CN114480862B - Method for recovering valuable elements from copper dross - Google Patents
Method for recovering valuable elements from copper dross Download PDFInfo
- Publication number
- CN114480862B CN114480862B CN202210085349.4A CN202210085349A CN114480862B CN 114480862 B CN114480862 B CN 114480862B CN 202210085349 A CN202210085349 A CN 202210085349A CN 114480862 B CN114480862 B CN 114480862B
- Authority
- CN
- China
- Prior art keywords
- leaching
- roasting
- slag
- copper
- copper dross
- 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
Links
- 239000010949 copper Substances 0.000 title claims abstract description 96
- 229910052802 copper Inorganic materials 0.000 title claims abstract description 75
- RYGMFSIKBFXOCR-UHFFFAOYSA-N Copper Chemical compound [Cu] RYGMFSIKBFXOCR-UHFFFAOYSA-N 0.000 title claims abstract description 68
- 238000000034 method Methods 0.000 title claims abstract description 29
- 238000002386 leaching Methods 0.000 claims abstract description 140
- 239000002893 slag Substances 0.000 claims abstract description 62
- AFVFQIVMOAPDHO-UHFFFAOYSA-N Methanesulfonic acid Chemical compound CS(O)(=O)=O AFVFQIVMOAPDHO-UHFFFAOYSA-N 0.000 claims abstract description 32
- 239000000779 smoke Substances 0.000 claims abstract description 24
- 229910052787 antimony Inorganic materials 0.000 claims abstract description 23
- 239000000428 dust Substances 0.000 claims abstract description 21
- 239000007788 liquid Substances 0.000 claims abstract description 20
- 238000000926 separation method Methods 0.000 claims abstract description 19
- WATWJIUSRGPENY-UHFFFAOYSA-N antimony atom Chemical compound [Sb] WATWJIUSRGPENY-UHFFFAOYSA-N 0.000 claims abstract description 18
- 239000003795 chemical substances by application Substances 0.000 claims abstract description 16
- 229910052797 bismuth Inorganic materials 0.000 claims abstract description 15
- 230000001590 oxidative effect Effects 0.000 claims abstract description 14
- 229940098779 methanesulfonic acid Drugs 0.000 claims abstract description 11
- 239000002253 acid Substances 0.000 claims abstract description 9
- JCXGWMGPZLAOME-UHFFFAOYSA-N bismuth atom Chemical compound [Bi] JCXGWMGPZLAOME-UHFFFAOYSA-N 0.000 claims abstract description 8
- 239000000843 powder Substances 0.000 claims abstract description 8
- 238000006243 chemical reaction Methods 0.000 claims description 28
- 229910052709 silver Inorganic materials 0.000 claims description 23
- 239000007787 solid Substances 0.000 claims description 20
- QAOWNCQODCNURD-UHFFFAOYSA-N Sulfuric acid Chemical compound OS(O)(=O)=O QAOWNCQODCNURD-UHFFFAOYSA-N 0.000 claims description 16
- 239000004332 silver Substances 0.000 claims description 16
- 150000007522 mineralic acids Chemical class 0.000 claims description 14
- VEXZGXHMUGYJMC-UHFFFAOYSA-N Hydrochloric acid Chemical compound Cl VEXZGXHMUGYJMC-UHFFFAOYSA-N 0.000 claims description 6
- 229910052745 lead Inorganic materials 0.000 claims description 6
- 239000004071 soot Substances 0.000 claims 2
- 229910052751 metal Inorganic materials 0.000 abstract description 6
- 239000011133 lead Substances 0.000 description 34
- 238000011084 recovery Methods 0.000 description 23
- BQCADISMDOOEFD-UHFFFAOYSA-N Silver Chemical compound [Ag] BQCADISMDOOEFD-UHFFFAOYSA-N 0.000 description 13
- 238000004064 recycling Methods 0.000 description 7
- 229910015902 Bi 2 O 3 Inorganic materials 0.000 description 6
- QVGXLLKOCUKJST-UHFFFAOYSA-N atomic oxygen Chemical compound [O] QVGXLLKOCUKJST-UHFFFAOYSA-N 0.000 description 5
- 229910052760 oxygen Inorganic materials 0.000 description 5
- 239000001301 oxygen Substances 0.000 description 5
- 238000004070 electrodeposition Methods 0.000 description 3
- XEEYBQQBJWHFJM-UHFFFAOYSA-N iron Substances [Fe] XEEYBQQBJWHFJM-UHFFFAOYSA-N 0.000 description 3
- 239000002184 metal Substances 0.000 description 3
- 238000002156 mixing Methods 0.000 description 3
- 239000000126 substance Substances 0.000 description 3
- 238000003723 Smelting Methods 0.000 description 2
- CDBYLPFSWZWCQE-UHFFFAOYSA-L Sodium Carbonate Chemical compound [Na+].[Na+].[O-]C([O-])=O CDBYLPFSWZWCQE-UHFFFAOYSA-L 0.000 description 2
- 238000002441 X-ray diffraction Methods 0.000 description 2
- 230000008901 benefit Effects 0.000 description 2
- 239000000571 coke Substances 0.000 description 2
- 229910052500 inorganic mineral Inorganic materials 0.000 description 2
