold from gold smelting slag chloridizing roasting leacheate - Google Patents

Method for recovering gold from gold smelting slag chloridizing roasting leacheate Download PDF

Info

Publication number
CN114934178B
CN114934178B CN202210706060.XA CN202210706060A CN114934178B CN 114934178 B CN114934178 B CN 114934178B CN 202210706060 A CN202210706060 A CN 202210706060A CN 114934178 B CN114934178 B CN 114934178B
Authority
CN
China
Prior art keywords
gold
organic phase
leacheate
smelting slag
chloridizing roasting
Prior art date
Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
Active
Application number
CN202210706060.XA
Other languages
Chinese (zh)
Other versions
CN114934178A (en
Inventor
王威
柳林
刘红召
曹耀华
王洪亮
Current Assignee (The listed assignees may be inaccurate. Google has not performed a legal analysis and makes no representation or warranty as to the accuracy of the list.)
Zhengzhou Institute of Multipurpose Utilization of Mineral Resources CAGS
Original Assignee
Zhengzhou Institute of Multipurpose Utilization of Mineral Resources CAGS
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Application filed by Zhengzhou Institute of Multipurpose Utilization of Mineral Resources CAGS filed Critical Zhengzhou Institute of Multipurpose Utilization of Mineral Resources CAGS
Priority to CN202210706060.XA priority Critical patent/CN114934178B/en
Priority to ZA2022/08377A priority patent/ZA202208377B/en
Publication of CN114934178A publication Critical patent/CN114934178A/en
Application granted granted Critical
Publication of CN114934178B publication Critical patent/CN114934178B/en
Active legal-status Critical Current
Anticipated expiration legal-status Critical

Links

Images

Classifications

    • 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
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/38—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds containing phosphorus
    • C22B3/387—Cyclic or polycyclic compounds
    • 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
    • C22B11/00—Obtaining noble metals
    • C22B11/04—Obtaining noble metals by wet processes
    • 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
    • C22B3/26—Treatment or purification of solutions, e.g. obtained by leaching by liquid-liquid extraction using organic compounds
    • C22B3/36—Heterocyclic compounds
    • C22B3/362—Heterocyclic compounds of a single type
    • 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
    • C22B1/00—Preliminary treatment of ores or scrap
    • C22B1/02—Roasting processes
    • C22B1/08—Chloridising roasting
    • 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

Landscapes

  • Engineering & Computer Science (AREA)
  • Chemical & Material Sciences (AREA)
  • Mechanical Engineering (AREA)
  • Manufacturing & Machinery (AREA)
  • Materials Engineering (AREA)
  • Metallurgy (AREA)
  • Organic Chemistry (AREA)
  • Geochemistry & Mineralogy (AREA)
  • Geology (AREA)
  • General Life Sciences & Earth Sciences (AREA)
  • Environmental & Geological Engineering (AREA)
  • Life Sciences & Earth Sciences (AREA)
  • Manufacture And Refinement Of Metals (AREA)

Abstract

The invention discloses a method for recovering gold from gold smelting slag chloridizing roasting leacheate, and relates to the technical field of secondary resource comprehensive utilization and metallurgy. The method comprises the following steps: mixing and extracting gold smelting slag chloridizing roasting leacheate and an organic phase containing an extracting agent, and separating a loaded organic phase; mixing the loaded organic phase with an oxalic acid solution for back extraction to obtain sponge gold; the organic phase containing the extracting agent consists of etheramic acid functional ionic liquid and imidazole ionic liquid; the method has excellent selectivity and extraction capability on hardware under the high-acid condition, and the ionic liquid is not easy to volatilize and combust, so that the safety of the operation space can be improved. The method has the advantages of high efficiency, environmental protection, simple operation, low comprehensive cost and the like, can realize the high-efficiency and green recovery of gold in gold smelting slag chloridizing roasting leacheate, and has wide application prospect.

