Atmospheric Leaching vs Pressure Leaching: Sulfur Decides

Last updated: 2026-11-06

The fate of sulfur — sulfate in solution or elemental sulfur locked in the solids — decides whether atmospheric leaching or autoclave pressure leaching is the right flowsheet. Full sulfate oxidation of a 28 % S concentrate makes roughly 3 kg of sulfuric acid per kilogram of sulfur oxidized; keep sulfur as S⁰ and the acid and its neutralization plant disappear, but so does the fast kinetics that pressure buys.

If you are still mapping the route options themselves, start with the leaching process overview before comparing these two oxidation paths in detail.

Key Takeaways

• Sulfur oxidation state, not temperature alone, decides between atmospheric leaching and autoclave flowsheets.
• Full sulfate oxidation generates acid that must be neutralized with lime and discarded as gypsum.
• Elemental sulfur routes halve oxygen demand but risk passivation and molten-sulfur agglomeration above 115 °C.
• Atmospheric leaching runs in 316L tanks; pressure leaching demands titanium-clad vessels and oxygen plants.
• Compare oxygen, lime, and disposal accounts per ton of concentrate — not just metal recovery.

What Separates Atmospheric Leaching From Pressure Leaching?

Atmospheric leaching is a hydrometallurgical oxidation route that dissolves metal sulfides or oxides in agitated tanks open to the air, at temperatures below the boiling point of the slurry (typically 80–105 °C) and with oxygen supplied either from air sparging or from enriched-air injection. Pressure leaching — usually called autoclave leaching in plant literature — is a high-temperature oxidation technique that confines the slurry in a pressure vessel so water can be heated to 150–250 °C and oxygen partial pressures of 0.5–3.0 MPa can drive reactions tens of times faster.

Between those two brackets sits a third concept that the rest of this page depends on. The degree of sulfur oxidation is a mass-balance metric that tracks where each atom of feed sulfur ends up: sulfate (S⁶⁺) in the liquor, elemental sulfur (S⁰) in the solids, or unreacted sulfide (S²⁻) still locked in the concentrate. Oxygen partial pressure is the thermodynamic and kinetic lever that determines how far that oxidation proceeds; raising it shifts the endpoint from sulfide toward elemental sulfur and, at sufficient intensity, all the way to sulfate. The passivating sulfur layer is a coating of molten or solid elemental sulfur that forms on partially reacted sulfide particles and blocks fresh oxidant from reaching unreacted mineral surfaces.

Everything else — vessel metallurgy, residence time, oxygen plant size, neutralization reagent — follows from which sulfur endpoint the flowsheet targets. For a full treatment of the vessel side, see our pressure oxidation autoclave guide.

Which Parameters Actually Differ?

The two routes diverge on nearly every operating axis, not just pressure. Table 1 summarizes typical ranges, to be confirmed by test work for any specific feed.

Three of those rows deserve comment. Residence time is the direct price of low oxygen partial pressure: at 0.1 MPa total pressure the dissolved oxygen ceiling is so low that even enriched-air tanks need 4–24 h, which is why tank sizing dominates capital cost on atmospheric leaching projects and why our residence time calculation and agitated tank design pages get so much traffic from process engineers. Solids concentration runs the other way: hot acid systems tolerate 35–40 wt% solids, while autoclaves thin the slurry to 10–25 wt% because viscosity rises sharply as molten sulfur and iron hydrolysis products accumulate. Materials of construction follow temperature and chloride: 316L or duplex suffices below 105 °C in sulfate media, but above 150 °C only titanium cladding or acid brick survives sustained attack, which is the single largest line item in an autoclave budget.

Why Does the Sulfur Oxidation State Decide the Flowsheet?

Here is the core argument, and it is a reagent-account argument rather than a recovery argument.

The sulfate route manufactures acid whether you want it or not. When sulfide sulfur is oxidized fully to sulfate, each kilogram of sulfur generates about 3.06 kg of H₂SO₄ equivalent (98.1/32.0 = 3.066). Unless the gangue consumes every kilogram of that acid, the plant ends up with a net acid surplus that must be neutralized before tailings discharge. On the worked example below — 500 t/d of concentrate at 28 % S — the full-sulfate route makes 429 t/d of acid equivalent and needs 245 t/d of lime, producing 753 t/d of gypsum sludge. That is a permanent neutralization and disposal liability bolted onto the metal plant.

