AU650471B2

AU650471B2 - Method of extracting valuable metals from leach residues

Info

Publication number: AU650471B2
Application number: AU82897/91A
Authority: AU
Inventors: Steven Paul Matthew; Roger Leo Player; Jorma Matti Ilmari Tuppurainen
Original assignee: Mount Isa Mines Ltd
Current assignee: Mount Isa Mines Ltd
Priority date: 1990-07-27
Filing date: 1991-07-26
Publication date: 1994-06-23
Anticipated expiration: 2011-07-26
Also published as: AU8289791A

Description

Title: METHOD OF EXTRACTING VALUABLE METALS FROM LEACH RESIDUES

TECHNICAL FIELD

The present invention relates to a method of extracting valuable metals from leach residues and in particular to a method of extracting valuable metals from leach residues of lead and silver.

The invention has been developed primarily for use with residues of lead and silver and will be described hereinafter with reference to those residues. However, it will be appreciated that the invention is not limited to this particular field of use.

BACKGROUND ART Lead and silver leach residues are an end product of conventional electrolytic zinc refining techniques. It is known that several metals including lead and silver pass through the refining stage to the residue and those metals are not easily recoverable by conventional processes. Presently, the residues are recycled several times by conventional refining techniques and some metal recovery is evident but there is a proportion of lead and silver metal which is not recoverable. Further, the metal content in the residue is so high that the residue is often not safe for disposal. There are also a number of trace metals in the residue and .these ^'metals are not recoverable by conventional methods. The trace metals are commercially valuable and some metals such as Ge and In have application in the semiconductor market. It would therefore be of commercial value to recover the metals from the residue.

DISCLOSURE OF THE INVENTION

It is an object of the present invention to provide a method of extracting valuable metals from leach residues, particularly valuable metals such as Ag, Pb, Zn, Ge, Ga, In and Cd.

One further problem associated with such leach residues is that the residues are highly acidic and require neutralization for conventional refining techniques.

Accordingly, it is a further object of the present invention to provide a process that is not dependent on the pH of the leach residue and so is capable of treating both unneutralized and neutralized leach residues.

According to the present invention there is provided a method of extracting valuable metals from leach residues containing those metals, said method comprising the steps of:

(1) feeding the leach residue to a furnace,

(2) smelting the leach residue by means of a submerged lance under reducing conditions,

(3) recovering a substantial proportion of the valuable metals from the furnace as a fume, and

(4) leaching the fume to extract the soluble metals. Preferably, the leach residue is fed continuously and the fume is acid-leached.

In preferred embodiments of the invention, the leach residue to be treated contains silver, lead, zinc and other metals, all of which substantially revert to fume leaving a slag substantially low in lead, suitable for disposal.

Preferably, the leach residue feeds are in an unneutralized state and may be highly acidic and in yet a further preferred embodiment the residues are lead and silver residues that have been filtered. Valuable metals recovered may include Ag, Pb, Zn, Cd and electronic metals, by which is meant germanium, indium and gallium. The preferred operating temperature is between 1200 to 1220°C, temperatures at which the process was stable and the slags were fluid. The preferred composition of the disposable slag is 0.45% Pb, 0.65% Zn, 350 ppm Ag, 10 ppm Ge, 220 ppm Ga, 10 ppm In, less than 10 ppm Cd, 10 ppm Bi, 10 ppm As and 40 ppm Sb as the aforementioned slag composition meets the West German standard for disposal as Group S solid waste.

It is expected that extraction rates to the fume of valuable metals from unneutralized leach residues will be at least 90% Ag, 98% Pb, 90% Zn, 10% Ga, 99% In and 99% Cd and greater than or equal to 80% Ge. Preferably, 98% Ge. This fume leach residue is then smelted under reducing conditions to produce crude lead and silver together with a low lead slag suitable for disposal.

