1 Purpose of Concentration

Raw ore contains only a small percentage of the desired metal (e.g., Cu ores carry 0.5–2% Cu). Ore dressing (beneficiation) removes gangue and concentrates the valuable mineral, typically achieving 15–35% metal content in the concentrate before smelting.

2 Gravity Separation (Hydraulic Washing)

Exploits the density difference between the heavy ore mineral and the lighter gangue. Powdered ore is agitated with water or washed by a strong current. Heavier ore particles settle in grooves while lighter gangue is washed away.

Applications: Heavier oxide ores — hematite (Fe₂O₃), tinstone (SnO₂), and gold (Au).

Fig 1.3 Gravity separation (Hydraulic washing) — heavier ore settles in grooves; lighter gangue particles are washed away by the water current.

3 Magnetic Separation

Used when either the ore or its gangue is magnetic. Powdered ore passes over a conveyor belt; one roll is fitted with an electromagnet. Magnetic particles are attracted and fall separately from non-magnetic material.

Example: Tinstone (SnO₂) is non-magnetic but contains magnetic impurities — iron tungstate (FeWO₄) and manganese tungstate (MnWO₄) — which are removed magnetically.

Fig 1.4 Magnetic Separation — the electromagnetic roller attracts magnetic impurities, deflecting them into a separate collection bin while non-magnetic ore continues on the belt.

4 Froth Flotation

The most important concentration method for sulphide ores (galena PbS, sphalerite ZnS, chalcopyrite CuFeS₂). Exploits differential wettability of ore and gangue surfaces:

• Collectors (e.g., xanthates): Adsorb on mineral surfaces, rendering them hydrophobic so they attach to air bubbles.
• Frothers (e.g., pine oil, MIBC): Stabilise the froth layer.
• Depressants / pH modifiers (e.g., CaO): Selectively depress unwanted minerals for differential flotation.

Fig 1.5 Froth Flotation — sulphide ore particles coated with collector (hydrophobic) attach to rising air bubbles and are carried to the froth; gangue (hydrophilic) sinks and is withdrawn as tailings.

5 Chemical (Leaching) Method

The ore is dissolved in a selective chemical reagent, leaving insoluble impurities behind. Metal is then recovered by precipitation or electrolysis.

Example — Bayer process (Al): Bauxite digested with NaOH at 150–220 °C. Al₂O₃ dissolves as NaAlO₂; Fe₂O₃, TiO₂, SiO₂ remain as red mud. Al(OH)₃ precipitated from the cooled liquor and calcined to Al₂O₃.

1 Calcination vs Roasting

Calcination: Heating in limited air below the melting point. Drives off moisture, water of hydration, and CO₂. Example: Al(OH)₃ → Al₂O₃ + H₂O; CaCO₃ → CaO + CO₂.

Roasting: Heating in excess air below fusion. Converts sulphide to oxide, expels SO₂, dries ore, removes volatile impurities. SO₂ produced is used to manufacture H₂SO₄.

2 Smelting & Flux Chemistry

Smelting reduces the oxide (or calcine) to metal using carbon or other reducing agents, with flux to remove gangue as slag. Two immiscible layers: denser metallic/matte layer below; lighter slag above.

• Basic flux (CaO): removes acidic gangue → CaO + SiO₂ → CaSiO₃ (slag).
• Acidic flux (SiO₂): removes basic impurities → SiO₂ + 2FeO → Fe₂SiO₄ (slag).

3 Reduction Methods

• Carbon reduction: ZnO + C → Zn + CO; SnO₂ + 2C → Sn + 2CO.
• Metallothermic (aluminothermy): Cr₂O₃ + 2Al → 2Cr + Al₂O₃ — highly exothermic; for Cr, Mn.
• Self-reduction: Cu₂S + 2Cu₂O → 6Cu + SO₂ — applied to Cu, Pb, Hg sulphide ores.
• Electrolytic reduction: Active metals (Al, Mg, Na) — fused-salt electrolysis.

4 Purification & Refining

Liquation: Crude metal heated just above M.P. on sloping hearth; pure metal drains, leaving infusible impurities. Used for Sn, Pb.

Fig 1.6 Liquation — impure metal is poured onto the sloping hearth of a reverberatory furnace and heated just above the melting point. Pure metal drains out; infusible impurities remain on the hearth.

Poling: Green wood is stirred through molten Cu or Sn; hydrocarbons reduce Cu₂O impurity back to Cu.

Fig 1.7 Poling — green wood or bamboo poles are stirred in molten copper or tin. The hydrocarbons in the pole reduce residual Cu₂O impurity back to metallic Cu. Used for Cu and Sn refining.

Distillation: Volatile metals (Zn, Hg) are purified by distillation; non-volatile impurities remain.

