Investigation of particles with high crack density produced by HPGR and its effect on the redistribution of the particle size fraction in heaps

The application of comminution technology such as the high-pressure grinding rolls (HPGRs), which is able to generate a high density of cracks in the ore particles, is favourable for leaching processes. Extraction of metallic values by the heap leach process, can take place on the particles with partial exposure of mineral grains, if it can provide sufficient surface front for chemical attack by leaching solution. The aim of this study was to assess the benefits of high crack density in the ore particles produced using the HPGR and how it could diminish due to inadequate percolation of the leaching agent.A zinc ore was comminuted using HPGR at three different pressure settings and with a cone crusher for the control experiment. Subsamples from the (+23/−25, +14/−16, +5.25/−6.75 mm) size fractions were characterized and packed into leach reactors. The reactors were stopped from time to time to investigate the progress of crack and micro-crack growth and its effect on metal extraction using the X-ray computed tomography (CT). The results are validated with those obtained using traditional techniques such as SEM and QEMSCAN. Investigation of the leach reactors residue indicated significant changes in the particle size distribution (PSD) of initial feed toward the fine size fraction. The residues from the reactors leaching the material prepared using the HPGR product contained more fine particles than the reactors, which were fed by cone crusher product. These differences were up to 10.3%.     

Ferric sulphate/chloride leaching of zinc and minor elements from a sphalerite concentrate

Atmospheric leaching of a sphalerite concentrate in sulphate and chloride media was performed and the effect of several variables, such as solid/liquid ratio and oxidant (Fe(III)) concentration were investigated. The behaviour of minor elements, such as Cu, In, As, Sb, Bi, Sn and Pb, was also studied under different conditions. The results showed that using a solid/liquid ratio of 5% (w/v) it was possible to leach 95% of zinc after 2 h, with a solution of 0.5 M H₂SO₄ and Fe₂(SO₄)₃ at 80 °C. The minor elements As, Sb and Bi were also completely leached whereas copper leaching was favoured by the use of chloride medium. The oxidation of Fe(II) during the leaching tests was studied and an improvement of 20% zinc extraction was observed in an oxygenated system. Cross-current leach tests using two/three stages and a solid/liquid ratio of 10% (w/v) were performed to achieve 90% of zinc extraction. The electron microprobe analysis of the leaching residues showed no change on the sphalerite composition after the leaching, which indicates that the leaching of sphalerite involves the break down of the sulphide structure.     

Enhanced leachability of gold and silver in cyanide media: Effect of alkaline pre-treatment of jarosite minerals

Composite samples of tailings containing gold (1.35 g/t) and significant amounts of silver (155 g/t) were subjected to batchwise cyanide leaching to assess the feasibility of extracting gold and silver. The tailings are waste solids arising from flotation and leaching operations whereby the flotation product (sphalerite concentrate) is calcined and then solubilised into dilute sulphuric acid solution and eventually sequestered from the electrolyte by electrowinning. Silver and gold are part of the zinc refinery residue, flotation tailings and to a limited extent the calcine leach tailings. Mineralogical results showed that composite tailings are refractory in nature (44% quartz, 17% silico aluminates and 12% jarosites).The concept of enhancing gold and silver recovery from the tailings focused on firstly decomposing the jarosite minerals by alkaline pre-treatment and then secondly leaching with cyanide solution. These two steps ensured that free gold and silver found in the zinc refinery residue and in the jarosite minerals could be leached simultaneously. The composite tailings were treated with Ca(OH)₂ solutions and then heated to 90 °C for 2 h to decompose the silver-bearing mineral (Ag,PbFe₃(SO₄)₂(OH)₆). The alkaline pre-treated tailings were then subjected to cyanide leach tests at different NaCN dosages (2.5–10 kg/t) and particle size (96–200 μm). Without an alkaline pre-treatment stage, leach efficiencies achieved were 41% and 25% for gold and silver, respectively at 40 °C and 8 h mixing time. But, better leach efficiencies (55% for Au, 81% for Ag) were achieved after the feed was pre-treated with Ca(OH)₂. The leaching mechanism of gold was explained by the shrinking sphere model denoted by surface chemical reaction.     