The elemental sulfur route sells the sulfur back to you as a filterable solid. Stop the oxidation at S⁰ and the acid plant, the lime plant, and the gypsum pond all shrink toward zero. Oxygen demand drops by half as well, because taking sulfide sulfur from −2 to 0 needs two electrons (½ mol O₂ per mol S) instead of eight electrons (2 mol O₂ per mol S) for full sulfate. This is why acid balance leaching — deliberately starving the system of oxidant so sulfur reports as S⁰ — has been pursued for chalcopyrite and other refractory sulfides for decades.

But the S⁰ route introduces the two failure modes pressure chemistry was designed to avoid. First, elemental sulfur passivation: a sulfur film coats unreacted sulfide particles, and leaching stops at 30–60 % extraction unless the film is disrupted. Second, sulfur melts at 115–119 °C; between roughly 119 °C and 159 °C molten sulfur is viciously viscous and glues unreacted particles into agglomerates that never leach. A counterintuitive point that only shows up in operating plants: in the full-sulfate route, sulfur passivation is essentially a non-issue, because there is no S⁰ phase to coat the particles — every sulfide surface stays clean until it dissolves. The elemental sulfur route has exactly the opposite burden: it must hold the slurry below 115 °C, or above the viscous window, and it must mechanically abrade the sulfur layer, which is why those flowsheets lean on fine regrind, surfactant dosing, or discrete 80–95 °C tank stages.

Iron has to go somewhere too. Atmospheric leaching of iron-bearing sulfides leaves ferric iron in solution, and the acid-balance circuit must remove it by precipitation as jarosite, goethite, or hematite in controlled stages — each with its own acidity window and its own gypsum or alkali consumption. Pressure leaching at 180–230 °C hydrolyzes ferric iron directly inside the autoclave to hematite or basic ferric sulfate, so iron leaves with the residue and the downstream liquor is comparatively clean. That in-autoclave iron rejection is one of the genuine process advantages of autoclave leaching that no tank circuit can replicate economically.

The decision therefore reduces to a question the metallurgist can answer quantitatively: does the value of fast oxidation and in-vessel iron rejection exceed the combined cost of the oxygen plant, the titanium shell, and — where sulfate is the endpoint — the lime and gypsum accounts? For the neutralization side of that equation, our sulfide precipitation vs lime neutralization comparison and the general acid consumption calculation walk through the same arithmetic at circuit level.

How Do You Size the Oxygen and Neutralization Accounts for 500 t/d?

This is a synthetic worked example. Assumptions: concentrate feed 500 t/d at 28 % S; sulfur atomic mass 32.0 kg/kmol; complete sulfur extraction; H₂SO₄ 98.1 kg/kmol; CaO 56.1 kg/kmol; gypsum as CaSO₄·2H₂O 172.2 kg/kmol; oxygen-sourcing power 0.35–0.50 kWh/kg O₂ (take 0.40 for the point estimate). All figures are indicative, synthetic-typical basis, to be confirmed by test work.

