Stirred milling kinetics of siliceous goethitic nickel laterite for selective comminution

The objective of this study is to determine how grinding conditions affect the breakage rate with respect to the sample mass, major elements, and minerals present in siliceous goethitic (SG) nickel laterite. This information is helpful in determining the optimal grinding conditions for selective comminution and nickel upgrade. The kinetics of batch wet grinding of nickel laterites with feed sizes of 2.38–1.68, 1.68–1.18, 1.18–0.85, 0.85–0.6, 0.6–0.42, 0.42–0.3, 0.3–0.21, and 0.21–0.15 mm were determined using a Netzsch LME4 stirred mill under the following conditions: 1000 rpm, 50% charge volume, 150.0 g of solid. The grinding behaviour of the majority of the feed samples was non-first-order due to the fast breakage rate of soft minerals and the low breakage rate of hard minerals in the feed. Therefore, an enrichment of the soft mineral was obtained in the underscreen product by selective grinding. The effect of selective grinding on Ni upgrade was evaluated by looking at grinding time, feed size, and product size. Optimum grinding time with respect to Ni upgrade was 0.25 min for SG nickel laterite samples. Generally, grinding larger particles and/or collecting finer product size yielded better Ni upgrade results. The effect of selective grinding was evaluated by the changes of the major soft and hard minerals for the selected samples. Selective grinding was also examined with respect to the major element weight ratio (e.g. Si/Ni for SG nickel laterite). With respect to Ni upgrade, the best result was achieved from the 1.18–0.85 mm feed on the −400 mesh product after grinding for 0.25 min. The Ni grade increased from 0.73% to 1.30% (upgrade 76.8%), with 14.4% Ni recovery; the Mg grade increased from 1.30% to 3.96% (upgrade 205.6%); the Si grade decreased from 28.7% to 16.2%.     

Agglomeration and column leaching behaviour of goethitic and saprolitic nickel laterite ores

Processability of complex, low-grade nickel (Ni) laterite ores via heap leaching is very limited due to some intractable geotechnical and hydrological challenges such as poor heap porosity/permeability and structural stability. This work presents some investigations on laboratory batch drum agglomeration and continuous column leaching behaviour of saprolitic (SAP) and goethitic (G) Ni laterite ores as part of the quest for an effective ore pre-treatment process for enhanced heap leaching. As a focus, the effect of ore mineralogy/chemistry on the agglomeration and column leaching behaviour of −2 mm (crushed from −15 mm run-of-mine) G and SAP Ni laterite ores was examined. To produce ∼5–40 mm agglomerates in <15 min, the SAP ore required a higher H₂SO₄ (30 wt.%) binder dosage compared with the G ore, although both ores displayed substantially similar, coalescence-controlled agglomeration mechanism. The resulting G agglomerates were more robust than the SAP ones based upon their compressive strength and acidic solution soak test measurements. However, over 100 days of continuous column leaching, the structural stability of the SAP agglomerate bed was slightly greater than that of G agglomerates, reflecting a lesser slump of the former. The pregnant leach solution analysis revealed greater Ni/Co extraction rates from the SAP than the G agglomerates. Whilst the total mass of acid consumed per ton dry ore processed was greater for the SAP ore, the total kg acid per kg Ni extracted was markedly lower. Incongruent leaching of gangue minerals’ constituent elements (e.g., Fe, Mn, Mg, Al and Si) occurred and contributed significantly to the overall acid consumption. The findings show the relevance of agglomeration and column leaching tests for providing useful information for plant designing and optimization of Ni laterite heap leaching operations.     

Selective sulphidation and flotation of nickel from a nickeliferous laterite ore

The sulphidation of a nickeliferous lateritic ore was studied at temperatures between 450 and 1100 °C and for sulphur additions of 25–1000 kg of sulphur per tonne of ore. The experiments demonstrated that the nickel could be selectively sulphidized to form a nickel–iron sulphide. It was found that both the grade and the sulphidation degree largely depended upon the temperature and the sulphur additions, with temperatures above 550 °C exhibiting the highest nickel sulphidation degrees and grades. A DTA/TGA with mass spectrometer was used to further elucidate the nature of the phase transformations that occurred upon heating of the ore in the presence of sulphur.It was found that at low temperatures, the Fe–Ni–S phase was submicron in nature and heating to temperatures between 1050 and 1100 °C allowed for the growth of the particles, due to the increased sulphide mobility associated with the formation of a liquid sulphide matte phase, containing dissolved oxygen. Flotation studies conducted on 60 g samples showed that the sulphides responded to flotation with maximum grades of up to 6–7 wt.% nickel being achieved. Recoveries were approximately 50% on a sulphide basis and it was determined that the low nickel grades were due to the entrainment of magnetite fines.     