- 229910052742 iron Inorganic materials 0.000 description 2
- 238000002844 melting Methods 0.000 description 2
- 230000008018 melting Effects 0.000 description 2
- 150000002739 metals Chemical class 0.000 description 2
- 239000011707 mineral Substances 0.000 description 2
- 150000007524 organic acids Chemical class 0.000 description 2
- 229920006395 saturated elastomer Polymers 0.000 description 2
- 238000009825 accumulation Methods 0.000 description 1
- 150000007513 acids Chemical class 0.000 description 1
- 229910000410 antimony oxide Inorganic materials 0.000 description 1
- 230000009286 beneficial effect Effects 0.000 description 1
- 238000001354 calcination Methods 0.000 description 1
- 239000003245 coal Substances 0.000 description 1
- 239000012141 concentrate Substances 0.000 description 1
- 230000007547 defect Effects 0.000 description 1
- 238000005265 energy consumption Methods 0.000 description 1
- 238000007667 floating Methods 0.000 description 1
- PCHJSUWPFVWCPO-UHFFFAOYSA-N gold Chemical compound [Au] PCHJSUWPFVWCPO-UHFFFAOYSA-N 0.000 description 1
- 229910052737 gold Inorganic materials 0.000 description 1
- 239000010931 gold Substances 0.000 description 1
- 238000010438 heat treatment Methods 0.000 description 1
- 229910052739 hydrogen Inorganic materials 0.000 description 1
- 239000000463 material Substances 0.000 description 1
- 238000005272 metallurgy Methods 0.000 description 1
- 239000000203 mixture Substances 0.000 description 1
- 238000012986 modification Methods 0.000 description 1
- 230000004048 modification Effects 0.000 description 1
- 230000003647 oxidation Effects 0.000 description 1
- 238000007254 oxidation reaction Methods 0.000 description 1
- VTRUBDSFZJNXHI-UHFFFAOYSA-N oxoantimony Chemical compound [Sb]=O VTRUBDSFZJNXHI-UHFFFAOYSA-N 0.000 description 1
- 238000000746 purification Methods 0.000 description 1
- 239000002994 raw material Substances 0.000 description 1
- 238000007670 refining Methods 0.000 description 1
- 229910000029 sodium carbonate Inorganic materials 0.000 description 1
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
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/04—Working-up slag
-
- 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
- C22B15/00—Obtaining copper
- C22B15/0002—Preliminary treatment
- C22B15/001—Preliminary treatment with modification of the copper constituent
- C22B15/0013—Preliminary treatment with modification of the copper constituent by roasting
- C22B15/0015—Oxidizing roasting
-
- 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
- C22B15/00—Obtaining copper
- C22B15/0063—Hydrometallurgy
- C22B15/0065—Leaching or slurrying
- C22B15/0067—Leaching or slurrying with acids or salts thereof
-
- 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
- C22B30/00—Obtaining antimony, arsenic or bismuth
- C22B30/02—Obtaining antimony
-
- 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
- C22B30/00—Obtaining antimony, arsenic or bismuth
- C22B30/06—Obtaining bismuth
-
- 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
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—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
- C22B7/00—Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
- C22B7/006—Wet processes
- C22B7/007—Wet processes by acid leaching
Landscapes
- Engineering & Computer Science (AREA)
- Chemical & Material Sciences (AREA)
- Manufacturing & Machinery (AREA)
- Materials Engineering (AREA)
- Mechanical Engineering (AREA)
- Metallurgy (AREA)
- Organic Chemistry (AREA)
- Life Sciences & Earth Sciences (AREA)
- Environmental & Geological Engineering (AREA)
- General Life Sciences & Earth Sciences (AREA)
- Geology (AREA)
- Manufacture And Refinement Of Metals (AREA)
Abstract
The invention relates to a method for recovering valuable elements from copper dross, which comprises the steps of crushing copper dross to be treated to obtain copper dross powder; oxidizing and roasting copper dross powder to obtain roasting slag and smoke dust; acid leaching is carried out on the roasting slag; and leaching the smoke dust by taking methanesulfonic acid as a leaching agent, and carrying out solid-liquid separation to obtain a first leaching solution rich in bismuth and a first leaching slag rich in antimony. The method is simple and practical, has low cost, and can effectively separate and recover valuable metal elements in the copper dross.