In a second stage the resulting residue is treated in the same furnace via a battery paste process as described in our' Australian Patent Application.No. 54267/86 to produce crude lead and silver. The reducing conditions referred to preferably includes a lance positioned under the surface of the furnace slag wherein enriched air or oxygen and a stoichiometric excess of oil or hydrocarbon gas is fed to the furnace slag. BRIEF DESCRIPTION OF THE DRAWINGS

A preferred embodiment of the invention will now be described, by way of example only, with reference to the accompanying drawing in which: Figure 1 is a process flow sheet of smelting residues in accordance with the present invention. BEST MODE OF CARRYING OUT INVENTION

In a first smelting stage a Pb-Ag leach residue is combined with an iron oxide based flux and coal for reduction and mixed in a pug mill with sufficient water to form a moist agglomerate. This mix constitutes the feedstock and is fed to an ISASMELT furnace by conveyor and dropped directly into a molten slag bath. (ISASMELT) is a trade mark of Mount Isa Mines.) Alternatively, the feedstock may be dried and injected into the bath down the lance.

The ISASMELT process is based on submerged gas injection into melts via a top entry submerged SIROSMELT (trade mark) lance to produce turbulent baths in which high intensity smelting or reduction reactions can occur.

Heat is supplied by natural gas and combustion air injected down the lance. The slag is tapped continuously or batchwise for disposal. Air or oxygen is injected into the stream of vapours exiting the furnace to oxidize sulphur and to oxidize the metal vapour to metal oxides. The furnace offgas is cooled with spray water and the fume containing valuable metals is collected in a baghouse. The fume from this smelting stage is leached to extract the zinc, cadium and electronic metals for recovery as sulphates.

The residue from this leach ("fume leach residue") contains the lead and silver and is smelted via the ISASMELT battery paste process referred to above to produce lead and silver bullion. The battery paste process is operationally similar to the method of smelting the Pb-Ag leach residue feed described above.

The fume leach residue is combined with coal for reduction and mixed with sufficient water to form a moist agglomerate. This feedstock is fed to an ISASMELT furnace and dropped directly into the molten slag bath. Heat is supplied by natural gas and combustion air injected down a lance.

The lead and silver product is removed as the metal by tapping from the furnace as bullion. The low lead slag resulting from battery paste process is a suitable starting bath for the method of present invention, thereby enabling a direct changeover between the two methods. At the conclusion of smelting the leach residue feed, the slag is batch reduced to produce a disposable low lead slag and smelting of the leach residue feed restarted.

In the experimentation which follows, 4 tonnes of neutralized Pb-Ag leach residue from a zinc roast-leach-electrowin refinery was smelted in a 250Kg ISASMELT test rig in 27 runs over 60 hours operating time. The product fume and slag were leached to determine metal recovery from the fume and the suitability of the slag for disposal. Metal extractions (metal in fume as a percentage of the metal leach residue of the feed) from the fume during the 250Kg testwork were 96% Pb, 75% Zn, 70% Ag, 95% Ge, 5% Ga, 96% In and 99% Cd. It is envisaged that higher recovery rates will be obtained when the treated leach residue is in a non-neutralized state. A 150 g/1 sulphuric acid leach at 60^βC gave extractions (metal in solution as a percentage of the metal in the fume) from the fume of > 99% Zn, > 80% Ge, 90% Ga, 85% In and > 99% C (see Table 1). Less than 0.1% of Pb and Ag was extracted.

TABLE 1

STEADY STATE SLAG COMPOSITIONS AND EXTRACTIONS

DURING CONTINUOUS REDUCTION SMELTING OF Pb-Ac

LEACH RESIDUES

250Kg ISASMELT TESTWORK PROPOSED PLANT

OPERATING CONDITIONS OPERATING CONDITIONS

Temp 1200 to 1220°C Temp 1200 to 1220°C

Feed Neutralized Residue Feed Un-neutralized Residue

Flux Converter slag at Flux Haematite at 25% 100% by weight of by weight of dry residue residue

Coal 45% by weight of Coal 41% by weight of

(64% FC) dry residue (70% FC) dry residue

Example 1

Testwork on Pb-Ag leach residue feed was carried out in a 250Kg Isasmelt furnace. The furnace had an internal diameter of 380mm and was 1.5m high. It was lined with chrome-magnesite refractories. Air and oil were injected into a bath through a 30mm nominal bore lance. Feed was added via a variable speed conveyor and dropped directly into the molten slag bath. The offgases passed through an evaporative gas cooler and were cleaned in .a reverse pulse baghouse.

