Fire method - routine treatment of anode mud process by electrolysis

Fire method - Electrolysis is a conventional method for treating copper anode mud for many years and is still widely used at home and abroad. Current production process generally consists of the steps of: (1) except copper and selenium; (2) smelting reduction noble output lead alloy; (3) expensive lead oxide refining gold silver alloy, i.e., anode mud; (4) Electrolytic Silver (5) After the silver anode mud is subjected to some treatment, gold electrolytic refining is performed. The process flow is shown in Figure 1.

Figure 1 Copper anode mud fire method - electrolytic conventional process

1. Sulfuration roasting

The reason why copper anode mud should be first burned to remove selenium is because when the anode is muddled by fire method, the existence of selenium will lead to the formation of a high-concentration selenium ice hole between the metal and the slag. The silver in the hole needs to extend the blowing oxidation time to extend the production cycle. On the other hand, selenium is dispersed in slag, ice caves and precious lead, which makes it difficult to recover selenium.

The main purpose of the copper anode roasting roasting is to oxidize selenium into SeO ₂ to volatilize it, into H ₂ SeO ₃ in the aqueous solution of the absorption tower, and then reduce it by SO ₂ in the furnace gas to form elemental selenium powder; For the soluble CuSO ₄ , the sulphated roasting slag is subjected to water leaching (or dilute sulfuric acid) to remove copper. The copper removal slag enters the gold and silver smelting system, the copper immersion liquid replaces the silver with a copper plate, the crude silver powder is sent to the gold and silver system, and the copper sulfate is sent to the copper electrolysis workshop to recover the copper.

(1) Roasting equipment - rotary kiln and absorption tower

Rotary kiln daily treatment of anode mud (wet mud) 1.5t. The rotary kiln is made of 16mm boiler steel plate. Its size is Φ750mm×10800mm, the rotation speed is 65r/s, the inclination is not more than 2%, and the inner wall has no furnace lining. In order to prevent the furnace material from sticking to the wall, the kiln is equipped with Φ75mm round steel with dent Cage, flip the anode mud. A refractory brick furnace fire chamber topical, external heating method, i.e., the entire body is placed in the combustion chamber kiln, heating the coal gas (or heavy). The kiln and the absorption tower maintain a negative pressure with a water ring vacuum pump. Lead iron lining for the absorber tower, absorption tower dimensions Φ (1000 ~ 1200) mm × (600 ~ 800) mm. Generally, one tower is Φ1200mm×800mm, and the second and third towers are Φ1000mm×600mm.

(2) Processing

The copper anode mud (20% to 40% water) is sent to the stainless steel mixing tank, and the theoretical amount of chemical reaction required by Cu, Ag, Se, Te and sulfuric acid is about 130% to 140%, and concentrated sulfuric acid is added. The mixture was stirred into a paste and uniformly fed into a rotary kiln by a feeder for sulfation roasting. The rotary kiln is heated indirectly by gas or heavy oil, and the temperature gradually rises from the feed end to the discharge end. The feed end temperature is 220-300 ° C, mainly the drying zone of the charge; the middle part is 450-550 ° C, mainly the sulfuric acid reaction zone; the discharge end is 600-680 ° C, the sulfation reaction is complete, and the SeO _{2 is} volatilized. The negative pressure is maintained in the kiln, the feeding end is 300-500Pa, the material is in the kiln (staying) for about 3h, and the volatilization rate of selenium can reach 93%-97%. The kiln residue (de-selenium slag) flows into the storage hopper and is released periodically. The slag contains 0.1% to 0.3% of selenium. The gas containing SeO ₂ and SO ₂ enters the absorption tower through the gas pipe at the feed end. The absorption tower is divided into two groups, three groups in each group are connected in series, and the two groups are used interchangeably. The tower is filled with water, and SeO ₂ in the furnace gas is dissolved in water to form H ₂ SeO ₃ , and is reduced to a powdery element selenium by SO ₂ , and dried by water to obtain about 95% of crude selenium. The absorption rate of the first column is about 85%, the second column is about 7% to 10%, and the third column is about 2% to 6%. After the column liquid and the washing liquid are replaced with iron, the selenium is less than 0.05 g/L and discarded. Selenium-containing replacement slag is returned to the kiln for treatment The kiln residue after roasting in the rotary kiln is leached with water or dilute sulfuric acid to remove copper. Leaching the solid-liquid ratio of 1: 2 ~ 3, the temperature above 90 ℃, mechanical stirring 2 ~ 3h, CuSO ₄ and AgSO ₄ tellurium sulfate and partially dissolved in water, filtered off by washing copper slag, smelting should send payment system. The solution was transferred to a displacement tank, heated to 90 ° C, and replaced with Ag, Te (sulfate, barium sulfate) with a copper plate until the solution was added with hydrochloric acid without a white precipitate. The precipitate is washed and filtered, the crude silver powder is sent to the gold and silver smelting system, and the copper sulfate solution is pumped to the copper electrolysis workshop to recover copper. The solution contains 30-60 g/L of copper, and the leaching slag contains 1% to 3% of copper. The volatile selenium was calcined by sulfation, and the leaching components after leaching the copper were listed in Table 1.

