Heap leaching
Heap leaching is an industrial hydrometallurgical process used to extract metals such as copper, gold, uranium, and nickel from low-grade ores, wherein crushed or run-of-mine ore is stacked into large, aerated heaps on impermeable liners and irrigated with a chemical lixiviant solution that percolates through the pile to dissolve target metals, which are then recovered from the collected pregnant leachate via methods like solvent extraction or precipitation.[1][2][3]
The technique traces its roots to 16th-century copper extraction practices in Hungary and gained modern prominence with uranium applications in the 1950s and gold heap leaching pioneered in Nevada in 1969, enabling the economical treatment of vast, marginal ore deposits that conventional milling could not justify due to its minimal comminution requirements and low capital intensity.[4][5]
While heap leaching has revolutionized metal production by facilitating recovery from oxide and secondary sulfide ores—particularly sulfuric acid leaching for copper and nickel, and cyanide-based cyanidation for gold and silver—it poses environmental risks including potential leaks from liners, generation of acid rock drainage, and toxicity from reagents like cyanide, necessitating rigorous containment, neutralization, and reclamation to mitigate groundwater contamination and ecological harm.[6][7][8]
History
Origins in Precious Metals Extraction
Heap leaching for precious metals extraction originated in the late 1960s in northern Nevada, United States, as a method to economically recover gold from low-grade, disseminated ores unsuitable for conventional milling due to their fine particle size and low concentrations, often below 0.1 ounces per ton. This technique involved stacking run-of-mine ore on impermeable pads and percolating dilute cyanide solutions through the heaps to dissolve and collect the metal, addressing the limitations of agitated cyanidation which required finer grinding and higher costs.[9] The development was empirically driven by the Carlin-type deposits discovered in the region, where gold occurs as submicroscopic particles refractory to traditional methods, making heap leaching a practical alternative for marginal resources.[4] The first commercial-scale gold heap leach operation began at the Cortez mine in 1969, marking the birthplace of modern precious metals heap leaching and utilizing cyanide lixiviant on uncrushed ore heaps.[4] [9] Prior small-scale trials occurred at Carlin Gold Mining Company sites in the late 1960s, but Cortez's implementation demonstrated viability for larger operations.[10] Rising gold prices post-1971, following the end of the Bretton Woods fixed-price era, further incentivized adoption by enhancing profitability for these low-grade feeds uneconomic at prior $35 per ounce levels.[11] Initial recovery rates in these early gold heaps ranged from 50% to 70%, significantly lower than the 80-95% from milling and tank leaching, yet the process's low capital investment—often under $10 per ounce recovered—and operational simplicity enabled extraction from ores as low as 0.02 ounces per ton.[12] Silver heap leaching followed similar principles and timelines, often co-extracted with gold in Nevada operations, though early applications focused primarily on gold due to its higher value and prevalence in the disseminated deposits.[9] These pioneering efforts laid the groundwork for heap leaching's expansion, prioritizing empirical validation over theoretical optimization in response to real-world ore characteristics and economic pressures.Commercial Expansion to Base Metals
The adaptation of heap leaching to base metals, particularly copper, began in the late 1960s with the commercialization of sulfuric acid leaching for oxide ores, marking a shift from earlier precious metals applications due to the technology's scalability for low-grade deposits. The Ranchers Bluebird mine in Arizona, operated by Ranchers Exploration and Development Corporation, initiated the first modern commercial copper heap leach operation in 1968, processing oxide copper ores with dilute sulfuric acid to achieve viable metal extraction.[13][14] This operation demonstrated empirical feasibility, producing cathode copper via solvent extraction and electrowinning (SX/EW), and ran until approximately 1975, validating heap leaching as an alternative to energy-intensive smelting for marginal ores.