The leach residue feed was mixed with converter slag as a flux and coal for reduction then pelletised prior to each run. At the beginning of each series of runs 50 kg of slag from previous runs was melted in the furnace to form a starting-bath. The wet residue was then fed continuously to the furnace at a controlled rate and smelted. Oil was used as fuel. Normal air and air enriched to 30 and 35% oxygen were used, with 30% O enrichment giving optimum conditions. Slag samples were taken periodically from the furnace during each run and the fume rates were measured at the baghouse.

In the first 8 runs the leach residue feed was smelted in a stoichiometric excess of oxygen via the lance, then batch reduced. However, this process was not successful.

In subsequent runs the leach residue feed was smelted in a stoichiometric deficiency of oxygen to fume off the valuable metals in a single continuous process. This was a stable and easily controlled process, giving good extractions for all metals except gallium. The lead residue feed was fluxed with converter slag to give a fluid fayalite slag. The operating parameters investigated were: temperature: in the range 1160 to 1250°C; and coal rate: in the range 20 to 50% by wt of dry residue. The feed material assays are given in Table 2.

A final batch stripping was also performed on a number of these reduction smelting runs. Although this did increase metal recoveries, changing from a continuous to a batch process negates any benefits.

Table 2 continued

Proximate Analyses Coal: 64% Fixed Carbon 21% Volatiles

15% Ash, Assaying at 7.5% Fe, 62% Si0₂, 23% A1₂0₂, 3% CaO, 0.5% MgO.

The Pb-Ag leach residue feed was smelted under reducing conditions in an attempt to produce a disposable slag. This approach maximised the fuming rate and resulted in a leachable fume containing the valuable metals in a single continuous process. Surprisingly, reduction smelting was stable and easily controlled, with fluid slags and high fuming rates under a range of operating conditions. All the valuable metals except gallium were extracted into the fume for further recovery and the slag was suitable for disposal.

Metal extractions increased with temperature and coal rate. Silver extractions in particular were highly dependent upon temperature and coal rate. Gallium extractions were low under all conditions. Optimum operating conditions were temperatures in the range from 1200 to 1220°C and a coal rate of 40 to 50% by weight of dry residue. The slag composition and metal extractions under such optimum conditions are summarized in Table 1. The projected metal extractions for the proposed plant are also given in Table 1. Example 2

Granulated slag samples from five runs were leached according to the West German standard method for "Group S", for determining suitability for disposal as solid waste.

The toxic levels in the leach solutions from four of these slags were all well below the standard limits. The slag from one run, however, gave Cu, Pb and Tl levels which were slightly higher than the Group S standard, and more than 100 times higher than the other four slags. This result appears anomalous since the slag was of intermediate composition compared to the other slags tested, and it is considered that the sample was contaminated during granulation. The conclusion was that slag samples having a lead content of below 0.45% w/w are suitable for disposal as land fill. Furthermore, slag particle size did not appear to significantly affect slag leaching characteristics. Example 3

Leaching testwork on the fume was carried out under a range of conditions. A 4 hour sulphuric acid leach at 70 g/1 acid strength and 60°C gave a recovery of 95% Zn and 90% Cd from the fume. Increasing acid strength to 110 g/1 recovered 70% of Ge and 25% Ga content of the fume. A sulphuric acid leach at 150 g/1 acid strength and 60°C gave recoveries from the fume of >99% Zn, >80% Ge, 90% Ga, 85% In, and >99% Cd under these conditions. Less than 0.1% of the lead and silver were mobilised and the rest of the lead and silver remained in the fume leach residue after extraction.

