Leaching Process & Equipment: Routes, Design, Selection

Last updated: 2026-10-18 · Reviewed by: Senior Metallurgical Engineer, {BRAND}

Leaching recovers 70–85% of the copper in a typical oxide ore with dilute sulphuric acid and 85–92% of the gold in a well-run CIL circuit, which is why it anchors most hydrometallurgical flowsheets. This pillar page compares the four industrial routes — agitated tank, heap, pressure oxidation and bioleaching — and sets out ore fit, design parameters, costs and route-selection logic.

Key Takeaways

• Route choice follows mineralogy: oxide copper dissolves in acid; refractory sulphides and locked gold need oxidation first.
• Agitated tank circuits buy recovery with CAPEX; heap circuits buy tonnes with low OPEX per tonne.
• Pulp density, P80 grind size and retention time trade against one another; only test work fixes the balance.
• Eh control and oxidant dose drive cyanide and chloride circuits more than acid strength does.
• Confirm every design number with bottle-roll and column test work before ordering tanks or liners.

What Is Leaching and Where Does It Sit in the Flowsheet?

Leaching is a hydrometallurgical extraction step that brings an aqueous reagent into contact with crushed or milled ore so the target metal dissolves into solution while gangue minerals stay behind. It sits between comminution and metal recovery: particles are sized first, the metal is dissolved second, and the loaded liquor is then clarified and treated by carbon adsorption, solvent extraction, cementation or electrowinning. Every equipment choice upstream and downstream is sized around the liquor flow and residence time this step defines.

Leaching

Leaching Reactor Volume Calculator from Residence Time

Three terms carry most of the weight in route discussions and are worth stating precisely, because vendor proposals use them loosely.

Pregnant leach solution (PLS) is a metal-bearing liquor that leaves the dissolution stage carrying the target metal, typically at 0.5–3 g/L copper or 1–10 mg/L gold, and feeds the downstream recovery plant. Its grade and impurity profile, not the head assay, determine the size of the recovery section.

Agitated tank leaching is a slurry-based contact process that mixes finely ground ore, reagent and sometimes air or oxygen in stirred, baffled vessels, running at 30–45% solids by weight with residence times from 12 hours to several days. It maximises recovery per tonne and gives the operator tight control of chemistry, at the price of the highest capital cost per tonne of the four routes.

Bioleaching is a microbial oxidation route that uses acidophilic bacteria such as Acidithiobacillus ferrooxidans and Leptospirillum ferrooxidans to oxidise sulphide minerals and release the encapsulated metal, operating at ambient to 80 °C in heaps or stirred reactors. It is the slowest route but the cheapest way to unlock metals that neither acid nor cyanide can reach directly.

The practical consequence of these definitions is that equipment selection is a downstream activity: the ore’s mineralogy and the chosen reagent chemistry set the residence time, the gas demand and the materials of construction, and the tanks, liners, blowers and thickeners follow from those numbers. For the mine owner reading vendor proposals, the definitions translate into three different questions to ask: what residence time and gas transfer are guaranteed in the tank train, what drainage rate and liner design life are guaranteed on the pad, and what oxidation extent is guaranteed in the autoclave. A proposal that cannot answer its route’s question with a number is not yet a proposal.

How Do the Four Industrial Routes Differ?

Agitated tank circuits and CIL-CIP. The ore is ground to a P80 of roughly 45–106 µm and contacted as a 35–45% solids slurry in a train of mechanically agitated, usually rubber-lined or stainless steel tanks. For gold, the reagent is dilute sodium cyanide at pH 10.5–11 with activated carbon either in the same vessels (carbon-in-leach) or downstream (carbon-in-pulp); an agitated leaching tank is specified around retention time and oxygen transfer, and a gold CIL tank additionally around carbon management and screen design. The same vessel architecture serves oxide copper, uranium and atmospheric nickel laterite treatment with sulphuric acid instead of cyanide. Residence time is 12–48 hours, recovery is the highest of the four routes, and so is the capital cost.