Investigating heap bioleaching: Effect of feed iron concentration on bioleaching performance

This paper describes an investigation into the effect of iron concentration in the leach solution on the bioleaching of a low grade copper ore, where chalcopyrite was the dominant copper sulphide. The concentration of dissolved iron is primarily controlled by pH and the relative proportion of ferric to ferrous iron, with significant jarosite precipitation occurring above pH ≈ 1.8 in a highly oxidised system. The solution pH may be increased by the dissolution of acid soluble gangue and when iron oxidation is significantly higher than sulphur oxidation. The study was approached using two experimental systems. In the former, the leach solution was recycled through an ore bed of low aspect (reactor height divided by diameter) ratio for a portion of the experiment. During the recycle phase, no acid was added to the system and acid consumption by gangue material led to a pH increase (1.6–2.2). The resulting jarosite precipitation reduced soluble iron from 2.5 g/l to less than 250 mg/l. Copper recovery decreased, but not in proportion to the decrease in iron. This was partly attributed to adsorption on, or entrainment within, the jarosites. To study the effect of reduced iron concentration on leach performance under more controlled conditions, bioleaching was performed in packed bed column reactors with feed iron concentrations ranging from 5 g/l to 200 mg/l. Observations indicated an initial decreased rate of copper liberation with reduced iron concentration in the feed. The relationship between available Fe³⁺ concentration and copper liberation was not proportional. However, with time, the liberation of copper became independent of iron concentration in the percolation liquor. Further, the specific rate of copper liberation was consistently below the theoretical value on a basis of ferric iron concentration. The highest values of copper liberation were reported at the lowest iron concentrations. In summary, while increased iron concentration in solution may enhance the initial rate of leaching, mineral availability appears to dominate CuFeS₂ leach kinetics through the majority of the leach. Furthermore, high iron concentrations in solution aggravate jarosite formation with concomitant retention of copper in the ore bed.     

Effect of water quality on the leaching of a low-grade copper sulfide ore

Copper extractions from a low-grade, ground copper sulfide ore (0.7% Cu) leached in three media were freshwater < seawater > double-strength seawater and pH 1.5 ⩾ pH 2; 84% extraction was achieved in pH 1.5 seawater in 28 days at 23 °C. Cu-oxide and carbonate dissolved completely and chalcocite was altered to secondary covellite, some of which persisted in all media for the duration of the 28-day experiment. Chalcopyrite and bornite were both oxidised more readily in saline water. Iron, sodium, potassium and sulfur (sulphate) concentrations in leach solutions diminished and the amounts of insoluble iron(III) reaction products increased with increased salinity and increased solution pH. While, overall, silicate dissolution was small, the amounts of poorly crystalline phases (both iron(III) and silica-rich phases) increased with increased salinity and were greater in pH 1.5 media. In the context of heap leaching, the increased amounts of secondary precipitates formed if saline water was used could result in lower extraction efficiency and the increased total dissolved solids, density and viscosity could result in increased energy costs for solution management at operations.The software package Geochemist’s Workbench was evaluated by modelling the synthetic seawater – pH 2 test. It was possible to predict the evolution of the solution composition, the main species and phase boundaries at the start and end of leaching, and the formation of three reaction products in accord with experimental data by applying the React sliding function.The tests were conducted using a pulverised ore sample to increase dissolution reaction kinetics, particularly for chalcopyrite. Future tests should be conducted using ore particle sizes appropriate to heap leaching. The copper distribution within particles indicated that the test ore may not be suited to heap leaching because the surface exposure of copper sulfide grains is limited. Therefore reactor designs better suited to smaller sized particles with/without pre-treatment should be considered.     