1. Convert the sulfur feed to a molar basis. Sulfur in: 500 t/d × 0.28 = 140 t S/d = 140,000 kg/d. Moles: 140,000 ÷ 32.0 = 4,375 kmol S/d.
2. Size oxygen for the full-sulfate route. Sulfide sulfur to sulfate is an eight-electron oxidation: S²⁻ + 2 O₂ → SO₄²⁻. Oxygen demand = 2 × 4,375 = 8,750 kmol O₂/d × 32.0 kg/kmol = 280 t O₂/d.
3. Size the acid make for the same route. One mole of acid per mole of sulfur: 4,375 kmol × 98.1 kg/kmol = 429 t H₂SO₄-equivalent/d (≈858 kg per ton of concentrate fed).
4. Size the neutralization account. Neutralizing to gypsum, CaO + H₂SO₄ → CaSO₄·2H₂O on a 1:1 molar basis: lime = 4,375 kmol × 56.1 = 245 t CaO/d; gypsum = 4,375 kmol × 172.2 = 753 t CaSO₄·2H₂O/d. If your site instead precipitates metals from that acidic liquor first, compare the options in our sulfide precipitation vs lime neutralization analysis.
5. Size oxygen for the elemental sulfur route. Sulfide to elemental sulfur is a two-electron oxidation: S²⁻ + ½ O₂ → S⁰. Stoichiometric minimum = 0.5 × 4,375 × 32.0 = 70 t O₂/d. Design practice carries an allowance for oxygen-transfer inefficiency and partial over-oxidation to sulfate, taking the basis to 1.0 kg O₂ per kg S → 140 t O₂/d. Acid make ≈ 0; lime ≈ 0; gypsum ≈ 0.
6. Convert both to power. At 0.40 kWh/kg O₂: full-sulfate route = 280,000 kg × 0.40 = 112,000 kWh/d; elemental sulfur route = 140,000 kg × 0.40 = 56,000 kWh/d. Across a year that is roughly 41 GWh versus 20 GWh — in remote sites with diesel or weak-grid power, this single line often decides the flowsheet before metallurgy does.

One caution on step 5: the 1.0 kg O₂/kg S basis is a design allowance, not chemistry. Plants that underdeliver oxygen see extraction sag; plants that overdeliver watch their carefully balanced elemental sulfur route drift back toward sulfate and the acid account reappears. The oxygen control loop is the flowsheet.

Which Failures Are Specific to Each Route?

These are the seven failure modes we see most often in commissioning and troubleshooting reviews. Each is listed as symptom → cause → response.

1. Elemental sulfur passivation (S⁰ routes). Symptom: extraction plateaus at 30–60 % despite fresh reagent and ample residence time. Cause: a passivating sulfur layer of S⁰ coats unreacted sulfide grains and shuts off oxidant transport. Response: fine regrind of middle-stage solids, surfactant or lignosulfonate dosing to disperse molten sulfur, and staged leaching so each stage sees fresh surfaces.
2. Molten sulfur agglomeration above 115 °C (S⁰ routes). Symptom: sand-like agglomerates in tank bottoms and residue filters; extraction drops simultaneously. Cause: sulfur melts at 115–119 °C and is highly viscous up to ≈159 °C, cementing particles together. Response: hold operating temperature below 115 °C or push past the viscous window; install temperature interlocks on agitators and heating coils.
3. Chloride stress corrosion cracking of autoclave internals (pressure route). Symptom: cracking of agitator shafts, thermowell bosses, and baffle welds within months. Cause: chloride concentrated in the acidic, high-temperature liquor combined with residual stress in austenitic internals. Response: specify titanium or specified duplex internals in the metallurgy review; control chloride in feed water and reagents; our chloride leaching reactor selection page covers the material decision tree.
4. Calcium sulfate and jarosite scale (both routes, worst in hot-acid tanks). Symptom: falling heat-transfer coefficient on coils and a shrinking effective tank volume. Cause: CaSO₄ and jarosite supersaturate as temperature and acidity cycle. Response: online descaling ports, acid wash cycles, and a targeted acid consumption calculation that includes the scale-forming calcium balance, not just the acid balance.
5. Oxygen supply interruption (both routes, critical in autoclaves). Symptom: extraction sag and, in autoclaves, a full depressurization and shutdown cycle costing a shift or more. Cause: air separation unit trip or compressed-air failure; autoclaves cannot tolerate hypoxia without quench procedures. Response: redundant ASU trains or liquid-oxygen buffer storage sized for at least the autoclave quench-and-restart sequence.
6. Pressure safety valve erosion (pressure route). Symptom: PSV fails its seat-tightness test at the first overhaul; chatter and wire-drawing marks on the seat. Cause: two-phase sulfide slurry flashing through the valve during relief events. Response: rupture-disc/PSV combination, clean-gas cushioning, and an inspection interval matched to the actual relief frequency, per ASME Section VIII pressure-vessel rules.
7. Gypsum scaling in the neutralization plant (sulfate routes). Symptom: lime slurry lines and neutralization tanks plug; discharge pH drifts. Cause: 753 t/d of gypsum in the worked example has to nucleate and settle somewhere, and it prefers your pipework. Response: high-solids recycle seed circuits, liner selection on lime lines, and a dedicated gypsum thickener/filter rather than letting the tailings pond do the clarifying.