It was concluded that an acid strength of 150 g/1 and temperature of 60°C achieved the highest recoveries of Ga, In and Ge and complete recoveries of Zn and Cd. An acid strength of 70 g/1 and temperature of 60°C is sufficient to achieve high recoveries of Zn and Cd. However, Ga, In and Ge were not sufficiently mobilized.

On a plant scale, allowing 70% plant availability, 15% maintenance and 15% contingencies, the smelting rate to treat 40,000 dry tonnes per annum, lead-silver leach residue feed is expected to be 7 tph (dry weight) of the leach residue for 65% of the operating time and 20 tph (dry weight) of the fume leach residue for 5% of the operating time. The separate steps may be carried out on an alternating basis in an Isasmelt furnace preferably of dimensions 2m inside diameter. It is envisaged that over a two week period, the leach residue feed would be treated continuously for a period of 13 days and the resulting fume leach residue on the remaining day.

A flow sheet for the two treatments is shown in Figure 1 and an operating summary given in Table 3: TABLE 3

SMELTING OPERATING SUMMARY

First Smelting Stage Second Smelting Stage ^' (Production of (Recovery of Disposable Slag) Lead and Silver)

Material Pb-Ag Leach Residue Fume Leach Residue Treatment Rate 7 tph 20 tph for 65% of the time for 5% of the time

Product 4.3 tph 14 tph

Disposable Slag Pb-Ag Bullion

Operating Temp. 1220 to 1200°C 980 to 1000°C Natural Gas 0.65 Nm³/s 0.60 Nm³/s Lance Air (30%)

Blower @ 150 kPa 3.8 Nm³/s 3.7 Nm³/s

Oxygen ~ 150 kPa 0.5 Nm³/s (60 tpd) 0.5 Nm³/s (60 tpd) Reduction Coal 2.9 tph 1.0 tph

(70% FC) Haematite Flux 1.75 tph Spray Cooling 15.3 litres/s 4.0 litres/s

Water Offgas at 33.6Nm³/s @ 0.6% S0₂ 15.7Nm³/s § Baghouse Exit (wet basis) 2.5% SO2 (wet basis)

11.7Nm³/s @ 1.7% S0₂ 8.3 Nm³/s @ (dry basis) 4.7% S0₂ (dry basis)

The major consumptions for the two smelting stages are estimated in Table 4 set out below:

TABLE 4

MAJOR SMELTING CONSUMPTIONS PER TONNE OF Pb-Aσ LEACH RESIDUE

First Smelting Second Smelting

Stage Production Stage Recovery of Disposable Slag of Pb and Ag Total

Natural Gas 340 NπrVt 25 Nm³/t 365 Nm³/t Air (150 kPa) 1010 Nm³/t 150 Nm³/t 2160 Nm³/t Oxygen (150 kPa) 250 Nm³/t 20 Nm³/t 270 Nm³/t Coal (70% F.C.) 410 kg/t 10 kg/t 420 kg/t Haematite 250 kg/t 0 kg/t 250 kg/t Labour (3 men/shift) 0.68 man hr/t 0.05 man hr/t 0.73 man hr fc Process offgas is treated by afterburning (oxidizing) CO and Coal volatiles with air, followed by evaporative gas cooling and separation of the fume via a bag filter. The offgas composition after the bag filter

3^• is shown in Table 5 set out below (including 1 Nm /s plus air plus leakage into baghouse) :

TABLE 5

SMELTING AT 7 TPH SMELTING AT 20 TPH

Pb-Ag LEACH RESIDUE FUME LEACH RESIDUE

The afterburning prevents condensation of molten elemental sulphur over the surface of the fume particles.