Description

Method for recovering gold from gold smelting slag chloridizing roasting leacheate
Technical Field
The invention relates to the technical field of comprehensive utilization of secondary resources and metallurgy, in particular to a method for recovering gold from gold smelting slag chloridizing roasting leacheate.
Background
At present, the gold concentrate acidizing roasting-acid washing-cyaniding process becomes a main method for treating refractory metallurgical concentrate, and is also one of the main processes adopted by gold production.
The chloridizing roasting process is an important research direction for recovering valuable components such as gold in gold smelting slag. In the chloridizing roasting process, valuable components such as gold and the like are recovered from roasting leaching slurry, most of the adopted processes adopt a reduction method for precipitating gold by adopting gold-sodium sulfite in sodium chlorate leaching ore pulp, but the gold content in the chloridizing roasting leaching solution is lower (less than 10 g/m) 3 ) The consumption of sodium sulfite reagent is high, the gold precipitation efficiency is not high, and a large amount of bubbles and acid gas are generated in the technical process, so that the operating environment is poor. There is a need to develop a more efficient and green gold recovery method, and to realize efficient recovery of gold.
Most of researches on recovering gold in an acidic solution by a solvent extraction method adopt dibutyl carbitol, methyl isobutyl ketone, trioctylamine, tribenzylamine, methyl sulfide and the like (gold, 2016, 37 (7): 56-60.) is researched by a solvent extraction method, and sponge gold is obtained by adopting an oxalic acid reduction mode after extraction, so that the problem of extraction agent volatilization recovery exists. Ionic liquid is taken as a green solvent and is widely concerned in metal extraction and separation, so Jie Wen and the like report a method for extracting and separating Au (III) in high-value waste circuit board leachate by imidazole ionic liquid (Yanjie Wen, pandean, libingyi. Research on extracting and separating Au (III) in high-value waste circuit board leachate by imidazole ionic liquid. Nonferrous metal engineering, 2020,10 (3): 41-45.), an extraction system adopts dibutyl carbitol and imidazole ionic liquid to extract gold in an acidic solution, and then uses an oxalic acid precipitation method to back extract gold, but the effective back extraction of cash can be realized by 3 times of back extraction under the condition that O/A (organic phase volume/aqueous phase volume) is 1/10 (the 1 time of back extraction efficiency is only 32%). At present, no related report of recovering gold from gold smelting slag chloridizing roasting leacheate by adopting an extractant method exists.
Disclosure of Invention
The invention aims to provide a method for recovering gold from gold smelting slag chloridizing roasting leacheate, which solves the problems in the prior art and has the advantages of environmental friendliness, simple process, high efficiency and easiness in recycling and reusing of a solvent.
In order to achieve the purpose, the invention provides the following scheme:
one of the technical schemes of the invention is a method for recovering gold from gold smelting slag chloridizing roasting leacheate, which comprises the following steps:
step 1, mixing and extracting gold smelting slag chloridizing roasting leacheate and an organic phase containing an extracting agent, and separating a loaded organic phase;
step 2, mixing the loaded organic phase with an oxalic acid solution for back extraction to obtain sponge gold;
the organic phase containing the extracting agent consists of ether amide functional ionic liquid and imidazole ionic liquid; the structural general formula of the etheramide functional ionic liquid is as follows:
Figure BDA0003706199570000021
in the formula, X - Is trifluoromethanesulfonimide (Tf) 2 N - ) Or hexafluorophosphate (PF) 6 - ) And n is 7, 9, 11 or 12. Anion selective trifluoromethanesulfonimide (Tf) 2 N - ) Or hexafluorophosphate (PF) 6 - ) Based on the fact that the ionic liquid containing the anions has better hydrophobicity, and the loss of an extracting agent in the extraction process can be reduced. N is selected to be 7, 9, 11 or 12 because the method is selected from a wide range of alkyl groups and the steric structure formed between the alkyl group and the ether amide functional group facilitates selective complexation of gold.
When X is present - Is trifluoromethanesulfonimide (Tf) 2 N - ) When the corresponding ether amide functional ionic liquid is [ BIMDGA ]][Tf 2 N]、[BIMDECGA][Tf 2 N]、[BIMDODECGA][Tf 2 N]Or [ BIMDTEDCGA ]][Tf 2 N]
When X is - Is hexafluorophosphate radical (PF) 6 - ) When the corresponding ether amide functional ionic liquid is [ BIMDGA ]][PF 6 ]、[BIMDECGA][PF 6 ]、[BIMDODECGA][PF 6 ]Or [ BIMDTEDCGA ]][PF 6 ]
The imidazole ionic liquid used as the organic phase diluent has the characteristics of difficult volatilization and nonflammability, and improves the quality and safety of the operating environment.
Further, in the step 1, the hydrochloric acid concentration in the gold smelting slag chloridizing roasting leacheate is 10-20 wt%, and the gold content is 4g/m 3 -15g/m 3 2-5 wt% of iron, 5-15 wt% of copper, 3-8 wt% of lead and 2-6 wt% of zinc。
The ratio of the gold smelting slag chloridizing roasting leacheate to an organic phase containing an extracting agent is 5-10.
