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Battery Breaking Blast Furnace

Rotary Kiln Short Rotary Furnace Separation Stage 1 Top Lance Slag Bath Reactor

Reverberatory Furnace Separation Stage 2 Paste Desulfurisation Short Rotary Furnace Leaching & Electrowinning Kettle Melting Short Rotary Furnace Plastics Metallics Blast Furnace Electric Arc Furnace Pastes Paste + Metallics Whole Drained Batteries

Slags for secondary

smelting Slag recycle

Slag recycle

Because the furnace conditions are not strongly reducing, a PbO-rich slag will be present and antimony, arsenic and tin tend to be oxidised and remain in the slag as per Equation 11.6. The lead bullion produced is thus relatively free of alloying components, although any copper, bismuth and silver present in the feed mix is captured by the bullion.

Pb + PbO₂= 2PbO (11.1) 4Pb + PbSO₄= 4PbO + PbS (11.2) PbS + PbSO₄= 2Pb + 2SO₂ (11.3) PbS + 2PbO = 3Pb + SO₂ (11.4) 2PbO + C = 2Pb + CO₂ (11.5) 2Sb + 3PbO = 3Pb + Sb₂O₃ (11.6)

Furnace gases are cooled either by direct water injection or by indirect heat exchange to below 200°C and are then cleaned in a bag filter. Collected dusts are recycled to the feed mix. Although most secondary smelters using this technology in the past discharged filtered gas directly to atmosphere, the SO₂content no longer makes this permissible in most locations. Hence gas scrubbing to remove SO₂using lime or ammonia solutions is necessary for the application of this technology.

Slag production is roughly one tonne for every four tonnes of bullion produced and typical slags contain around 75 per cent Pb and close to 90 per cent of the antimony in the feed material, with bullion commonly reporting around 0.1 per cent Sb. The reverberatory furnace is not efficient for reduction and the slag is normally treated in separate equipment with a higher reduction potential. The blast furnace has commonly been used, but also the short rotary furnace and electric furnace are used for this duty. The bullion produced from reduction of reverberatory slag is enriched in alloying elements such as antimony and tin and requires refining or careful blending to produce suitable alloys.

In some instances first run reverberatory slag can be re-run through the furnace on a campaign basis. The slag is mixed with coal at around five per cent of feed and more than half the lead content

Feed

Arched Roof Exhaust ports Exhaust gas

Burner

Feed port Working ports Drying feed

Slag Bullion Hearth

can be recovered as bullion. The resulting second run slag is enriched in impurity elements and reduced in volume for separate treatment. However, this does leave an intermediate bullion stream with somewhat higher impurity levels and will possibly need to be refined.

THE BLAST FURNACE

The blast furnace is used in two modes, one for the smelting of whole batteries and the second for the reduction of reverberatory slags. A key feature of the blast furnace is its requirement for bulk feed to maintain an open structure in the shaft so that gas flow is not impeded. As a result it is not able to process fine material and application to separated battery paste reduction is impractical.

Secondary blast furnaces are much smaller than primary smelter furnaces and are often circular in cross-section rather than rectangular as with larger furnaces. Furnace shaft diameter at the tuyeres ranges from 0.7 to 1.5 m, and working height is from 2.5 to 3.8 m. Larger furnaces, mainly used for whole battery smelting, tend to be of rectangular cross-section similar to the primary smelter units (see Chapter 5).

Whole battery smelting

This is typified by the Varta process implemented by Varta Batterie AG at Hannover, Germany and as used by Boliden-Bergsoe AB at Landskrona, Sweden. The batteries are drained and charged with fluxes (lime, silica and shredded scrap iron), coke and an amount of recycle lump slag. The slag has a melting point above 1000°C and maintains an open structure in the shaft until it nears the tuyere zone with a temperature of around 1150°C.

Coke use is around 180 kg per tonne of crude lead bullion.

Larger furnaces of rectangular cross-section are generally required in this case and the Varta furnace is 4 m2in hearth area and 7 m high. The furnace used at Britannia Lead in the United Kingdom (UK) was similar at 1.1 m wide by 4 m long at the tuyere level with eight tuyeres per side and 4.7 m high (Koch and Niklas, 1989; Taylor and Moore, 1980).