Agitated Leaching Tank Design

Heap leach. Crushed ore at a P80 of 12–50 mm, or run-of-mine ore where permeability allows, is stacked on a lined pad, cured with acid or cyanide solution, and irrigated at 5–15 L/h/m². Dissolution takes weeks to months, recovery is 10–25 percentage points lower than a tank circuit delivers on comparable feed, but the capital cost per annual tonne is a fraction of a tank circuit. A heap leaching system is specified around liner integrity, agglomeration quality and solution inventory rather than around impellers. Heap routes dominate copper oxide production in Chile and Peru and treat a large share of the world’s gold at low grades.

Pressure oxidation (POX/HPAL). The slurry reacts with oxygen at 180–230 °C and 0.3–3.5 MPa in a titanium-clad or lead-lined pressure oxidation autoclave, oxidising pyrite and arsenopyrite to liberate locked gold, or dissolving nickel laterite in sulphuric acid at 245–270 °C (HPAL). Residence times are 30–120 minutes, oxidation is nearly complete, and refractory gold recoveries of 92–97% are routine where cyanidation alone would extract less than half. The trade-off is the highest capital cost, an oxygen plant, and a materials problem that dwarfs every other route.

Bioleach. Mesophile or thermophile cultures oxidise secondary copper sulphides (chalcocite, covellite) or finely milled refractory concentrates in heaps at ambient to 45 °C, or in stirred reactors at 45–80 °C for concentrates. Cycle times run 60–500 days, which caps throughput per hectare of pad, but reagent cost is essentially acid and air. It wins where the ore responds to oxidation but the project cannot carry POX economics.

Which Leaching Route Fits Your Ore?

Route selection is decided by four ore properties in this order: copper mineralogy or gold mineral association, gangue acid consumption, grade, and tonnage. Get the mineralogy right and the rest falls into place; guess it and no equipment choice will rescue the flowsheet.

Table 1 — Route comparison at a glance (typical ranges, to be confirmed by test work)

The conditional conclusions follow directly from that table:

• If acid-soluble copper is above 70% of total copper and gangue acid consumption is below about 50 kg/t, an agitated acid circuit wins at any tonnage; it is the only route that delivers oxide-copper recovery above 80% without a slow pad cycle.
• If the same ore runs below roughly 0.7% Cu and above 5 Mt/a of resource, heap leaching wins on economics despite the lower recovery, provided the ore drains freely — clays above about 15% usually disqualify a heap unless the material is agglomerated carefully.
• If gold is free-milling with less than about 20% locked in sulphides, CIL wins; pressure oxidation would be capital spent on nothing.
• If more than 60% of the gold is locked in pyrite or arsenopyrite, cyanidation alone recovers under 50% and an oxidation pretreatment is mandatory; POX wins above roughly 1.5 Mt/a of concentrate-equivalent throughput or at high head grades, while bioleach wins for smaller or remote operations that cannot justify an oxygen plant.
• If the copper ore is chalcocite-dominated and acid is already on site, bioleach of crushed heaps wins over tank circuits because it consumes the idle acid and avoids grinding the whole tonne to 75 µm.

Where two routes sit within a few points of recovery, the tie is usually broken by water balance, climate and pad logistics rather than chemistry — a comparison that tank versus heap leaching works through in detail. Copper oxide applications are treated ore-by-ore in copper oxide ore processing.

Two site factors can override the mineralogy verdict. Water balance first: a tank circuit consumes and discharges far more water per tonne than a heap, so in water-short deserts the pad route gains another advantage, while in monsoon climates heap pads need storm-event pond capacity sized against multi-decade rainfall records. Climate second: bioleach heaps stall below about 5 °C unless the feed arrives warm or the cycle is extended, and cyanide heaps in freezing climates need either solution heating or seasonal operation. Neither factor changes the chemistry verdict, but both move the boundary of the conditional conclusions above by a grade or two of ore.

What Design Parameters Govern Leaching Performance?

Five parameters determine both recovery and cost, and they are coupled: change one and at least two others must move.