Electrolytic refining: Impure metal cast as anode; pure metal deposited on cathode in appropriate metal-salt electrolyte. Noble impurities collect as anode slime (Au, Ag, Pt). Purity: up to 99.99%.

Fig 1.8 Electrolytic Refining — impure metal (anode) dissolves into the electrolyte; pure metal deposits on the cathode. Less electropositive metals remain in solution; noble metals (Au, Ag) settle below the anode as "anode mud".

1 Properties & Minerals of Copper

At. No. 29 | At. Wt. 63.54 | Density 8.94 g/cm³ | M.P. 1083 °C | B.P. 2595 °C.

2 Concentration & Roasting

A natural Cu sulphide ore contains only 0.5–2% Cu. After crushing and grinding (~40 μm), froth flotation using CaO (pH control) and xanthate (collector) produces concentrate with 15–35% Cu, 15–35% Fe, 25–35% S, 3–15% gangue.

Roasting (if high FeS₂ content, <25% Cu): partial oxidation at ~550 °C in multiple-hearth or fluidised-bed roaster. SO₂ → H₂SO₄ plant.

Fig 2.1 Pyrometallurgical Extraction of Cu from Sulphide Ores — complete flowsheet showing the conventional route (left: roasting → smelting → converting → refining) and the newer route (right: flash smelting → continuous smelting).

3 Smelting to Matte

Calcine is smelted with siliceous flux at ~1250 °C. Two immiscible layers form: slag (SG 2.8–3.8, iron silicate, floats) and matte (SG 5.0–5.5, Cu₂S–FeS mixture, 35–45% Cu). Key exchange reaction: FeS + Cu₂O → Cu₂S + FeO (O₂ has greater affinity for Fe than Cu per Ellingham diagram).

Fig 2.2 Multiple Hearth Roaster — concentrate fed at top, moves downward through successively hotter hearths. Conventional roasting at 550 °C. SO₂ produced at ~5% concentration.

Fig 2.3 Fluidized Bed Roaster — concentrate particles fluidised by air blast at 550 °C. Autogeneous; produces rich SO₂ stream (8–15%) suitable for contact H₂SO₄ plant.

4 Converting

A side-blown Peirce-Smith converter (4 m dia., 9 m long, 100–200 t capacity) receives molten matte and blows O₂-enriched air through 40 tuyeres (5 cm dia.). Two stages:

• Slagging stage: FeS oxidised; FeO + SiO₂ → converter slag (2–9% Cu); skimmed off.
• Blister-formation stage: Cu₂S (white metal) oxidised → blister Cu (~98.5–99.5%) + SO₂. Cu₂S + O₂ → 2Cu + SO₂.

Fig 2.4 Converter Operation — showing the tilting converter with tuyeres submerged in the matte bath during the slagging stage. Slag (2–9% Cu, 40–50% Fe) is poured off by tilting.

Fig 2.5 Operating Principles of the Side-Blown (Peirce-Smith) Converter — showing the two-layer bath (matte + blister Cu) during the blister formation stage and the tuyere air injection.

5 Refining

Fire refining (reverberatory furnace, 400 t): oxidise S, Fe, Se, Zn → skim. Poling with green wood reduces Cu₂O. Purity: 99.97%.

Electrolytic refining: Crude Cu anode in CuSO₄/H₂SO₄ electrolyte at 50–60 °C. Cu deposits at cathode (99.99%); Fe, Co, Ni go into solution; Au, Ag, Pt, Se, Te → anode mud.

Zone Refining (for ultra-high purity): A narrow molten zone traverses a bar slowly; impurities concentrate in the melt and migrate to one end.

Fig 2.10 Zone Refining — a narrow molten zone is moved along the bar. Impurities preferentially dissolve in the liquid zone and are swept to one end, producing ultrapure metal at the other. Used for semiconductor metals and high-purity Cu.

1 Why Modern Processes?

Conventional reverberatory furnace smelting has low SO₂ concentration in off-gas, high fuel consumption, and batch operation. Modern processes use autogeneous (self-heating) reactions with O₂-enriched air, generating rich SO₂ gas and enabling continuous operation.

2 Flash Smelting (Outokumpu)

Dry concentrate is blown with preheated O₂-enriched air into a reaction shaft. Sulphide particles combust in suspension at ~1300 °C, releasing SO₂ and producing high-Cu matte (70% Cu) in a single step.

• Autogeneous: exothermic sulphide oxidation provides all process heat.
• High SO₂ (20–70%): ideal for H₂SO₄ production.
• Products: Matte (70% Cu, 8% Fe, 22% S) + slag (40% Fe).

Fig 2.6 Flash Smelting (Outokumpu) — dry concentrate + O₂-enriched air are injected through the concentrate burner into the reaction shaft where combustion produces matte and slag, collected in the settler. Rich SO₂ off-gas exits to the acid plant.