Water is evaporated in the upper part of the shaft, followed by decomposition of the plastic components as the temperature rises to 500°C. Lead also melts and reacts with PbO₂to form PbO. This is reduced by CO to lead further down the shaft, and sulfates are reduced to PbS. A feature of the Varta process is the use of metallic iron to react with PbS and form lead and FeS as a matte, thus capturing the majority of the sulfur in feed.

The crucible at the base of the furnace contains lead bullion and a mixed slag–matte phase. Lead bullion is usually tapped continuously via the lead well into a ladle, and slag may be tapped intermittently into slag pots where the matte separates. The solidified slag and matte can be broken up and mechanically separated. The lump slag is crushed to an appropriate size and part is returned to the furnace. The matte phase contains excess iron and is represented as (FeS + 0.35Fe). It contains 25 to 26 per cent sulfur and six to nine per cent Pb as entrained metal. Matte provides the main outlet for sulfur and is usually disposed of in landfill.

Furnace gases pass through a fuel fired afterburner to raise gas temperature to 1000°C in order to combust organics distilled from decomposing plastics in the upper shaft. After burning the gases pass through a drop-out chamber and are then cooled and filtered through a bag house before discharge to atmosphere. Although most of the sulfur is captured as matte this will not be complete, and the advent of tighter emission standards may make scrubbing of final gas necessary. The destruction of organics by afterburning requires added fuel in addition to the CO content of the gas and is a significant inefficiency in the energy requirements of this process.

Chlorine from PVC separators forms volatile lead chloride, which vaporises in the lower shaft and condenses in the upper shaft, forming a recycle loop as well as wall accretions. Some lead chloride reports to flue dusts, again causing accretions in the gas handling system, and collected dusts must be leached for removal of chlorine before recycling to the furnace feed.

Another difficulty with the processing of whole batteries is the tendency of the cases to melt in the upper part of the shaft to form a sticky mass bridging across and blocking the furnace.

The blast furnace is also a high energy consumer for various reasons, not the least being the requirement to melt a high circulating load of slag. Furnace heat is supplied by the burning of coke, which is a costly fuel. Together with the complications of consuming plastic materials, the bulk of which can be a valuable by-product, and the environmental issues with disposal of lead-rich mattes, these problems have rendered the blast furnace smelting of whole batteries an unattractive approach.

Blast furnace smelting of reverberatory slags

The furnaces used in this case are relatively small and commonly circular in cross-section. A schematic of a typical small-scale furnace is shown in Figure 11.3. Construction of the shaft is water jacketed mild steel, sitting on a brick crucible fitted with lead well and siphon. The top is sealed using a cone seal through which the charge is periodically dropped. The furnace feed is crushed reverberatory furnace slag and it is fuelled with coke at around ten per cent of the charge. Lime, silica and shredded scrap iron are added as fluxes, making up about 20 per cent of the charge.

Charge Hopper Gas Offtake Cone seal Water jacketed shaft Tuyere Slag spout Bullion spout Crucible Shaft Feed Dam

In some cases the bullion produced can be rich in antimony and to achieve high antimony recoveries under these circumstances it is necessary to avoid speiss formation, and hence keep the iron input relatively low. In turn this means that slags are lower in iron than normal primary smelter slags with FeO:SiO₂ratios at 0.8:1 and CaO:SiO₂at 0.9:1. This compares with typical primary smelter slag ratios of 1.3 and 0.7 respectively. The lead content of slag is typically between one and three per cent.

Where lower antimony slags are processed and recovery is not a critical issue, then more typical lead blast furnace slags can be used. Where feed has a significant sulfur content, iron can be added to capture the sulfur as a matte, and in this case relatively high iron slags can be produced with FeO:SiO₂ ratios above 2.0 and CaO:SiO₂ratios below 0.5.

Blast furnace gases pass through an afterburner to convert CO to CO₂, and are then cooled by water injection and cleaned in a bag filter before discharge to atmosphere. Collected dusts are recycled to the reverberatory furnace.

It is usual to operate the blast furnace on a campaign basis, stockpiling feed until sufficient has been accumulated for a reasonable campaign run of from one to four weeks. It is also possible to re-run blast furnace slags under different conditions, particularly if it is desired to concentrate antimony or other alloying elements into a small quantity of final bullion. This mode of operation provides considerable flexibility and allows the furnace to be cleaned between runs, minimising operational difficulties due to accretion build-up and maximising the efficiency of the furnace (Pike, 1990).