Description
Technical Field
The invention relates to a method for recovering valuable elements from copper dross, belonging to the field of nonferrous metal metallurgy.
Background
Copper dross is produced in the refining process of the lead bullion, and at present, the copper dross is generally treated in a reverberatory furnace by a soda-iron scrap method in industry, namely, corresponding iron scraps, sodium carbonate, coke and a small amount of lead concentrate are added according to the composition of the copper dross, and reduction smelting is carried out to produce the lead bullion and copper matte. However, the above-mentioned treatment process has a problem that the melting operation time is long, and even a treatment time of 12 hours or more is required, the melting yield is low, and a large amount of lead-containing matte is still produced. The lead matte contains a large amount of gold and silver, meanwhile, the slag and the matte are not well separated, copper is contained in the slag, the copper is contained in the matte, most of valuable metals in the slag and the matte are difficult to recycle, a large amount of accumulation is caused, funds are accumulated, and the benefit of enterprises is influenced.
Disclosure of Invention
Aiming at the defects of the prior art, the invention provides a method for recycling valuable elements from copper dross, so as to better realize separation and recycling of various valuable metal elements in the copper dross.
In order to solve the technical problems, the technical scheme of the invention is as follows:
a method for recovering valuable elements from copper dross, comprising the steps of:
s1, crushing copper dross to be treated to obtain copper dross powder;
s2, oxidizing and roasting the copper dross powder obtained in the step S1 to obtain roasting slag and smoke dust;
s3, carrying out acid leaching on the roasting slag obtained in the step S2;
and leaching the smoke dust obtained in the step S2 by taking methanesulfonic acid as a leaching agent, and carrying out solid-liquid separation to obtain a first leaching solution rich in bismuth and a first leaching slag rich in antimony.
As one of the implementation modes of the invention, when the content of Ag in the copper dross is more than 1000g/t, the content of Pb is more than 70wt%, and the content of Cu is 2-10wt%, in S2, the temperature of oxidizing roasting is 400-750 ℃, further 500-700 ℃, further 550-650 ℃, and the roasting time is 1-4 hours, further 2-3 hours;
s3, performing solid-liquid separation after performing primary leaching on the roasting slag by using inorganic acid as a leaching agent to obtain second leaching slag and second leaching liquid rich in copper; then, performing secondary leaching on the second leaching slag by taking methanesulfonic acid as a leaching agent, and performing solid-liquid separation to obtain third leaching slag rich in silver and third leaching liquid rich in lead;
wherein the inorganic acid is sulfuric acid and/or hydrochloric acid. Thus, the main component of the obtained roasting slag is CuO, pbO, ag, and the main component of the obtained smoke dust is Bi 2 O 3 、Sb 2 O 3 The method comprises the steps of carrying out a first treatment on the surface of the The copper can be recovered after entering the second leaching solution by leaching with inorganic acid, lead and silver are reserved in the second leaching slag, and silver (mainly silver simple substance) is reserved in the third leaching slag by leaching with methylsulfonic acid, and lead is enriched in the third leaching solution to generate (CH) 3 SO 3 ) 2 Pb, thereby realizing the separation, enrichment and recovery of copper, lead and silver. Thus, roasting at lower temperature and acid leaching of specific kind of acid are combinedThe method can realize separation and recovery of Pb, cu, bi, sb, ag and other elements in the floating slag, and is simple, practical and low in energy consumption. The Cu leaching rate can reach more than 99 percent, and the Pb leaching rate can reach more than 98 percent. The second leaching solution can be repeatedly used for leaching, and the solution is sent to a copper purification and electrodeposition system when meeting the copper electrodeposition concentration requirement. The main component of the third leaching slag is simple substance Ag, the slag rate is less than 1%, and the Ag content is more than 80%.
According to the invention, the smoke dust is leached by the methylsulfonic acid, so that bismuth is enriched in the first leaching solution, and antimony is remained in the first leaching residue (the antimony mainly exists in the form of antimony oxide), thereby realizing separation, enrichment and recovery of bismuth and antimony. The first leaching residue can be further sent to an antimony smelting system for recycling antimony.