Further, in the step 1, the extraction time is 5-20min.
Further, in the step 2, the phase ratio of the loaded organic phase to the oxalic acid solution is 2-10.
Further, in step 2, the back extraction is specifically: back extraction at 80-100 deg.c for 2-6 hr.
Further, in the step 2, the concentration of the oxalic acid solution is 1-5mol/L.
Further, the concentration of the etheramide functional ionic liquid in the organic phase containing the extractant is 0.1-0.5mol/L.
Further, the imidazole ionic liquid is 1-alkyl-3-methylimidazole hexafluorophosphate and/or 1-alkyl-3-methylimidazole bistrifluoromethylsulfonyl imide salt.
Further, the 1-alkyl-3 methylimidazole hexafluorophosphate salt is 1-butyl-3 methylimidazole hexafluorophosphate salt and/or 1-octyl-3 methylimidazole hexafluorophosphate salt; the 1-alkyl-3-methylimidazole bistrifluoromethylsulfonyl imide salt is 1-butyl-3-methylimidazole bistrifluoromethylsulfonyl imide salt and/or 1-octyl-3-methylimidazole bistrifluoromethylsulfonyl imide salt.
According to the second technical scheme, the gold is recovered from the gold smelting slag chloridizing roasting leacheate by using the method.
The invention discloses the following technical effects:
the invention develops a method for recovering gold from gold smelting slag chloridizing roasting leacheate aiming at the conditions of poor operating environment, complex process, low efficiency and the like of the gold recovering process in the existing gold smelting slag chloridizing roasting process. The novel gold extractant system (the etheramide functional ionic liquid and the imidazole ionic liquid) has excellent selectivity and extraction capacity on gold under the high-acid condition, and the ionic liquid is not easy to volatilize and combust, so that the safety of an operation space can be improved. And the decomposition temperature and boiling point of the ionic liquid are higher than those of the conventional organic matters, so that the oxalic acid precipitation operation is carried out under the heating condition without causing the loss of an organic phase. The method has the advantages of high efficiency, environmental protection, simple operation, low comprehensive cost and the like, can realize the high-efficiency and green recovery of gold in gold smelting slag chloridizing roasting leacheate, and has wide application prospect.
Drawings
In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the drawings needed in the embodiments will be briefly described below, and it is obvious that the drawings in the following description are only some embodiments of the present invention, and it is obvious for those skilled in the art to obtain other drawings without creative efforts.
FIG. 1 is a process flow diagram for recovering gold from gold smelting slag chloridizing roasting leacheate.
Detailed Description
Reference will now be made in detail to various exemplary embodiments of the invention, the detailed description should not be construed as limiting the invention but as a more detailed description of certain aspects, features and embodiments of the invention.
It is to be understood that the terminology used herein is for the purpose of describing particular embodiments only and is not intended to be limiting of the invention. Further, for numerical ranges in this disclosure, it is understood that each intervening value, between the upper and lower limit of that range, is also specifically disclosed. Every smaller range between any stated value or intervening value in a stated range and any other stated or intervening value in a stated range is encompassed within the invention. The upper and lower limits of these smaller ranges may independently be included or excluded in the range.
Unless defined otherwise, all technical and scientific terms used herein have the same meaning as commonly understood by one of ordinary skill in the art to which this invention belongs. Although only preferred methods and materials are described herein, any methods and materials similar or equivalent to those described herein can be used in the practice or testing of the present invention. All documents mentioned in this specification are incorporated by reference herein for the purpose of disclosing and describing the methods and/or materials associated with the documents. In case of conflict with any incorporated document, the present specification will control.
It will be apparent to those skilled in the art that various modifications and variations can be made in the specific embodiments of the present disclosure without departing from the scope or spirit of the disclosure. Other embodiments will be apparent to those skilled in the art from consideration of the specification. The description and examples are intended to be illustrative only.
As used herein, the terms "comprising," "including," "having," "containing," and the like are open-ended terms that mean including, but not limited to.
The term "normal temperature" as used herein means 15 to 30 ℃ unless otherwise specified.
The "ratio" in the present invention means a volume ratio unless otherwise specified.
The percentages stated in the examples of the invention are, unless otherwise specified, based on mass percentages.
The technological process chart of the invention for recovering gold from gold smelting slag chloridizing roasting leacheate is shown in figure 1.