Atmospheric acid leaching of siliceous goethitic Ni laterite ore: Effect of solid loading and temperature

In this study, atmospheric acid leaching behaviour of siliceous goethitic nickel (Ni) laterite ore is investigated. Specifically, the effect of −200 μm feed solid loading (30 vs. 45 wt.%) and temperature (70 vs. 90 °C) on leach kinetics, acid consumption capacity and Ni and cobalt (Co) extraction was studied under isothermal, batch (4 h) leaching conditions at pH 1. Incongruent leaching was observed for constituent elements reflecting slow but steady release of value (Ni and Co) and some of gangue metals such as Fe, Mg and Al accompanied by faster and sharp release of Na and Si. Higher temperature and lower pulp solid loading, both led to a 40–50% increase in overall Ni and/or Co extraction and higher acid consumption. At 70 °C and 45 wt.% solid loading, Ni/Co extraction after 4 h was the lowest (∼14/16%) whilst the highest extraction (∼67/56%) was observed at 90 °C and 30 wt.% solid loading. Temperature appeared to have dramatic influence on Ni/Co and other impurity metals’ extractions revealing the chemical reaction controlled nature of the leaching. Higher solid loading and longer leaching time also both slowed down the leach kinetics. A two-stage chemical reactions-controlled leaching mechanism involving a faster initial leaching kinetics followed by a slower leaching at lower rate constants and higher activation energies was established for release of Ni, Co, Fe and Mg. The mechanism reflects the fast leaching of reactive host mineral phases (e.g., clays and Mg–silicates) during first 30 min followed by slow leaching of more refractory mineral phases (e.g., goethite and quartz) during the rest of leaching period. The findings provide a greater understanding for enhanced atmospheric acid leaching process of siliceous goethitic laterite ores.     

Acid leaching and rheological behaviour of a siliceous goethitic nickel laterite ore: Influence of particle size and temperature

This study investigates the isothermal, batch, H₂SO₄ acid leaching behaviour of siliceous goethitic (SG) nickel (Ni) laterite ore and its links to pulp rheology. Specifically, the effect of feed ore particle size (−0.2 vs −2.0 mm), leaching temperature (70 vs 95 °C) and pulp rheology on Ni and pay metal, cobalt (Co) extraction kinetics and yield was studied for 4 h on 40 wt.% solid dispersions at pH 1. The leaching behaviour was distinctly incongruent, reflecting the disproportionate proliferation of major gangue mineral’s constituent elements (e.g., Fe, Al, Mg, Na, Si) alongside Ni and Co in the pregnant leach solution. At 70 °C, Ni/Co extraction rates were notably lower (<20%) in contrast with 95 °C where a significant increase in Ni/Co extraction to 78/77% and 74/77%, respectively, for the −0.2 and −2.0 mm feeds occurred. The slurries displayed a non-Newtonian, shear thinning Bingham plastic rheological behaviour of which the viscosity and shear yield stress increased markedly in the course of 4 h leaching. The pulp viscosity and shear yield stress were greater at lower temperature than at higher temperature and they were also greater in slurries with finer than coarser feed particles. The dynamic pulp rheology, however, had no marked effect on the overall Ni/Co extraction rates. Whilst the feed ore particle size had no remarkable impact on overall Ni/Co extraction, it led to noticeably higher acid consumption and enhanced slurry rheology in the finer sized ore. The mechanism of leaching the SG ore followed a two-stage, first order chemical reaction-controlled shrinking core model, the kinetics of which gave higher rate constants and lower activation energies for the release of Ni, Co, Fe and Mg in the first stage. A faster leaching process involving more reactive minerals during the first 30 min is envisaged to be followed by leaching of the more refractory minerals.     