When Is Pressure Leaching the Wrong Answer?

For low-sulfide, oxidized ores such as oxide copper and zinc carbonates, pressure leaching is preferred to be avoided — more precisely, acid tank leaching at atmospheric conditions is preferred because the gangue, not sulfide oxidation, controls acid consumption and the extra oxygen buys no kinetics. For low-grade clay-type deposits where solids handling dominates, agitated atmospheric leaching is preferred because thin autoclave slurries (10–25 wt% solids) make vessel volumes punitive at clay tonnages. For remote sites with unstable grid or diesel-generated power, atmospheric leaching is preferred because the elemental sulfur or mild-oxidation route halves the oxygen plant load — 56,000 versus 112,000 kWh/d in the worked example — and an interrupted 4 MPa oxygen supply is not a nuisance, it is a shutdown. Where those conditions match your deposit, review the tank vs heap leaching comparison before committing to a vessel.

When Is Atmospheric Leaching the Wrong Answer?

For high-sulfide pyritic concentrates, atmospheric leaching is the wrong answer because 4–24 h residence times at air-side oxygen pressures cannot burn through the sulfide fast enough, and the sulfuric acid generated at scale (429 t/d in the worked example) overwhelms most gangue acid consumption. For chalcopyrite concentrates where >90 % copper recovery is the target, atmospheric leaching alone is the wrong answer because the passivating sulfur layer and the stability of chalcopyrite at low temperature cap recovery far below autoclave performance. For sites that lack both a lime source within economic haul distance and a gypsum disposal area, the full-sulfate atmospheric route is the wrong answer because 245 t/d of CaO consumption and 753 t/d of gypsum have nowhere to go. Where sulfate-neutralization capacity is the binding constraint, our zinc calcine leaching and nickel HPAL application pages show how the acid account is engineered rather than merely tolerated.

Three Operating Snapshots

All figures are indicative, synthetic-typical basis, anonymized and representative of published operating envelopes.

Snapshot A — zinc concentrate, sulfate route, temperate coastal site. A 380 t/d sphalerite concentrate circuit ran atmospheric leaching at 90–95 °C in eight 316L tanks, 30 wt% solids, 14 h total residence time. Zinc extraction held at 96–97 %, and the acid make was consumed almost entirely by calcine in a matched roast-leach balance; the residual neutralization load was about 60 t/d CaO equivalent.

Snapshot B — chalcopyrite concentrate, pressure route, high-altitude site. A 150 t/d concentrate autoclave operated at 220 °C and 3.2 MPa total pressure with 0.9 MPa O₂, 0.8 h residence, 14 wt% solids. Copper extraction reached 95–96 % with iron rejected as hematite in-vessel, and the site’s oxygen plant drew close to 60,000 kWh/d for the single line.

Snapshot C — mixed sulfide, elemental sulfur route, remote diesel-grid site. A 220 t/d circuit attempted acid-balance leaching at 95–105 °C across six agitated tanks, 32 wt% solids, 18 h residence. Extraction stabilized near 72 % until a surfactant-and-regrind retrofit broke the passivating sulfur layer, after which copper extraction moved to 84–86 % and the measured sulfur-to-S⁰ fraction stayed above 70 %, keeping the lime account under 25 t/d.

FAQ

Does higher temperature alone justify choosing pressure leaching?
No. Temperature and oxygen partial pressure buy kinetics, but kinetics only matter if sulfur’s destination — sulfate versus elemental sulfur — fits your reagent and disposal accounts. A plant that cannot absorb 429 t/d of acid equivalent per 140 t/d of sulfur should not choose the sulfate route regardless of how fast it extracts metal. Judge the flowsheet on the oxygen, lime, and gypsum columns of your own mass balance, with recovery as the tiebreaker rather than the headline.

