An example of the process scheme of the hydrochloric acid-FeCl3 leaching process of Yunxi Sanye

Cloud tin Metallurgical process flow shown below, and the operation indicators are as follows:

Tuyun tin company solder anode anode mud acid wet method comprehensive recycling process

Leaching of hydrochloric acid-FeCl3:

(1) Wet grinding and screening: The anode mud is slurried and ground in a ball mill . The pulp concentration is up to 50% and ground to a particle size of 80 mesh.

(2) Leaching: carried out in a stirring leaching tank. The tank is made of ¢8m×1.7m steel shell, lined with rubber and ceramic tile, and steam is directly heated. Leachate composition (g / L): 170 ~ 180HC1, 20 ~ 40FeC13; liquid to solid ratio of 4:1; temperature 85 ~ 90 ° C; stirring time 4h; stop stirring, add a small amount of coagulant, clarified cooling for 4h.

(3) Treatment of leaching products: the supernatant containing tin, antimony and bismuth is pumped to the high level tank; the lead and silver precipitates are slurried, washed, filtered and sent to the lead removal process, the composition of which is: 4.5% to 5% Ag, 29% to 41% Pb.

Hot water leaching:

(1) Hot water leaching (preliminary lead removal): liquid to solid ratio of 30:1, pH>3, steam is directly heated to 95 ° C, and boiled for 2 h.

(2) The supernatant containing PbCl2 was extracted while hot, and the slag was washed twice in the same tank.

(3) Composition of boiled slag: silver is increased to 15% to 18%, lead is reduced to 5% to 7%, and others are 3% to 5% Sn, 0.5% As, 2% Sb, and 0.5% Bi. Gold and silver into the slag of 96% to 98%.

Replacement - flotation:

(1) After the residue boiled with iron powder will be replaced with a sponge silver AgCl, in order to select the floating silver enamel reaction pot.

(2) Flotation separation of lead silver: silver or gold is collected by butylamine black medicine or amyl yellow medicine, and silver concentrate of 35% to 45% Ag is produced. Control tailings less than 0.25% silver, silver recovery rates of 96% to 97%. In hexametaphosphate, sodium carboxy methyl cellulose phosphate or lead suppressed, so that the lead tailings, concentrate output lead chloride containing 45% ~ 50% Pb and Pb recovery rates higher than 97%.

Recycling silver:

(1) Silver concentrate composition (%): Ag35 to 45, Au35 to 45 g/t, Pb8 to 12, Sn1 to 2, As0.5 to 1, Sbl to 2, Bi0.5 to 1, CI-3 to 4. Among them, Cl- is mainly brought in by PbCl2.

(2) Iron powder displacement dechlorination: carried out in a stirred leaching tank. First, the silver concentrate is slurried, and then the pH is adjusted to 1 to 2 with sulfuric acid, the temperature is higher than 90 ° C, and the iron powder is added to replace the C1- in the PbC12 into the FeC12 into the solution.

(3) Silver nitrate leaching: The silver concentrate after dechlorination is added to the 4~4.5mo1/LHNO3 solution, stirred, and the silver becomes AgNO3 dissolved in water. The resulting Pb(NO3)2 reacts with the residual sulfate in the concentrate to form PbSO4 into the leach residue. The slag still contains 3% to 6% of silver and 250 to 320 g/t of gold, which is a raw material for gold extraction. The silver leaching rate is 97% to 98%. The NO2 produced in the operation is washed by a venturi tube, and the resulting eluent is returned to the leaching.

(4) Silver sulphate: High-purity AgCl is precipitated by adding hydrochloric acid to a silver nitrate solution. The rate of sinking silver is higher than 99%. The mother liquor is discharged after treatment.

(5) Hydrazine hydrate reduction silver: Hydrazine hydrate (N2H4·H2O) is a strong reducing agent, which can reduce AgCl to silver powder in alkaline bismuth. The reaction is:

4AgCI+N2H4+4NH4OH=4Ag↓+N2↓+4NH4Cl+4H2O

This work is carried out in a stirred leaching tank. Add a small amount of water to the tank, heat it directly to 50-60 ° C, and add 20% ammonia to a liquid-solid ratio of 3:1. Add a small amount of hydrazine hydrate to adjust the solution to pH = 9 ~ 10; stir again, slowly (several times) add a predetermined amount of AgCl. The supernatant is taken from the tank and added to the hydrazine hydrate reaction until no precipitation is reached. This reaction is fast and the reduction rate is as high as 99%. The mother liquor contains Ag below 0.00 lg/L. Lkg silver powder consumes 20% ammonia water 1~1.5kg, 40% hydrated 肼 0.45kg.

A white spongy silver powder was produced, and the composition (%) was: 99.983 Ag, 0.002 Pb, 0.0006 Cu, 0.004 Sb, 0.0025 Bi, 0.0075 Fe.