A study of atmospheric acid leaching of a South African nickel laterite

Nickel and cobalt acid leaching from a low-grade South African saprolitic laterite using sulphuric acid was studied. Ore characterisation was performed by XRD and XRF. Batch agitation leaching tests were conducted at atmospheric pressure investigating main parameters: particle size and percent solids at 25 °C and 90 °C. Ore characterisation showed that the ore is a saprolitic laterite with nickel present in lizardite. Leaching tests showed that nickel and cobalt could be leached from the ore at atmospheric pressure. Nickel was found to be more leachable from the coarser −106 + 75 μm fraction, with 98% Ni being extracted at 90 °C after 480 min. Cobalt was not favoured by variation in particle size and increased percent solids. Increasing ore percent solids improved nickel extraction at 25 °C however at 90 °C extraction decreased due to a diffusion layer build-up as a result of amorphous colloidal silica. The co-dissolution of magnesium and iron was elucidated. Nickel leaching data at increased temperature and percent solids fit the shrinking core model equation, k_dt = 1−2/3x − (1 − x)^2/3 showing that nickel leaching reaction was diffusion controlled under the set conditions.     

The effect of sodium sulphate on the hydrogen reduction process of nickel laterite ore

A series of nickel laterite ores with different calculated amounts of anhydrous sodium sulphate were prepared by physical blending or sodium sulphate solution impregnation. The reduction of the prepared nickel laterite ore by H₂ was carried out in a fluidised-bed reactor with provisions for temperature and agitation control, and the magnetic separation of the reduced ore was performed using a Davis tube magnetic separator. The mineralogical properties of the raw laterite ore, reduced ore and magnetic concentrate were characterised using ICP, TG–DSC, N₂ adsorption, X-ray diffraction and optical microscopy. The catalytic activity of sodium sulphate was also studied by using Hydrogen temperature-programed reduction. The experimental results indicate that Na₂SO₄ could overcome the kinetic problems faced by the laterite ore and that it exhibited noticeable catalytic activity only if the temperature reached at least 750 °C. This high temperature accelerated the crystal phase transition of the silicate minerals and increased the utilisation of H₂. In comparing the results from the two different methods for adding Na₂SO₄, the nickel content and recovery of the magnetic concentrate were increased by using the impregnation method rather than the physical blending method and the increasing amount of sodium sulphate assisted in the further beneficiation of nickel. The partial pressure of H₂ and the reducing time also affected the reduction process of the iron oxides. The results of the microscopic study indicated that the formation of a Fe–S solid solution, which was derived from the SO₂ sulphide reduction of FeO, was conducive to mass transfer and accelerated the coalescence of metallic ferronickel particles. For the nickel laterite ore, under the synergistic effect of sodium sulphate and hydrogen, a nickel content and nickel recovery of 6.38% and 91.07% were obtained, respectively, with high product selectivity.     

Kinetics of saprolitic laterite leaching by sulphuric acid at atmospheric pressure

Mineralogical analyses of the saprolitic laterite material have been characterized by X-ray fluorescence (XRF), X-ray diffraction (XRD), Fourier transform infrared (FTIR) spectroscopy, thermal analysis, scanning electron microscope (SEM) and energy dispersive X-ray analysis (EDAX). Results showed that the saprolitic laterite material consists mainly of nickel-substituted lizardite showing the pebble-like morphology and traces of magnetite and phlogopite. Leaching results showed that as much as 84.8% nickel could be leached under the experimental conditions of 10% (v/v) H₂SO₄, 90 °C reaction temperature, leached within 5 min, particle size d₅₀ = 25 μm, stirring at 500 rpm and liquid to solid ratio 3:1. The kinetics of nickel and magnesium leaching from the saprolitic laterite material have been investigated in a mechanically stirred reactor and the activation energies were determined to be 53.9 kJ mol^?1 for nickel and 59.4 kJ mol^?1 for magnesium respectively, which are characteristic for a chemical reaction controlled process. The similarity of the activation energies of nickel and magnesium leaching from the saprolitic laterite material by sulphuric acid means that nickel in lizardite is loosely bound within the octahedral layer and almost all of the nickel could be leached simultaneously with magnesium but without complete decomposition of the silicate structure.     