Further, during one-stage leaching, H in the reaction system is controlled + And Cu is in an initial molar ratio of 2-4:1, the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h; during the second leaching, CH in the reaction system is controlled 3 SO 3 The initial molar ratio of H to Pb is 2-4:1, the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
Further, in S3, when the smoke dust is leached, CH in the reaction system is controlled 3 SO 3 The initial molar ratio of H to Bi is 3-5:1, the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
As a further embodiment of the present invention, when the content of Ag is 500-1000g/t, the content of Pb is 30-50wt%, and the content of Cu is 10-40wt% in the copper dross, the temperature of the oxidizing roasting is 900-1200 ℃ and the roasting time is 1-4 hours in S2;
s3, when the Cu content in the copper dross is less than or equal to 20wt%, leaching the roasting slag by using methanesulfonic acid as a leaching agent, and performing solid-liquid separation to obtain a fourth leaching solution rich in copper and silver; thus, MSA leaching can realize leaching of Cu and Ag, so that Cu and Ag enter the solution. Basically no leaching slag, and a very small amount of unreacted leaching slag can be used as a return material to return to carry out leaching reaction. After the solution is saturated, copper and silver can be recovered, and the leaching rate of Cu and Ag can reach more than 96%.
When the Cu content in the copper dross is more than 20wt%, leaching the roasting slag by using inorganic acid as a leaching agent, and then carrying out solid-liquid separation to obtain a fifth leaching slag rich in silver and a fifth leaching liquid rich in copper;
the inorganic acid is sulfuric acid and/or hydrochloric acid, so that high-efficiency enrichment of copper can be realized, and meanwhile, effective separation of Cu and Ag in roasting slag can be realized. The mineral acid is preferably sulfuric acid, and the main component of the fifth leaching residue is Ag 2 SO 4 The slag rate is less than 1%, the Ag content is 40%, and the Cu leaching rate can be more than 99%.
Further, in S3, when inorganic acid is used as a leaching agent to leach the roasting slag, H in a reaction system is controlled + And Cu in an initial molar ratio of 2-4:1, and a liquid-solid ratio of 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
Further, in S3, when the methylsulfonic acid is used as a leaching agent to leach the roasting slag, the total molar quantity of Cu and Ag and H in a reaction system are controlled + The ratio of the initial molar quantity of (2) to (5) is 1:1-5, and the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
Further, in S3, when the smoke dust is leached, the total molar quantity of Bi and Pb and CH in a reaction system are controlled 3 SO 3 The initial molar ratio of H is 1:3-6, and the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
Further, in S1, copper dross was crushed to-200 mesh >85%.
Further, the oxidizing roasting is performed in the rotary kiln, so that a higher roasting efficiency can be obtained.
Further, the oxidizing roasting is performed in an air atmosphere or oxygen-enriched air having an oxygen concentration of 23 to 26 vol%.
Further, in the oxidizing roasting process, the temperature can be obtained by heating in an electric furnace or adding coal or coke.
Further, the copper dross contains 26.8% of Cu, 42.1% of Pb, 500g/t of Ag, 7.6% of Sb, 3.2% of Bi, and usually, cu, pb, ag, sb, bi exists mainly in the form of simple substance.
In the present invention, the reaction that may occur during the oxidative calcination is as follows:
2Cu+O 2 =2CuO (1)
Pb+O 2 =2PbO (2)
4Bi+3O 2 =2Bi 2 O 3 (3)
4Sb+3O 2 =2Sb 2 O 3 (4)
4Ag+O 2 =2Ag 2 O (5)
when leaching is carried out with mineral acids containing sulfuric acid, the possible reactions are as follows:
CuO+H 2 SO 4 =CuSO 4 +H 2 O (6)
PbO+H 2 SO 4 =PbSO 4 ↓+H 2 O (7)
Bi 2 O 3 +3H 2 SO 4 =Bi 2 (SO 4 ) 3 ↓+3H 2 O (8)
Ag 2 O+H 2 SO 4 =Ag 2 SO 4 ↓+H 2 O (9)。
the reaction that may occur when leaching with methanesulfonic acid is as follows:
CuO+2CH 3 SO 3 H=(CH 3 SO 3 ) 2 Cu+H 2 O (10)
PbO+2CH 3 SO 3 H=(CH 3 SO 3 ) 2 Pb+H 2 O (11)
Bi 2 O 3 +6CH 3 SO 3 H=2(CH 3 SO 3 ) 3 Bi+3H 2 O (12)
Ag 2 O+2CH 3 SO 3 H=2(CH 3 SO 3 )Ag+H 2 O (13)
PbSO 4 +2CH 3 SO 3 H=(CH 3 SO 3 ) 2 Pb+H 2 SO 4 (14)
Bi 2 (SO 4 ) 3 +6CH 3 SO 3 H=2(CH 3 SO 3 ) 3 Bi+3H 2 SO 4 (15)
Ag 2 SO 4 +2CH 3 SO 3 H=2(CH 3 SO 3 )Ag+H 2 SO 4 (16)。
the invention adopts the process of copper dross oxidation roasting combined acid leaching to recover copper, lead and silver, and has the advantages of simple process, low cost, short period and high copper leaching rate.