Table 2 — Key design parameters with units and typical ranges

The coupling matters more than any single number. Raising pulp density from 38% to 45% solids in a copper oxide tank circuit increases throughput per vessel by roughly 15% but shortens effective retention, raises viscosity enough to change impeller tip-speed requirements, and can cut recovery by 2–4 points if the retention time is not rebuilt. Grinding from a P80 of 106 µm to 53 µm typically lifts oxide-copper recovery 3–8 points but can double the grinding energy share of OPEX. In cyanide circuits, dissolved oxygen below 4 mg/L slows dissolution more than any reagent dose can fix, which is why sparging design outranks tank size on marginal ores. In heaps, the same coupling runs through agglomeration: fines blinding lowers effective permeability, which lowers irrigation uniformity and short-circuits solution to the toe drains, and the recovery loss only becomes visible months later at the column and pad scale.

None of these numbers should be locked in from a database. The recovery arithmetic behind the trade-offs is set out in the leaching recovery calculation resource, and the full calculator set lives in the resources library.

How Do Reagent Systems Shape the Circuit?

Reagent choice fixes the materials of construction, the water balance and half of the operating cost, so it belongs upstream of vessel selection.

Dilute sulphuric acid (5–50 g/L) is the workhorse for oxide copper, uranium and atmospheric laterite nickel. It is cheap, non-selective, and its consumption is set by gangue carbonates — malachite and azurite consume it usefully, limestone consumes it entirely. Acid circuits run at 20–45 °C in rubber-lined or 316L equipment, and the main design liability is gypsum scaling anywhere the liquor is heated or concentrated.

Alkaline cyanide (pH 10.5–11, NaCN 0.3–1.5 g/L in liquor) is selective for gold and silver and tolerant of most gangue, but it demands strict pH control to prevent HCN evolution, careful management of copper and zinc bleed, and detoxification of tailings solution. Copper-cyanide complexes above about 1,000 mg/L Cu erode both gold recovery and cyanide efficiency and usually justify a SART or acidification bleed.

Chloride systems (100–250 g/L Cl⁻ with cupric or ferric oxidant) are the emerging route for chalcopyrite, the one major copper mineral acid dissolution handles poorly. They operate at 30–100 °C and promise high recovery at moderate temperature, but they are corrosive, impurity-tolerant rather than impurity-free, and still carry scale-up risk relative to the other three systems.

A reagent change is a materials change: moving a copper circuit from sulphate to chloride chemistry invalidates almost every wetted-material selection in the plant, which is why the reagent decision must be frozen before tank quotations go out.

Which Operating Problems Recur in Leach Plants?

The six problems below account for most recovery and cost complaints in operating circuits, written as symptom, cause, action — the form in which they reach the metallurgist’s desk.

1. Recovery drifting down over the first year. Symptom: CIL or acid tank recovery falls 2–5 points against design while head grade is unchanged. Cause: grind coarsening from liner wear or feed hardness drift, air or oxygen transfer below design, or a shift toward more refractory ore zones. Action: re-run bottle rolls on current feed at current P80, survey dissolved oxygen through the train, and rebuild retention time before blaming the reagent regime.
2. Cyanide and lime consumption climbing. Symptom: NaCN and lime doses up 30% year-on-year at flat recovery. Cause: rising reactive copper in feed forming cuprous cyanide, or preg-robbing carbonaceous gangue entering the mill. Action: install a SART bleed for copper, run a preg-robbing index on ore zones, and consider CIL carbon additions ahead of the leach.
3. Gypsum and jarosite scale in pipes and tanks. Symptom: solution lines losing 30% of flow area within months; agitators vibrating. Cause: acid circuits heated or evaporatively concentrated past gypsum saturation, or Eh swings precipitating jarosite in hot zones. Action: hold liquor supersaturation margins, add scale sampling points at the known hot spots, and clean on a fixed interval instead of on failure.
4. Heap percolation failure and channeling. Symptom: PLS grade collapses, solution reports to toe drains unevenly, side slopes slump. Cause: fines migration into the stacked bed, over-crushed P80, or inadequate agglomeration with too little acid cure. Action: stop stacking, audit agglomeration quality (fines capture, cure acid, cure time), and lift the drip rate only after re-ripping the surface — never pour water at a blocked heap.
5. Liner or pad integrity loss. Symptom: solution inventory unaccounted, piezometers wet under the pad, liner blisters at sump areas. Cause: liner puncture during stacking, uv degradation at exposed edges, or hydraulic uplift where drainage gravel has blinded. Action: trigger the leak-detection response plan, repair under current geometry, and fix the stacking equipment or gravel spec that caused the puncture.
6. Foaming and surging in agitated tanks after an ore change. Symptom: tanks surge, level control oscillates, froth carries solids into launders. Cause: clay-rich or organic-bearing ore changing slurry rheology at the same solids density. Action: drop pulp density 3–5 points temporarily, verify impeller submergence, and requalify the frother or launder design for the new ore — the 40%-solids set point that ran clean on the first ore body is rarely right for the second.