[15] Economic pressures, including declining average ore grades—from over 1% copper in the mid-20th century to below 0.6% by the 1970s—and rising smelting energy costs amid oil price shocks, drove adoption. Heap leaching offered capital efficiencies, with lower upfront investments compared to milling and flotation, enabling processing of ores uneconomic by conventional methods; recoveries for oxide copper typically reached 70-80% under optimized conditions, far surpassing dump leaching yields.[16][17] First-principles scalability stemmed from the process's reliance on simple percolation dynamics and acid consumption tied directly to mineralogy, allowing modular expansion without proportional infrastructure scaling. By the mid-1980s, the technology proliferated globally, particularly in major copper producers like Chile, where large-scale operations commenced with the Lo Aguirre mine in 1980, treating oxide ores on engineered pads to support low-cost production amid surging demand.[18] This expansion to permanent heaps in arid regions facilitated heap sizes exceeding millions of tons, with irrigation rates tuned for percolation efficiency, cementing heap leaching's role in over 20% of global copper output by decade's end and enabling extraction from vast low-grade resources previously bypassed.Technological Evolution Post-1980s
Following the commercial expansion of heap leaching in the 1970s, the 1980s marked a pivotal shift toward process refinements aimed at addressing limitations in ore types previously unsuitable for effective percolation, such as those with high clay content or fines. Agglomeration techniques, involving the binding of crushed ore particles with cement or lime to form stable nodules, emerged as a key innovation to mitigate channeling and pooling of leach solutions. The first commercial agglomeration-heap leach operation commenced in 1980 for precious metals extraction, with 36 such facilities operational by 1983, demonstrating improved solution flow and metal yields in challenging feeds.[19][20] Parallel advancements integrated heap leaching with solvent extraction-electrowinning (SX-EW) for base metals, particularly copper, enabling scalable recovery from low-grade oxide ores. This coupling, operationalized in major Chilean projects from 1980 onward, facilitated cathode-grade copper production without smelting, with heap leaching contributing over 20% of global copper output by the late 1980s through enhanced selectivity and reduced energy demands.[21][22] Concurrently, high-pressure grinding rolls (HPGR) gained recognition in the late 1980s for generating denser, micro-fractured particles that accelerated leaching kinetics compared to conventional crushing.[23] The 1990s introduced bacterial-assisted methods for refractory sulfide ores, expanding heap leaching to previously uneconomic deposits via bio-oxidation pretreatments that liberated encapsulated metals. U.S. Patent 5,196,052, issued in 1993, detailed a process using acidophilic bacteria to enhance solubilization in clayey or fine refractory ores, minimizing fines migration and improving overall heap integrity. By the 2000s, these bio-oxidation heaps achieved sulfide oxidation levels sufficient for subsequent cyanidation or acid leaching recoveries of 80-90% for gold and copper, as validated in pilot and commercial applications treating whole-ore sulfides.[24][25]Fundamental Principles
Ore Preparation and Heap Formation
Ore preparation for heap leaching begins with crushing to enhance permeability and increase surface area for solution contact, typically reducing particle sizes to a range where 80% passes 75 mm, though finer crushing to minus 10-20 mm is applied for low-permeability ores to mitigate poor percolation.[5][26] Run-of-mine ore may be used directly for competent, coarse materials, achieving recoveries up to 70% in northern Nevada operations without crushing.[5] For ores with high fines content exceeding 10-15% below 0.1 mm, agglomeration is essential to bind particles, control fines migration, and prevent channeling that could bypass significant ore volumes during leaching.[27] This process involves mixing crushed ore with 2-10 kg per tonne of Portland cement or lime in rotary drums or on conveyors, forming stable nodules that maintain heap porosity above 10-15% even under compaction.[20][28] Empirical tests confirm that proper agglomeration reduces preferential flow paths, with drum agglomeration preferred over conveyor methods for uniform binder distribution in both precious and base metal operations.[27] Heap formation entails stacking prepared ore in sequential lifts of 5-10 meters high on engineered pads to optimize drainage gradients of 1-2% and facilitate solution collection.[26] Permanent heaps use high-density polyethylene (HDPE) geomembrane liners over compacted clay or soil barriers for impermeability, with leak detection systems to monitor containment integrity.[6] Conventional pad designs provide uniform stacking on flat terrain, while valley-fill configurations leverage natural topography for heights exceeding 100 meters in phases, though they demand geotechnical assessment to ensure slope stability and even drainage without ponding.[26][29] Heap geometry influences internal oxygen diffusion rates, critical for oxidative processes, with wider bases and moderate slopes enhancing gas exchange over narrow valley profiles.[26]Leaching Chemistry and Solution Dynamics