(6) Sponge silver casting: After the sponge silver is dried, it is filled with No. 120 graphite crucible and placed in a m0.5m×0. 8m diesel crucible furnace or a medium frequency induction electric furnace for melting. The temperature is raised to 1200 ° C and naturally oxidized and refined. When the impurities such as bismuth and bismuth in the silver powder are high, oxygen blowing may be appropriately performed to ensure that the Ag content of the fine silver is higher than 99.95%. The yield of silver refining is higher than 99%. The direct yield from silver concentrate to fine silver is 95%.

Recovery gold:

(1) The slag after leaching of silver nitrate is enriched with gold, and the components (%) are: Ag3-6, Au250-320g/t, Pb3-7, Sn5-6, Bil-2, Sb6-8, As2-3, Sel. The method of recovering gold from the slag may be a thiourea leaching-iron replacement method or an aqueous solution chlorination-oxalic acid reduction method. Both are carried out in a stirred tank.

(2) Thiourea leaching-iron replacement method: The solution contains thiourea (CS(NH2)2) 30 g/L, the liquid-solid ratio is 10:1, and the pH is adjusted to 1.5 with sulfuric acid. After immersing for 3 h at 40 ° C, the silver leaching rate was 80% to 85%, and the gold leaching rate was 95% to 96%. Replacement with iron powder, the replacement slag contains gold up to 3%.

(3) Aqueous solution chlorination-oxalic acid reduction method: the slurry is slurried, and then chlorine chloride is chlorinated, or gold is leached with sodium hypochlorite (NaClO3 + NaCl) to make gold into AuC13 or AuOCI into the solution. The gold leaching rate is over 98%. The control slag contains Au below 2 g/t and Ag below 2%. The solution is reduced with gold oxalate to control the gold powder, and the control powder contains Au higher than 99.9%.

Recycling tin:

(1) The supernatant component (g/L) of the anode mud leached with hydrochloric acid and ferric chloride is: 20 to 25 Sn, 0.1 to 0.15 Ag, 2 to 2.5 Pb, 10 to 13 As, 18 to 20 Sb, and 8 to 12 Bi. , 3 ~ 5Cu, 1.5 ~ 2.2H +. After removing the As, Sb, Bi, and Cu by the iron filing replacement method, the tin concentrate is produced by the lime neutralization method, or the metal tin is produced by the electrowinning method.

(2) Discharge of As, Sb, Bi, and Cu by iron powder: The operation is carried out in a sealing tank of ×1.8×1.7 m, and a good air suction device is required to keep the inside of the tank as a negative pressure. The solution was directly heated to 45 to 50 ° C with steam, stirred with compressed air, and the operation was completed within 4 hours. Replacement rate: arsenic is higher than 85%, strontium is higher than 90%, strontium is higher than 95%, and tin is lower than 3%. Most of the tin in the form of SnCl2 remains in the solution.

(3) Zhonghe method Shenxi: Neutralize the SnCl2 solution with lime milk to PH=4~4.5, which can produce tin concentrate containing more than 40% tin, and the tin recovery rate is higher than 90%. The concentrate is Sn(OH)2·xH2O, which is calcined by drying and then smelted into a metal.

(4) Electrowinning method: tin is made by using SnCl2 solution as electrolyte, iron plate as anode, fine tin sheet as cathode, and electrowinning in plastic electrolytic cell. The current density is controlled to be 80 to 100 A/m 2 and the cell voltage is 0.5 to 0.6 V. The cathode tin produced contains 75% to 85% Sn, 3% to 50% Pb, 1% to 3% Bi, and 0.2% to 0.4% Sb. The tin recovery rate is up to 94% and the current efficiency is 75% to 80%. The power consumption is 225 kW · h / t cathode tin.

Recovery of arsenic:

(1) The content (%) of the replacement slag when the tin is recovered is 11 to 17 As, 21 to 27 Sb, 12 to 25 Bi, 1 to 2 Sn, 0.2 to 0.3 Pb, 0.15 Ag, and 6 Fe. The slag should be stored in a thin layer to naturally oxidize and convert arsenic and antimony into oxides. The slag is treated regularly every year by first leaching the oxidized slag with sodium sulfide solution, converting arsenic and antimony into thioarsenate and thiodecanoate into the solution; and neutralizing arsenic and antimony with sulfuric acid. The sulfide is precipitated from the solution; then the sulphide sulfide is volatilized by dry distillation to leave the sulphide slag.

(2) Sodium sulfide leaching, arsenic arsenic: leaching, the solution is Na2S+NaOH. Its response is

(Sb, As) 2O3 X6Na2S+3H2O=2Na3 (Sb,As)S3+6NaOH

As2O3+6NaOH=2Na3AsO3+3H2O

The replacement slag was dried and ground to -80 mesh, and added to the stirring leaching tank in a 1:1 weight ratio with sodium sulfide. The liquid-solid ratio is 8:1, the steam is heated to 96-98 ° C, and stirred for 2 h. The leaching rate of strontium can reach 82-85%, and the arsenic leaching rate is >96%. Copper and copper are left in the leaching residue.