The preparation method of the etheramide functional ionic liquid used in the embodiment of the invention comprises the following steps:
weighing 50g of diglycolic anhydride, dissolving the diglycolic anhydride in 300mL of dichloromethane, adding 95g of di-n-octylamine, reacting for 48 hours at normal temperature, washing the obtained solution with a dilute hydrochloric acid solution, dehydrating with anhydrous sodium sulfate, distilling, and drying in vacuum to obtain dioctyl diglycolic amic acid; dissolving 50g dioctyl diglycol amic acid, 16g aminopropyl imidazole, 31g dicyclohexyl carbodiimide and 20g 1-hydroxybenzotriazole in 500mL chloroform, reacting at normal temperature for 12h under the protection of argon, washing with sodium carbonate solution to remove residual 1-hydroxybenzotriazole, vacuum distilling, and performing silica gel column chromatography to obtain 2- [2- (aminopropyl imidazole-2 oxo)]-N, N-dioctylacetamide; weighing 50g2- [2- (aminopropylimidazole)-2 oxo)]Dissolving N, N-dioctyl acetamide and 19g of bromobutane in 500mL of acetonitrile, stirring and reacting for 48h at 80 ℃, washing with N-hexane to remove excessive bromobutane, and performing rotary evaporation to remove acetonitrile to obtain etheramide functionalized imidazole bromide salt; dissolving 50g of etheramide functionalized imidazole bromide in 400mL of acetonitrile, mixing with 43g of acetonitrile solution of lithium bis (trifluoromethanesulfonyl) imide, reacting for 24h at normal temperature, washing with water to remove residual lithium bis (trifluoromethanesulfonyl) imide, and performing rotary evaporation on acetonitrile and water under vacuum conditions to obtain etheramide functional ionic liquid 1-butyl-3-2- [2- (aminopropylimidazole-2 oxo) with bis (trifluoromethanesulfonyl) imide as an anion]-N, N-dioctylacetamidimidazolbistrifluoromethanesulfonylimide, labelled [ BIMDGA][Tf 2 N]。
Weighing 50g of diglycolic anhydride, dissolving the diglycolic anhydride in 300mL of dichloromethane, adding 117g of didecylamine, reacting for 48 hours at normal temperature, washing the obtained solution with a dilute hydrochloric acid solution, dehydrating with anhydrous sodium sulfate, distilling, and drying in vacuum to obtain didecyl diglycolic amic acid; 50g didecyl diglycol amic acid, 14g aminopropyl imidazole, 28g dicyclohexyl carbodiimide, 18g 1-hydroxy-benzotriazole were dissolved in 500mL chloroform, reacted at room temperature under the protection of argon gas for 12 hours, washed with sodium carbonate solution to remove the residual 1-hydroxy-benzotriazole, vacuum distilled, and subjected to silica gel column chromatography to give 2- [2- (aminopropyl imidazole-2 oxo)]-N, N-didecyl acetamide; weighing 50g2- [2- (aminopropylimidazole-2-oxo)]Dissolving N, N-didecyl acetamide and 18g of bromobutane in 500mL of acetonitrile, stirring and reacting for 48h at 80 ℃, washing with N-hexane to remove excessive bromobutane, and performing rotary evaporation to remove acetonitrile to obtain etheramide functionalized imidazole bromide salt; dissolving 50g of etheramide functionalized imidazole bromide in 400mL of acetonitrile, mixing with 32g of acetonitrile solution of potassium hexafluorophosphate, reacting for 24h at normal temperature, washing with water to remove residual potassium hexafluorophosphate, and rotationally evaporating acetonitrile and water under vacuum condition to obtain etheramide functional ionic liquid taking hexafluorophosphate as anion, 1-butyl-3-2- [2- (aminopropylimidazole-2-oxo)]-N, N-didecyl acetamide imidazole hexafluorophosphate, [ BIMDECGA [ ]][PF 6 ]。
Weighing 50g of diglycolic anhydride, dissolving the diglycolic anhydride in 300mL of dichloromethane, adding 139g of didodecylamine, reacting for 48 hours at normal temperature, and reactingWashing the obtained solution with a dilute hydrochloric acid solution, dehydrating with anhydrous sodium sulfate, distilling, and vacuum drying to obtain didodecyl diglycol amic acid; dissolving 50g of didodecyl diglycol amic acid, 13g of aminopropylimidazole, 28g of dicyclohexylcarbodiimide and 15g of 1-hydroxybenzotriazole in 500mL of chloroform, reacting at normal temperature for 12h under the protection of argon, washing with sodium carbonate solution to remove residual 1-hydroxybenzotriazole, vacuum distilling, and performing silica gel column chromatography to obtain 2- [2- (aminopropylimidazole-2 oxo)]-N, N-didodecylacetamide; weighing 50g2- [2- (aminopropylimidazole-2-oxo)]Dissolving N, N-didodecylacetamide and 19g of bromobutane in 500mL of acetonitrile, stirring and reacting at 80 ℃ for 48h, washing by using N-hexane to remove excessive bromobutane, and removing the acetonitrile by rotary evaporation to obtain etheramide functionalized imidazole bromide salt; dissolving 50g of etheramide functionalized imidazole bromide in 400mL of acetonitrile, mixing with 35g of acetonitrile solution of bis (trifluoromethanesulfonylimide) lithium, reacting for 24h at normal temperature, washing with water to remove residual bis (trifluoromethanesulfonimide) lithium, and performing rotary evaporation on acetonitrile and water under vacuum conditions to obtain etheramide functional ionic liquid 1-butyl-3-2- [2- (aminopropylimidazole-2-oxo) with bis (trifluoromethanesulfonimide) as an anion]-N, N-dioctylacetamidimidazolbistrifluoromethanesulfonylimide, labelled [ BIMDODECGA][Tf 2 N]。