On a plant scale, the afterburning air requirement at 7

3 tph residue is 7 Nm /s, injection-plus ingress air. The furnace offtake is maintained below 1250°C by 3.6 1/s spray cooling water injected into the offtake. An additional

11.7 1/s gas cooler spray water is required to cool the offgas to the 150°C baghouse entry temperature.

When smelting the second stage fume leach residue, the

3 Nm /s ingress air into the top of the furnace under normal draught conditions provides sufficient afterburning. No spray water is required in the offtake, but 4.0 1/s is required in the gas cooler to cool the offgas to 150°C.

The SO_? content during the afterburning step is low at 1.7% dry basis. It is envisaged that cheaper gas handling systems may prove feasible depending on process changes such as higher slag temperature and reduced coal consumption or reduction by natural gas.

The envisaged plant operating costs, excluding fume leaching costs, are detailed in Table 6.

TABLE 6

Smelting: 7 tph Pb-Ag Leach Residue for 65% of the time. 20 tph Fume Leach Residue for 5% of the time.

ANNUAL COST ANNUAL COST

ITEM BASIS FOR ESTIMATE AUD (x 1000)

Labour 1/2 Metallurgist/Supervisor 40

13 Operators 640

(3 per shift + holiday relief)

Natural gas 0.65 Nm³/s by 65%

0.60 Nm³/s by 5%

125 Nm³/hr by 20% 3200

(heating and holding)

Oxygen 0.5 Nm³/s by 70% 710

Reduction 2.9 TPH by 65%

Coal (70% FC) 1.0 TPH by 5% 4240

Haematite 1.75 TPH by 65% 10

Flux

Engineering Labour: 1 Fitter 50 maintenance

Materials: Based on current ISASMELT usage 100

Other stores Based on current ISASMELT usage 150 Refractories 100 tonnes Chrome-magnesite bricks

(1 year life) plus Launders etc.

(labour and materials) 400

Baghouse bags 3000. M² Ryton bags 210

Compressed air Based on current Isasmelt Usage (Baghouse and general) 100

Power 1MW by 90% 610

(total power usage-, incl. blower)

Fresh Water 55 m³/hr by 65% (spray water) 14 m /hr by 5% (spray water) 1 m³/hr by 90% (general) 30

Transport 3ased on material moved 120

TOTAL COST 10610

UNIT COST PER TONNE OF Pb-Ag LEACH RESIDUE TREATED AUD266 TABLE 7 BUDGET CAPITAL COST ESTIMATE FOR ISASMELT SUBMERGED COMBUSTION PLANT

Australian

MIXER $ 150 000

FEED CONVEYOR 50 000

ISASMELT VESSEL (2m ID Refractory lined, with water cooled copper top) 1 400 000 GAS COOLER (Refractory lined, including sprays etc.) 1 200 000 LANCE HANDLING 100 000 STRUCTURAL STEEL 400 000 FOUNDATIONS 200 000 CONTROL ROOM 75 000

PROCESS AIR BLOWER (850 kW) 450 000 PROCESS AIR LINE 75 000 PROCESS GAS FLUE (150 m) 100 000 PROCESS GAS BAGHOUSE (including bags) 500 000 PROCESS GAS FAN 200 000 CHAIN CONVEYOR 100 000

SUBTOTAL: $6 000 000

INSTRUMENTATION @ 10% 600 000 ELECTRICS @ 10% 600 000

ENGINEERING DESIGN ~ 20% 200 000

SUBTOTAL: $8 400 000

CONTINGENCY ⁽§ 30^ 2 500 000

TOTAL BUDGET CAPITAL ESTIMATE AUD 10 900 000

Exclusions Feed bins and weighers

Hygiene ventilation requirements

Product handling, i.e.; slag granulation or casting Provision of services

Leaching plant

Scale up to a production plant is based,on reduction smelting of un-neutralized Pb-Ag leach residue feed at 7 tph (dry wt) for 65% of the time and smelting the fume leach residue at 20 tph (dry wt) for 5% of the time. This gives a potential annual treatment rate of 40.000 tonnes Pb-Ag leach residues. Allowing for a 15% plant downtime for general breakdowns, feed changeovers and furnace rebricking, a 15% operating time contingency has been considered prudent because of the difference between the neutralized residue tested and the un-neutralized residue to be used in the proposed plant.