In the invention, the smoke dust is leached and separated by methanesulfonic acid. After leaching, the main component of the leaching slag is Sb 2 O 3 The Sb accounts for more than 78 percent, the Bi is enriched in the obtained leaching solution, and the Bi leaching rate is more than 98 percent.
Compared with the prior art, the invention has the following beneficial effects:
(1) The method is simple and practical, has low cost, and can effectively separate and recover valuable metal elements in the copper dross.
(2) According to the method, the technological conditions of oxidizing roasting can be flexibly controlled according to the content of Pb, cu, bi, sb, ag and other elements in the copper dross, and the high-efficiency separation and the separate recovery of valuable metals in the copper dross can be realized through the subsequent specific wet recovery procedure.
(2) According to the invention, according to the chemical reaction characteristics of the inorganic acid and the methylsulfonic acid and the saturated solubility of each element, the inorganic acid and the organic acid are leached or the inorganic acid is leached or the organic acid is leached out from different roasting products, the recovery rate of Pb, cu, bi, sb, ag is high and can reach more than 98%, bi and Sb in smoke dust are thoroughly separated, and the subsequent recovery process of Bi and Sb is simplified.
Drawings
Fig. 1 is a flow chart of embodiment 1 of the present invention.
Fig. 2 is a flowchart of embodiment 2 of the present invention.
Fig. 3 is a flowchart of embodiment 3 of the present invention.
FIG. 4 is an XRD pattern of the roasting slag of example 1 of the present invention.
Fig. 5 is an XRD pattern of smoke of example 1 of the present invention.
Detailed Description
The invention will be described in detail below with reference to the drawings in connection with embodiments. Unless otherwise stated, the relevant percentages refer to mass percentages.
Example 1
In this example, the method for recovering valuable elements from copper dross is as follows:
copper dross (Cu 8.8%, pb72.1%, ag1200g/t, sb5.4%, bi 2.9%) is crushed, ball milled and ground to-200 mesh>85vol% of the copper dross is then oxidized and roasted in an electric furnace at 600 ℃ for 2.5 hours in an air atmosphere. Roasting slag and smoke dust are obtained after roasting. The main phase of the detected roasting slag is CuO, pbO, ag. The smoke and dust mainly comprises Bi 2 O 3 、Sb 2 O 3 The volatilization rates of Bi and Sb are respectively 95% and 97%.
Mixing the roasting slag with sulfuric acid solution, reacting, and controlling H in a reaction system + The initial molar ratio of Cu is 3:1, liquid-solid ratio is 4mL:1g and reacting for 3 hours at 50 ℃ to obtain leaching solution and leaching slag, wherein the Cu content of the leaching solution is 21.7g/L, and the Cu recovery rate is 98.5%.
Leaching residue and methylsulfonic acid (CH) 3 SO 3 H) Reaction, control CH in reaction system 3 SO 3 H. The initial molar ratio of Pb is 3:1, the liquid-solid ratio is 4mL:1g, the temperature is 60 ℃, the reaction is carried out for 3 hours, the obtained leaching solution contains 109.9g/L of Pb, the lead recovery rate is 99.2%, the leaching slag is Ag-rich slag, the silver content is 85%, and the recovery rate reaches 99.6% after being sent to an Ag system for recovery.
The smoke dust reacts with the methylsulfonic acid to control CH in the reaction system 3 SO 3 H. The initial molar ratio of Bi is 4:1, the liquid-solid ratio is 4mL:1g, the reaction is carried out for 3 hours at 60 ℃, and the obtained leaching solution contains 76.2g/L of Bi and the recovery rate of Bi is 98.7 percent. The leaching slag is antimony-rich oxide (containing 80% of Sb), and is sent to an antimony system for recycling, and the antimony recycling rate reaches 99.1%.
Example 2
In this example, the method for recovering valuable elements from copper dross is as follows:
copper dross(Cu26.8%, pb42.1%, ag500g/t, sb7.6%, bi3.2%) is crushed, ball milled and ground to-200 meshes>85vol% of the copper dross was then oxidized and roasted in an electric furnace at 1150℃for 2 hours in an oxygen-enriched air atmosphere (oxygen concentration: 23 vol%). Roasting slag and smoke dust are obtained after roasting. The main phase of the detected roasting slag is CuO and Ag 2 O. The main component of the smoke is PbO and Bi 2 O 3 、Sb 2 O 3 The volatilization rates of Bi and Sb are 95.6 percent and 97.5 percent respectively.