Innovative technology for processing saprolitic laterite ores by hydrochloric acid atmospheric pressure leaching

An innovative technology for processing saprolitic laterite ores from the Philippines by hydrochloric acid atmospheric leaching and spray hydrolysis is proposed. The factors that affect the hydrochloric acid atmospheric leaching of the laterite ores and spray hydrolysis of the atmospheric acid leach solution are investigated. Experimental results show that the leaching of Ni, Fe, and Mg is 98.9 wt%, 97.8 wt%, and 80.9 wt%, respectively, under optimal acid leaching conditions. The hydrolysis of Ni and Fe by the atmospheric acid leach solution approaches 100 wt% at the temperature range of 450–500 °C. Characterization results show that a serpentine mineral, nominally Mg₃Si₂O₅(OH)₄, is the major component and goethite, FeO(OH), is the minor one in the laterite ores. Treatment by hydrochloric acid atmospheric leaching breaks the mineral lattices of the laterite ores and makes amorphous silica the primary product in the atmospheric acid leach residue. The grade of Ni in the hydrolyzate increases to 4.55%. The hydrolyzate with high Ni content can be utilized for ferro-nickel production.     

Use of oxidation roasting to control NiO reduction in Ni-bearing limonitic laterite

Use of limonitic laterite as an iron source in conventional ironmaking is restricted due to its gangue composition and small particle size. Even direct reduction cannot effectively produce direct reduced iron (DRI) because NiO would be reduced together with iron oxide to form Fe–Ni. A small amount of Ni (about 2 wt.%) in DRI degrades the physical properties of final steel products. The current study investigated how oxidation roasting of limonitic laterite ores affected NiO reduction, with the goal of producing Ni-free DRI and Ni-bearing slag. Ni-bearing slag can be a good secondary Ni resource. Oxidation roasting made NiO inert under H₂ reduction at 900 °C by forming Ni-olivine. Optimum roasting temperature was proposed by examining phase transformation of limonitic laterite ores during heating and by FactSage calculation of the equilibrium Ni fraction in Ni-bearing phases. Furthermore, the effect of Mg–silicate forming additives on the control of NiO reducibility was clarified to maximize the suppression of NiO reduction. Among various additives such as MgSiO₃, Mg₂SiO₄ and Fe–Ni smelting slag, Ni-free olivine-typed flux was found to be the most effective form of Ni-olivine because Ni–Mg ion exchange between Ni-bearing phase and Ni-free olivine occurs more readily than other Ni-olivine formation schemes. Finally, the mechanism of Ni-olivine formation during roasting was studied using a diffusion couple test. Calculated diffusivity values of Ni in Mg₂SiO₄ indicated that the two major routes of Ni-olivine formation while roasting limonitic laterite ore are (1) Ni partitioning within Mg–Ni silicate before crystallization and (2) Ni diffusion from spinel to Ni free olivine after crystallization.     

Preparation of layered nickel aluminium double hydroxide from waste solution of nickel

The separation of nickel has been carried out from a waste solution containing 3.18 g/L Ni with other impurities such as Fe, Zn, Cu and As. Iron was removed by precipitation and Cu and Zn were removed by solvent extraction using LIX 622N and NaTOPS-99, respectively. After removal of all these impurities nickel was extracted by 1.5 M NaTOPS-99 in two counter-current stages at A:O ratio of 3:1 and the loaded organic was stripped with 30 g/L H₂SO₄ at phase ratio of unity. The strip solution of nickel was treated with Al₂(NO)₃ · 9H₂O for co-precipitation by increasing the pH of solution with 1 M NaOH up to 10. The Ni–Al layered double hydroxide was confirmed through XRD characterization.     

印度尼西亚某红土镍矿还原焙烧-磁选试验

印度尼西亚某低品位红土镍矿含镍1.57%、含铁21.67%,其中镍主要以硅酸镍形式存在。为将该矿石的镍含量提高到6%以上以符合印度尼西亚政府对出口红土镍矿的规定,以硫酸钠和碳酸钠为助熔剂,进行了还原焙烧-弱磁选试验。试验结果表明,当煤用量为25%、硫酸钠+碳酸钠的配比和总用量分别为3∶1和20%、焙烧温度为1 200 ℃、焙烧时间为60 min、磨矿细度为-0.074 mm占85%、磁场强度为96 kA/m时,可获得产率为22.06%、镍品位为6.05%、镍回收率为85.03%、铁品位为65.74%、铁回收率为66.92%的镍铁精矿,其镍品位超过印度尼西亚出口红土镍矿的品位下限。     

Selective pressure leaching of Fe (II)-rich limonitic laterite ores from Indonesia using nitric acid

