AU632650B2

AU632650B2 - Conversion of zinc sulphide to zinc

Info

Publication number: AU632650B2
Application number: AU68934/91A
Authority: AU
Inventors: Steven Paul Matthew; Roger Leo Player
Original assignee: Mount Isa Mines Ltd
Current assignee: Mount Isa Mines Ltd
Priority date: 1989-12-05
Filing date: 1990-12-04
Publication date: 1993-01-07
Anticipated expiration: 2010-12-04
Also published as: WO1991008317A1; AU6893491A; IE904367A1

Description

TITLE: "ZINC SMELTING"

The present invention relates to a method of producing zinc from a composition containing zinc sulphide, for example from a zinc concentrate or a bulk concentrate containing both zinc sulphide and lead sulphide. BACKGROUND OF THE INVENTION

In the prior art, several methods are known for the recovery of zinc including the electrolytic zinc process and the use of a chemical reducing agent.

The electrolytic zinc process consumes a large quantity of electrical energy and produces a toxic heavy metal waste (jarosite sludge) as a by-product. The by-product presents a problem in waste disposal and it would be advantageous to produce zinc from zinc concentrate without the toxic by-product.

The use of a chemical reducing agent (instead of electricity) eliminates the toxic product and overcomes the problems associated with toxic by-product disposal, but is in itself an expensive process due to the necessary inclusion of a pre-treatment sintering step and the large amount of coke required. A well-known example of such a process is the Imperial Smelting Process, which includes sintering of the concentrate after roasting of the concentrate to eliminate sulphur.

A flash smelting process has been proposed by Davey in which a bone dry finely divided zinc sulphide concentrate and a fuel are mixed with, and subjected to combustion in, gaseous oxygen (reaction I).

The chemical equilibrium for reaction I requires that very high temperatures are employed in order to transform a high percentage of the zinc from zinc sulphide to zinc vapour. In this flash smelting process Davey proposed to combine oxygen with a diluent in a selected ratio as a means for obtaining an acceptable yield at temperatures as low as 1500°K. Shock chilling of the reaction gas stream was proposed as a means of suppressing reversion (reaction II).

According to Davey, optimum yields would be obtained by using from 20 to 30 grams of coal (dry, ash free) per mole of zinc and from 2.5 to 4.0 moles of oxygen per mole of zinc at a temperature in excess of 1500°K and preferably in excess of 1600°K. However, it is difficult in practice to provide satisfactory refactory materials at 1600°K. The design of equipment for the combustion of fuel and concentrate in a stream of diluted oxygen with subsequent separation of products from the reaction gases by shock chilling, without prior appreciable loss of temperature, as envisaged by Davey, has been regarded as daunting and has prevented commercial development of the process.

To avoid the difficulties of a single step flash combustion smelting process a two-bath furnace process has been proposed. In the first furnace, zinc sulphide is fully oxidized in a molten bath at 1350°C with industrial oxygen to produce zinc oxide slag (reaction III) .

2ZnS + 30 2 - 2Zn 0 + 2SO2 (^VIII)'

The zinc oxide bearing slag is transferred to a second furnace where the zinc oxide in the bath is subjected to strongly reducing conditions (reaction IV) to produce a fume of metallic zinc which is collected in a splash condenser.

ZnO + C - Zn + CO (IV)

More recently it has been proposed to carry out both above reactions (III) and (V) in adjacent interconnected reaction zones in a single furnace. Zinc sulphide is injected with an oxidizing gas by lance into the first of two zones in a furnace to produce zinc oxide in a slag and remove sulphur dioxide from a first flue. The slag layer containing the dissolved zinc oxide is then transferred to a second zone wherein the zinc oxide is reduced to zinc metal which is removed as a vapour from a second flue. The zones are separated by a fluid cooled wall which extends into the slag layer and divides the gas space above the bath.

However, the two zone smelting apparatus and method creates, by its own design, problems associated with circulation and mixing of the slag between the two zones. It is essential to the efficiency of the two zone smelting apparatus and process that proper circulation and mixing of the slag occurs between the two zones as proper circulation and mixing ensures transfer of zinc oxide from the first zone to the second zone. It is not clear how the control of this circulation would be achieved.