Weighing 50g of diglycolic anhydride, dissolving the diglycolic anhydride in 300mL of dichloromethane, adding 150g of isomeric ditridecylamine, reacting for 48 hours at normal temperature, washing the obtained solution with a dilute hydrochloric acid solution, dehydrating with anhydrous sodium sulfate, distilling, and drying in vacuum to obtain isomeric ditridecyldiglycolic amic acid; dissolving 50g of isomeric ditridecyldiglycolic amic acid, 12g of aminopropylimidazole, 25g of dicyclohexylcarbodiimide and 1lg of 1-hydroxybenzotriazole in 500mL of chloroform, reacting at normal temperature for 12h under the protection of argon, washing with a sodium carbonate solution to remove residual 1-hydroxybenzotriazole, vacuum distilling, and performing silica gel column chromatography to obtain 2- [2- (aminopropylimidazole-2 oxo)]-the N, N-isomeric ditridecylacetamides; 50g of 2- [2- (aminopropylimidazole-2-oxo) are weighed]Dissolving N, N-isomeric ditridecyl acetamide and 15g of bromobutane in 500mL of acetonitrile, stirring and reacting at 80 ℃ for 48h, washing with N-hexane to remove excess bromobutane, and rotary evaporating to remove ethylNitrile to obtain ether amide functionalized imidazole bromide salt; dissolving 50g of etheramide functionalized imidazole bromide in 400mL of acetonitrile, mixing with 28g of acetonitrile solution of potassium hexafluorophosphate, reacting for 24h at normal temperature, washing with water to remove residual potassium hexafluorophosphate, and performing rotary evaporation on acetonitrile and water under vacuum conditions to obtain etheramide functional ionic liquid taking hexafluorophosphate as anion, 1-butyl-3-2- [2- (aminopropylimidazole-2-oxo)]-N, N-didecyl acetamide imidazole hexafluorophosphate, [ BIMDTEDCGA [ ]][PF 6 ]。
Example 1
Step 1, gold smelting slag chloridizing roasting leacheate (the concentration of hydrochloric acid is 10 percent, and the gold content is 4 g/m) 3 Iron content of 2%, copper content of 5%, lead content of 3%, zinc content of 2%) and an organic phase containing an extractant (concentration of 0.2mol/L, wherein the extractant is [ BIMDGA ]][Tf 2 N]The organic solvent is 1-butyl-3-methylimidazolium hexafluorophosphate) is extracted according to the proportion that the ratio of the water phase (leacheate) to the organic phase is 5, and a loaded organic phase is separated; the extraction time is 10min; the extraction rate of gold is 97.5%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the organic phase is compared with a water phase (oxalic acid solution) to be 5, and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 1mol/L, the back extraction temperature is 90 ℃, the back extraction time is 2h, and the gold back extraction rate is 96.6%.
Example 2
Step 1, gold smelting slag chloridizing roasting leacheate (the concentration of hydrochloric acid is 15 percent, and the gold content is 4 g/m) 3 Iron content of 4%, copper content of 8%, lead content of 4%, zinc content of 2%) and an organic phase containing an extractant (concentration of 0.3mol/L, wherein the extractant is [ BIMDGA ]][PF 6 ]The organic solvent is 1-octyl-3-methylimidazolium hexafluorophosphate) is extracted according to the proportion that the ratio of the water phase (leacheate) to the organic phase is 8, and a loaded organic phase is separated; the extraction time is 10min; the extraction rate of gold is 98.5%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the ratio of the organic phase to the aqueous phase (oxalic acid solution) is 8, and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 2mol/L, the back extraction temperature is 100 ℃, the back extraction time is 3h, and the gold back extraction rate is 98.2%.
Example 3
Step 1, chloridizing roasting leacheate of gold smelting slag (the concentration of hydrochloric acid is 20%, the content of gold is 10g/m & lt 3 & gt, the content of iron is 4%, the content of copper is 6%, the content of lead is 4%, and the content of zinc is 3%), and an organic phase (the concentration is 0.5mol/L, wherein the extractant is [ BIMDECGA ]][Tf 2 N]The organic solvent is 1-butyl-3-methylimidazole bistrifluoromethanesulfonylimide) according to the proportion that the ratio of a water phase (leacheate) to an organic phase is 10, and a loaded organic phase is separated; the extraction time is 15min; the extraction rate of gold is 99.4%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the organic phase is compared with a water phase (oxalic acid solution) to be 8, and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 5mol/L, the back extraction temperature is 100 ℃, the back extraction time is 6h, and the gold back extraction rate is 99.8%.
Example 4
Step 1, chloridizing roasting leacheate of gold smelting slag (the concentration of hydrochloric acid is 20%, the content of gold is 10g/m & lt 3 & gt, the content of iron is 4%, the content of copper is 6%, the content of lead is 4%, and the content of zinc is 3%), and an organic phase (the concentration is 0.5mol/L, wherein the extractant is [ BIMDECGA ]][PF 6 ]The organic solvent is 1-butyl-3-methylimidazolium hexafluorophosphate) is extracted according to the proportion that the ratio of the water phase (leacheate) to the organic phase is 10, and a loaded organic phase is separated; the extraction time is 20min; the extraction rate of gold is 99.2%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the ratio of the organic phase to the aqueous phase (oxalic acid solution) is 8, and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 5mol/L, the back extraction temperature is 100 ℃, the back extraction time is 6h, and the gold back extraction rate is 99.7%.