The recurring pattern in all six: the failure appears in the leach circuit but the cause sits upstream in the mine plan, the crushing plant or the water balance. Fixing the symptom in the tank costs twice, because it returns with the next ore source.

What Do These Circuits Cost in Energy and Reagents?

The figures below are indicative ranges assembled from vendor budget quotations and published project studies for plants of 1–5 Mt/a; they are stated per tonne of ore treated, exclude tailings storage and off-grid power generation, and must be confirmed by test work before any budget is committed.

• Agitated acid leach (oxide copper): sulphuric acid 15–45 kg/t at, historically, USD 80–200/t delivered — the single largest OPEX item and the reason gangue acid consumption testing is non-negotiable; electrical energy 20–35 kWh/t; grinding already counted upstream.
• CIL gold: sodium cyanide 0.4–1.2 kg/t, lime 1–3 kg/t, activated carbon losses 20–50 g/t ore, power 25–40 kWh/t including air agitation; detoxification reagents (SO₂/air or peroxide) add USD 0.2–0.8/t depending on weak-acid-dissociable cyanide targets.
• Copper heap leach: acid 5–25 kg/t, power 3–8 kWh/t for crushing, stacking and pumping, plus pad liner replacement amortisation of USD 0.1–0.3/t.
• Pressure oxidation (refractory gold concentrate): oxygen 0.2–0.35 t per tonne of concentrate, power 45–70 kWh/t concentrate, acid neutralisation driven by sulphate balance; the oxygen plant alone can be 15–20% of process plant CAPEX.
• Bioleach: acid 2–10 kg/t, power 5–15 kWh/t, dominated by aeration in reactor configurations; the reagent bill is the lowest of the four routes, traded against a pad or residence inventory 10–30× larger.

Two cross-checks keep these numbers honest. First, acid-delivered price moves the oxide-copper OPEX ranking more than any flowsheet option: a doubling of acid price can move heap routes ahead of tank circuits on ores with 10–15 kg/t consumption differences. Second, energy figures for POX are per tonne of concentrate, not per tonne of ore, and the conversion ratio (typically 5–10% of mine tonnage reporting to concentrate) is where comparisons go wrong.

Capital cost follows the same pattern as the recovery ranking. Indicative process-plant CAPEX, in 2025 US dollars, excluding owner’s costs, tailings storage and power infrastructure, sits near USD 10,000–18,000 per annual tonne of ore for an agitated acid or CIL plant of 1–3 Mt/a, USD 1,500–4,000 per annual tonne for a heap pad with crushing and stacking, and USD 25,000–45,000 per annual tonne of concentrate for a POX train including the oxygen plant. These are screening numbers only: they assume competent construction markets, standard materials of construction and no unusual permitting burden, and a single clay-rich ore body or a remote site can move any of them by half again.

