Module - 6

Module - 6

Learning Objectives

• Complex nature of sulphides and sulphide metallurgy

• Pyrometallurgical extraction process for copper, zinc, lead, nickel etc.

• Hydrometallurgical extraction processes

• Concept of Process Fuel Equivalent ( PFE)

Methods of metal recovery

• Thermal decomposition• Roasting and subsequent reduction• Controlled roasting matte metal• Flash smelting matte metal• Metallothermic reduction of sulphide• Hydrometallurgical processing• Chlorination chlorides metal• Electrolytic refining of matte to pure metal

(Commercial process for Ni only)

Extraction of copper

Things to Note During roasting iron sulphides are first oxidized in preference to copper sulphides.If Fe2O3 forms then this cannot be slagged. Hence roasting is left incomplete, no copper oxide is formed. Even some FeS is left behind to ensure that Fe3O4 and Fe2O3 are not formed.

During smelting, residual Fe2O3, if any, is reduced.10Fe2O3 + FeS 7Fe3O4 +SO2

3Fe3O4 + FeS 10FeO +SO2

The aim is to produce only a mixture of copper and iron sulphides. FeO is removed in slag.During converting copper metal is produced only through oxidation of sulphides.Cu2S(l) + 2Cu2O (l) 6Cu(l) +SO2 (g)

Roasting reactions (~5500 C) 2CuFeS2 + 6.5O2 2CuO + Fe2O3 + 4SO2

CuFeS2 + 4O2 CuSO4 + FeSO4

2CuFeS2 + O2 Cu2S +2FeS+SO2

2CuFeS2 + 4.5O2 Cu2S + Fe2O3 + 2SO2

2CuFeS2 + 7.5O2 CuO.CuSO4 +2FeSO4 + 2FeSO4 +SO2

3CuFeS2 +9.5O2 3Cu2O + Fe3O4 + 6SO2

6CuFeS2 +13O2 3Cu2S +2Fe3O4 + 9SO2

Smelting Reactions ( ~12500 C) with fluxes Product 2 layers ( clear cut separation)

Upper Slag Layer Lower ‘Matte’ i.e. metallic sulphides

6CuO+4FeS 3Cu2S+4FeO+SO2

2CuSO4+2FeS Cu2S+2FeO+3SO2

Everything is moltenCu2O + FeS Cu2S + FeO10Fe2O3+FeS 7Fe3O4+SO2

3Fe3O4+FeS 10FeO+SO2

Converting (In side blown converters)

To remove Fe, S and other impurities from matteSlagging – Exothermic

2FeS+3O2 2FeO+2SO2

2FeO+SiO2 2FeO,SiO2(Fayalite)Slag – 1-5% (CaO + MgO), Fe – 40–50%

Cu – 2–9%, SiO2 – 20–30%Blister formation

2Cu2S (l) + 3O2 (g) 2Cu2O + 2SO2 (g)

Cu2S (l) + 2Cu2O (l) 6Cu(l) + SO2 (g)

Overall: 3Cu2S + 3O2 6Cu + 3 SO2

Conventionally, a smelting operation is carried out in reverberatory furnaces fired with either coal or oil. A typical reverberatory furnace is shown in Fig. Smelting has also been carried out in electric furnaces. A typical electric furnace is shown in Fig. An electric furnace is more advantageous than a reverberatory furnace if hydroelectric power is available freely and inexpensively because the generation of a large volume of combustion gases is avoided. This facilitates both the recovery of SO2 and the cleaning of the furnace gases, which is generally carried out by an electrostatic precipitator in order to recover the copper-bearing dust. However, an electric furnace consumes a large amount of energy, when fossil fuel is burnt especially to generate electricity. It has now given way to more energy-efficient process, namely, flash smelting and continuous smelting, which are described later .

In the early days of copper concentrate smelting, the average capacity of the reverberatory furnace was 100 tons per day, but, at present, the reverberatory or electric furnace can smelt over 1000tons a day.

Converting

The purpose of converting is to remove iron, sulphur and other impurities from matte. For this, the molten matte produced as a result of smelting is charged into a side-blown converter which is a cylindrical vessel with a capacity of 100-200 tons of matte. A typical vessel is 4 m in diameter and 9 m in length and is lined with a layer of chrome-magnesite refractory (about 40 cm thick). A typical side-blown converter is shown in Fig.

In the converter, the atmosphere is highly oxidizing compared with the neutral or mildly oxidizing atmosphere during smelting. Air or oxygen-enriched air (up to a maximum limit of 32 vol % oxygen in the blast) is injected into the molten matte through tuyeres. Each tuyere is about 5 cm in diameter, and there are about 40 tuyeres in a converter. The total volume of gas flowing through these tuyeres is about 600 m3 / min. The products of the converter are slag and blister copper.

The relative volumes of the two layers can be determined by the lever rule. When the sulphur level eventually level reaches 1.2 per cent, only the metallic copper phase remains. At this stage, care to be exercised to ensure that the metal is not overixidised to Cu2O.