The roasting slag reacts with sulfuric acid, and H in a reaction system is controlled + The initial molar ratio of Cu is 2:1, reacting for 2 hours at 60 ℃ with a liquid-solid ratio of 5mL to 1g to obtain leaching solution and leaching slag, wherein the Cu content of the leaching solution is 52.9g/L, the copper recovery rate is 98.7%, the slag rate is 0.5%, the leaching slag is silver-rich slag and is sent to an Ag system for recovery, and the Ag recovery rate is 99.5%.
The smoke dust reacts with the methylsulfonic acid to control CH in the reaction system 3 SO 3 H. The initial molar ratio of (Bi+Pb) was 3:1, the liquid-solid ratio is 3mL:1g, and the obtained leaching solution contains 18.6g/L Bi, 252g/L Pb, 98.2% bismuth recovery rate and 98.9% lead recovery rate, and can be used as a lead electrodeposition raw material after being reacted for 3 hours at 60 ℃. The leaching slag is antimony-rich oxide, contains 78% of Sb, and is sent to an antimony feeding system for recycling, and the antimony recovery rate reaches 99.2%.
Example 3
In this example, the method for recovering valuable elements from copper dross is as follows:
copper dross (Cu16.5%, pb49.1%, ag500g/t, sb7.6%, bi3.2%) is crushed, ball milled and ground to-200 mesh>85vol% and then taking 1kg of ground copper dross, oxidizing and roasting at 1000 ℃ in an electric furnace, and roasting for 2 hours in an air atmosphere with 23% oxygen enrichment. Roasting slag and smoke dust are obtained after roasting. The main phase of the detected roasting slag is CuO and Ag 2 O. The main component of the smoke is PbO and Bi 2 O 3 、Sb 2 O 3 The volatilization rates of Bi and Sb are 95.2 percent and 97.3 percent respectively.
Mixing roasting slag with methylsulfonic acid, and controlling CH in a reaction system 3 SO 3 H. The initial molar ratio of (cu+ag) is 3:1, the liquid-solid ratio is 5mL to 1g, and the reaction is carried out for 3 hours at 75 ℃, and the solid is liquidSeparating to obtain leaching solution and leaching residue, wherein the obtained leaching solution contains Cu90.8g/L, ag0.10 g/L, copper recovery rate of 98.8%, silver recovery rate of 99.1%, and Cu and Ag can be recovered from the leaching solution later.
Mixing smoke dust with methylsulfonic acid, and controlling CH in a reaction system 3 SO 3 H. The initial molar ratio of (Bi+Pb) was 3:1, controlling the liquid-solid ratio to be 4 mL/1 g, reacting for 2 hours at 75 ℃, and carrying out solid-liquid separation to obtain leaching liquid and leaching residues. The leaching solution contains 12.7g/L of Bi and 141.8g/L of Pb, the leaching solution recovers Bi and Pb, the recovery rate of bismuth is 98.9%, and the recovery rate of lead is 99.2%. The leaching slag is antimony-rich oxide, contains 79% of Sb, and is sent to an antimony system for recycling, and the antimony recovery rate is 99.2%.
The foregoing examples are set forth in order to provide a more thorough description of the present invention, and are not intended to limit the scope of the invention, since modifications of the invention in various equivalent forms will occur to those skilled in the art upon reading the present invention, and are within the scope of the invention as defined in the appended claims.
Claims (8)
1. A method for recovering valuable elements from copper dross, comprising the steps of: s1, crushing copper dross to be treated to obtain copper dross powder;
s2, oxidizing and roasting the copper dross powder obtained in the step S1 to obtain roasting slag and smoke dust;
s3, carrying out acid leaching on the roasting slag obtained in the step S2;
leaching the smoke dust obtained in the step S2 by taking methanesulfonic acid as a leaching agent, and carrying out solid-liquid separation to obtain a first leaching solution rich in bismuth and a first leaching residue rich in antimony; wherein, in the copper dross, the content of Ag is more than 1000g/t, the content of Pb is more than 70wt%, the content of Cu is 2-10wt%, in S2, the temperature of oxidizing roasting is 400-700 ℃, and the roasting time is 1-4h;
s3, performing solid-liquid separation after performing primary leaching on the roasting slag by using inorganic acid as a leaching agent to obtain second leaching slag and second leaching liquid rich in copper; then, performing secondary leaching on the second leaching slag by taking methanesulfonic acid as a leaching agent, and performing solid-liquid separation to obtain third leaching slag rich in silver and third leaching liquid rich in lead;
wherein the inorganic acid is sulfuric acid.