Zinc is usually found in nature as a sulphide. The ore is commonly concentrated prior to further treatment. Zinc is often found in combination with lead and an increasing amount of concentrate is becoming available which contains both zinc and lead in a form which renders difficult the separation of zinc and lead from the ore. For example, "Low Grade Middlings" (LGMs) or "bulk concentrate" assays from 25% to 35% of zinc and from 20% to 10% of lead.

An object of the present invention is to provide a process for the recovery of zinc from a composition containing zinc sulphide which avoids or at least ameliorates disadvantages of the prior art. An object of a preferred embodiment is to provide an improved process whereby zinc and lead can be recovered from a bulk concentrate as elemental metals. DISCLOSURE OF THE INVENTION

According to a first aspect, the present invention consists in a method for producing zinc from a composition containing zinc sulphide comprising the steps of:

(1) establishing a molten slag bath containing ferrous and ferric oxides at a bath temperature of from 1150°C to 1300°C;

(2) adding said composition to the slag bath whereby to dissolve the zinc sulphide;

(3) oxidizing the dissolved zinc sulphide by reaction with ferric oxides in the bath to produce zinc metal vapour, sulphur dioxide gas and ferrous oxide;

(4) allowing the zinc metal vapour with the sulphur dioxide gas to leave the bath and then separating the zinc from the sulphur dioxide gas; and

(5) controlling the ratio of ferric to ferrous iron in the bath within a predetermined range so as to optimize the production of zinc metal vapour.

The composition may be a concentrate containing both zinc sulphide with lead sulphide. In that case both sulphides dissolve in the bath and are oxidized by the ferric oxide. Both zinc metal vapour and lead metal vapour leave the bath with the sulphur dioxide gas. Some lead may remain as a molten lead layer below the slag depending upon the initial lead 'content of the composition.

The zinc and, if present lead, metal vapours may be collected directly as metals by rapid quenching or by suitable means. Alternatively, the metal vapours may be oxidized in the gas stream for example by addition of an oxygen containing gas such as air to precipitate a fine metal oxide powder which may be separated from the gas stream by conventional methods from which the valuable metals may be recovered by conventional electrochemical means.

For preference, the ferric to ferrous iron ratio in the bath is maintained within the range of 0.01 to 0.10 at 1200°C or 0.10 to 0.25 at 1300°C and by linear interpolation at intermediate temperatures by adjusting the rate of injection of oxygen and or of addition of carbon. The bath is vigorously mixed (by the injection of the oxygen bearing gases at below the surface of the slag bath) to ensure a substantially uniform distribution of ferric iron in the slag.

Preferred embodiments of the invention permit zinc containing concentrate (or a combined zinc and lead bulk concentrate) to be fed into a molten slag bath continuously and enable a high yield of zinc, or lead and zinc, to be obtained at a commercially viable temperature. The method eliminates the sintering step of previous chemical reducing agent methods and, advantageously, coke is not required. The process can be conducted efficiently in a single furnace of relatively simple construction. In preferred embodiments of the invention the process is conducted at from 1150°C to 1250°C to reduce energy costs and minimize refractory wear. DESCRIPTION OF PREFERRED EMBODIMENTS OF THE INVENTION

The invention will now be more particularly described by way of example only with reference to the accompanying drawings wherein:

Figure 1 is a schematic diagram showing an apparatus for conducting the method of the invention.

Figure 2 is a schematic diagram showing another apparatus for conducting the method of the invention.

An important aspect of the present invention is the control of the oxygen activity of a one step oxidation of ZnS to give metallic zinc vapour and sulphur dioxide as follows:

ZnS+O2 - Znvapour + SO2 (V) '

At normal smelting temperatures, zinc is above its boiling point and so exists as a metallic vapour. The oxygen activity for this reaction can be calculated from chemical thermodynamics and is less than 10^~ atmospheres.

Due to this low oxygen activity, it is difficult to add gaseous oxygen directly to particles of zinc sulphide concentrate at smelting temperatures without the formation of a solid layer of zinc oxide (ZnO) or zinc ferrite (ZnO.Fe 2O3) around a core of unreacted zinc sulphide.