Table 1 Composition of slag after roasting and leaching of copper (%)

Leaching and displacement were carried out in stainless steel tanks. The leaching tank is Φ1200mm×600mm and mechanically stirred. The displacement tank is Φ1500mm×1600mm.

Calcination in addition to selenium is usually carried out by oxidizing roasting and soda sintering roasting. Oxidation roasting has been proven to be related to dust collection equipment, and the metal copper powder contained in the furnace gas (from the anode mud), the unburned soot and SO ₂ react with selenite, and a series of vice reactions occur. The reaction reduces the selenous acid to metal selenium or forms an insoluble selenide precipitate, which reduces the recovery of selenium, and the roasting soot often leads to the loss of precious metals. Therefore, oxidative baking has not been used for a long time. The recovery rate of selenium in soda sintering roasting is above 90%. However, most of the strontium strontium is formed by sodium citrate. When leaching with hot water, strontium will enter the solution together with selenium, which makes it difficult to separate strontium selenium, and it is difficult to obtain high-purity selenium. Therefore, the soda sintering roasting method is not suitable for treating strontium-containing strontium. Anode mud.

Second, reduction smelting yields precious lead alloy

The smelting of copper anode leaching copper leaching slag has used a reverberatory furnace or a flat furnace in the past. At present, converters or electric furnaces are widely used at home and abroad. The leaching residue is added with a reducing agent and a flux, and is reduced and smelted to produce an alloy of precious metal and lead containing gold and silver in a total amount of 30% to 40% (commonly known as noble lead). Therefore, the metallurgical furnace of the smelting operation is commonly known as the precious lead furnace.

(1) Equipment - cylindrical horizontal converter

The reduction smelting is carried out in a converter. The size of the furnace is shown in Table 2. For example, the converter of the second plant, the outer casing is welded with a 16 mm thick boiler steel plate, the size is 562400 mm × 4200 mm, the hearth area is 5.5 m ² , and the outlet is 620 mm × 520 mm. Bed capacity (treatment of anode mud) 1 ~ 1.2t / (m ² · d), mechanical transmission, the number of revolutions 12r / min. Since the muffle furnace shell lined with two layers of 10mm asbestos board, the whole radial upright furnace masonry brick layer of magnesium aluminate, magnesia powder with a bottom, a mixture of refractory clay elevate coke powder 400mm, the life of the furnace 200 heats above, the use of the furnace It should be baked and washed before. Newly built converters or repairs or discontinued production should be carried out in an oven to gradually increase the temperature of the furnace to protect the masonry in the furnace and extend the age of the furnace. To wash the furnace, add waste lead or lead oxide soot to the furnace (adding coke dust, reducing agent and flux such as sodium carbonate and fluorite to the soot washing furnace), so that the bricks in the furnace are filled with lead to improve gold and silver. Direct yield. After the furnace is finished, lead is released from the ingot for reuse.

Table 2 Example of main size of converter

(2) Smelting of leaching slag after selenium removal and copper removal

Smelting de-selenium and copper leaching slag is to add leaching slag to the furnace after washing, and to produce precious lead ingot by reduction smelting. For example, the composition of copper and selenium leaching slag in a plant is: H ₂ O30%, Au1%~1.5%, Ag10%~15%, Pb15%~20%, Cu<5%, Se<0.3%, Te0.4 %, SiO ₂ <5%. When smelting, 8% to 15% sodium carbonate, 3% to 5% fluorite powder, 6% to 10% coke breeze (or pulverized coal), and 2% to 4% iron filings are added. The amount of soda added may also be added in an amount of 1.8 times or a slight excess of the SiO ₂ content. In the smelting process, if the furnace junction is too thick or there is too much slag, the amount of soda is appropriately increased (if the slag is too much, it is reduced). Since the precious lead smelting is carried out in a micro-reducing atmosphere, the amount of reducing agent (crushing or pulverized coal) should be added according to the needs of copper, nickel and part of lead contained in the reduced leaching slag (actual production is based on Produce actual experience ingredients), do not make it excessive. If the excess is too large, a large amount of impurities will be reduced together into the precious lead, and the content of gold and silver in the precious lead will be lowered.