(3) Sulphuric acid neutralization and precipitation of arsenic arsenic: the reaction is

3Na3 (As,Sb)S3+3H2SO4=(As,Sb)2S3+3Na2SO4+3H2S

Neutralize at room temperature, control pH = 2 ~ 2.5. The cerium precipitation rate is 98%, and the arsenic precipitation rate is 95%. The arsenic slag component (%) is: 35-40Sb, 6-8As, and a suction device is required on the stirring leaching tank for neutralization to prevent the H2S gas from escaping. The extracted gas is passed through a venturi, circulated with a NaOH solution, and the Na2S is recovered for leaching.

(4) Recovery of arsenic and arsenic bismuth by dry distillation of arsenic sulphide sulphide: arsenic slag is denitrified by low temperature dry distillation and arsenic is recovered in the form of white arsenic. The reaction is:

2AsS (gas) + 7/2O2 (gas) → As2O3 + 2SO2

The dry distillation operation was carried out in an electric stainless steel rotary kiln, and the temperature was controlled at 330 °C. The volatilized AsS gas is oxidized to white arsenic (As2O3) through the condensation chamber and the baghouse, and the grade is 70% to 80%. After one rectification, the As2O3 content is higher than 98%, which is the finished product.

The remaining sulphurized slag residue, which contains more than 50% bismuth, is the raw material for the production of fine sputum.

Recycling copper:

(1) The Na2S leaching residue is As, Sb, Bi, and Cu slag, and contains (%): 18 to 21 Bi, 2 to 3 Cu, 0.7 to 1.0 As, 6 to 8 Sb, and 0.25 to 0.3 Ag. After the slag is naturally oxidized, the copper ruthenium is leached with hydrochloric acid to make it into a chloride solution, and then the copper ruthenium is replaced by iron powder to become a sponge metal, which is roughened by sulfur removal, and the copper slag is used as copper. raw material.

(2) Hydrochloric acid leaching of copper bismuth: Copper and bismuth in the slag after natural oxidation are easily dissolved by hydrochloric acid into BiC13 and CuCl2, and most of AgCl and arsenic bismuth remain in the leaching slag. When the content of strontium is high, HCI+FeC13 may be used for leaching, or a small amount of nitrate as a oxidant may be added to the hydrochloric acid leaching solution to increase the leaching rate of hydrazine. The leaching operation controls the liquid-solid ratio of 7:1, and the solution contains HC 165-70 g/L, and is immersed for 6 hours at room temperature. The leaching rate is higher than 95%. The leaching slag contains Ag 0.6% to 1.2%, and is returned to the anode mud for leaching to recover Ag, Au.

(3) Iron powder replacement bismuth copper: The leaching solution containing bismuth copper is heated to 50-70 ° C in a sealed tank with a suction device, and the sponge metal is replaced by iron powder. The composition (%) is: Bi>70 , Cu3 to 7, Sb2 to 3, Sn1 to 2, and As0.2 to 0.3.

(4) Recovery of copper and bismuth copper from sponge metal: firstly, the sponge metal is melted in the refining pot by alkali, melted at 700 °C, and then oxidized to remove arsenic arsenic. The temperature is lowered to 550 °C to remove arsenic slag and cooled to 320. °C plus sulfur to remove copper. The work was carried out under stirring, and sulfur was slowly and uniformly added. At the end, the slag was black powder and then dropped to 280 ° C. The copper sulfide slag contains 13% to 15% Cu, 8% to 9% S, and can be used as a raw material for producing copper sulfate.

The metal after decoppering is coarse, containing 97%~98% Bi, 0.5%~0.7%Sb, 0.1%-0.3% Cu, 0.05%-0.06% Ag, from arsenic bismuth copper slag to coarse output. The yield of hydrazine can reach 90% to 91%. The crude crucible is dezincified by adding zinc , and the lead chloride is removed by chlorine gas to produce a refined product containing Bi higher than 99.99%.

Recycling lead:

The PbCl2 tailings produced by flotation separation of silver and lead contain 40% to 50% of lead and Ag2000 to 2500g/t. The tailings are slurried in a stirring tank, adjusted to pH 2 by adding hydrochloric acid, heated to 95 ° C, and then iron powder is added and stirred for 2 hours to produce sponge lead, which contains Pb higher than 75%. The lead replacement rate can reach 97%.

Sponge lead powder has high impurity content and is easily oxidized when stored. Therefore, it must be melted into high-tin bismuth crude lead and sent to hydrofluoric acid for refining.