3 WORCRA Process

Combines smelting and converting in a single continuous reactor using counter-current movement of gas and concentrate. Features: directly produces blister Cu (not matte); no separate converter needed; continuous SO₂ extraction. Disadvantages: refractory lining not durable; high operating cost.

Fig 2.7 Vertical sectional diagram of the straight-form WORCRA reactor — concentrate enters one end and blister Cu is withdrawn continuously at the other. Counter-current gas and charge movement gives large reaction surface area and effective Fe removal.

4 Noranda Process

High-grade Cu matte or blister Cu forms directly from sulphide concentrate in a horizontal reactor. Air blown through tuyeres oxidises matte. Three layers: Cu (bottom), matte (middle), slag (top).

Fig 2.8 Schematic diagram of the Noranda Process — concentrate + air injected through tuyeres produce three distinct layers: Cu at the bottom, matte in the middle, and slag at the top. Slag contains higher % Cu than WORCRA and requires cleaning.

5 Mitsubishi Process

Three furnaces in cascade connected by gravity launders — smelting furnace → slag-cleaning furnace → converting furnace. Each stage operates continuously:

• Smelting furnace: wet concentrate + flux + O₂ → matte (60–65% Cu) + slag.
• Slag-cleaning furnace: slag gravity-separated; matte passes to converter.
• Converting furnace: matte + O₂-enriched air + limestone → blister Cu + lime-ferrite slag.

Fig 2.9 Mitsubishi Continuous Smelting Process — three-furnace cascade (S-furnace → CL-furnace → C-furnace) connected by gravity launders. Fully continuous; each furnace optimised for its specific reaction.

Fig 2.9(b) Mitsubishi Process — detail view showing the lance injection of O₂-enriched air and the gravity flow of matte between furnace stages.

1 Properties & Minerals of Zinc

At. No. 30 | At. Wt. 65.38 | Density 7.13 g/cm³ | M.P. 419.5 °C | B.P. 910 °C.

Minerals: Sphalerite (ZnS, zinc blende) — chief ore; Zincite (ZnO); Smithsonite (ZnCO₃); Willemite (Zn₂SiO₄); Franklinite. Indian deposits: Zawar mines (Rajasthan), HZL.

2 Roasting & Sintering

ZnS cannot be smelted directly (does not melt even at 1500 °C). Concentrate (~55% Zn) is roasted to ZnO in a fluidised-bed roaster (autogeneous, high SO₂). ZnO is sintered for handling. Hearth roaster used for initial volatilisation of Pb impurities.

3 Retort Processes & Electrolytic

Fig 2.11 Extraction of Zinc by Reduction of ZnO by Carbon — flowsheet showing roasting, sintering, and the three retort routes (horizontal, vertical, electrothermic) leading to crude zinc (spelter) and then fractional distillation for purification.

• Horizontal retort: 2 m long, 25 cm dia. clay retort; ZnO + coke breeze at 1400 °C; Zn volatilises and condenses. Low capacity, batch.
• Vertical retort: 10 m tall, 70 cm dia., SiC walls (5× conductivity of horizontal); 7–10 t Zn/day.
• Electrothermic: 15 m vertical retort, 100 t Zn/day; graphite electrodes; off-gas contains 40–45% Zn vapour.
• Electrolytic process: Highest purity (~99.95%); ZnO dissolved in H₂SO₄; electrolysed with Al cathodes and Pb-Ag anodes. Si rectifiers reduce power consumption.

All retort-derived Zn (spelter) further refined by fractional distillation: Zn B.P. 907 °C, Cd 780 °C, Pb 1620 °C — wide separation facilitates purification.

4 Imperial Smelting Process (ISP)

The ISP blast furnace handles mixed Zn–Pb concentrates simultaneously — unique process that recovers both metals. Preheated coke charged from top; sinter via double-bell system. ZnO + C → Zn vapour (at >1120 °C); condensed in molten-lead splash condenser. Pb bullion (70–80% Pb + Au, Cu, Sb) collected at hearth.

Fig 2.12 Imperial Smelting Blast Furnace — preheated coke and sinter charged from the top; hot blast through tuyeres; Zn vapour exits top to the lead splash condenser; Pb bullion and slag tapped at the hearth.

Fig 2.13 Zinc Recovery System in the ISP — Zn vapour from the blast furnace dissolves in the lead splash condenser; on cooling the Zn-rich lead is separated by gravity (Zn floats); Zn refined by fractional distillation to SHG (special high grade, 99.95%).

5 Slag Fuming & Uses

Slag fuming process (COMINCO): Slag from the lead blast furnace (15–18% Zn) is treated with pulverised coal + air through tuyeres. Zn is volatilised and oxidised to ZnO fume, collected in a bag house. Fume dispatched to electrolytic Zn plant.

Fig 2.14 Zinc Elimination during Slag Fuming — molten slag is treated with coal/air through tuyeres. Zn is volatilised (as ZnO fume), while the slag is solidified and discarded. The Zn fume is collected and sent to the electrolytic plant.