The copper anode slime is leached with slag after de-separation of copper, with lime, soda, fluorite, iron scrap as flux, pulverized coal or coke powder as reducing agent, uniformly mixed, and then sent into the converter through a belt conveyor. The furnace is kept under negative pressure (30 ~ 100Pa). Heavy oil is used as fuel, heavy oil is preheated to above 60 °C, and air with a pressure of 16 kPa or more is sent to the furnace for atomization combustion. The melting period temperature is maintained at 1200 to 1300 ° C, and the oxidation period temperature is maintained at 700 to 900 ° C.

After the charge enters the furnace, the temperature is gradually increased to remove moisture, and the oxides (As, Sb, Pb, etc.) are successively volatilized to enter the furnace gas. The charge begins to melt. And slagging reaction occurs:

Na ₂ COa=Na ₂ O+CO ₂

Na ₂ O+As ₂ O ₅ =Na ₂ O·As ₂ O ₅

Na ₂ O+Sb ₂ O ₅ =Na ₂ O·Sb ₂ O ₅

Na ₂ O+SiO ₂ =Na ₂ O·SiO ₂

PbO+SiO ₂ =PbO·SiO ₂

CaO+SiO ₂ =CaO·SiO ₂

At the same time, a reduction reaction also occurs:

2PbO+C=2Pb+CO ₂

PbO+Fe=Pb+FeO

PbSO ₄ +4Fe=Fe ₃ O ₄ +FeS+Pb

PbS+Fe=Pb+FeS

Ag ₂ S+Fe=2Ag+FeS

The gold and silver in the anode mud are trapped by the reduced metal lead melt to form noble lead, and the reaction can be expressed by the following formula:

Pb+Ag+Au=Pb(Ag+Au)

The precious lead melt and the slag do not dissolve each other, and the density difference is large, so the slag floats on the surface of the molten pool, and the precious lead sinks in the lower layer of the molten pool. In order to improve the grade of gold and silver in precious lead, the slag is discharged, and air is continuously blown into the noble lead melt to oxidize impurities such as As, Sb, Cu, Bi, etc., and As, Sb form a low-oxide oxide. Volatilized into the furnace gas (4As + 3O ₂ = 2As ₂ O ₃ ↑), (4Sb + 3O ₂ = 2Sb ₂ O ₃ ↑). If further oxidized to form a high-valent oxide (2Sb ₂ O ₃ + 2O ₂ → 2Sb ₂ O ₅ ), it can be slag with the basic oxide (Na ₂ O+2Sb ₂ O ₅ =Na ₂ O·Sb ₂ O ₅ ).

The whole furnace operation time is 18 to 24 hours. The lead production rate is 30% to 40%, and the composition is: Au 0.2% to 4.0%, Ag 25% to 60%, Bi 10% to 25%, Te 0.2% to 2.0%, Pb 15% to 30%, As3. %~10%, Sb5%~15%, Cu1%~3%. The output rate of dilute slag is 25%~35%, including Au<0.01%, Ag<0.2%, Pb15%~45%, sent to lead smelting system to recover Pb, or sent to blast furnace for enrichment and then into precious lead furnace to smelt copper and silver. alloy. The slag and the oxidized slag (late slag) contain Au and Ag, which are returned to the reduction smelting. The flue gas is vented after being dusted by the wet method, and the obtained soot is used as a raw material for extracting As and Sb.