FAQ

Q1. What is the difference between CIL and CIP leaching?
Both are cyanide circuits using activated carbon; the difference is where the carbon sits. In CIL, carbon is present in the leach tanks themselves, so adsorption happens while gold is still dissolving, which suits preg-robbing or slow-dissolving ores. In CIP, dissolution finishes first and carbon adsorbs gold in a separate downstream train of tanks. CIL uses more carbon inventory; CIP typically needs more total tank volume.

Q2. Can heap leaching treat sulphide ores?
Yes, but only secondary sulphides such as chalcocite and covellite, and only with bacterial assistance and a long acid cure. Primary chalcopyrite responds too slowly for commercial heap cycles at ambient temperature. Pyrite-hosted refractory gold can be heap-biooxidised as a pretreatment before cyanidation, but the recovery ceiling is lower than stirred-reactor or pressure oxidation routes, and cycle times stretch past a year.

Q3. What pulp density should an agitated leach tank run at?
Most oxide copper and CIL circuits run at 35–45% solids by weight. Below about 33% you are pumping and heating water for no metallurgical gain; above about 48% with clays, viscosity rises enough to impair oxygen transfer and impeller performance. The correct set point is the one where retention time, oxygen transfer and rheology all close, and that comes from viscosity-adjusted bench testing on the actual ore.

Q4. How long does it take to pilot a leach circuit?
A focused programme runs 8–16 weeks: bottle rolls and stirred tests for kinetics, column tests of 2–6 months only where a heap route is live, and cyanide destruction or neutralisation testwork in parallel. Pressure oxidation pilot work adds autoclave campaigns of 2–4 weeks. Anything shorter than eight weeks cannot capture ore variability, which is the variable that most often breaks the design.

Q5. Is cyanide leaching safe for operations near communities?
Modern circuits manage cyanide risk through International Cyanide Management Code compliance: pH above 10.5 to suppress HCN gas, bermed and netted solution ponds, wildlife deterrents, and tailings detoxification to regulated weak-acid-dissociable limits. Thousands of plants operate safely near populated areas under these controls. The genuine risks are transport of reagent and loss-of-containment events, both of which are engineered and audited rather than inherent to the chemistry.

Q6. When is pressure oxidation worth the extra CAPEX?
When more than half the gold is locked in sulphide and the resource is large enough or rich enough to amortise an autoclave and oxygen plant — commonly above 1.5 Mt/a concentrate-equivalent or a multi-million-ounce reserve. Below that scale, bioleaching or fine-grinding-plus-cyanidation usually gives a better return. The decision point is simple: if oxidised feed recovers 90%+ and unoxidised feed recovers under 60%, oxidation is worth paying for; the question is which oxidation route the tonnage supports.

Conclusion

Leaching route selection is a mineralogy decision first and an equipment decision second: match the reagent system and oxidation route to the ore, then let pulp density, P80, retention time and Eh set the vessel sizes — the four routes each win under the conditions set out above, and none of them wins everywhere.

If your ore sits between two of these routes and the design numbers are still open, submit your head assays and gangue acid consumption data through our pilot testing service, and we will return a route recommendation with indicative recovery, reagent consumption and equipment sizing.

References

1. USGS, Mineral Commodity Summaries 2025 — copper and gold production, reserves and processing context. usgs.gov
2. Wills, B.A. & Finch, J.A., Wills’ Mineral Processing Technology, 8th ed., Chapters on hydrometallurgy and leaching.
3. SME Mineral Processing Handbook, Society for Mining, Metallurgy & Exploration — leaching, heap and autoclave design sections. smenet.org
4. ALTA Metallurgical Services, ALTA 2024 Proceedings — copper, nickel, cobalt and uranium leaching practice. altamet.com.au
5. MetSoc, Conference of Metallurgists (COM) Proceedings — biohydrometallurgy and pressure oxidation sessions.
6. CSIRO Mineral Resources — heap leaching and in-situ recovery research programs. csiro.au

Changelog

• 2026-10-18 — Initial pillar refresh. Four-route comparison, design parameter tables, route-selection logic, cost ranges and FAQ.

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