Simulated small-scale pilot plant heap leaching of low-grade oxide zinc ore with integrated selective extraction of zinc

The leaching of low-grade oxide zinc ore and simultaneous integrated selective extraction of zinc were investigated using a small-scale leaching column and laboratory scale box mixer-settlers. Di-2-ethylhexyl phosphoric acid (D2EHPA) dissolved in kerosene was used as an extractant. The results showed that it was possible to selectively leach zinc from the ores by heap leaching. The zinc concentration of the leach liquor in the first leaching–extraction circuit was 32.57 g/L, and in the 16th cycle the zinc concentration was 8.27 g/L after the solvent extraction. The leach liquor was subjected to solvent extraction, scrubbing and selective stripping for the enrichment of zinc and the removal of impurities. The pregnant zinc sulfate solution produced from the stripping cycle was suitable for zinc electrowinning.     

Phase distribution identification in the column leaching of low grade ores using MRI

Heap bioleaching is gaining importance as an approach for the recovery of valuable metals (e.g. Cu²⁺) from low grade ores. In this process iron and/or sulfur oxidising microorganisms are used to aid the oxidation of base metal sulfides in the ore, thereby liberating the metal ions into solution. Leach performance is strongly influenced by the contacting of the leach solution and the ore particles. In order to better understand the distribution of the leaching solution on the pore scale in these heaps, Magnetic Resonance Imaging (MRI) was used to acquire images non-invasively of a section of an irrigated ore bed. This was made possible by the use of specialist MRI acquisition sequences suited to the magnetically heterogeneous environment as presented by the ore material. From the images we were able to determine the pore-occupancy of the liquid and gas phases and to provide novel measurement of the interfacial area between air, leach solution and ore.     

Electrochemical passivation of sphalerite during bacterial oxidation in the presence of galena

Sphalerite (ZnS), a mineral of particular hydrometallurgical interest, is generally found associated with galena (PbS). As such, the effect of galena on the leaching of sphalerite is of importance, and has been studied here for bacterial and ferric sulphate leaching. During both leaching processes galena was selectively oxidised to anglesite in favour of the dissolution of sphalerite. It is believed that this selective behaviour is due to galvanic interactions between the two minerals, whereby galena is sacrificed and sphalerite is passivated. This theory is consistent with the order of measured rest potential values of both minerals in solution, being 325 mV for galena and 375 mV for sphalerite (versus a standard hydrogen electrode). During bacterial oxidation, sphalerite passivation was observed across a range of mixed sphalerite/galena samples, including a mineral species and four ore specimens of varying grades. From the bacterial oxidation of a mixed mineral species, sphalerite was found to leach in the presence of lead sulphate precipitate, though the presence of this precipitate is believed to have caused diffusion limitation.     

低品位氧化铜矿柱浸试验研究

以某低品位氧化铜矿为研究对象, 通过实验室小型柱浸试验, 对柱浸前不同熟化加酸量和柱浸喷淋强度的影响进行了研究。结果表明: 喷淋强度和柱浸前的熟化是影响柱浸效果的关键因素。当喷淋强度为15 L/(h·m²)时, 浸出剂在柱中很好地扩散并与矿石发生反应, 此时Cu浸出率和浸出速率都较高;柱浸前的熟化不仅可提高Cu总浸出率, 还可显著加快柱浸前期Cu浸出速率。熟化过程中浓硫酸与矿石中某些组分发生化学反应, 增加了矿石中酸溶铜含量, 同时导致矿石裂隙扩大, 孔隙数量增加, 使得一些被包裹的难以与硫酸接触的铜裸露出来, 从而提高浸出率。     

表面活性剂对锌精矿在浆萃取浸出率的影响

研究了几种不同的表面活性剂对锌精矿有机溶剂在浆萃取过程中锌浸出率的影响 ,并对表面活性剂的用量对锌精矿在有机溶剂中的行为进行了探讨 ,测定了经表面活性剂处理后矿粉在水溶液中的粘度以及浸出溶液表面张力变化情况 ,结果表明加入 0 .0 5～ 0 .2 g/L表面活性剂十二烷基磺酸钠改变了矿粉在水溶液中的润湿性 ,降低了浸出溶液的表面张力 ,有效地提高了锌精矿浸出率     