(3) Electric furnace smelting of anode mud leaching slag from Hitachi Smelter

In order to improve the direct yield of gold and silver, the Japanese mining company Hitachi smelter reduced the intermediate product, shortened the smelting working hours and the backlog of liquidity, switched to the electric furnace smelting anode mud, copper leaching slag and using new electric furnace ingredients to make the electric furnace to silver. The intermediate products that need to be returned to the furnace during the smelting process are reduced to three kinds (electric furnace copper, lead oxide, lead oxide, copper oxide, lead oxide, silver slag, sodium nitrate sodium slag) to three types (lead oxide silver slag) Etc), and greatly reduce the gold and silver content of each intermediate product. According to the statistics after the improvement of the ingredients, the gold and silver grades of the charge and the quantity, grade and recovery rate of the products are greatly improved, and the recovery rate of gold and silver is over 99.3%.

Third, the oxidation of precious lead

The precious lead containing gold and silver obtained by reduction smelting is generally in the range of 35% to 60%, and the remainder is impurities such as Pb, Cu, As, Sb, etc., and oxidizing and refining is carried out under the condition of a converter temperature of 900 to 1200 ° C, and air is added and a flux is added. An oxidizing agent or the like oxidizes most of the impurities into oxides insoluble in gold and silver, enters the soot and forms slag, and obtains an anode plate containing more than 90% of gold and silver, which is suitable for silver electrolysis.

In the process of oxidative refining of precious lead, the oxidation order of various metals in precious lead is: Sb, As, Pb, Bi, Cu, Te, Se, Ag. Precious lead generally contains more lead and is also prone to oxidation. Therefore, in oxidative refining, PbO is mainly used as an oxygen transfer agent to oxidize arsenic and antimony (2Pb+O ₂ =2PbO, 2Sb+3PbO=Sb ₂ O ₃ +3Pb, 2As+3PbO= As ₂ O ₃ +3Pb), these arsenic, antimony low-oxides and part of PbO are easily volatilized into the flue gas, and the soot obtained after dust collection in the bag is returned to the smelting treatment. As ₂ O ₃ and Sb ₂ O ₃ can be further oxidized to high-valent oxides (Sb ₂ O ₅ , As ₂ O ₅ ), and slag with basic oxides (PbO, Na ₂ O, etc.), or directly formed into sub- Lead arsenic (or bismuth) (3PbO+Sb ₂ O ₅ =3PbO·Sb ₂ O ₅ , 2As+ 6PbO=3PbO·As ₂ O ₃ +3Pb, 2Sb+6PbO=3PbO·Sb ₂ O ₃ +3Pb), lead arsenite Lead in the presence of excess air can also form lead arsenate (3PbO·As ₂ O ₃ +O ₂ =3PbO·As ₂ O ₅ ).

Since the decompression pressure of As ₂ O ₅ is lower than that of Sb ₂ O ₅ , most of them enter slag in the form of arsenate, and most of them are volatilized into the furnace gas. When the arsenic bismuth oxidation is almost completed (ie, no white smoke is emitted), the surface blowing is changed to oxidative refining, and the lead can be completely oxidized and removed.

A metal which is hard to be oxidized such as Cu, Bi, Te, or Se is difficult to be oxidized by PbO. However, when As, Sb, and Pb are oxidized and removed, oxidative refining is continued, and cerium is oxidized (4Bi+3O ₂ = 2Bi ₂ O ₃ ) to form a slag containing some impurities such as copper, silver, arsenic, antimony, etc., and precipitated and smelted. In order to reduce the amount of silver, it is used as a raw material for the recovery of ruthenium.

When the alloy in the furnace reaches Au+Ag=80% or more, add 5% of Na ₂ CO ₃ and 1% to 3% of NaNO ₃ with a precious lead amount, and thoroughly agitate the copper, bismuth and selenium by artificial stirring (2NaNO _{3 ).} =Na ₂ O+2NO ₂ +[O], 2Cu+[O]=Cu ₂ O, Me ₂ Te+8NaNO ₃ =2MeO+8NO ₂ +TeO ₂ +4Na ₂ O, MeSe ₂ +8NaNO ₃ =2MeO+8NO ₂ +SeO ₂ ). TeO ₂ forms sodium citrate with the added Na ₂ CO ₃ , that is, a so-called soda slag (TeO ₂ + Na ₂ CO ₃ = Na ₂ TeO ₃ + CO ₂ ↑) is formed, which is used as a raw material for recovering ruthenium.

Finally, when Au+Ag reaches 95% or more, it is cast into an anode plate for silver electrolytic refining to obtain silver and further extract gold and platinum palladium .