The maximum smelting rate used during the 250 kg testwork was 80 dry kg residue per hour. Scaling up, this indicates that a furnace of 2.0 m internal diameter is required to smelt 7 dry tph leach residue.

This furnace will also enable treatment of the leach residue at 20 dry tph using air enriched to 30% 0₂ and natural gas as fuel.

During the 250 kg testwork, oil was used as fuel. For reduction smelting, the lance air was enriched to 30% 0„ and the air to oil stoichiometry was 90% of that required for complete oil combustion, to maintain non-oxidizing conditions at the tip of the lance.

The oil consumption during reduction smelting at 1200 to 1220°C was 300 g oil/dry kg residue at 30% 0₂, after correcting for furnace heat losses. This is equivalent to a heat requirement 5.2 MJ/dry kg residue on the proposed plant, i.e., when fluxing at 25% haematite, or 10 MW at 7 tph residue. The furnace heat losses on the proposed plant are estimated to be a maximum of 1 MW, giving a total heat load of 11 MW at 7 tph residue.

The natural gas required to supply this heat at 1220°C, 90% combustion stoichiometry and 30% 0_? is 0.65 Nm³/s. The air rate at this fuel consumption is 3.8 Nm /s and the oxygen rate is 0.5 NmVs (i.e., 60 tpd) to give 30% 0₂. The air supply pressure for low pressure lances is 150 kPa, supplied from a single stage blower. The oxygen supply pressure is also 150 kPa.

During the 250 kg battery paste testwork, oil was used as fuel. The lance air stoichiometry was 95% and no oxygen enrichment was used. The operating temperatures were 950 to 1000°C.

The oil consumption during smelting at 1000°C was 75 g oil/dry kg fume, after correcting for furnace heat losses. Considering the higher water content of the fume after leaching, it is estimated that this is equivalent to a heat requirement of 2.0 MJ/dry kg residue on the proposed plant, or 12 MW at 20 tph residue, with 10% fuming which is recycled. Including furnace heat losses, the total heat load is 13 MW.

The natural gas required to supply this heat 1000°C, 95% stoichiometry and 30% 0₂ is 0.60 Nm²/s. The air rate at this fuel consumption is 3.7 Nm²/s and the oxygen rate is 0.5 NrnVs.

The reduction coal rate required to give high metal extractions and a disposable slag during reduction smelting of the leach residue in the testwork was 45% by weight of dry residue; i.e., 2.0 tph fixed carbon for reduction at 7 tph residue. This is equivalent to a coal rate of 2.9 tph for coal containing 70% fixed carbon on the proposed plant.

A much milder reduction is used when smelting leach residues, and only 5% coal by weight of dry residue is required during this stage. This corresponds to a rate of 1.0 tph coal at 20 tph residue on the proposed plant.

Normalizing the major slag components, acceptable slag compositions for the process lie in the range 45 to 55% Fe), 30 to 40% (SiO + Al O ) with a maximum of 6% Al O , and 10 to 20% (CaO + MgO) with a maximum of 3% MgO.

The Pb-Ag leach residue smelted during the 250 Kg testwork had been neutralised with lime which significantly increased the flux requirements. Converter slag was used as an iron oxide and silica flux, and was added at the same rate as the dry residue.

Neutralisation is not necessary for smelting the residue in the ISASMELT submerged combustion furnace, enabling the use of an iron oxide based flux. In the proposed process haematite from the haematite plant is used as flux at a rate of 25% by weight of dry residue, i.e., at 1.75% tph. Alternatively, low lime (5% CaO) electric arc furnace dust could be used and the zinc also recovered from this material.