Bioleaching and chemical leaching as an integrated process in the zinc industry

This work sought to integrate bioleaching and chemical leaching as a cost-effective process to treat zinc sulphides. The continuous bioleaching of a sphalerite concentrate, assaying 51.4% Zn, 1.9% Pb, 31.8% S and 9.0% Fe with mesophile iron and sulphur-oxidizing bacteria followed by chemical leaching of the bioleaching residue were assessed. In the bioleaching step, the first reactor was used to produce Fe(III) concentrations as high as 20 g/L. This solution was fed to the subsequent bioleaching reactors to oxidize sphalerite. It was possible to achieve 30% zinc extraction for 70 h residence time. In chemical leaching experiments, carried out with the residue of the bioleaching step, the effects Fe_total and acidity on zinc extraction were studied. It was noticed that Fe(III) concentrations over 12 g/L did not affect zinc recoveries. Furthermore, the higher the acidity, the larger the zinc recovery, for experiments carried out up to 181 g/L sulphuric acid. The results have demonstrated that it is possible to devise a new process capable of achieving 96% zinc extraction, similarly to the conventional roasting–leaching–electrolysis process.     

Complex treatment of mining and metallurgical wastes for recovery of base metals

A complex process for the recovery of copper and zinc from mining and metallurgical wastes has been investigated and proposed. It includes sulfuric acid leaching of old pyrite flotation tailings to produce ferric containing leach solution; followed by ferric leaching of copper converter slag flotation tailings with the leach solution. A sample of old pyrite flotation tailings from the concentrator containing 0.36% of copper and 0.23% of zinc was leached with 10% sulfuric acid in the column. Recovery of copper and zinc reached 47.1% and 47.2%, respectively. The pregnant leach solutions contained 15.9 g/L of ferric iron. The subsequent ferric leaching of copper converter slag flotation tailings containing 0.53% copper and 2.77% zinc with the pregnant leach solution was conducted. The effects of various process parameters on the leaching dynamics of metals under batch conditions were investigated. Under the best conditions (temperature 70 °C, pulp density 30%, ferric iron concentration 15.9 g/L, initial pH of the pulp 0) the recovery of copper and zinc reached 79.6% and 43.7%, respectively. It was concluded that acid leaching of base metals from old pyrite flotation tailings with pregnant leach solution for the ferric leaching of copper converter slag flotation tailings is a prospective and promising technique for the complex treatment of mining and metallurgical wastes.     

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.     

Thiosulphate leaching of gold in the presence of ethylenediaminetetraacetic acid (EDTA)

The addition of low levels of ethylenediaminetetraacetic acid (EDTA) in the ammoniacal thiosulphate gold leach system lowered the catalytic cupric/cuprous redox equilibrium potential, hence the mixed solution potential and reduced the consumption of thiosulphate. In the leaching of pure gold, gold dissolution was enhanced in the presence of EDTA at a relatively low concentration, but excessive EDTA decreased gold dissolution. Raman analysis of the leached gold foil indicated that the stabilisation of thiosulphate by EDTA decreased the formation of the passivation layers of elemental sulphur and copper sulphide at the gold surface. In the leaching of a sulphide ore, the leaching kinetics and overall extractions of gold and silver were enhanced substantially, while the consumption of ammonium thiosulphate was reduced from 9.63 kg/t to 3.85 kg/t in the presence of 2.0 mM EDTA after 24 h leaching. This beneficial effect became more pronounced at a higher EDTA concentration. The enhanced gold and silver extractions by EDTA were attributed to the increase in the dissolution of gold and silver bearing sulphides, the stabilisation of copper and thiosulphate in leach solutions, the prevention of leaching passivation and the decrease in the interference of foreign heavy metal ions. The use of EDTA in the ammoniacal thiosulphate leaching system makes it practical to achieve satisfactory gold extraction over extended periods of leaching under low reagent concentrations, where the consumption of thiosulphate is low.     