Example 5
Step 1, chloridizing roasting leacheate of gold smelting slag (the concentration of hydrochloric acid is 20%, the gold content is 10g/m & lt 3 & gt, the iron content is 4%, the copper content is 6%, the lead content is 4%, and the zinc content is 3%), and an organic phase containing an extracting agent (the concentration is 0.5mol/L, wherein the extracting agent is [ BIMDODECGA ]][PF 6 ]The organic solvent is 1-butyl-3-methylimidazole bistrifluoromethanesulfonylimide) according to the proportion that the ratio of a water phase (leacheate) to an organic phase is 10, and a loaded organic phase is separated; the extraction time is 10min; the extraction rate of gold is 99.5%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the organic phase is 10 to the water phase (oxalic acid solution), and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 5mol/L, the back extraction temperature is 100 ℃, the back extraction time is 6h, and the gold back extraction rate is 99.6%.
Example 6
Step 1, chloridizing roasting leacheate of gold smelting slag (the concentration of hydrochloric acid is 20%, the content of gold is 10g/m & lt 3 & gt, the content of iron is 4%, the content of copper is 6%, the content of lead is 4%, and the content of zinc is 3%), and an organic phase (the concentration is 0.5mol/L, wherein the extractant is [ BIMDODECGA ]][Tf 2 N]The organic solvent is 1-butyl-3-methylimidazole bistrifluoromethanesulfonylimide) according to the proportion that the ratio of a water phase (leacheate) to an organic phase is 10, and a loaded organic phase is separated; the extraction time is 5min; the extraction rate of gold is 99.5%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the organic phase is compared with a water phase (oxalic acid solution) to be 10, and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 5mol/L, the back extraction temperature is 100 ℃, the back extraction time is 6h, and the gold back extraction rate is 99.5%.
Example 7
Step 1, chloridizing roasting leacheate of gold smelting slag (the concentration of hydrochloric acid is 20%, the content of gold is 10g/m & lt 3 & gt, the content of iron is 4%, the content of copper is 6%, the content of lead is 4%, and the content of zinc is 3%) and an organic phase (the concentration of which is 0.5mol & gt/ml & lt/ml & gt) containing an extracting agentL, wherein the extractant is [ BIMDTEDCGA ]][Tf 2 N]The organic solvent is 1-butyl-3-methylimidazole bistrifluoromethanesulfonylimide) according to the proportion that the ratio of a water phase (leacheate) to an organic phase is 10, and a loaded organic phase is separated; the extraction time is 10min; the extraction rate of gold is 99.4%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the organic phase is 10 to the water phase (oxalic acid solution), and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 5mol/L, the back extraction temperature is 100 ℃, the back extraction time is 6h, and the gold back extraction rate is 99.7%.
Example 8
Step 1, chloridizing roasting leacheate of gold smelting slag (the concentration of hydrochloric acid is 20%, the content of gold is 10g/m & lt 3 & gt, the content of iron is 4%, the content of copper is 6%, the content of lead is 4%, and the content of zinc is 3%), and an organic phase (the concentration is 0.5mol/L, wherein the extractant is [ BIMDTEDCGA ]][PF 6 ]The organic solvent is 1-butyl-3-methylimidazole bistrifluoromethanesulfonylimide) according to the proportion that the ratio of a water phase (leacheate) to an organic phase is 10, and a loaded organic phase is separated; the extraction time is 20min; the extraction rate of gold is 99.8%, and the extraction rates of copper, lead, zinc and iron are all below 2%.
Step 2, mixing the loaded organic phase obtained in the step 1 with an oxalic acid solution according to the proportion that the organic phase is 10 to the water phase (oxalic acid solution), and back-extracting gold in the loaded organic phase by using the oxalic acid solution; the concentration of oxalic acid is 5mol/L, the back extraction temperature is 100 ℃, the back extraction time is 6h, and the gold back extraction rate is 99.4%.
The extraction system of the invention can selectively extract gold in gold smelting slag chloridizing roasting leacheate, which is probably because gold is in hydrochloric acid system in the form of chloroauric acid radical ions [ AuCl ] 4 ] - The metal ions such as copper, lead, zinc, iron and the like mostly exist in a cation form, and under a strong acid system, the extraction system extracts gold into an organic phase in a gold chloroauric acid form in a complex extraction mode.
The organic phase after the back extraction can be reused for extracting gold in gold smelting slag chloridizing roasting leacheate after being washed by water, so that the recycling of the organic phase containing the extractant is realized.
The above-described embodiments are merely illustrative of the preferred embodiments of the present invention, and do not limit the scope of the present invention, and various modifications and improvements of the technical solutions of the present invention can be made by those skilled in the art without departing from the spirit of the present invention, and the technical solutions of the present invention are within the scope of the present invention defined by the claims.