Uses of Zn: Galvanising steel (cathodic protection); die-casting alloys; Cu-Zn brasses (30–37% Zn); ZnO pigments; Zn spraying for corrosion protection.

1 Properties & Minerals of Lead

At. No. 82 | At. Wt. 207.21 | Density 11.34 g/cm³ | M.P. 327.4 °C | B.P. 1737 °C.

Minerals: Galena (PbS) — chief ore; Cerussite (PbCO₃); Anglesite (PbSO₄). Indian: HZL (Tundoo, Bihar) and Visakhapatnam (15,000 t/yr).

2 Roasting, Sintering & Blast Furnace

Galena concentrate is roasted to oxides then sintered in a Dwight-Lloyd sintering machine (S reduced from 16–18% to 1–2%). Up-draught sintering preferred (higher capacity, better SO₂ recovery).

Fig 2.15 Flow Sheet of Production of Lead Bullion — from galena ore through roasting, sintering, blast furnace reduction, and the series of melt-separation steps (slag, matte, speiss, lead bullion).

In the blast furnace, sinter is reduced with coke and siliceous flux. Four distinct layers form by specific gravity at the hearth:

• Slag (SG ~3.6): iron silicate — discarded.
• Matte (SG ~5.2): Cu₂S + PbS — returned for Cu recovery.
• Speiss (SG ~6): Fe-As-Ni alloy (when As + Fe → speiss); contains recoverable Cu.
• Lead bullion (SG ~10.8): crude Pb tapped continuously.

Fig 2.16 Molten Lead Blast Furnace — showing the four-layer separation (slag / matte / speiss / lead bullion) at the hearth. The siphon-top device maintains constant Pb level; bosh retards charge descent in front of tuyeres.

3 Refining of Lead Bullion

Lead bullion contains Sb, Sn, As, Bi, Cu, Ag, Au. Cascade refining:

• Liquation (drossing): Heated at 350 °C → Cu₂S dross skimmed; raised to 500–550 °C → Cu eliminated as Cu₂S.
• Softening: Sb, Sn, As oxidised and skimmed in a reverberatory furnace — gives soft, malleable Pb with improved corrosion resistance.

Fig 2.17 Lead-Zinc Phase Diagram — used to understand the Parkes desilverising process; Zn has limited solubility in Pb, and the Zn layer solidifies first on cooling to carry Ag and Au into the Zn-rich phase.

Fig 2.18 Refining of Lead Bullion — showing the sequence of steps: drossing → softening → Parkes desilverising → dezincing → debismuthising → pure lead.

4 Parkes Desilverising & Dezincing

Parkes Process: Zn added to molten Pb; Zn has greater affinity for Ag and Au than Pb. Zn layer (higher M.P.) solidifies first, carrying Ag/Au. Cupellation recovers Ag: Zn-Pb-Ag alloy heated in reverberatory furnace with air blast → Pb and Zn oxidise to litharge/ZnO; pure Ag remains in the cupel.

Dezincing: Residual Zn removed by: (a) Cl₂ treatment — ZnCl₂ skimmed off; (b) Vacuum dezincing at 540 °C — Zn reduced from 0.6% to 0.002% in Pb.

Fig 2.19 Dezincing Process — Cl₂ is blown into an enclosed chamber containing molten desilverised lead. Zn reacts to form ZnCl₂ scum which is skimmed off. At the end point, white PbCl₂ crystals appear on the surface signalling complete Zn removal.

Debismuthising: Ca and Mg added to form insoluble Ca-Mg-Bi intermetallic compound, skimmed off.

5 Electrolytic Refining

Pb bullion anode in H₂SiF₆ + PbSiF₆ electrolyte. Pure Pb deposits at cathode; Bi and precious metals → anode slime. Purity: 99.99%.

Uses of Pb: Lead-acid batteries; cable sheathing; Pb pigments (white lead 2PbCO₃·Pb(OH)₂, red lead Pb₃O₄); radiation shielding; Pb-Sb alloys for bearings.

1 Properties & Minerals of Nickel

At. No. 28 | At. Wt. 58.69 | Density 8.98 g/cm³ | M.P. 1452 °C | B.P. 2900 °C.

Minerals: Pentlandite [(NiFe)₉S₈] — chief ore; Violarite [Ni₂FeS₄]. Ni occurs alongside Cu in sulphide ores. Sudbury (Canada) INCO — 80% world Ni. Indian deposits: Sukinada (Odisha), 100 million tons at 0.85% Ni.

