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

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

Figure   Yunnan Tin Company's solder anode anode mud acid wet process comprehensive recycling process

Leaching of hydrochloric acid- FeCl 3 :
( 1 ) Wet grinding and screening: The anode mud is slurried and ground in a ball mill . The pulp concentration is 50 % and ground to a particle size of 80 mesh.
( 2 ) Leaching: carried out in a stirred leach tank. Slot is ¢ 8m × 1.7m Steel shell, lined with rubber and ceramic tile, steam heated directly. Leachate component (g / L) as: 17 0 ~ 180HC1, 2 0 ~ 40FeC1 3; liquid-solid ratio of 4: 1; temperature 8 5 ~ 9 0 ℃; stirring time 4h; After stirring was stopped adding a small amount of flocculant to clarify the cooled 4h .
( 3 ) Treatment of leaching products: the supernatant containing tin, antimony and bismuth is pumped to a high level tank; the lead and silver precipitates are slurried, washed, filtered and sent to a 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 , boil for 2h .
( 2 ) The supernatant containing PbCl 2 was extracted while hot , and the slag was washed twice in the same tank.
( 3 ) The composition of the boiled slag: the silver is increased to 15 % to 18 %, the lead is reduced to 5 % to 7 %, and the others are 3 to 5 % Sn , 0.5 % As , 2 % Sb , 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 or amyl yellow, 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 (%) is: Ag3 5 ~ 45 , Au35 ~ 45g /t , Pb 8 ～ 12 , Sn l ～ 2 , As0. 5 ～ 1 , Sb l ～ 2 , Bi0. 5 ～ 1 , CI - 3 ～ 4 . Wherein Cl - is mainly brought in by PbCl 2 .
( 2 ) Desalination of iron powder displacement: carried out in a stirred leach tank. First, the silver concentrate is slurried, and then the pH is adjusted to 1 to 2 with sulfuric acid , and the temperature is higher than 9 0 °C , Replacement of iron powder was added in PbC1 2 C1 - FeC1 2 became into solution.
( 3 ) Nitric acid immersion silver: The dechlorinated silver concentrate is added to the 4 ～ 4.5mo1/LHNO 3 solution, stirred, and the silver becomes AgNO 3 dissolved in water. The resulting Pb(NO 3 ) 2 reacts with the residual sulfate in the concentrate to form PbSO 4 into the leach residue. Slag still containing silver 3% to 6% gold 25 0 ~ 320g / t, is the raw material of gold. The silver leaching rate is 97 % to 98 %. The NO 2 produced during the operation is washed by a venturi tube, and the resulting eluent is returned to leaching.
( 4 ) Hydrogenation of hydrochloric acid: Hydrochloric acid is added to the silver nitrate solution to precipitate high purity AgCl . The silver rate is higher than 99 %. The mother liquor is discharged after treatment.
( 5 ) Hydrazine hydrate reduction silver: Hydrazine hydrate (N 2 H 4 · H 2 O) is a strong reducing agent, which can reduce AgCl to silver powder in alkaline bismuth. The reaction is:

4AgC I + N 2 H 4 + 4NH 4 OH = 4Ag ↓+ N 2 ↓+ 4NH 4 Cl + 4H 2 O

This work is carried out in a stirred leaching tank. To add a small amount of water in a tank, directly to the steam heating 5 0 ~ 6 0 ℃, plus 20% aqueous ammonia to liquid to 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 0.00lg/L . Lkg silver powder consumption 20 % ammonia 1 ~ 1 .5 kg , 40 % hydrazine 0.45kg .
The 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, put the No. 120 graphite crucible into the crucible. 0.5 m × 0. 8m Melted in a diesel cooker or medium frequency induction furnace. Warm up to 1200 °C Natural oxidative refining. When the impurities such as strontium and barium 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 %. [next]
Recovery gold:
(1) residue after enrichment with silver nitrate immersion gold, the component (%) is: Ag 3 ~ 6, Au250 ~ 320g / t, Pb 3 ~ 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: solution containing thiourea ( CS ( NH 2 ) 2 ) 30g /L , liquid to solid ratio 10 : 1 , adjust the pH to 1.5 with sulfuric acid . in 4 0 °C The temperature was immersed for 3 h , the silver leaching rate was 80 % to 85 %, and the gold leaching rate was 95 % to 96 %. Replace with iron powder, the replacement slag contains gold up to 3 %.
( 3 ) Aqueous solution chlorination - oxalic acid reduction method: the slurry is slurried, then chlorinated with chlorine gas, or gold is leached with sodium hypochlorite (NaClO 3 + NaCl) to make gold into AuC1 3 or AuOCI into the solution. The gold leaching rate is over 98 %. Control slag contains Au lower than 2g /t , Ag is less than 2 %. The solution is reduced with gold oxalate to control the gold powder to contain 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 25Sn , 0.1 to 0.15Ag , 2 to 2.5Pb , 10 to 13As , 1 8 to 20Sb , 8 to 12Bi , 3 to 5Cu , 1. 5 to 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) replacement of iron removal As,, Sb, Bi, Cu : job ¢ 1. 8 × 1.7m In the sealing groove, a good air suction device is required to keep the tank under negative pressure. Solution was heated to steam directly 4 5 ~ 5 0 ℃, stirred with compressed air, the control operation is completed within 4h. Replacement rate: arsenic is above 85 %, strontium is above 90 %, strontium is above 95 %, and tin is below 3 %. Most of the tin in the form of SnCl 2 remains in the solution .
( 3 ) Neutralization method Shenxi: Neutralize the SnCl 2 solution with lime milk to PH= 4 ～ 4.5 , which can produce tin concentrate with tin content higher than 40 %, and the tin recovery rate is higher than 90 %. This concentrate is Sn (OH) 2 · xH 2 O , calcined by drying, and then smelted into a metal.
( 4 ) Electrowinning method: using SnCl 2 solution as electrolyte, iron plate as anode, fine tin sheet as cathode, and electrowinning in plastic electrolytic cell. Controlling the current density 80 ~ 100A / m 2, the cell voltage 0. 5 ~ 0.6V. 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 225kW · h/t cathode tin.
Recovery of arsenic:
(1) replacement of slag composition (%) of recovering tin: 11 ~ 17As, 21 ~ 27Sb , 12 ~ 25Bi, 1 ~ 2Sn, 0. 2 ~ 0.3Pb, 0.15Ag, 6Fe. 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 Na 2 S+NaOH . Its response is