Heap leaching relies on specific chemical reactions tailored to the target mineral, where lixiviants dissolve metals into a mobile phase for extraction. For refractory precious metals like gold and silver in oxide ores, alkaline cyanide solutions enable complexation, with gold dissolution governed by the Elsner equation: $4[Au](/page/Au) + 8NaCN + O_2 + 2H_2O \rightarrow 4Na[Au(CN)_2] + 4NaOH.[30] This oxygen-dependent oxidation forms the stable aurocyanide ion [Au(CN)_2]^-, which remains soluble under controlled conditions, allowing percolation through the ore matrix. Silver follows analogous chemistry, forming [Ag(CN)_2]^-, though with higher cyanide consumption due to competing reactions with sulfide impurities.[30] In base metal extraction, particularly copper from oxide ores, sulfuric acid serves as the lixiviant, driving dissolution reactions such as CuO + H_2SO_4 \rightarrow CuSO_4 + H_2O for malachite or azurite, and Cu_2O + H_2SO_4 \rightarrow CuSO_4 + Cu + H_2O for cuprite.[31] For secondary sulfide ores like chalcocite, initial acid attack produces copper sulfate, but primary sulfides such as chalcopyrite require oxidative conditions for efficient leaching: CuFeS_2 + 4Fe^{3+} \rightarrow Cu^{2+} + 5Fe^{2+} + 2S^0.[32] These processes generate copper sulfate in solution, with acid consumption varying by mineralogy—typically 10-50 kg/t ore for oxides.[3] Bioleaching enhances sulfide oxidation in low-grade heaps through chemolithoautotrophic bacteria, notably Acidithiobacillus ferrooxidans, which oxidizes Fe^{2+} to Fe^{3+} ($4Fe^{2+} + O_2 + 4H^+ \rightarrow 4Fe^{3+} + 2H_2O) and elemental sulfur to sulfuric acid, regenerating ferric ions as the primary oxidant.[33] This microbial catalysis accelerates kinetics in acidic environments, with optimal activity at pH 1.5-2.5 and temperatures of 20-40°C, though bacterial consortia including Leptospirillum ferrooxidans improve resilience to high metal loads.[34] Reaction rates depend on bacterial density, often reaching 10^8-10^9 cells/g ore in mature heaps.[35] Control of pH and redox potential (Eh) is critical for reaction selectivity and efficiency. Cyanide leaching maintains pH 9.5-11 to suppress hydrogen cyanide formation (below pH 9.5, free HCN increases toxicity and consumption) while ensuring adequate gold recovery (above pH 11, hydroxide precipitation hinders dissolution).[9] Acid leaching targets pH 1-2 for proton donation and bacterial viability, with Eh >600 mV (vs. Ag/AgCl) favoring ferric regeneration over passivation by elemental sulfur.[36] Deviations, such as Eh drops from ferrous accumulation, slow oxidation kinetics, necessitating oxidant additions like air sparging.[17] The pregnant leach solution (PLS) dynamics involve metal ion buildup and lixiviant recycling, where solution chemistry dictates loading capacities—copper PLS typically reaches 1-6 g/L Cu before extraction, limited by solubility and precipitation risks at higher concentrations.[3] Gold PLS concentrations are lower, often 0.5-5 mg/L Au, due to slower kinetics and dilution from recycle flows.[5] Fluid dynamics influence speciation and transport, with advective flow delivering reagents to reactive sites while diffusive processes govern ion exchange; uneven channeling can reduce effective contact, but agglomeration mitigates this by enhancing uniform percolation and reaction fronts.[37] Barren solution recycle replenishes lixiviant while minimizing freshwater use, though buildup of impurities like iron requires periodic bleeding to prevent gyp-sum precipitation or bacterial inhibition.[38]Operational Mechanics
Solution Application and Percolation