2 INCO Process — Extraction Flow

The ore is a mixed Cu-Ni sulphide with nearly equal Cu and Ni:

• Grinding + differential froth flotation → Cu concentrate, Ni concentrate, pyrrhotite concentrate.
• Roasting (fluidised bed, 550–600 °C) → 40% S oxidised; autogeneous; high SO₂ output.
• Reverberatory smelting + siliceous flux → matte (20% Cu, 7% Ni) + slag.
• Pierce-Smith converting (1150 °C) → Ni-enriched matte (50% Ni, 25% Cu, 21.5% S).
• Slow cooling (melt → 400 °C over 3 days) → Three phases: Cu₂S (top), Cu-Ni alloy (700 °C), Ni₃S₂ (575 °C).
• Magnetic separation + flotation → Cu₂S and Ni sulphide streams separately.

Fig 2.20 Nickel Extraction — INCO Process flowsheet showing the complete route from Cu-Ni sulphide ore through flotation, smelting, converting, slow cooling, magnetic separation, and the split into Cu₂S and Ni sulphide refining streams.

3 Carbonyl Refining — Mond's Process

At 40–90 °C, metallic Ni combines with CO to form gaseous nickel carbonyl [Ni(CO)₄]. At 150–300 °C, Ni(CO)₄ decomposes to give pure Ni metal and CO (recycled).

Ni + 4CO → Ni(CO)₄  |  Ni(CO)₄ → Ni + 4CO

Fe(CO)₅ and Co₂(CO)₈ are also volatile but separated by fractional condensation. Cu does not form a volatile carbonyl. INCO Pressure Carbonylation: at 180 °C and 70 atm, carbonyls of Ni, Fe, and Co form; Ni(CO)₄ recovered by fractional distillation → Ni pellets or powder. Final purity: 99.97% Ni.

4 Electrolytic Refining

Ni oxide reduced by coke in electric furnace → cast as anode. Electrolyte: 60 g/L Ni²⁺, 95 g/L SO₄²⁻, 35 g/L Na⁺, 55 g/L Cl⁻, 16 g/L H₃BO₃ at 60 °C.

Fig 2.21 Electrolytic Refining of Ni — Ni oxide anodes dissolve into the NiSO₄-rich electrolyte; pure Ni deposits at the cathode. Cu removed by cementation with active Ni powder; Fe and others removed by aeration; Co by chlorine oxidation. Electronickel purity: 99.93–99.95%.

Uses of Ni: Ni-Cu alloys (Monel) for chemical/oil industry; German silver (Ni-Cu-Zn); stainless steel; nuclear reactors; Ni-based superalloys for jet engines.

1 Properties & Minerals of Magnesium

At. No. 12 | At. Wt. 24 | Density 1.738 g/cm³ | M.P. 650 °C | B.P. 1107 °C. The lightest structural metal.

Minerals: Dolomite (MgCO₃·CaCO₃); Magnesite (MgCO₃); Brucite [Mg(OH)₂]; Carnallite (MgCl₂·KCl·6H₂O); sea water (MgCl₂, MgSO₄). Indian: NML Jamshedpur; CECRI Karaikudi (Tamil Nadu).

2 Dow Process (Sea Water)

Sea water (~1.3 kg/m³ Mg as MgCl₂): Sea water + CaO → Mg(OH)₂ ↓ + Ca²⁺. Mg(OH)₂ filtered; dissolved in 10% HCl → MgCl₂. Concentrated and dried → anhydrous MgCl₂. Electrolysed → Mg (cathode) + Cl₂ (anode).

Fig 2.22 Production of Mg from Sea Water (Dow Process) — sea water is treated with lime to precipitate Mg(OH)₂, which is converted to anhydrous MgCl₂ by treatment with HCl and drying. The MgCl₂ is then electrolysed in fused-salt cells.

3 Electrolysis of MgCl₂

Aqueous electrolysis would evolve H₂ before Mg deposits — fused-salt electrolysis required. Bath: MgCl₂ + NaCl + KCl + CaCl₂ flux (reduces viscosity, increases conductivity). Cathode: steel; Anode: graphite. Mg (SG 1.74) floats on denser salt bath. Cl₂ collected in hood; recycled to dry incoming MgCl₂.

Fig 2.23 Electrolytic Cell for Mg Production — graphite anodes and steel cathodes in a fused MgCl₂ salt bath. Mg metal floats on the bath surface (SG 1.74 < salt bath SG ~1.8); Cl₂ is collected under a hood and recycled.

4 Pidgeon Process (Silicothermic)

Calcined dolomite (CaO·MgO) + ferrosilicon (75% Si) in briquettes, heated at 1100–1200 °C under vacuum (0.1 mmHg) in a Nichrome-steel retort:

2(CaO·MgO) + Si(Fe) → Ca₂SiO₄ + 2Mg↑ (vapour)

Mg vapour condenses outside the retort. 120 t dolomite + 12 t ferro-Si → 10 t Mg (endothermic — external heat required). Small CaF₂ or MgF₂ addition acts as a catalyst.