The present invention proceeds from the discovery that reaction (V) can be readily controlled in a molten ferrous silicate slag bath, the slag bath having three functions:

(1) The slag acts as a buffer to control the oxygen activity, via the redox reaction:

PO 2 = K x (aFeO1-5)⁴ (VII)

(aFeO)⁴ wherein "PO " denotes partial pressure of oxygen, "K" is the equilibrium constant and "a" denotes the activity of the indicated species. Thus the oxygen activity in the slag can be controlled directly by control of the ferric to ferrous oxide ratio.

(2) The slag also acts as an oxygen transfer medium from the point of gaseous oxygen injection into the melt to the point of addition of zinc sulphide containing concentrate to the melt, so isolating the sulphides from direct contact with oxygen gas. (3) Additionally, the slag fluxes any high melting point phases that may be formed e.g. ZnO or.

The sequence of smelting reactions can be written as:

(A) Air or oxygen enriched air is injected down a lance, or via tuyeres, and reacts with ferrous oxide in the slag to form ferric oxide (reaction VIII) :

0 2 + 4FeO - 4FeO1.5 (VIII)

(B) Zinc sulphide concentrate or bulk lead zinc concentrate is added to the slag, dissolves, and is oxidised by the ferric oxide in the bulk of the slag to yield zinc metal vapour, sulphur dioxide and ferrous oxide (equation IX) :

ZnS + 4FeO • 5 - Zn(vap) + SO2 + 4FeO (IX)

Similarly, for lead sulphide, the oxidation reaction is (reaction X) :

PbS + 4FeO, _c - Pb + SO- + 4FeO (X) Other sulphides in the concentrate (e.g. FeS) are similarly oxidized. The zinc vapour and sulphur dioxide leave the bath, while the ferrous oxide is available to continue the redox reaction (reaction VIII). Metallic lead, if present, forms both vapour and molten lead as a separate layer below the slag. The proportions of lead in the vapour and molten phase depend on the initial lead content of the concentrate. (C) Coal is added to the furnace to provide heat and act as a reductant via reactions of the type.

?Fe>Ω

1.5 + C - 2FeO + CO (XI)

The ferric to ferrous iron ratio of the slag is determined by, and indicative of, the balance between reactions VIII to XI. That is to say the ratio is determined by the rate of supply of oxygen to the slag forming ferric oxide, and the rate of reaction of the ferric oxide with the concentrate plus the rate of transfer of oxygen from the slag (via the ferric iron) to the reductant. By maintaining the ferric to ferrous iron ratio within a predetermined range the production of zinc as vapour may be optimized. Thus, if the ferric to ferrous iron ratio is above a desired value in the predetermined range, the rate of reductant addition is increased or the rate of oxygen injection decreased. If the ratio is below the desired value, the rate of oxygen injection is increased or the rate of reductant addition ■ decreased. By this means, the reaction is controlled to produce the metal vapour in high yield direct from the bath. For the particular case where lump coal floating on the surface of the slag is used as a reductant to control the ferric to ferrous iron ratio, it has been found that, for coal with 65% fixed carbon, it is desirable to use a minimum specific coal addition rate

2 of a 130 Kg/hr of coal per m of the undisturbed slag surface area of the slag bath. For a coal addition rate above about 150 Kg/(hr.m 2), the extraction of zinc from the slag is largely independent of the coal addition rate. In that case the ferric/ferrous iron ratio can be controlled simply by increase or decrease in oxygen injection rate.

With reference to Figure 1 there is shown schematically by way of example a furnace or reactor 1 containing a molten slag bath 2 containing ferric and ferrous iron. The slag in the present examples is a ferrous silicate slag but can be a calcium ferrite or other' ferrite slag. Zinc concentrate is added continuously at feed port 3 together with lump coal. The concentrate dissolves in bath 2 while the lumps of coal tend to float on the surface of the bath in a layer 4. Air or oxygen enriched air is injected to below the surface of the slag bath 2 via a submerged smelting lance 5.* Introduction of the oxygen below the surface of the slag minimizes direct reaction of oxygen with the carbonaceous material floating in layer 4 on top of slag 2 and vigorously agitates the bath.

Fuel is added to the reactor by either injecting a suitable fuel (gas, oil or pulverised coal) together with the injected oxygen via lance 5 or by adding lump coal via feed port 3 to the top surface of the slag bath as described above. The process is not sensitive to the fuel rate and excessive amounts of fuel do not result in a loss of yield.