The selective extraction of nickel and cobalt over iron from an Indonesian limonitic laterite was investigated using nitric acid pressure leaching (NAPL). The mineralogical analysis showed that the major minerals were goethite and magnetite, and the content of the divalent iron was as high as 7.06%. Nickel and cobalt were mainly distributed in these two minerals; however, the distribution was non-uniform. A series experiments were conducted to examine the basic parameters and propose the optimal conditions for the extraction. When the ore was treated via HPAL under the optimal condition, the extracted nickel and cobalt were less than 75%, and the iron concentration in the leach liquor was over 12.5 g/L. By contrast, over 85% of nickel and cobalt were extracted and about 1.8 g/L iron was achieved using NAPL. The loss of nickel and cobalt can be mainly attributed to the undissolved magnetite and manganese minerals. The leaching process of NAPL is a dissolution–oxidation–precipitation mechanism, and in this process nitric acid acts as both a lixiviant and an oxidant. The formation of hematite results in a low iron concentration in the leach liquor without oxygen injected. Meanwhile, the oxidation and the precipitation of dissolved divalent iron results in a calculated savings in acid consumption of about 120 kg nitric acid per ton of ore can be obtained, which is equal to over 93 kg of sulfuric acid per ton of ore. Moreover, lower residual acid (20 g/L nitric acid) is a significant advantage of NAPL. The iron residues had a high iron content (>56 wt%) with no sulfur, making it suitable as raw materials for ironmaking.     

红土镍矿深度还原-磁选试验研究

采用深度还原-弱磁-强磁工艺对低品位红土镍矿进行了开发利用研究,重点研究了深度还原合适的温度、还原时间、配碳系数、料层厚度、强磁精矿返回量等参数。研究表明,适宜的深度还原条件为：还原温度1 275 ℃、还原时间50 min、配碳系数2.5、料层厚度25 mm、强磁精矿返回量占原矿量的25%,还原产物经弱磁选（场强为130 kA/m）,可获得镍、铁品位分别为6.96%、34.74%,镍、铁总回收率分别为94.06%、80.44%的优质镍铁精矿产品;同时富含大量细小镍铁颗粒的强磁精矿是红土镍矿深度还原的优质成核剂。     

Thermodynamic analysis of the Fe–Ni–Co–Mg–Si–O–H–S–C–Cl system for selective sulphidation of a nickeliferous limonitic laterite ore

An increasing percentage of nickel is being extracted from the laterite ores for which there exists considerable potential for new process development. In the current research, a thermodynamic analysis of the Fe–Ni–Co–Mg–Si–O–H–S–C–Cl system has been performed in order to establish the possible sulphidation conditions for the upgrading of a limonitic laterite ore. The study was performed using the equilibrium module of HSC Chemistry® 6.1, and the data available in the literature were utilised to determine the activity coefficients of the various species employed in the calculations. The effects of variables such as temperature, amount of sulphur and carbon and chlorine additions on both the nickel grade and recovery in the monosulphide solid solution (mss) were determined. Alternative sulphidizing agents were also considered. Nickel recoveries of about 85% could be achieved at a grade of 10% nickel in the mss. Limiting factors in the sulphidation process were the formation of iron sulphide (FeS) and the relatively high stability of nickel oxide in the nickel ferrite (NiFe₂O₄), in comparison to the other phases, which resulted in excessive sulphur requirements of over 100 kg of S/tonne of ore. Methods to overcome these restrictions were discussed and the thermodynamic results were compared to the published experimental values.     

还原-磨选从高镁低品位红土镍矿中回收镍铁

采用还原-磨选工艺对高镁低品位红土镍矿制备镍铁合金粉进行了研究。考察了还原温度、还原时间、原料粒度区间、还原剂用量、添加剂用量等因素对镍直收率的影响。研究结果表明, 合适的还原制度为: 原料粒度0.09～0.12 mm、还原剂用量3%、添加剂用量2.5%, 还原温度1 300 ℃, 还原时间3.0 h。还原产物经球磨、磁选后, 获得镍品位为7.0%以上的镍铁合金粉, 镍直收率87%以上, 实现了从高镁低品位红土镍矿中回收镍铁的目的。     