The completion of blow can be determined by casting a small sample of the copper and examining the fracture of this sample. The blistery appearance of this sample lends the name blister copper to this product. In Industrial practice, the blister produced contains 0.02 – 0.05 per cent S along with 0.2 – 0.5 per cent dissolved oxygen.

Attempts made in the early days to produce blister copper in the bottom-blown converter used in the steel industry ended in failure. This is because , after point b is reached, a layer of copper rich liquid will be formed at the bottom in contact with the tuyers. There will be very little heat generation due to Cu2S oxidation in this layer, though heat is generated due to oxidation of copper.4Cu+02 = 2 Cu2OAlthough there is not much difference in the heat generated per mole of oxygen for copper oxidation and that for Cu2S oxidation, the efficiency of copper oxidation is much lower. Consequently, the temperature drops rapidly in the tuyer region of the Bessemer converter. This leads to the clogging of the tuyers with solid and the stoppage of the conversion of the matte to metal. It was only after the introduction of the side-blown Bessemer converter that the conversion of white metal to blister copper in a converter became possible.

The oxidation taking place during roasting and that resulting from the leakage of air into the furnace during smelting determine the extent of oxidation of the iron sulphide in the charge to the slag. Generally , the object is to produce a matt that contains 35-45 per cent Cu, 20-22 per cent S, and 25-35 per cent Fe. This not only minimizes the loss of copper to the slag but also provides a matt with a sufficient quantity of iron sulphide for use in the next stage, i.e. converting, where iron sulphide oxidation provides all the heat required to ensure an autogenous converting operation. The relationship between the percentage Cu in the slag and that in the matte is given in Fig. for various reverberatory furnace operations around the world.

Fire refining of blister copper

Sulphur is removed from liquid copper by slow oxidation.

Residual oxygen is eliminated by hydrocarbons using poling i.e. stirring by green tree branches. These are done in 400 t. capacity reverberatory furnaces over 12-16 hours, door kept open to mild air blast

Slow oxidation removes S, Fe, Se and Zn, solid oxides being skimmed off poling is done at last. Though the method is crude it is still the most common method. Finally Copper produced is 99.7 per cent pure.

Electrolytic refining is done in concrete or wooden tanks (B5 x 1.1 x 1 m) using 250-320 Kg copper anode Electrolyte is copper sulphate (35 g/L) in H2SO4 (2009 g/l) with some additives (glue, alcohol)

Newer Process

Flash smelting

(combines roasting and

smelting in one unit)

Produces Matte

WORCA Process: Derived by combining the first three alphabets of invertor’s name (H.K. Worner) and CRA (Conzino Riotinto of Australia Ltd.)

The process seeks to maximize the conservation of energy obtained from smeling and converting by integrating to a high degree continuous unit operations.

Process Fuel Equivalent (PFE)

The concept was first introduced by Kellogg (1974) while comparing energy consumption of different copper smelting processes.

PFE = F + E + S – B

F = Quantity of fuel directly consumed by process

E = The fuel equivalent of electricity (2650 kcal / kwh) which is the normal energy need to generate power from fossil fuel

S = The fuel equivalent of major supplies used in smelting as reagents, oxygen and flukes

B = The fuel equivalent of by products and useful surplus heat

Hydrometallurgy of copper

~ 85 per cent of world copper production is by pyrometallurgy. Hydrometallurgy can be employed for oxidized ores or low grade sulphide ores.

Ferric chloride is an ideal leaching agent.

CuFeS2 + 3FeCl CuCl + 4FeCl2 + 2S

Not attractive commercially because energy requirment is high compared to pyrometallurgy.

Fig. Ferric Chloride Leaching of Copper Concentrate

Sulphation roasting

Figure shows stability regions in the Cu-O-S system. In a fluid bed roaster a roasting temperature higher than 6500C is required to achieve good roasting kinetics. 7000C is optimum when partial pressure of SO2 lies between 0.04 – 0.09 and Po2 lies between 0.04 – 0.10

During the subsequent electrowinning of CuSO4 solution (obtained from the dilute acid leaching of the roasted concentrate), H2SO4 is generated. The reaction is

Cu2+SO42-+H2O Cu + H2SO4+1/2O2

The sulphuric acid in the spent electrolyte is normally used for the vat leaching of copper oxide ores. If such copper ores are not available, H2SO4 is neutralized with either lime or limestone and rejected as gypsum.