Claims (9)

1. A method for recovering gold from gold smelting slag chloridizing roasting leacheate is characterized by comprising the following steps:
step 1, mixing and extracting gold smelting slag chloridizing roasting leacheate and an organic phase containing an extracting agent, and separating a loaded organic phase;
step 2, mixing the loaded organic phase with an oxalic acid solution for back extraction to obtain sponge gold;
the organic phase containing the extracting agent consists of etheramic acid functional ionic liquid and imidazole ionic liquid; the structural general formula of the etheramic acid functional ionic liquid is as follows:
Figure DEST_PATH_IMAGE001
；
in the formula, X - Is trifluoromethanesulfonimide or hexafluorophosphate, and n is 7, 9, 11 or 12;
in the step 1, the hydrochloric acid concentration in the gold smelting slag chloridizing roasting leacheate is 10-20 wt%, and the gold content is 4g/m 3 -15g/m 3 The iron content is 2-5 wt%, the copper content is 5-15 wt%, the lead content is 3-8 wt%, and the zinc content is 2-6 wt%.
2. The method of recovering gold from gold smelting slag chloridizing roasting leachate according to claim 1, wherein the ratio of the gold smelting slag chloridizing roasting leachate to the extractant-containing organic phase is 5-10.
3. The method for recovering gold from gold smelting slag chloridizing roasting leacheate according to claim 1, wherein in the step 1, the extraction time is 5-20min.
4. The method for recovering gold from gold smelting slag chloridizing roasting leacheate according to claim 1, wherein in step 2, the ratio of the loaded organic phase to the oxalic acid solution is 2-10.
5. The method for recovering gold from gold smelting slag chloridizing roasting leacheate according to claim 1, wherein in the step 2, the back extraction is specifically as follows: back extraction at 80-100 deg.c for 2-6 hr.
6. The method for recovering gold from gold smelting slag chloridizing roasting leacheate according to claim 1, wherein in the step 2, the concentration of the oxalic acid solution is 1-5mol/L.
7. The method for recovering gold from gold smelting slag chloridizing roasting leacheate according to claim 1, wherein the concentration of the etheramic acid functional ionic liquid in the organic phase containing the extractant is 0.1 to 0.5mol/L.
8. The method for recovering gold from gold smelting slag chloridizing roasting leacheate according to claim 1, wherein the imidazole-based ionic liquid is 1-alkyl-3 methylimidazole hexafluorophosphate and/or 1-alkyl-3 methylimidazole bistrifluoromethanesulfonylimide salt.
9. The method of recovering gold from gold smelting slag chloridizing roasting leacheate of claim 8, wherein the 1-alkyl-3 methylimidazolium hexafluorophosphate is 1-butyl-3 methylimidazolium hexafluorophosphate and/or 1-octyl-3 methylimidazolium hexafluorophosphate; the 1-alkyl-3-methylimidazole bis-trifluoromethanesulfonimide salt is 1-butyl-3-methylimidazole bis-trifluoromethanesulfonimide salt and/or 1-octyl-3-methylimidazole bis-trifluoromethanesulfonimide salt.
CN202210706060.XA 2022-06-21 2022-06-21 Method for recovering gold from gold smelting slag chloridizing roasting leacheate Active CN114934178B (en)