In practice, it is found that the process can be operated with acceptable yield at bath temperatures in the preferred range of 1150°C to 1300°C, more preferably 1200°C to 1250°C. This temperature range is just above the melting point of the ferrous silicate slag but below the temperature at which the usual magnesite chrome refractories lose strength thus minimising furnace refractory wear.

Metallic zinc vapour is evolved from slag 2, permeates with sulphur dioxide gas through carbon layer 4, mixes with the fuel combustion products in the space 6 above the bath and leaves the reactor still substantially at the preferred smelting temperature via furnace offtake throat 7. If the gases are allowed to cool slowly reversion reactions of the type:

3Zn + SO → ZnS + 2ZnO (XII) and

may occur. Thus desirably the gases are shock cooled to avoid or minimize reversion.

Shock cooling is achieved for example by bringing the hot gases into direct contact and heat exchange relationship with a fluidized bed 8 of cooled particles for example silica or zinc particles. Bed 8 is selected to be of a^' large thermal mass in comparison with that of the gas stream in heat exchange relationship with the bed. Bed 8 may be cooled by water injection or by other cooling means not illustrated in Figure 1 to below the condensation temperature and preferably to below the melting point of zinc in a rapid quenching action. The metal vapour is thus rapidly condensed and captured in the fluidized bed as zinc metal and the sulphur dioxide is separated from the solids, leaving via offtake 15. A portion of the fluidized bed containing captured zinc is drawn off from seal 13.

With reference to Figure 2 there is shown schematically another system for shock cooling the zinc vapour. Parts of Figure 2 corresponding in function to parts illustrated in Figure 1 are identified by corresponding numerals.

The arrangement of Figure 2 differs from that of Figure 1 in that the fluidized bed 8 is recirculated between a first chamber 9 and a second chamber 10. Particles 8 enter the furnace offgas stream at port 11 of chamber 9 and are carried by the gas stream into chamber 10 where the gas is cyclonically separated from the solids and removed at 12. The bed 8 particles are recovered, a proportion removed via seal 13 and the balance recirculated to chamber 9 via seal 14. The particles are cooled prior to entry into the gas stream at opening 11 by direct addition of coolant and/or by heat transfer devices in chamber 9 or. in chamber 10 or both. By means such as shown in Figure 1 or Figure 2 the metal vapour of the gas stream is rapidly cooled, from at or above the bath temperature (above 1,140°C) to below the metal condensation temperature and preferably to below the metal melting point (419°C in the case of zinc, 327°C in the case of lead). Moreover, the cooling takes place rapidly, typically in less than 100 milliseconds and preferably in 10 milliseconds or less.

The fluidized bed of figures 1 and 2 has a particle loading in excess of 200 Kg/m 3, preferably greater than 400 Kg/m 3 and as much as 1600 Kg/m3 or more.

Preferably the thermal mass (mass multiplied by specific heat) of particles in the fluidized bed is such that if the bed were uncooled the rate of heat temperature rise of the bed in heat exchange relationship with the furnace gas would be less than 100°C/sec. and desirably less than 20°C/sec. Desirably, the particles of the bed have an average grain size of less than 2 mm in diameter and preferably 90% of the particles of the fluidized bed have a grain size of less than 0.5 mm in diameter.

In order to minimize reversion prior to quenching it is desirable to control the oxygen concentration, in the vapour-bearing gas stream above the bath and to maintain the temperature of the gas stream at or above the bath temperature.

Oxygen concentration in the vapour stream at the entrance to the fluidized bed can be controlled by introducing inert gases to the furnace, or by maintaining a positive pressure to prevent air ingress, or by adjusting feeds to consume a predetermined ratio of oxygen, or by use of after burners located between the bath and the fluidized bed. Likewise, the temperature of the vapour entering the fluidized bed may be controlled at an elevated level by use of after burners or other heaters and/or by furnace design to minimize heat loss prior to quenching.

Instead of recovering the metal vapour by condensation, the metal may be recovered by oxidation. In that case, air is admitted to the space above the bath or to the furnace offgas stream so as to oxidize the zinc and lead metal vapours, whereby to precipitate the zinc as zinc oxide and lead as lead oxide. The metal oxides may then be separated from the gas stream by means of cyclones, bag house filters or the like.