Fig 2.24 Retort for the Pidgeon Process — a 25 cm dia., 3 m long horizontal Nichrome-steel retort heated externally by gas or electricity. Briquettes of ferro-Si + calcined dolomite (1:6 ratio) are charged; Mg vapour condenses in the cooler outer portion of the retort.

5 Magnotherm Process & Uses

Similar to Pidgeon but operates at ~1500 °C with the bath in molten state, stabilised by Al₂O₃ addition to form molten slag. Higher throughput but more energy-intensive.

Uses of Mg: Structural — automotive (lightest structural metal, ⅓ weight of Al); non-structural — alloying element in Al alloys; desulphurisation and deoxidation of molten metal; Kroll process reductant for Ti; incendiary flares; die-casting alloys.

1 Properties & Minerals of Aluminium

At. No. 13 | At. Wt. 27 | Density 2.702 g/cm³ | M.P. 660.4 °C | B.P. 2467 °C.

Minerals: Bauxite (mixture of gibbsite Al₂O₃·3H₂O and diaspore Al₂O₃·H₂O) — chief ore; contaminated with Fe₂O₃, TiO₂, SiO₂. India: HINDALCO, NALCO, ALIND, MALCO. East Coast bauxite — among world's largest deposits.

2 Bayer's Process — Alumina Production

Bauxite crushed and ground with NaOH in ball mills → digestion at 150–220 °C and 5–25 atm:

Al₂O₃·3H₂O + 2NaOH → 2NaAlO₂ + 4H₂O (gibbsite)  |  Al₂O₃·H₂O + 2NaOH → 2NaAlO₂ + 2H₂O (diaspore)

Fe₂O₃, TiO₂, SiO₂ remain insoluble → red mud settled/filtered. Clarified liquor cooled → Al(OH)₃ precipitated by seeding. Calcined in rotary kiln at 1400 °C → anhydrous Al₂O₃.

Fig 2.25 Flow Sheet of Bayer's Process — showing bauxite crushing, ball-mill grinding with NaOH, pressure digestion (autoclave), red mud separation, clarification, seeded precipitation of Al(OH)₃, thickening, and calcination in a rotary kiln to produce Al₂O₃.

3 Hall-Héroult Process

Al₂O₃ dissolved in molten cryolite (Na₃AlF₆) at ~980 °C (up to 15% Al₂O₃ dissolves). Additives: CaF₂ + NaF improve conductivity. Cell: rectangular steel box (5 m × 2 m × 1 m), C lining as cathode, consumable graphite anodes. 5–7 V DC, 100–300 kA. Energy: 13–15 kWh/kg Al.

Cathode reaction: Al³⁺ + 3e⁻ → Al (liquid, SG 2.3 — sinks below cryolite bath SG 2.1). Anode reaction: 2O²⁻ → O₂ + 4e⁻; O₂ + C → CO₂ (graphite consumed at ~0.5 kg C/kg Al).

Fig 2.26 Synthesis of Cryolite (Na₃AlF₆) — HF reacts with sodium aluminate solution to produce synthetic cryolite. CaF₂ + H₂SO₄ → HF + CaSO₄; HF + NaAlO₂ + H₂O → Na₃AlF₆ + Al(OH)₃. Cryolite bath conductivity enhanced by CaF₂ and NaF additions.

Fig 2.27 Electrolytic Reduction Cell for Aluminium — rectangular refractory-lined steel box (cathode); consumable graphite anodes hang into the cryolite-Al₂O₃ bath at 980 °C; liquid Al collects at the bottom and is tapped periodically; operating voltage 5–7 V.

4 Anode Effect

When Al₂O₃ content drops below 2%, the anode-bath contact is disrupted by a gas film → cell voltage spikes (10–40 V), current drops, normal operation ceases — called the anode effect. Corrected by vigorous agitation and immediate Al₂O₃ addition.

Causes: Thermal (local overheating → gas film) or electrostatic (charged gas bubbles form adherent film). Factors: Higher temperature, greater electrolyte purity, and graphite anodes reduce critical current density.

5 Hoopes Electro-refining & Alternative Processes

Hoopes three-layer process refines Al to 99.99% based on density stratification:

• Bottom (anode): impure Al-Cu alloy (SG > middle) — anode dissolves.
• Middle (electrolyte): AlF₃ (36%) + BaF₂ (18%) + CaF₂ (16%) — intermediate SG.
• Top (cathode): pure Al (SG 2.3) — deposits as cathode product.

Fig 2.28 Hoopes Three-Layer Electro-refining Process — impure Al-Cu anode alloy (bottom), fluoride electrolyte (middle), and pure Al cathode (top) stratify by density. Current causes impure Al to dissolve at the bottom and pure Al to deposit at the top. Purity: 99.99%.