THE ELECTRIC ARC FURNACE

The electric arc furnace has been applied as a replacement for the blast furnace by RSR Corporation for the purpose of treatment of reverberatory furnace slags (Eby, 1990).

RSR recycle first run slags through the reverberatory furnaces to produce a second run slag enriched in alloying elements such as antimony. The second run slag reports 35 to 50 per cent Pb, five to ten per cent Sb, one to three per cent As and one to three per cent Sn and is fed to the electric arc furnace mixed with coke fines at five to seven per cent, limestone at four to seven per cent and iron turnings at zero to five per cent of the charge weight respectively.

Mixed feed is dried in a rotary dryer using furnace off gases and air to burn CO and is dropped through a port in the furnace roof. Lead oxide is reduced by carbon to metal along with antimony and tin oxides to produce a bullion containing these elements. A residual slag floats on the surface of the bullion and is maintained at a minimum depth of at least 150 mm to provide a resistance path for the electrodes. Any shallower depth has the danger of electrical shorting through the underlying bullion. During operation slag depth will increase to 450 mm.

Slag temperature is relatively high and high iron levels are not required to maintain a fluid slag as is the case in the blast furnace. Hence FeO:SiO₂ratios can be in the range 0.15 to 0.4, with CaO:SiO₂ ratios around 1.0. This will avoid the formation of a speiss phase.

Sulfur in the charge will form a matte phase usually composed of CaS and FeS.

The lead content of slag is between 0.5 and two per cent and the total of lead, antimony, arsenic and tin is generally less than 4.5 per cent, compared with six to 12 per cent for the blast furnace. The slag is disposed of in landfill.

The RSR furnace is 4.8 m internal diameter by 4 m high and is fitted with three 355 mm diameter graphite electrodes, rated at 4.0 MVA. Electrode consumption is around 7 kg/t of slag smelted. Construction is a water-cooled steel shell lined with chrome magnesite brick, with a suspended domed roof. Electrodes are sealed into the roof.

The furnace is operated by slowly adding feed to the furnace and building up a layer of solid charge to a set level. Feed is then stopped and power is increased to fully melt the charge, and at completion the slag is tapped and run down to the minimum level. During the above cycle of about eight hours from one slag tap to another, bullion is also periodically tapped. Energy consumption per cycle is 12 000 to 14 000 kWh. Lead production per cycle is understood to be of the order of 30 tonnes.

ROTARY FURNACE SMELTING

Two broad types of rotary furnace are used – the rotary kiln and the short rotary furnace. Rotary kilns are characterised by large length to diameter ratios of the order of 10 to 15:1, whereas for the short rotary furnace the length to diameter ratio rarely exceeds 1.5:1.

One important aspect of the operation of these furnaces is the slag regime used. Traditionally, a high sodium slag has been used since it had the characteristics of a relatively low melting point and low viscosity, allowing for good separation of slag and bullion. It also provided a means of capturing sulfur into the slag rather than reporting to the smelter gas and the need for costly gas scrubbing. However, there are significant costs in the use of soda ash and there are environmental problems with the disposal of soda slags (see below). As a result of these issues the lime–iron–silica slags used in primary smelting have become more favoured.

Slag regimes

The use of soda slags is primarily a means of capturing sulfur from the smelting operation and the slag is essentially a sulfide or matte – usually a mixture of FeS and Na₂S. The normal fluxes used are shredded iron scrap as cast iron chips or steel turnings and soda ash (sodium carbonate). Small amounts of lime and/or silica might be added depending on the composition of feed material, but is generally not needed with pure battery feeds.

For the smelting of battery pastes the reactions given by Equations 11.7 to 11.14 take place:

PbO + CO = Pb + CO₂ (11.7)

CO₂+ C = 2CO (11.8)

PbSO₄+ 4CO = PbS + 4CO₂ (11.9)

PbS + CO + Na₂CO₃= Pb + Na₂S + 2CO₂ (11.10)

PbS + Fe = FeS + Pb (11.11)

PbO₂+ Pb = 2PbO (11.12)

PbO₂+ 2CO = 2CO₂+ Pb (11.13)

The gas phase mainly contains CO and CO₂, but some SO₂can be present due to limited reaction according to Equation 11.14 – ‘the roast reaction’:

It should be noted that CO₂ generated from sodium carbonate will react according to Equation 11.8 and thus will lead to increased fuel consumption. For this reason and because of the cost of soda ash, the balance between the addition of iron and soda ash should favour a greater proportion of iron.