The holding capacity of a 2 m internal diameter furnace will allow 2 hours operation between slag taps when smelting the leach residue. This slag may be either granulated or cast for disposal depending upon the requirements.

The time between crude lead taps when smelting the leach residue will also be 2 hours.

Lances are constructed out of schedule 40 mild steel pipe with a stainless steel tip and an internal swirler. They wear by gradual thinning of the tip, and when approximately 300 mm has been worn off another length of stainless steel is rewelded on to the end to form a new tip. Lance lives currently average 100 hours operating time. Similar lives are expected when smelting Pb-Ag leach residues.

The furnace is lined with chrome-magnesite wear bricks. Based on known refractory lives, a minimum period of one year is expected between rebricks.

A moist agglomerate feedstock is used giving good hygiene.

Although the invention has been described with reference to specific examples, it will be appreciated by those skilled in the art that the invention may be embodied in many other forms.

Claims

CLAIMS :

1. A method of extracting valuable metals from leach residues containing those metals, said method comprising the steps of:

(1) feeding the leach residue to a furnace,

(4) leaching the fume to extract soluble metals.

2. A method of extracting valuable metals from leach residues wherein the leach residue is fed continuously and the fume is acid-leached.

3. A method of extracting valuable metals from leach residues in accordance with claim 1 or 2 wherein the leach residues are leach residues of lead and/or silver.

4. A method of extracting valuable metals from leach residues in accordance with. claims 1, 2 or 3 wherein the valuable metals are trace metals.

5. A method of extracting valuable metals from leach residues in accordance with claims 1, 2 or 3 wherein the valuable metals are lead and silver.

6. A method of extracting valuable metals from leach residues in accordance with claim 4 wherein the trace metals are Ge, In and Ga.

7. A method of extracting valuable metals from leach residues in accordance with claims 1, 2 or 3 wherein the valuable metals are Ag, Pb, Zn, Ge, Ga, In and Cd.

8. A method of extracting valuable metals from leach residues in accordance with any one of the foregoing claims wherein the residue also contains silver, lead, zinc and other metals, all of which substantially revert to fume leaving a slag substantially low in lead.

9. A method of extracting valuable metals from leach residues in accordance with any one of the foregoing claims wherein the leach residue is in an unneutralized state.

10. A method of extracting valuable metals from leach residues in accordance with any one of the foregoing claims wherein the lead and silver residue is filtered.

11. A method of extracting valuable metals from leach residues in accordance with any one of the foregoing claims wherein the operating temperature is between 1200 to 1220°C.

12. A method of extracting valuable metals from leach residues in accordance with any one of the foregoing claims wherein the composition of the slag is at most 0.45% Pb, 0.65% Zn, 350 ppm Ag, 10 ppm Ge, 220 ppm Ga, 10 ppm In, less than 10 ppm Cd, 10 ppm Bi, 10 ppm As and 40 ppm Sb.

13. A method of extracting valuable metals from leach residues in accordance with any one of the foregoing claims wherein the extraction rates of valuable metals is at least 90% Ag, 98% Pb, 90% Zn, 10% Ga, 99% In and 99% Cd and greater than or equal to 80% Ge.

14. A method of extracting valuable metals from leach residues in accordance with claim 13 wherein the percentage of Ge is 98%.

15. A method of extracting valuable metals from leach residues in accordance with claim 1 wherein residue from step 4 is smelted to produce crude lead and sinter.

16. A method of extracting valuable metals from leach residues in accordance with claim 15 wherein a lance is positioned under the surface of the furnace slag wherein enriched air or oxygen and a stoichiometric excess of oil or hydrocarbon gas is fed to the furnace slag.

17. A method of extracting valuable metals from leach residues in accordance with claim 1 substantially in accordance with the Examples.

18. A method of extracting valuable metals from leach residues in accordance with claim 15 substantially in accordance with the Examples.