Large particle effects in chemical/biochemical heap leach processes - A review

The application of heap leach technology to recovery of economically important metals, notably copper, gold, silver, and uranium, is wide-spread in the mining industry. Unique to heap leaching is the relatively coarse particle size, typically 12-25 mm top size for crushed and agglomerated ores and larger for run-of-mine dump leaching operations. Leaching from such large particles is commonly assumed to follow shrinking core type behaviour, although little evidence for the validity of this assumption exists. This review investigates the current state of knowledge with respect to the understanding of the characteristics and mineralogy of large particles and how these influence leaching in a heap context and the tools to characterize these. This includes the study of ore and particle properties, visualization techniques for ore characterization, the connection between comminution and leaching behaviour, as well as particle models within heap leach modelling. We contend that the economics of heap leaching are strongly governed by the trade-off between the slow rate and limited extent of leaching from large particles and the cost of crushing finer. A sound understanding of the underlying large particle effects will therefore greatly inform future technology choices in the area of heap leaching.     

热压浸出法从黑白钨精矿中浸出钨的试验研究

进行了热压浸出法处理黑白钨矿浸出钨的试验,重点考察了碱矿比、苛性钠配比、液固比、氧气压力、浸出温度与浸出时间对钨浸出率的影响。研究结果表明,碱矿比从20%逐渐增加到45%时,钨的浸出率随碱量的增加逐渐提高;苛性钠添加量在高于390 g时,钨浸出率在93%以上,且钨浸出率随着苛性钠用量的增加而增加;随液固比的增大,钨的浸出率由87.43%提高至95%以上,但最佳液固比为3∶1时,能提高精矿的处理能力;通氧不利于钨的浸出,最佳浸出温度为160℃、浸出时间为1 h时,钨的浸出率达95.55%。     

Particle damage and exposure analysis in HPGR crushing of selected copper ores for column leaching

In mining operations, jaw and gyratory crushers are generally used for primary crushing, and cone crushers are used for secondary crushing. During the past decade, however, high-pressure grinding rolls (HPGR) are being considered due to potential processing benefits such as energy savings, improved exposure/liberation and particle weakening. At this time there is no detailed quantification of particle damage and downstream benefits from HPGR crushing are uncertain. In the present research, copper ores (copper oxide ore and copper sulfide ore) were crushed by a jaw crusher and by HPGR and the products were evaluated for particle damage and copper grain exposure by X-ray computed tomography. Column leaching was done to determine the rate and extent of copper recovery.X-ray computed tomography analysis and laboratory column leaching experiments for copper oxide ore revealed that products from HPGR crushing have more particle damage and higher copper recoveries when compared with products of the same size class from jaw crusher crushing. Generally the copper recovery from column leaching of the oxide ore was found to be dependent on the extent of grain exposure, which increases with a decrease in particle size.In the case of the copper sulfide ore, copper recovery was found to be independent of the crushing technique despite the fact that more particle damage was observed in products from HPGR crushing. This unexpected behavior for the copper sulfide ore might be due to the high head grade and strong leach solution. Column leaching results also show that about 80–90% of the copper was recovered from the copper sulfide ore in a relatively short leaching time irrespective of crushing technique. As expected, copper recoveries improved with a decrease in the particle size of the copper sulfide ore as exposure of copper mineral grains increased.     

On the use of lignin-based biopolymer in improving gold and silver recoveries during cyanidation leaching

In the present paper, influence of a lignin-based biopolymer in improving gold and silver recoveries during cyanidation leaching is discussed. Series of laboratory cyanidation leaching testworks at the absences and the presences of the biopolymer were performed under variations of the biopolymer concentration, NaCN concentration, slurry density and particle size distribution. The ore samples used were from Pongkor Gold Mine of PT. Antam, Tbk. in Indonesia. The biopolymer additive was found to reduce slurry viscosity that leads to a higher dissolved oxygen concentration in the leach solution. The testwork results show that the use of biopolymer improves gold and silver recoveries as well as leaching kinetics. At biopolymer dosage of 800 mg/L, the increases of gold and silver recoveries were 4.5% and 6.4%, respectively. The lignin-based biopolymer is able to disperse fine solid particles and stabilizes the ore suspension in the leach-slurry that results in a better contact between the ore and the leaching agents. The effect of biopolymer on the gold and silver recoveries was found to be more significant at lower cyanide dosages. The addition of biopolymer resulted in the increase of cyanide consumption, lowering final pH and higher dissolved iron which is associated with gold-bearing pyrite dissolution. Enhanced dissolution of gold-bearing pyrite at higher dissolved oxygen concentration was found as the major factor that improves gold recovery at the presences of the biopolymer in addition to lowering slurry viscosity that leads to higher dissolved oxygen concentration and effect of fine particles dispersion.     