In heap leaching operations, the lixiviant solution is delivered to the top of the ore stack primarily through drip emitters or overhead sprinkler systems to promote uniform wetting and contact with mineral particles.[39][40] Drip systems, favored for their precision and reduced evaporation, apply solution at rates typically ranging from 5 to 10 liters per square meter per hour, minimizing channeling and ensuring even distribution across the heap surface.[39] Sprinkler irrigation, operating at low pressures of 10-20 psi, is employed in scenarios with minimal wind or for higher flow requirements to complete leach cycles efficiently.[41] Application cycles vary by mineral type, with gold heaps often requiring 60-120 days of intermittent irrigation—such as a 70-day primary cycle applying approximately 1.3 tonnes of solution per tonne of ore—while copper operations extend to 90 days or up to several years due to slower dissolution kinetics in oxide or sulfide ores.[26][3][5] Percolation of the lixiviant through the heap occurs under gravity-driven flow, governed by Darcy's law, which quantifies the volumetric flow rate as proportional to the hydraulic gradient and intrinsic permeability of the ore bed while inversely related to fluid viscosity.[42] This model aids in predicting solution dynamics, but real-world heterogeneities in ore packing and particle size can cause uneven flow, resulting in preferential channels, localized flooding that impedes aeration, or dry zones that limit mineral exposure.[43][44] To mitigate these hydraulic inefficiencies, operators employ empirical strategies such as constructing heaps in modular lifts—stacking ore in sequential layers of 5-10 meters—to allow interim drainage, re-agglomeration, and adjusted irrigation for improved saturation uniformity.[27] Operational monitoring focuses on collecting and analyzing pregnant leach solution (PLS) samples from collection points at the heap base, typically at intervals of 8 hours or daily, to track metal concentrations (e.g., gold grades in g/t or copper in g/L) and detect impurity accumulation such as iron or silica that could foul downstream processes.[45] These measurements inform adjustments to irrigation rates or solution chemistry, ensuring percolation efficiency without referencing post-collection recovery steps.[46]Metal Recovery Techniques
In heap leaching operations, metal recovery from the pregnant leach solution (PLS) employs downstream hydrometallurgical techniques optimized for specific metals, focusing on selective separation and purification to maximize yield while minimizing impurities. For copper, the dominant method is solvent extraction-electrowinning (SX-EW), where the PLS undergoes phase separation to isolate copper ions before electrolytic deposition. Solvent extraction uses organic extractants, such as aldoxime-ketoxime mixtures in kerosene diluents, to selectively load copper from the acidic PLS into the organic phase, achieving separation factors over 1000:1 against iron and other impurities. The loaded organic is stripped with dilute sulfuric acid to yield a high-purity electrolyte (typically 40-50 g/L Cu), which feeds electrowinning cells operating at 200-250 A/m² and 1.9-2.1 V, depositing LME-grade copper cathodes at 99.99% purity.[47] [36] SX-EW recovery from PLS routinely exceeds 90-95%, with stage efficiencies of 93-98% per extraction circuit, enabling overall heap leach copper extractions of 70-85% from oxide ores when combined with upstream dissolution kinetics.[48] [3] Similar SX-EW variants apply to nickel, using extractants like Cyanex 272 for selective recovery from sulfate or chloride PLS, producing cathodes via electrowinning at efficiencies above 90% under controlled pH (4-6).[49] For gold and silver in cyanide PLS, activated carbon adsorption predominates, often via carbon-in-columns (CIC) for clarified solutions or carbon-in-pulp (CIP) adaptations for slurries. In CIC, PLS flows through 3-5 m diameter columns packed with 6x16 mesh coconut-shell activated carbon, adsorbing aurocyanide complexes (Au(CN)₂⁻) via physical and chemisorption, with loading capacities of 1000-5000 g/t Au under ambient conditions. Loaded carbon undergoes elution with 1-2% NaOH at 110-140°C and 50-100 kPa, followed by electrowinning in stainless-steel cells at 2-5 V, yielding doré bullion at 90-99% recovery per cycle. CIC/CIP efficiencies from PLS reach 95-99%, supporting total heap recoveries of 60-85% for amenable ores.[50] [51] The Merrill-Crowe process serves as an alternative precipitation method for precious metal PLS, particularly where organic fouling limits carbon use or silver predominates. After filtration and deaeration to <0.5 mg/L O₂, zinc dust (finely powdered with lead promoters) is added to precipitate metals via cementation:This exothermic reaction drives >99% depletion of Au/Ag from PLS at pH 10-11, with the zinc precipitate filtered, fluxed, and smelted to bullion; recoveries exceed 95% but require high reagent consumption (0.5-2 kg Zn per kg Au).[52] [53] Compared to carbon methods, Merrill-Crowe excels in high-cyanide or impurity-laden solutions but incurs higher operating costs due to zinc handling.[49]