The oxidizing refining is heated by heavy oil, and the operation per furnace is 45 to 72 hours. The converter is made of 12mm boiler steel plate, the outer casing size is Φ600mm×2240mm, the hearth area is 1.5m ² , the bed capacity (precious lead) is 1.6t/(m ² ·d), the bottom of the furnace is 100mm high, and the floor is set up in the radial direction. Magnesia bricks.

4. Recovery of platinum and palladium

After the gold electrolyte is used for a period of time (about 2 to 3 months), the impurities accumulate and cannot be used any more. The gold is reduced and precipitated with ferrous sulfate, oxalic acid or sulfur dioxide, and is cast into an anode plate to return to gold electrolysis. The solution contains Pt 5 ~ 15g / L, Pd 15 ~ 30g / L, sent to recover platinum, palladium. First, ammonium chloroplatinate is obtained by sinking platinum with NH ₄ CL, and calcined to obtain a platinum concentrate. The solution was replaced with a zinc sheet to obtain a palladium concentrate. After the platinum and palladium concentrates are refined and purified, pure sponge platinum and palladium are obtained.

V. Comprehensive recovery of other valuable components

In addition to precious metals, copper anode mud has some valuable components that must be recycled comprehensively. In general, the focus is on the recovery of antimony, antimony and selenium. In addition to the value of arsenic and deodorization, it is more important to eliminate their pollution to the environment, so it must be recycled.

(1) Recovery of strontium: The soda slag produced in the later stage of oxidative refining of precious lead fire contains 碲5% to 15%, and the remaining components are: Se0.2%~1.0%, Cu3%～10%, Pb3%～8 %, Bi 10% to 20%, and SiO ₂ 5% to 15%.

The soda slag wet grinding liquid solid ratio is 2 ~ 3, grinding at room temperature for 6h, to -80 mesh; water dilution 4 ~ 5 times leaching, heating to 80 ° C above clarified filtration; solution with Na ₂ S, CaCl ₂ purification after slag ball milling The solution is neutralized with dilute H ₂ SO ₄ to pH=5 (>80° C.), and clarified and filtered to contain more than 65% of TeO ₂ ; TeO _{2 is} washed with water and then dissolved in NaOH to prepare an electrolyte (NaOH 90-100 g/ L, Te 150 ~ 300g / L, Pb <0.1g / L, Se < 1.5g / L) electrolysis, the cathode ruthenium (content > 98%), and then ingot, yielding 99% ~ 99.9% of bismuth.

(2) Recovery of antimony: The composition of cerium oxide slag produced by oxidative refining of gold and silver is: Bi14%～35%, Pb15%～25%, Cu10%～20%, Sb10%～14%, As<0.005%, Ag1% to 3%, and SiO ₂ 15% to 25%.

The cerium oxide slag is reduced and smelted in the converter for 20 to 24 hours. The ingredients are generally: 3% to 4% soda, 20% to 30% iron sulfide, 3% to 4% fluorite, and <3% pulverized coal. 5 to 6t. The composition of the obtained niobium alloy is: Bi 50% to 65%, Pb 9% to 10%, Cu 9% to 25%, Sb 2% to 4%, Au + Ag 3% to 4%, and Fe amount. The straight yield is 80% to 90%. In the casting pot (Φ1000mm × 900mm) treatment, in order to remove various impurities, you can get No.1 and No.2.

(III) Recovery of arsenic: The smelting dust collected by wet dust collection, the general composition is: As10%～25%, Sb20%～35%, Pb8%～12%, Fe1%, Bi2%～4%, Te0.2 % to 0.4%, Au < 0.001%, Ag 0.2% to 0.4%, and H ₂ O 25% to 35%.

The smelting soot is mixed with soda roasting-leaching-filtration-concentration of the filtrate to obtain a sodium arsenate product. The sodium arsenate component is: As12% to 17.6%, Sb<0.1%, Fe<0.01%, Na ₂ CO ₃ 25% to 30%, Pb trace, and Bi trace amount. The crystallization efficiency is 88% to 90%.

(4) Recovery of strontium: After leaching arsenic by smelting dust, the composition is: As1.7%～3.0%, Sb40%～60%, Pb13%～20%, H ₂ O30%～40%, Na ₂ CO ₃ 5 % to 7%.

After immersing arsenic, the slag is mixed with pulverized coal and soda. After reduction and smelting, it is oxidized and volatilized, and then refined and refined to obtain fine sputum.