ALCOA Process: Al₂O₃ + 2Cl₂ + 3C → 2AlCl₃ + 3CO (chlorination at 700–900 °C) → AlCl₃ electrolysed in fused AlCl₃-NaCl-LiCl at 700 °C → Al (cathode) + Cl₂ (anode, recycled).

Fig 2.29 Flow Sheet of the ALCOA Process — Al₂O₃ from Bayer process is chlorinated with C at 700–900 °C to produce AlCl₃ vapour, condensed as solid, then electrolysed in a fused-chloride cell to give liquid Al and Cl₂ (recycled to chlorinator).

Fig 2.30 ALCOA Bipolar Electrolytic Cell — multiple bipolar electrodes reduce the total cell voltage and energy consumption. Al is collected at the bottom cathode; Cl₂ at the top anode is recycled.

Uses of Al: Electrical conductors; aerospace structures (Al-Cu, Al-Mg, Al-Zn alloys); automotive lightweighting; food packaging; construction.

1 Titanium — Kroll's Process

At. No. 22 | Density 4.54 g/cm³ | M.P. 1668 °C. Minerals: Rutile (TiO₂), Ilmenite (FeO·TiO₂). TiO₂ cannot be reduced by carbon (forms TiC). The Kroll process:

Step 1: TiO₂ + 2Cl₂ + 2C → TiCl₄ + 2CO (chlorination, 800–1000 °C)

Step 2: TiCl₄ + 2Mg → Ti (sponge) + 2MgCl₂ (800 °C, Ar atmosphere)

Ti sponge recovered by: (a) leaching MgCl₂ + excess Mg with dilute HCl, or (b) vacuum distillation of MgCl₂ and Mg. Max yield 96%. MgCl₂ electrolysed to regenerate Mg + Cl₂ (closed loop).

Fig 2.32 Production of Graded Titanium Sponge — flowsheet showing TiO₂ chlorination, TiCl₄ purification (fractional distillation), Kroll reduction with Mg in an inert-atmosphere reactor, and sponge recovery by vacuum distillation or HCl leaching. Ti sponge is melted by double vacuum arc remelting (VAR).

2 Titanium — Hunter Process

TiCl₄ reduced by sodium instead of Mg: TiCl₄ + 4Na → Ti + 4NaCl (highly exothermic). Reduction proceeds through sub-chlorides (TiCl₄ → TiCl₃ → TiCl₂ → Ti). The reaction in fused NaCl is electrochemical — Na oxidation at anodic sites, Ti sub-chloride reduction at cathodic sites.

Hunter process produces denser Ti with lower O, N, C contamination in some configurations. Uses: Jet engine compressor blades; airframes; missiles; spacecraft; medical implants. Indian extraction: Tuticorin (Tamil Nadu), Kerala (beach sand ilmenite).

3 Extraction of Zirconium

At. No. 40 | Density 6.45 g/cm³ | M.P. 1852 °C. Mineral: Zircon (ZrSiO₄). Key challenge: Zr/Hf separation (chemically nearly identical). Separation methods: differential oxidation of tetrachlorides; lower chloride formation; differential reaction of double chlorides with alkali chlorides. Then Kroll reduction of ZrCl₄ by Mg. Consolidated by double vacuum arc remelting.

Fig 2.31 Production of Zirconium Sponge — ZrSiO₄ (zircon) is processed to separate Zr from Hf; ZrCl₄ is produced by chlorination; Kroll reduction with Mg gives Zr sponge + MgCl₂; sponge recovered by vacuum distillation. Used for nuclear reactor cladding.

Uses: Nuclear reactor cladding (low neutron capture cross-section); non-nuclear flash bulbs; ZrO₂ ceramics. Indian sites: Tarapur (Maharashtra), Kota (Rajasthan), Kalpakkam (Tamil Nadu).

4 Extraction of Tin

At. No. 50 | Density 7.31 g/cm³ | M.P. 231.9 °C. Mineral: Cassiterite (SnO₂). Carbothermic reduction feasible at moderate temperature. However, FeO is more stable than SnO₂ above 1200 °C — selective Sn reduction from Sn-Fe concentrate is difficult. Three-stage smelting:

• First stage: 99% Sn recovered; almost all Fe in slag (reverberatory or rotary furnace).
• Second and third stage: Slag reduced further; products purified by liquation.

Pyrometallurgical refining: Impure Sn cooled just above 232 °C; intermetallics (Cu₂Sb, FeSn₂) of higher M.P. float and are skimmed. S removes Cu as CuS; Al removes As, Sb; NH₄Cl removes excess Al.

Fig 2.33 Smelting of Tin Concentrate — three-stage carbothermic reduction in reverberatory or rotary furnace. First stage gives ~99% Sn recovery; second and third stage slags are further reduced and liquated to recover residual tin.

Fig 2.34 Flow Sheet of Production of Refined Tin — from crude Sn through pyrometallurgical refining (deleading, decopperising by S addition, Al treatment for As/Sb) and electrolytic refining (cresol sulphonic acid electrolyte) to 99.95% Sn cathode sheets.