Can atmospheric leaching reach the same recovery as autoclave leaching?
For oxide feeds and many zinc calcines, yes, at comparable or better cost. For chalcopyrite and other refractory sulfides, usually not: the passivating sulfur layer and slow low-temperature kinetics cap extraction at roughly 60–85 % in most tank circuits. The gap narrows with fine regrind, surfactants, and staged hot-acid designs, but closing it fully typically costs more than the autoclave would have.

Why is elemental sulfur a problem at high temperature if pressure leaching avoids it?
Because the two routes sit on opposite sides of 115 °C. Pressure leaching runs at 150–250 °C, hot enough that any elemental sulfur is molten and dispersed, and at high oxygen partial pressure most sulfur is pushed to sulfate anyway. Elemental sulfur routes deliberately operate at 80–105 °C, just below the melting point, so the flowsheet must police the 115 °C boundary continuously to prevent molten sulfur agglomeration.

How much lime does the sulfate route really consume?
Stoichiometrically, 1 kg CaO per 1.749 kg H₂SO₄ neutralized, which on fully oxidized sulfur means about 1.75 kg CaO per kg of sulfur. For the 500 t/d worked example that is 245 t CaO/d. Real plants consume 10–30 % more, depending on gangue acid consumption, iron precipitation chemistry, and how much acid is recycled to the leach.

Is the oxygen plant really half the size for elemental sulfur routes?
The stoichiometric ratio is 0.5 kg O₂ per kg S versus 2.0 kg O₂ per kg S, so chemistry says one quarter. Design practice inflates the elemental sulfur figure to roughly 1.0 kg O₂ per kg S for transfer inefficiency and over-oxidation control, giving the 140 versus 280 t/d comparison in the worked example. Size your air separation plant from measured oxygen utilization in test work, not from the stoichiometric minimum.

What do I do with the gypsum?
Gypsum from neutralized sulfate liquors (753 t/d in the worked example) is typically thickened, filtered or centrifuged, and stacked in a dedicated cell. Some cements or wallboard markets take high-purity gypsum, but hydrometallurgical gypsum usually carries metal traces that close those doors. Assume a lined disposal facility in every sulfate-route feasibility study, and cost its closure in year one.

Which route is cheaper to operate per ton of concentrate?
It depends on three site-specific numbers, so build the comparison rather than assume it: oxygen power (112,000 vs 56,000 kWh/d at 500 t/d), neutralization reagent (245 t CaO/d or near zero), and maintenance (titanium autoclave internals vs 316L tank internals). Send us your feed assay, tonnage, and site power cost through the request-a-quote page and we will run both accounts for your case.

Conclusion

Sulfur decides. If your flowsheet oxidizes sulfur fully to sulfate, you are buying kinetics with an oxygen plant and paying again in lime and gypsum; if you stop at elemental sulfur, you halve the oxygen account and eliminate neutralization, but you inherit the passivating sulfur layer and the 115 °C melting constraint that define atmospheric leaching practice. Run the sulfur mass balance before the vessel study — the 500 t/d worked example above shows how quickly 429 t/d of acid and 753 t/d of gypsum change a project’s economics. When you are ready to compare both routes against your own concentrate, share the feed data through our request-a-quote page and our process team will return a sized oxygen and neutralization account for each option.

References

1. U.S. Geological Survey, Mineral Commodity Summaries (latest edition). https://www.usgs.gov/
2. Society for Mining, Metallurgy & Exploration, SME Mineral Processing Handbook — chapters on hydrometallurgical reactor selection and pressure oxidation.
3. Wills, B.A. & Finch, J.A., Wills’ Mineral Processing Technology (9th ed.) — leaching and oxidation chapter.
4. TMS / MetSoc of CIM, proceedings on pressure hydrometallurgy — autoclave leaching of chalcopyrite and nickel sulfide concentrates.
5. Dreisinger, D., “Baghals and ferric arsenate: industrial pressure oxidation practice review,” TMS pressure hydrometallurgy session papers.
6. CSIRO Minerals Resources, research notes on acid balance and elemental sulfur control in sulfide leaching. https://www.csiro.au/

Changelog

• 2026-11-06: Initial publication.

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