2. The method according to claim 1, wherein the initial molar ratio of H to Cu in the reaction system is controlled to be 2-4 during the first leaching: 1, the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h; during the second leaching, CH in the reaction system is controlled 3 SO 3 The initial molar ratio of H to Pb is 2-4:1, the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
3. The method according to claim 1 or 2, wherein in S3, the CH in the reaction system is controlled when leaching the soot 3 SO 3 The initial molar ratio of H to Bi is 3-5:1, the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
4. A method for recovering valuable elements from copper dross, comprising the steps of: s1, crushing copper dross to be treated to obtain copper dross powder;
s2, oxidizing and roasting the copper dross powder obtained in the step S1 to obtain roasting slag and smoke dust;
s3, carrying out acid leaching on the roasting slag obtained in the step S2;
leaching the smoke dust obtained in the step S2 by taking methanesulfonic acid as a leaching agent, and carrying out solid-liquid separation to obtain a first leaching solution rich in bismuth and a first leaching residue rich in antimony;
wherein, in the copper scum, the content of Ag is 500-1000g/t, the content of Pb is 30-50wt%, the content of Cu is 10-40wt%, and in S2, the temperature of oxidizing roasting is 900-1200 ℃ and the roasting time is 1-4h; s3, when the Cu content in the copper dross is less than or equal to 20wt%, leaching the roasting slag by using methanesulfonic acid as a leaching agent, and performing solid-liquid separation to obtain a fourth leaching solution rich in copper and silver; when the Cu content in the copper dross is more than 20wt%, leaching the roasting slag by using inorganic acid as a leaching agent, and then carrying out solid-liquid separation to obtain a fifth leaching slag rich in silver and a fifth leaching liquid rich in copper;
wherein the inorganic acid is sulfuric acid and/or hydrochloric acid.
5. The method according to claim 4, wherein in S3, when the roasting slag is leached by using inorganic acid as a leaching agent, the initial molar ratio of H to Cu in the reaction system is controlled to be 2-4:1, and the liquid-solid ratio is controlled to be 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
6. The method according to claim 4, wherein in S3, when the roasting slag is leached by using methanesulfonic acid as a leaching agent, the ratio of the total molar amount of Cu and Ag to the initial molar amount of H in the reaction system is controlled to be 1:1-5, and the liquid-solid ratio is controlled to be 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
7. The method according to any one of claims 4 to 6, wherein in S3, the total molar amount of Bi and Pb and CH in the reaction system are controlled when leaching the soot 3 SO 3 The initial molar ratio of H is 1:3-6, and the liquid-solid ratio is 2-6mL:1g, leaching temperature is 50-95 ℃ and leaching time is 1-4h.
8. The method according to claim 1, characterized in that in S1 the copper dross is crushed to-200 mesh >85%.
Priority Applications (1)
Application Number | Priority Date | Filing Date | Title |
---|---|---|---|
CN202210085349.4A CN114480862B (en) | 2022-01-25 | 2022-01-25 | Method for recovering valuable elements from copper dross |
Applications Claiming Priority (1)
Application Number | Priority Date | Filing Date | Title |
---|---|---|---|
CN202210085349.4A CN114480862B (en) | 2022-01-25 | 2022-01-25 | Method for recovering valuable elements from copper dross |
Publications (2)
Publication Number | Publication Date |
---|---|
CN114480862A CN114480862A (en) | 2022-05-13 |
CN114480862B true CN114480862B (en) | 2024-01-30 |
Family
ID=81475500
Family Applications (1)
Application Number | Title | Priority Date | Filing Date |
---|---|---|---|
CN202210085349.4A Active CN114480862B (en) | 2022-01-25 | 2022-01-25 | Method for recovering valuable elements from copper dross |
Country Status (1)
Country | Link |
---|---|
CN (1) | CN114480862B (en) |
Citations (7)
Publication number | Priority date | Publication date | Assignee | Title |
---|---|---|---|---|
GB1525175A (en) * | 1977-02-02 | 1978-09-20 | Isc Smelting | Extraction of metals |
CN104846202A (en) * | 2015-05-15 | 2015-08-19 | 西北矿冶研究院 | Method for producing electrodeposited copper from multi-metal copper slag |