A hydrochloric acid process for nickeliferous laterites

The development of a hydrometallurgical hydrochloric acid based leaching process for the treatment of nickeliferous laterites of various types is described. Work carried out in the Department of Mining and Mineral Engineering, Leeds University, U.K. over a number of years, on the development of a route for treating all types of lateritic nickel ore, is reported in fairly general terms. This paper presents an update on the work carried out and a discussion of a possible final process. Conceptually the process comprises leaching of the ores in azeotropic HCl; counter-current washing to maximise nickel recovery; separation of iron either by precipitation as goethite or by solvent extraction; cobalt recovery by solvent extraction with a tertiary amine; recovery of nickel by solvent extraction followed by electrowinning, hydrogen pressure reduction or pyrohydrolysis and regeneration of hydrochloric acid by pyrohydrolysis with by-product recovery of magnesia some of which might be used for acidity control in the process. Attempts to leach successive batches of ore in a cyclic or counter-current manner to build up metal concentration were unsuccessful due to passivation by formation of a product layer around the ore particles. A key process step is concentration of the nickel in the leach liquor (1–5 g/L) and its separation from magnesium and/or iron to a tenor compatible with recovery methods. This can be achieved using solvent extraction with Cyanex®301, Cyanex®302 or Versatic 10, the last of which displayed the best extraction characteristics. Cyanex®301 showed excellent selectivity for nickel but stripping required high concentrations of acid whilst Cyanex®302 showed appreciable co-extraction of magnesium. Whilst some of the process steps have been thoroughly researched, others remain to be proven and it was not possible to test the process as a whole or to carry out an energy balance, so that nothing can yet be stated definitively regarding the economics of the route but some factors regarding energy consumption in pyrohydrolysis are discussed. Relevant up to date literature on chloride based hydrometallurgical processes for laterites has been briefly reviewed. The contributions of individual research workers are acknowledged at appropriate points in the text.     

Leaching of limonitic laterite ore by acidic thiosulfate solution

Cobalt is usually recovered as a by-product of copper and nickel processing, and only a small amount of cobalt is derived from laterites although a vast majority of cobalt resources in them. The exploitation of limonitic laterite containing high content of cobalt is becoming increasingly important. The mineralogy of a limonitic laterite ore was characterized by environmental scanning electron microscope (ESEM) and X-ray diffraction (XRD) in this paper. The results show that nickel occurs in goethite mainly, while cobalt is predominantly associated with manganiferous minerals. Thiosulfate is found to be able to selectively leach cobalt from limonitic laterite in the presence of sulfuric acid, and 91% Co, 22% Ni, 10% Fe are leached from an ore containing 0.13% Co, 1.03% Ni within the first 5 min at 90 °C under the conditions of 10 g/L sodium thiosulfate, 8% (w/w) sulfuric acid and 10:1 L/S ratio. The leaching kinetics of Mn and Co by acidic sodium thiosulfate solution can be characterized by the Avrami equation. In acidic solution, thiosulfate readily decomposes into sulfur and sulfur dioxide as intermediary reagents to reduce pyrolusite (MnO₂) and goethite (FeOOH); therefore, nickel and cobalt associated with goethite and pyrolusite respectively are extracted due to reduction dissolution. Furthermore, cobalt is selectively leached over iron and nickel because pyrolusite is preferentially reduced by acidic thiosulfate rather than goethite. The novel process may give an alternative method to selectively recover cobalt as the primary product from limonitic laterites at atmospheric pressure.     

Solubilization of heavy metals from gold ore by adjuvants used during gold phytomining

Adjuvants such as cyanide, thiocyanate, and thiosulfate are applied to gold-bearing ores to increase gold solubility and plant uptake of gold during phytomining. The influence of these three adjuvants, added at rates comparable to field application (1 g kg DW^?1 ore), on the leaching of heavy metals from gold ore from the Davis stockpile at the Stawell Gold Mines was examined in batch and column studies. The overall goal was to provide data that specifically addressed the possible collateral impacts that could occur during adjuvant-assisted phytomining in terms of heavy metal solubilization and leaching. Leachate from the cyanide-amended ore had the highest concentrations of Au, Co, Cu, and Ni, with the Au concentration representing ～22% of the total gold present. Thiosulfate-amended ore leached the highest Fe concentration while thiocyanate-amended ore leached the highest Al and Zn concentrations. The results suggest that for this ore, while these adjuvants are present, there is the potential for the mobilization of undesirable elements and Au out of the ore and potentially into groundwater unless adequate hydraulic controls are implemented following adjuvant addition.