Periodically, slag samples are withdrawn from the bath and the ratio of ferric:ferrous iron is determined.

The preferred range for the ratio is a function of temperature. At 1200°C the ratio should be between 0.01 and 0.10. At 1300°C the ratio should be within 0.10 and 0.25. Intermediate ratios apply at intermediate temperatures by linear interpolation. Provided the ratio is within a predetermined range having regard to the bath temperature the process is within control. If necessary, the rate of injection of oxygen via the lance is increased or decreased to control the ferric to ferrous iron ratio in the slag bath or the rate of carbon addition is adjusted. Alternatively, the "magnetite" content or "magnetic coefficient" of the slag can be measured and used as an approximate control.

If the ferric to ferrous ratio in the slag falls below the preferred range, the sulphur is inadequately oxidized and thus inadequately removed from the bath.

If the ferric to ferrous ratio exceeds the preferred range, the zinc oxide content of the slag becomes too high to maintain the fluidity of the slag at the temperatures indicated.

It is preferable vigorously to mix the bath to ensure a substantially uniform slag composition.

In accordance with a preferred embodiment of the present invention, the carbonaceous material added to the surface of the bath is low volatile coal having a high percentage of fixed carbon. It is believed that char resulting from fixed carbon in the coal floats on top of the surface of the slag and reacts with the slag while the volatiles leave the system substantially unreacted.

Depending upon the ash and gangue contents of the concentrate and fuel, it may be necessary to add fluxes to produce a fluid slag, as in normal metallurgical practice. It could be expected that minor amounts of lime and silica would need to be added.

Zinc containing concentrate is added to the surface of the bath, together with sufficient additional oxygen in stoichiometric quantities to oxidise all the sulphur in the concentrate to sulphur dioxide and the iron to iron oxides while maintaining the ferric to ferrous oxide ratio referred to earlier. The metallic zinc vapour is rapidly cooled so that zinc metal may be recovered in the presence of sulphur dioxide or is oxidized to facilitate separation.

It is not economical to reduce the zinc content of the slag to a level suitable for disposal under continuous smelting conditions at commercially viable rates .and temperatures. The furnace is therefore operated in a cyclic manner with a 2 to 3 hour total cycle time, involving a continuous smelting stage followed by batch stripping and tapping. During smelting, concentrate, fluxes and coal are fed continuously to the furnace. Smelting air is also added to contr.ol the ferric to ferrous ratio in the slag, as well as fuel oil and combustion air as required to maintain furnace temperatures of 1200 to 1250°C. During smelting, zinc in slag concentrations of 10 to 20% are maintained and approximately 85% to 95% of the zinc in the concentrate is fumed off. After smelting for from 1 1/2 hours to 2 1/2 hours the furnace is filled with slag and at this point the concentrate, fluxes and smelting air additions are stopped. Coal addition continues, together with fuel oil and combustion air injection to maintain 1200 to 1250°C furnace temperatures, and the slag is batch reduced to fume the residual zinc from the slag to a level suitable for disposal. Finally the furnace is tapped leaving sufficient slag in which to resume smelting and smelting is then resumed.

This invention will now be more particularly described with reference to examples and figures from testwork performed at 250 Kg and 3 tonne scales. Example 1

A zinc concentrate feed mixture containing 66% coal and 6% limestone flux by weight of dry concentrate (assays given in Table 1) was prepared by mixing 200 Kg of zinc concentrate containing 6.8% water, 123 Kg of coal containing 0.2% free water and 11.2 Kg of limestone containing 0.2% water with 8 Kg of additional water and pelletising.

Prior to the run, the furnace was heated to 1200°C and 200 Kg of ferrous silicate was added and melted over 4 hours. During this time, the furnace temperature was raised to 1300°C and thermally equilibrated. Air and oil were injected into the bath through a 40 mm diameter submerged lance. The ferrous silicate slag was then partially tapped out, leaving a 150 mm deep layer of slag at the bottom of the furnace as a starting bath. The above pelletised feed mix was then smelted at a rate of 165 Kg/hr, this being equivalent to a feed rate of 90 Kg/hr dry concentrate. Smelting air was injected into the bath through a 40 mm diameter submerged lance at a rate of 315 Nm /hr, this being equivalent to 4.27 mole oxygen per mole of zinc in the feed. An additional 108 NmVhr of air and 10.2 Kg/hr of fuel oil were also injected down the lance and burnt to maintain the operating temperature at 1300°C. The reaction chemistry was monitored by means of periodic slag samples taken from the furnace, giving the results shown in Table 2.