5 Extraction of Niobium

At. No. 41 | Density 8.57 g/cm³ | M.P. 2750 K. Minerals: Columbite [(Fe,Mn)(Nb,Ta)₂O₆]; Pyrochlore. Extracted jointly with Ta — separated by liquid-liquid extraction (MIBK or tributyl phosphate). Nb₂O₅ reduced aluminothermically or by C/H₂; consolidated by electron-beam or vacuum arc melting.

Fig 2.35 Extraction of Niobium and Tantalum — columbite ore is fused/leached to separate Nb and Ta by solvent extraction; Nb₂O₅ is then reduced to metal by aluminothermy or carbothermy; consolidated by electron-beam melting for high-purity applications.

Uses of Nb: HSLA steel microalloying (0.04% Nb raises yield strength by ~100 MPa); jet engine alloys; superconducting magnets (Nb-Ti, Nb₃Sn) for MRI and particle accelerators; Nb₂O₅ in high-refractive-index optical glasses.

1 Properties & Minerals of Uranium

At. No. 92 | At. Wt. 238.03 | Density 19.1 g/cm³ | M.P. 1405 K | B.P. 4404 K. Uranium is the densest naturally occurring element. Mineral: Uraninite (UO₂). Major producers: Kazakhstan, Canada, Australia.

2 Extraction of Uranium

• Mining and crushing → ore ground to expose uraninite grains.
• Leaching: Acidic leach (H₂SO₄) or alkaline leach (Na₂CO₃) dissolves U as UO₂²⁺ or [UO₂(CO₃)₃]⁴⁻.
• Solid-liquid separation: Counter-current decantation (CCD) or filtration removes gangue (tailings).
• Purification: Solvent extraction (tributyl phosphate/kerosene) or ion-exchange resins selectively extract U from Fe, Th, Ra impurities.
• Precipitation: U stripped from solvent → ammonium diuranate ("yellow cake" — (NH₄)₂U₂O₇) by adding NH₃.
• Conversion: Yellow cake → UF₆ (for isotope enrichment of ²³⁵U) or UO₂ pellets (for natural-U CANDU reactors).
• Enrichment: Gaseous diffusion or centrifuge increases ²³⁵U from 0.7% to 3–5% for LWR fuel.

Fig 2.36 Extraction of Uranium Ingot — complete flowsheet from uranium ore through crushing, leaching (H₂SO₄ or Na₂CO₃), solvent extraction or ion exchange, precipitation of yellow cake, conversion to UO₂ or UF₆, enrichment, and fabrication of nuclear fuel elements.

3 Uses of Uranium

• Nuclear fuel: Enriched U (3–5% ²³⁵U) as UO₂ pellets in light-water reactors; natural U in CANDU heavy-water reactors.
• Naval propulsion: Highly enriched U (>90% ²³⁵U) powers nuclear submarines and aircraft carriers.
• Depleted uranium (DU): Very dense (19.1 g/cm³); ballast for ships; armour-piercing ammunition; radiation shielding.
• Breeder reactors: ²³⁸U → ²³⁹Pu — multiplies nuclear fuel supply.

4 Environmental Pollution from Extraction

• Particulate emissions: Dust from crushing, grinding, transfer points, casting operations.
• Gaseous emissions: SO₂ (major — from sulphide roasting and converting); CO, CO₂; As₂O₃ vapour; HF (from Al smelting cells).
• Noise pollution: Crushers, ball mills, fans.
• Waste water: Acid plant effluent, electrolyte spillage, cooling water containing heavy metals.
• Solid waste: Tailings, slag, red mud (Al), spent refractories, anode stubs.
• Fugitive emissions: From conveyor belts, stockpiles, open vessels, ladle transfers.

5 Prevention and Control Methods

• SO₂ capture: Rich SO₂ from flash smelting and fluid-bed roasters → double-contact double-absorption H₂SO₄ plant — >99.7% SO₂ capture.
• Primary gas treatment: Venturi scrubbers or waste-heat boilers cool furnace off-gas; cyclones as spark separators before bag filters.
• Secondary fume capture: Canopy hoods and evacuation systems over converters, ladles, and casting bays → bag house filters or electrostatic precipitators (ESP).
• Wet scrubbers: Capture water-soluble gases (SO₂, HCl, HF) — need additional effluent treatment facilities.
• Bag house filters / ESP: High particulate efficiency for sub-micron particles; in Al smelters, HF and Al₂O₃ fume captured by dry Al₂O₃ scrubbing (fume returned to pots).
• Stockpile management: Indoor/covered stockpiles; water spray systems; dust suppressants; enclosed conveyors; sealed silos for bulk powder.
• Effluent treatment: Neutralisation, precipitation, and clarification of acid mine water before discharge.