CN105779789A (en) * | 2016-03-11 | 2016-07-20 | 中南大学 | Wet method for separating bismuth from antimony |
CN106676270A (en) * | 2017-01-05 | 2017-05-17 | 中南大学 | All-wet method for extracting lead from lead plaster and lead sulfide concentrate |
CN110396608A (en) * | 2019-08-02 | 2019-11-01 | 中南大学 | A kind of methane sulfonic acid system bismuth sulfide concentrate oxygen leaching method |
CN112813278A (en) * | 2021-01-15 | 2021-05-18 | 昆明冶金研究院有限公司 | Recovery processing method of copper dross |
CN112877543A (en) * | 2021-01-14 | 2021-06-01 | 华中科技大学 | Method for recovering lead from lead slag |
Family Cites Families (1)
Publication number | Priority date | Publication date | Assignee | Title |
---|---|---|---|---|
US9322104B2 (en) * | 2012-11-13 | 2016-04-26 | The University Of British Columbia | Recovering lead from a mixed oxidized material |
-
2022
- 2022-01-25 CN CN202210085349.4A patent/CN114480862B/en active Active
Patent Citations (7)
Publication number | Priority date | Publication date | Assignee | Title |
---|---|---|---|---|
GB1525175A (en) * | 1977-02-02 | 1978-09-20 | Isc Smelting | Extraction of metals |
CN104846202A (en) * | 2015-05-15 | 2015-08-19 | 西北矿冶研究院 | Method for producing electrodeposited copper from multi-metal copper slag |
CN105779789A (en) * | 2016-03-11 | 2016-07-20 | 中南大学 | Wet method for separating bismuth from antimony |
CN106676270A (en) * | 2017-01-05 | 2017-05-17 | 中南大学 | All-wet method for extracting lead from lead plaster and lead sulfide concentrate |
CN110396608A (en) * | 2019-08-02 | 2019-11-01 | 中南大学 | A kind of methane sulfonic acid system bismuth sulfide concentrate oxygen leaching method |
CN112877543A (en) * | 2021-01-14 | 2021-06-01 | 华中科技大学 | Method for recovering lead from lead slag |
CN112813278A (en) * | 2021-01-15 | 2021-05-18 | 昆明冶金研究院有限公司 | Recovery processing method of copper dross |
Non-Patent Citations (1)
Title |
---|
氧化焙烧―酸浸法从铜浮渣中提取铜;吴跃东;范兴祥;董海刚;赵家春;李博捷;行卫东;付光强;刘杨;秦庆彦;景小宇;;有色金属(冶炼部分)(08);第5-6页 * |
Also Published As
Publication number | Publication date |
---|---|
CN114480862A (en) | 2022-05-13 |
Similar Documents
Publication | Publication Date | Title |
---|---|---|
US10106868B2 (en) | Process for extracting noble metals from anode slime | |
CN106011488B (en) | A kind of method of high arsenic-and copper-bearing cigarette ash comprehensively recovering valuable metal | |
CN109338111B (en) | Method for recovering valuable metal from zinc-tin-containing material | |
CN102586627B (en) | Method for recovering bismuth from bismuth slag | |
JP4461283B2 (en) | Recovery of non-ferrous metals from zinc residues | |
CN112063854B (en) | Method for comprehensively recovering bismuth, silver and copper metals by taking precious lead as raw material | |
JPH0643619B2 (en) | Method for leaching sulfide containing zinc and iron | |
JP2008081799A (en) | Method for recovering lead | |
CN106011489A (en) | Iron vitriol slag treatment method | |
EP0557312B1 (en) | Direct sulphidization fuming of zinc | |
EP4061972B1 (en) | Improved copper smelting process | |
CN114480862B (en) | Method for recovering valuable elements from copper dross | |
US4613365A (en) | Method for recovering precious metals | |
EP2417274B1 (en) | Method of refining copper bullion comprising antimony and/or arsenic | |
US4333762A (en) | Low temperature, non-SO2 polluting, kettle process for the separation of antimony values from material containing sulfo-antimony compounds of copper | |
CN103589880B (en) | Method for preparing crude bismuth by reducing bismuth oxychloride | |
Mirzanova et al. | Technology for processing industrial waste containing non-ferrous metals | |
CN112143908A (en) | Smelting process for treating complex gold ore | |
CN109943727A (en) | The extracting method of valuable metal in a kind of copper anode mud | |
CN114990337B (en) | Method for recovering tin in silver separating slag of copper anode slime by combining pyrogenic process and wet process | |
CN1167833A (en) | Integrated recovery of Pb and Sn as valuable metal from chloride slag | |
AU646510C (en) | Direct sulphidization fuming of zinc | |
CN116814964A (en) | Method for recycling noble metal by removing copper through fire refining of high-copper lead slag | |
CN115747481A (en) | Method for preparing crude silver from lead anode mud | |
CN115125395A (en) | Method for separating and extracting tin from silver separating residues of copper anode slime by microwave roasting and wet method |
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 |