The zinc extraction rate from the concentrate under these operating conditions was 94%. At the conclusion of the run the slag was partially tapped out, leaving a 150 mm deep layer at the bottom of the furnace as the starting bath for the next run which followed on immediately. Example 2

A zinc concentrate feed mixture containing 50% coal and 6.5% limestone flux by weight of dry concentrate (assays given in Table 1) was prepared by mixing 200 Kg of zinc concentrate containing 7.3% water, 93 Kg of coal containing 0.2% free water and 12 Kg of limestone containing 0.2% water with 8 Kg of additional water and pelletising. This run followed on immediately from another run which was only partially tapped to leave a 150 mm deep slag layer as the starting bath.

The above pelletised feed mix was smelted at a rate of 100 Kg/hr, this being equivalent to a feed rate of 60 Kg/hr concentrate. The operating temperature was 1200°C. Smelting air was injected into the bath through a 40 mm diameter lance at a rate of 145 Nm /hr, this being equivalent to 3 mole oxygen per mole of zinc in the feed. An additional 92 NmVhr of combustion air and 8.7 Kg/hr of fuel oil were also injected down the lance and burnt to maintain the operating temperature at 1200°C. The reaction chemistry was monitored by periodic slag samples taken from the furnace, giving the results shown in Table 3.

The zinc extraction rate from the concentrate under these operating conditions was 93%. Example 3

A zinc concentrate feed mixture containing 32% coal and 5.4% limestone flux by weight of dry concentrate (assays given in Table 1) was prepared by mixing 200 Kg of zinc concentrate containing 7.3% water, 60 Kg of coal containing 0.2% free water and 10 Kg of limestone containing 0.2% water with 8 Kg of additional water and pelletising. Prior to the run, the furnace was thermally equilibrated and a 150 mm deep ferrous silicate slag starting bath was established in the same manner as in Example 1 above except that the operating temperature was 1200°C. The above pelletised feed mix was then smelted at a rate of 90 Kg/hr, this being equivalent to a feedrate of 60 Kg/hr dry concentrate. Smelting air was injected into the bath through a 40 mm diameter lance at a rate of 169 NmVhr, this being equivalent to 3.5 mole oxygen per mole of zinc in the feed.

The process was fully fuelled by the coal in the feed and no additional fuel oil was required to maintain the operating temperature at 1200°C. The reaction chemistry was monitored by periodic slag samples taken from the furnace, giving the results shown in Table 4.

The zinc extraction rate from the concentrate under the operating conditions was 78%. Example 4

Smelting trials at rates of 2 and 3 tph zinc concentrate were conducted on a 3 ton Pilot Plant. A starting slag bath was produced from slag already in the furnace from the previous operation. This slag was batch reduced with coal. Silica and limestone fluxes were added to produce a reduced slag of similar composition to the required zinc smelting slag. During this slag conditioning period the furnace temperature was raised to 1250°C.

In this testwork zinc concentrate slurry was gravity fed from a stock tank and dewatered on a drum vacuum filter. The resulting filter cake was mixed with silica flux, limestone flux and coal in a twin shaft paddle mixer. This agglomerate, containing typically 12% water, was continuously fed by conveyor to the furnace where it dropped directly into the molten slag bath. Stable operation was achieved under two sets of conditions as tabulated in Table 5.

Smelting air was injected into the bath through a

100 mm nominal bore lance. At 2 tph zinc concentrate

₃ the smelting air rate was 4900 Nm /hr, which was equivalent to 3.0 mole oxygen per mole of zinc in the feed. To maintain the operating temperature at 1250°C,

270 litres per hour of fuel oil together with 2500

₃ Nm /hr of combustion air were also injected down the lance. At 3 tph concentrate the smelting air rate was 6750 Nm 3/hr, equivalent to 2.75 moles oxygen per mole of zinc in the feed, and 160 litres of fuel oil and 1500

₃ Nm /hr of combustion air were required to maintain

1250°C operating temperatures. The coal rate was kept constant at 700 Kg/hr throughout the trials to maintain a specific coal rate of 250 Kg per hour per square meter of bath surface at both 2 and 3 tph zinc concentrate.