Priority Applications (2)

Application Number Priority Date Filing Date Title
CN202210706060.XA CN114934178B (en) 2022-06-21 2022-06-21 Method for recovering gold from gold smelting slag chloridizing roasting leacheate
ZA2022/08377A ZA202208377B (en) 2022-06-21 2022-07-27 Method for recovering gold from gold smelting slag chloridizing roasting eluent

Applications Claiming Priority (1)

Application Number Priority Date Filing Date Title
CN202210706060.XA CN114934178B (en) 2022-06-21 2022-06-21 Method for recovering gold from gold smelting slag chloridizing roasting leacheate

Publications (2)

Publication Number Publication Date
CN114934178A CN114934178A (en) 2022-08-23
CN114934178B true CN114934178B (en) 2023-03-24

Family

ID=82867961

Family Applications (1)

Application Number Title Priority Date Filing Date
CN202210706060.XA Active CN114934178B (en) 2022-06-21 2022-06-21 Method for recovering gold from gold smelting slag chloridizing roasting leacheate

Country Status (2)

Country Link
CN (1) CN114934178B (en)
ZA (1) ZA202208377B (en)

Families Citing this family (1)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN116217802A (en) * 2023-03-08 2023-06-06 山东圳谷新材料科技有限公司 Imidazole type ionic liquid polymer and preparation method and application thereof

Family Cites Families (10)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN101748276A (en) * 2009-12-31 2010-06-23 中国科学院成都有机化学有限公司 Process for extracting, separating and purifying Ag, Au, Pd and Pt
CN104404569B (en) * 2014-11-04 2018-02-23 常州钇金环保科技有限公司 The preparation technology of high pure gold
CN105112677B (en) * 2015-09-16 2017-10-17 中国地质科学院郑州矿产综合利用研究所 Method for comprehensively recovering valuable metals in gold smelting slag
CN106148704B (en) * 2016-06-30 2018-06-29 华南理工大学 One kind recycles gold from waste mobile phone electronic component（Au）Method
CN108531746B (en) * 2017-03-02 2020-07-14 厦门稀土材料研究所 Extracting agent for separating noble metal and method for extracting and separating noble metal by using extracting agent
CN107287456B (en) * 2017-06-20 2019-01-18 江西省科学院应用化学研究所 A kind of extracting process of separating-purifying heavy rare earth
CN109837400B (en) * 2018-11-14 2021-08-06 中国地质科学院郑州矿产综合利用研究所 Method for extracting rare earth and niobium from niobium-rich slag
CN110042235B (en) * 2019-04-29 2020-10-02 山东大学 Extracting agent and extracting method for separating gold
JP2021075755A (en) * 2019-11-08 2021-05-20 国立研究開発法人産業技術総合研究所 Highly efficient extraction agent of gold ion, and separation recovery method of gold using the same
CN114436997B (en) * 2022-01-19 2023-06-30 山东大学 Amide ionic liquid, synthesis method thereof and application of amide ionic liquid in extraction and separation of gold

Also Published As

Publication number Publication date
ZA202208377B (en) 2022-11-30
CN114934178A (en) 2022-08-23

Similar Documents

Publication Publication Date Title
Ashiq et al. Hydrometallurgical recovery of metals from e-waste
CN101619399B (en) Method for selectively leaching voluble metal in copper refinery cinder
CN114934178B (en) Method for recovering gold from gold smelting slag chloridizing roasting leacheate
CN103526057B (en) Method of comprehensive recovery of rare precious metals in rare earth smelting slag
CN101660054A (en) Method for extracting metal indium from waste residues generated from lead and zinc smelting
CN106399686B (en) A kind of acidic ion liquid and the method that electrolysis separating-purifying rare earth or rare precious metal are coupled using molten extraction
CN105112674A (en) All-wet recovery process for waste circuit boards
CN106222430B (en) Method for recovering copper and cobalt from copper-cobalt slag by wet metallurgy
CN112981139B (en) Hydrophobic eutectic solvent for separating nickel and cobalt ions, preparation method thereof and method for separating nickel and cobalt ions
CN107488789A (en) A kind of method of the co-desorption gold thiosulfate anion complex ion from strong-base anion-exchange resin
CN106065434A (en) A kind of method of the wet method direct purifying gold of synthetical recovery silver anode slime
Davris et al. Leaching of rare earths from bauxite residues using imidazolium based ionic liquids
CN114134337B (en) Low-cost and high-efficiency method for recovering noble metal in circuit board components
Balomenos et al. Scandium extraction from bauxite residue using sulfuric acid and a composite extractant-enhanced ion-exchange polymer resin
Zhang et al. Separation and extraction of scandium from titanium dioxide waste acid
CN1966745A (en) Extraction and separation technology for reclaiming main impurity in indium in hydrochloric acid system
Liu et al. An integrated and sustainable hydrometallurgical process for enrichment of precious metals and selective separation of copper, zinc, and lead from a roasted sand
CN111218562B (en) Preparation of high-molecular extracting agent and method for efficiently recovering palladium by using high-molecular extracting agent
CN109837400B (en) Method for extracting rare earth and niobium from niobium-rich slag
CN105132695A (en) Technology for recovering silver from zinc hydrometallurgy leaching residue
CN114717420B (en) Method for recycling copper, silver and gold step by utilizing waste printed circuit board
CN103088218B (en) Silver, plumbous method is extracted from the smelting slag that copper anode mud pyrogenic attack produces
CN111346738B (en) S-alkoxy alkyl ester collecting agent and preparation and application thereof
CN109468472A (en) A method of separating cobalt from the material containing cobalt and nickel
CN109881004B (en) Method for reducing gold grade in crude antimony leaching solution

Legal Events

Date Code Title Description
PB01 Publication
PB01 Publication
SE01 Entry into force of request for substantive examination
SE01 Entry into force of request for substantive examination
GR01 Patent grant
GR01 Patent grant