硅酸盐矿物的硅酸盐细菌浸出工艺研究

采用5种硅酸盐细菌,对含不同硅酸盐矿物的5种矿样（包括铝土矿原矿和人工混合矿）进行细菌浸矿脱硅摇瓶试验和工艺试验。摇瓶试验表明：5种细菌均能较好地浸出矿样中的硅,但其中JXF-1菌的脱硅效果最好;细菌存活环境和硅酸盐矿物的晶格结构影响浸硅效果。工艺试验表明：连续浸出工艺的细菌浸硅效果明显优于摇瓶试验和半连续浸出工艺,并可将浸出时间缩短4 d;以JXF-F菌为浸矿细菌,以含2%蔗糖的阿什比基质矿物盐培养基为细菌培养液和浸出液,采用连续浸出工艺可从铝土矿原矿和含绿泥石人工混合矿获得铝硅比分别为12.53和17.74的精矿,硅浸出率分别达70.1%和83.6%。     

Evaluation of leaching parameters for a refractory gold ore containing aurostibite and antimony minerals: Part I – Central zone

A cyanidation study was conducted on a mild refractory gold ore sample from the Central zone of Clarence Stream Property, owned by Freewest Resources Canada, to develop a leaching strategy to extract gold. Gold, at a grade of 8.00 g/t, is present as native gold, electrum and aurostibite. The ore also contains 2.8% pyrrhotite, together with several antimony minerals (0.8% berthierite and gudmundite, 0.18% native antimony and stibnite). It also exhibits weak preg-robbing properties with 0.16% organic carbon. Aurostibite, a gold antimony compound, is particularly known to be insoluble in cyanide solution. The antimony dissolves in cyanide solution to form antimonates, which retards gold dissolution. Industrial practice of extracting gold from aurostibite generally consists of producing a flotation concentrate, which is leached in a pipe reactor at low alkalinity and high oxygen pressure with about 20 g/L cyanide.The proposed new approach is efficient and allows the extraction of gold directly from an ore at atmospheric pressure and a low cyanide concentration at pH 10.5. The effects of grinding, pre-treatment, lead nitrate, kerosene and cyanide concentrations have been investigated. The maximum gold extraction obtained on the ore was 87.9% using 800 ppm NaCN, 500 g/t lead nitrate, 30 g/t kerosene, DO (dissolved oxygen) 10 ppm and pH 10.5 in 168 h. The associated cyanide consumption was 1.3 kg/t. The additions of lead nitrate and kerosene increased gold extraction. In comparison to a P₈₀ of 74 μm, a P₈₀ of 30 μm significantly increased gold extraction. Gold in solid solution in gudmundite and arsenopyrite was believed to be responsible for the un-leached fraction until mineralogical analysis of hydroseparation concentrates of leach residues showed that most of the un-leached gold occurs as aurostibite, either as locked grains in sulphides/sulpharsenides or as grains with passivation rims of an Au–Sb–O phase. Coarse gold was also found. Gold extraction was not sensitive to cyanide concentration from 250 to 1200 ppm NaCN and high pH was detrimental. Decreasing the cyanide concentration reduced the cyanide consumption from 1.39 to 0.85 kg/t. The removal of coarse gold using a Knelson concentrator and a Mosley table prior to leaching increased the gold extraction to 90.4% (leach residue at 0.77 g/t).