The silica and limestone flux rates were controlled to give a SiO_/Fe ratio in the slag of 0.6 and Si0₂/Ca0 ratio of 3.

The reaction chemistry was monitored by periodic samples taken from the furnace, giving the result shown in Table 6 and Table 7.

A number of additional tests have been performed and the results have been summarised in Tables 8 and 9 for a range of coal and air rates.

A number of tests similar to those outlined in Examples 1, 2 and 3 are summarised in Tables 8 and 9. At a constant feed rate of 60 Kg/hr of concentrate to the 250 Kg test furnace it is seen that the residual zinc in slag remains substantially unchanged as the coal rate is increased from 18 to 42 Kg/hr i.e. as the specific coal rate increases from 150 to 350

₂ Kg/(hr.m ). Additionally from Table 9 it is seen that at a constant coal rate of 30 Kg/hr (a constant specific

₂ coal rate of 250 Kg/(hr.m ) the concentrate rate was increased from 60 to 240 Kg/hr with only a marginal increase in the zinc in slag while the quantity of coal expressed a fraction of the concentrate rate decreased from 50% to 12 1/2%. During these tests the smelting oxygen addition rate, expressed as a mole fraction of the zinc content of the concentrate, was adjusted depending upon the coal rate to control the ferric to ferrous iron ratio to maximize zinc extraction (Figure

3). Large scale smelting trials at 2 and 3 tph concentrate gave similar results (Table 9) at the same

2 specific coal rate of 250 Kg/(hr.m ).

It is therefore preferred to maintain a substantially constant specific coal addition rate and to control the ferric to ferrous iron ratio in the bath primarily by control of the oxygen injection rate.

The process and apparatus herein described may be altered to an extent which will be apparent to those skilled in the art from the teaching hereof or which may be determined by routine experiment based upon the disclosure herein contained and all such variations are deemed to be within the scope of the invention.

/

TABLE 1

TABLE 2

TABLE 3

TABLE 4

TABLE 5

TABLE 6

Smelting Rate = 2 tph Zinc Concentrate

TABLE 7

Smelting Rate = 3 tph Zinc Concentrate

TABLE 8

TABLE 9

Claims

CLAIMS : -

1. A method for producing zinc from a composition containing zinc sulphide comprising the steps of:

(1) establishing a molten slag bath containing ferrous and ferric oxides at a bath temperature of from 1140°C to 1300°C;

(2) adding said composition to the bath whereby to dissolve the zinc sulphide;

• (3) oxidizing the dissolved zinc sulphide by reaction with ferric oxides in the bath to produce zinc metal vapour, sulphur dioxide gas and ferrous oxide;

2. A method according to Claim 1 wherein the ratio of ferric to ferrous iron in the bath is controlled by adjusting the rate of oxygen injection below the bath surface.

3. A method according to Claim 1 or Claim 2 wherein the ratio of ferric to ferrous iron in the bath is controlled by adjusting the rate of addition of a carbonaceous fuel to the bath.

4. A method according to any one of the preceding claims wherein the ratio of ferric to ferrous iron in the bath is controlled at from 0.01 to 0.10 at 1200°C, at from 0.10 to 0.25 at 1300°C and at an intermediate ratio determined by linear interpolation at an intermediate temperature.

5. A method according to any one of the preceding claims wherein the zinc vapour and sulphur dioxide are combined in a gas stream and the zinc vapour is separated from the sulphur dioxide by oxidation of the zinc to zinc oxide and separation of the solid zinc oxide from the gas stream.

6. A method according to any one of claims 1 to 4 wherein the zinc vapour and sulphur dioxide are combined in a gas stream, said method further comprising the steps of:

(6) controlling the oxygen activity in and temperature of the gas stream so as to inhibit reaction of the zinc vapour with the sulphur dioxide,

(7) then shock cooling the gas stream to below the condensation temperature of the zinc whereby to condense the zinc to zinc metal, and

(8) separating the condensed zinc from the remainder of the gas stream.

