Question 3 of 7: Pyrometallurgical Processes (20 marks)

Question 3 — Pyrometallurgical Processes (20 marks)

Question text not reproduced: the examination questions are © Engineers and Geoscientists BC. Open the official past paper (linked at the top of this page) to read the question, then follow the worked solution below.

(a) Carbothermic reduction. Carbothermic reduction uses carbon, or the carbon monoxide it generates in situ, as the reducing agent to strip oxygen from a metal oxide and liberate the metal, exploiting the fact that carbon's oxidation to CO becomes thermodynamically more favourable (more negative ΔG) as temperature rises, eventually crossing below the oxide's own line on an Ellingham diagram. Tin smelting is the textbook direct-carbon example: $$\mathrm{SnO_2(s) + 2\,C(s) \longrightarrow Sn(l) + 2\,CO(g)}$$ Blast-furnace ironmaking is the indirect (CO-mediated) case that dominates industrial tonnage, where coke first forms CO and CO then does the reducing: $\mathrm{Fe_2O_3+3\,CO\rightarrow2\,Fe+3\,CO_2}$, with the CO₂ regenerated back to CO by the Boudouard reaction $\mathrm{CO_2+C\rightarrow2\,CO}$ on the descending coke.

(b) Zinc fuming. Zinc fuming recovers zinc left in a smelting slag (from copper, lead or tin operations) by injecting pulverised coal or coke and air/oxygen into molten slag held at 1200–1300 °C. The carbon reduces the slag's ZnO to metallic zinc vapour, which is volatile at furnace temperature and leaves the melt as gas rather than remaining dissolved in the (now zinc-depleted) discard slag: $$\mathrm{ZnO(slag) + C(s) \longrightarrow Zn(g) + CO(g)}$$ The zinc vapour is then deliberately reoxidised in the furnace freeboard with excess air, $\mathrm{Zn(g)+\tfrac12\,O_2\rightarrow ZnO(s)}$, and the resulting fume is captured in a baghouse as a saleable zinc-oxide product. Fuming is what makes it economic to discard a slag rather than retreat it in a zinc circuit.

(c) Magnetizing roast. A magnetizing roast is a mildly reducing (or reduction/reoxidation-controlled) roast that converts weakly magnetic haematite, Fe₂O₃, into strongly magnetic magnetite, Fe₃O₄, without removing further oxygen to metallic iron: $$\mathrm{3\,Fe_2O_3 + CO \longrightarrow 2\,Fe_3O_4 + CO_2}$$ The purpose is purely physical separation downstream: once haematite ore or a haematite-bearing tailing is converted to magnetite, it can be upgraded on an ordinary low-intensity magnetic separator, which is far cheaper per tonne than gravity or flotation concentration of fine, weakly magnetic haematite.

(d) Sulfating roast. A sulfating roast is run at lower temperature and higher oxygen partial pressure than a dead (oxidising) roast so that the sulphur is retained in the calcine as a water-soluble metal sulphate rather than driven off entirely as SO₂, preparing the material for a subsequent aqueous leach: $$\mathrm{ZnS + 2\,O_2 \longrightarrow ZnSO_4}$$ Copper and cobalt sulphide concentrates are sulfate-roasted for the same reason: the sulphate is soluble in dilute sulphuric acid, so the roast becomes the front end of a leach–solvent-extraction–electrowinning (SX–EW) flowsheet rather than a smelting one.

(e) Chloridizing roast. A chloridizing roast is conducted with a chloride source present — common salt (NaCl), CaCl₂ or chlorine gas — so the metal reports as a chloride, which may be leachable or, for some metals, volatile enough to separate by condensation from the off-gas. The classical salt roast of a lead sulphide converts it to leachable lead chloride while fixing the sulphur as sodium sulphate: $$\mathrm{PbS + 2\,NaCl + 2\,O_2 \longrightarrow PbCl_2 + Na_2SO_4}$$ Chloridizing roasting historically underpinned silver recovery from complex sulphide ores (the "patio" and related processes) and remains relevant for treating refractory concentrates and recovering non-ferrous values from flue dusts and residues; its main drawback is the corrosivity of the chloride-bearing off-gas toward refractory brick and ductwork.