Extraction of LeadUses : Antiknock compounds ( ~11%), batteries ( ~35%), sheets and pipes ( ~11%), sheathing cables ( ~17%), Miscellaneous ( rest)Common Ores : Galena (PbS) associated with ZnS, FeS, CuS , PbCO3 and several precious metalsProcess : Roast below 8000 C ( to avoid fusion) in Dwight- Lloyd sintering machine - sinters for blast furnace smelting using fluxes ( Limestone + Quartz0PbS + 3/2 O2 = PbO + SO2

SiO2 + 2PbO = 2PbO.SiO2

2PbSO4 + SiO2 = 2 PbO.SiO2 +2SO2

In lead blast furnace ( Temp <12000 C)PbO + C = Pb + CO2PbO + C = Pb + CO2

Scrap Iron is also charged in blast furnacePbSiO3 + Fe = FeSiO2 +PbPbO + Fe = FeO + Pb

In the blast furnace smelting produces four distinct layers First : Slag ( Sp. Gr ~ 8.6)Second - Matte containing copper ( Sp. Gr~5) and other elementsThird – Spiss FeAs4 + other impurities ( Sp. Gr ~ 6)Fourth – Lead ( Sp Gr ~11) called base bullionBase bullion contains many impurities which must be eliminated systematically.

Modern Developments in Lead SmeltingAt present, about 90 per cent of the world’s primary lead is provided by the conventional process, i.e. sintering followed by blast furnace reduction, and about 10 per cent by the Imperial Smelting process. Recent environment protection laws, which seek to control the emission of lead fumes into the atmosphere , coupled with a shortage of energy needed for lead production have necessitated the development of processes that are both cleaner and less energy consuming. In this regard, the one-step(continuous) smelting process appears attractive. Here PbS is oxidized to yield lead according to the reaction.PbS +O2 --- Pb + SO2

Figure shows the equilibrium phases in the Pb-S-O system at 12000C. At this temperature , metallic lead, with 1-3 per cent sulphur, can be produced in presence of pure SO2.

Extraction of ZincUses : Maximum use in protective coating ( galvanizing) because it forms and impervious ZnCO3 layer . Zinc protects irons by being more electropositive. Used widely as Cu-Zn brasses Zn is attractive because of low m.pt, high structural strength, good dimensional stability.Ideal for die castingRolled zinc plates are used in dry cell batteries.Also used in paints and pigments

Several methods of Production

Sources Sphalerite(ZnS), Zincite(ZnO), also othersLow grade ores need concentration. ZnS melts at 15000C. However concentrate ( 55 % Zn) can be roasted at 8000 C, then ground, agglomerated, sintered to provide feed for retorts. ZnO is reduced ~12000 C, Zinc distils off for collection in condensers.

During roasting one must avoid formation of ferrite ZnO. Fe2O3 which is not leachable ( but easily reducible)Since Zn is high is electrochemical series the electrolytic solution must be free of Cd, Pb, Cu, Fe, Ag and Ge or else Zn deposited on cathode will be contaminated. Hence purification is necessary prior to electrolysis. For Cu purification is done afterwards.

Important Points

Imperial Smelting Process (ISP)

• Simultaneously smelt low-grade complex mixed charges of Zn and Pb ores and concentrates to recover both Zn and Pb

• Overall thermal efficiency is higher, Zn recovery more economical

• No electricity is required• Variety of furnace sizes available• Furnace operation is continuous and fully automated• Mechanism is robust – can stand shut downs and

restarts

Advantages

This process involves the following basic steps :(1) The treatment of ores to obtain a concentrate rich in zinc.(2) The roasting of the concentrate to convert the zinc into a soluble form(3) The treatment of the roasted concentrate to form a zinc sulphate solution(4) The purification of the zinc sulphate solution by precipitation of impurities(5) The removal of zinc from purified solution by electrolysis(6) The melting of zinc sheets to form ingots

Extraction of NickelBoth oxidic and sulphidic ores are available. Ni is important as an alloying element – alloys in chemical processing, space research, nuclear reactor engineering. More than 3000 commercial alloys for mechanical properties and corrosion resistance. Ni-Cu alloys are known as Monel metals, Ni-Cu-Zn alloys make German silver.

Sulphide ores – Ni2FeS4, ( NiFe)9S3 etc.

Nickel from Oxide Ores

More abundant than sulphudic ores. Ores are Nickeliferrous laterites typically ~1% Ni and 40-50% Fe in ‘overburden’ , more Ni ( ~1.5%) in layers below in limonite or goethite, ( FeNi) O, (OH).nH2O. Ni in solid solution with FeO.In India this is there in chromite mines of Sukinda ( Orissa)Usual Procedure : Dry, reduce around 800-10000C, reduce Ni – Leach ( acid or ammonia) - treat by solvent extraction -- ppt. NiCO3--Ni or else leach – electrolyzeAt no stage should ore be heated beyond 10000C.

By keeping controlled temperature and pco one produces Ni and FeO ( easy to leach)In reduction smelting calcined ores are smelted in electric furnace ( 1550-16500C) with carbon to produce Ferro-nickel ( 25-40% Ni) Gangue silicates form a slag.One has to strike a balance between % Ni and total Nickel recovery.There is often Co( ~ 1/10 – 1/20 of Ni) associated with Ni