7. A method according to any one of the preceding claims wherein the composition is a concentrate containing zinc sulphide and lead sulphide, said method comprising the steps of:

(i) establishing a molten slag bath containing ferrous and ferric oxides at a bath temperature of from 1140°C to 1300°C, (ii) adding said composition to the bath whereby to dissolve the zinc sulphide and lead sulphide,

(iii)oxidizing the dissolved zinc sulphide and lead sulphide by reaction with ferric oxides in the bath to produce zinc metal vapour, lead metal vapour, sulphur dioxide gas and ferric oxide,

(iv) allowing the zinc metal vapour and lead metal vapour with the sulphur dioxide to leave the bath and then separating the zinc and lead from the sulphur dioxide gas.

8. A method according to Claim 7 wherein the lead vapour and zinc vapour are separated from the sulphur doxide by oxidation followed by separation of the zinc oxide and lead oxide from the gas stream.

9. A method according to Claim 7 wherein the zinc vapour, lead vapour and sulphur dioxide are combined in a gas stream; said method comprising the steps of:

- controlling the oxygen activity and temperature of the gas stream so as to inhibit reaction of the zinc vapour and lead vapour with the sulphur dioxide, and then shock cooling the gas stream to below the condensation temperature of zinc whereby to condense both the zinc and the lead, and separating the condensed zinc- and lead from the gas stream.

10. A method according to any one of the preceding claims wherein coal is added, to the bath surface.

11. A method according to any one of the preceding claims wherein the ratio of added fixed carbon to zinc is greater than 25g/mole zinc.

12. A method according to any one of the preceding claims wherein the ratio of fixed carbon to zinc is greater than 40g/mole zinc.

13. A method according to any one of the preceding claims wherein the bath is maintained in a state of vigorous agitation.

14. A method according to anyone of the preceding claims further including the addition of limestone to the concentrate or furnace.

15. A method according to Claim 4 wherein the coal is a low volatile coal.

16. A method according to Claim 6 or Claim 9 wherein the temperature is controlled at above 960°C prior to shock coolin and the vapour is shock cooled to below 500°C.

17. A method according to Claim 16 where the zinc metal vapour is shock cooled to below the melting point of zinc metal.

18. A method according to Claim 17 where the shock cooling to below the melting point of zinc metal takes place in less than 100 milliseconds.

19. A method according to any one of the preceding claims wherein the oxygen activity above the bath is controlled independently of the oxygen activity in the bath.

20. A method substantially as herein described with reference to any one of the examples.

21. Apparatus for smelting zinc comprising: a refractory lined furnace adapted to contain a molten slag bath at above 1140°C, the bath including ferric and ferrous iron; means for feeding zinc concentrate and carbon to the slag bath; means for injecting an oxygen containing gas at below the surface of the molten slag bath; a condenser for rapidly chilling a gas stream containing zinc vapour eminating from the bath to below the melting point of zinc, and means to vary the rate of injection of oxygen and/or of addition of carbon so as to maintain the ratio of ferric to ferrous iron within a predetermined range.

22. Apparatus according to Claim 21 wherein the gas stream which enters the condenser is cooled at a sufficiently rapid rate to condense the zinc substantially as zinc metal without reversion.

23. Apparatus according to Claim 21 or 22 wherein the condenser comprises means for bringing the gas stream into direct contact with a fluidized bed having a high thermal mass in comparison with the gas stream in heat exchange relationship with the bed; the bed containing an inventory of cool particles sufficient to condense the zinc vapour.

24. Apparatus according to any one of claims 21 to 23 further including means to control the temperature of the gas stream after leaving the bath and prior to entering the condenser at a temperature equal to or greater than the bath temperature.

25. Apparatus according to Claim 24 comprising at least one burner in the gas stream above the bath.

26. Apparatus according to any one of claims 21 to 25 wherein the means for feeding carbon adds the carbon onto the bath surface.

27. Apparatus according to Claim 23 wherein the bed is a recirculating fluidized bed.

28. Apparatus according to Claim 27 wherein the hot gas stream is brought into contact with the fluidized bed particles in one chamber and the fluidized bed particles are cooled in a second chamber prior to being recirculated into the gas stream.

29. Apparatus according to any one of claims 21 to 29 wherein the means for feeding zinc and coal are adapted for continuous operation.

30. Apparatus substantially as herein described with reference to any one of the drawings.