The slag resulting from the above is basically Na₂S and FeS with a small amount of PbO and possibly CaO and SiO₂. A phase diagram for the Na-Fe-S system is given in Figure 11.4 and shows that between 25 and 75 per cent FeS the melting point is relatively low. The lowest melting point eutectic is 640°C at 36 per cent FeS and the highest for a liquid slag is 760°C at 60 per cent FeS.

A high FeS content in the matte tends to increase the lead content and reduce lead recovery to bullion by reducing the reaction given by Equation 11.11. The equilibrium constant (k) for Equation 11.11 is given by Equation 11.15:

Log k = -1610/T + 2.388 (11.15)

where:

T is the temperature in K

At 1000°C, k = 13.3, so that the activity of PbS in the matte is the activity of FeS divided by 13.3. A high FeS content of the matte also has more tendency to precipitate Fe or magnetite (Fe₃O₄) at lower reduction potentials. These are solid phases at normal operating temperatures and will raise the viscosity of the matte or slag mix, tending to increase the loss of lead as entrained metal. An optimum Na₂S:FeS ratio in matte is regarded as 1:1.86.

If significant slag forming components are present, such as in residue smelting, a separate slag phase will form in equilibrium with the matte. The Na₂O activity coefficient in slag is low compared

CHAPTER 11 – Secondary Smelting Methods

1200 1000 800 600 400 0 20 40 60 80 100 Na2S Composition % FeS T emperature °C Liquid 5Na2S .2FeS ( 730°) 3 Na2S.4FeS 760°C 640°C 650°C 695°C

with matte and will cause sodium to transfer from the matte to the slag until activities are equal. The sodium composition of slag will thus be higher than matte and the sodium contained in slag is essentially wasted as a means of capturing sulfur. Consequently, for cases where a significant amount of slag is produced the aim should be for a low Na₂S:FeS ratio in the matte of little more than 1:4. A reasonable level of sodium is needed to maintain low slag viscosity and high levels unnecessarily consume soda ash as a costly reagent. Some Na₂S will also dissolve in slag and an optimum level of sodium in slag is regarded as 25 per cent as (Na₂S + Na₂O).

An iron–soda matte–slag mix tends to be emulsified and is difficult to separate (Queneau, Cregar and Mickey, 1989).

Although soda slags provide a ready means of capturing sulfur and provide a low viscosity slag/ matte at relatively low temperature for the efficient separation from bullion, there are significant associated problems, which may be listed as follows:

•

Soda ash is a costly reagent.

•

Soda slags are very aggressive to chrome magnesite brickwork and to steel if exposure occurs.

•

Soda slags are hygroscopic and can expand and break down in landfill. Under these conditions the FeS can be oxidised to Fe₂O₃and So, generating heat and raising the temperature sufficiently to ignite the sulfur and the remaining sulfides.

•

The slags contain water soluble heavy metal compounds, which can be leached in landfill situations, particularly when breakdown occurs as above.

Because of the above, disposal of soda slags are classified as hazardous wastes with significant disposal difficulties and are unacceptable in many locations, hence there has been significant incentive to change to a more conventional silicate slag regime as used in primary smelting (see Chapter 5). This will require an increase in operating temperature to 1150 to 1200oC, which can be assisted by the use of oxygen enrichment. A matte can still be generated using metallic iron fluxing, but sulfur has more potential in this case to report to the gas phase. The solubility of sulfur in the silica slag is also limited, and in general the sulfur should be less than one per cent in the feed material. This will essentially require desulfurisation of the battery pastes prior to smelting with the use of silica slag. Target silica slags are generally of the olivine structure with melting points in the range of 1100 to 1200°C and composition as follows:

•

ten to 15 per cent CaO

•

40 per cent SiO₂

•

45 to 50 per cent FeO

Fluxing materials required are silica, shredded iron scrap and lime, with coal as the reductant.

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