(Sb , As ) 2 O 3 十 6Na 2 S + 3H 2 O = 2Na 3 (Sb , As ) S 3 + 6NaOH

As 2 O 3 + 6NaOH = 2Na 3 AsO 3 + 3H 2 O

The replacement slag was dried and ground to -80 mesh, and added to the stirring leaching tank in a weight ratio of 1 : 1 with sodium sulfide . Liquid to solid ratio 8 : 1 , steam heated to 9 6 ~ 98 °C , stir for 2h . The leaching rate can reach 8 2 ～ 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

3Na 3 (As , Sb)S 3 + 3H 2 SO 4 =( As , Sb ) 2 S 3 + 3Na 2 SO 4 + 3H 2 S

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 composition (%) is: 35 ~ 40Sb , 6 ~ 8As , and a suction device should be provided on the stirring leaching tank for neutralization to prevent the H 2 S gas from escaping. The extracted gas was passed through a venturi, circulated with a NaOH solution, and Na 2 S was recovered for leaching.
( 4 ) Recovery of arsenic and arsenic arsenic by dry distillation of arsenic sulphide sulphide: arsenic slag is arsenic removed by low temperature dry distillation and arsenic is recovered in the form of white arsenic. The reaction is:

â–³

(Sb , As)S (solid)

→

SbS (solid) + AsS (gas)

2 AsS (gas) + 7/2O 2 (gas) → As 2 O 3 +2SO 2

The dry distillation operation is carried out in an electric stainless steel rotary kiln to control the temperature. 33 0 °C . The volatilized AsS gas is oxidized to white arsenic (As 2 O 3 ) through the condensation chamber and the baghouse , and the grade is 70 % to 80 %. After one rectification, the As 2 O 3 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. [next]
Recycling copper:
(1) Na 2 S leaching residue is A s, S b, B i , Cu slag, containing (%): 1 8 ~ 21Bi , 2 ~ 3Cu, 0.7 ~ 1.0As, 6 ~ 8Sb, 0.25 ~ 0.3Ag. 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 BiC1 3 and CuCl 2 , and most of AgCl and arsenic bismuth remain in the leaching slag. When the content of strontium is high, HCI+FeC1 3 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 HC16 5 to 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) Copper bismuth substituted iron: bismuth-copper-containing leach solution in a sealed vessel with a ventilation device, with steam heated to 50 ~ 70 °C , Plus a metal sponge iron obtained displacement, which component (%) is: B i> 70, Cu 3 ~ 7, Sb2 ~ 3, Snl ~ 2, As0.2 ~ 0.3.
( 4 ) Sponge metal and sulfur removal of copper and beryllium copper: firstly, the sponge metal is melted in the refining pot, 70 0 °C After melting, oxidize and remove arsenic and arsenic, and cool down to 55 0 °C Remove the arsenic slag and cool down to 32 0 °C Add sulfur to remove copper. The work is carried out under agitation, the sulfur is added slowly and evenly, and the slag is black powder at the end, and then falls to 28 0 °C Residue. The copper sulfide slag contains 13 % to 15% Cu and 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 %. After adding zinc crude desilvering bismuth, lead and zinc chlorine-off after the output of the product-containing refined bismuth Bi is higher than 99.99%.
Recycling lead:
The PbCl 2 tailings produced by flotation separation of silver and lead contain 40 % to 50 % lead , Ag200 0 to 2500g/t . This immersion tank tailings slurry was stirred, hydrochloric acid was added to adjust to pH 2, heated to 9 5 °C Then add iron powder and stir for 2h to produce sponge lead with 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.

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