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By S. Tavani, R. E. Browner, R. Dunne

Lead is used in CIP/CIL circuit to increase the leaching rate of gold from ores. At least four mechanisms have been proposed for the action of lead in gold leaching circuits: (I) precipitation of sulphides as highly insoluble lead sulphides preventing consumption of cyanide and oxygen; (2) su,face reaction on sulphide particles forming an insoluble lead sulphide layer peventing dissolution of those sulphides; ( 3) precipitation of lead sulphide, which undergoes further oxidation to soluble lead sulphate recycling the lead to solution,· ( 4) contact reduction of lead onto the gold szaface which then acts as a secondary oxidising agent. Rotating disc studies using a pure gold disc show that small amounts (5ppm) of sulphide ions are sufficient to significantly reduce the leaching rate of gold. Lead increased the leaching rate at small concentrations but at higher levels decreased the rate of leaching. The detrimental effects of lead were only found for a pure solution when leaching the gold disc. Ores mediate the amount of lead available by ads01ption/reaction, thus, reduced leaching rates from excess lead are not serious even when 500ppm of lead ions are added to the sluny. The detrimental effects of sulphide ions could be countered by lead addition during the leaching process. A penalty, in the form of reduced leaching rate, will occur the longer lead addition is delayed. A time lag was noted between the addition of lead and the increased leaching rate when a large concentration of sulphide ions was present.

Jan 1, 2000

By Rodney H. Sergent

After being forgotten or ignored for nearly 100 years, the ability of bromine chemistry to extract gold from ore has been rediscovered. While bromine chemistry is similar to chlorine, there are some important differences. Bromine in its elemental form is a dense, fuming, red liquid. Elemental chlorine is a gas and requires a substantial amount of equipment to handle, store and deliver to solution. Bromine has an advantage in its ability to rapidly and completely oxidize heavy metals to water soluble halide salts. The first patent for use of bromine chemistry for gold extraction was issued in 1881 (1). By 1895, work done in Colorado concluded well roasted ores required approximately 1-1/2 lbs/ton of elemental bromine for superior extraction of metals (2)). At that time, bromine was deemed to be cost effective vs. chlorine even though it was more expensive on a pound basis than chlorine. The advent of cyanide technology virtually eliminated both chlorine and bromine from consideration for leaching of ores. Cyanide was so inexpensive that it made deposits requiring unique technologies unattractive for mining. Only since the deregulation of gold price has bromine/chlorine technology been considered for those ores not amenable to cyanide. While chlorine has found its way into the pretreatment steps, bromine is just beginning to realize its niche in the marketplace.

Jan 1, 1998

By M. Misra, J. Nanor

The removal and stability of arsenic over a wide pH range (3.0-10.0) from synthetic solutions and industrial waste effluents (including mining waste tailing slurry) can be accomplished with addition of impure lanthanum chloride alone or in conjunction with ferric sulfate. The long term stability of these correspond­ing precipitated compounds is critical from the disposal stand­point. In this paper test results of this new process for arsenic removal and stabilization from the above mentioned systems are presented. Also, precipitated arsenic compounds (lanthanum arsenate, lanthanum-ferric arsenate, ferric-lanthanum arsenate and ferric arsenate) from synthetic solutions have been charac­terized. The characterization technique used was XRD. XRD patterns which shows that the lanthanum arsenate and lan­thanum-ferric arsenate formed in acidic and alkaline pH at room temperature are crystalline. On the contrary, ferric arsenate and ferric-lanthanum arsenate compounds formed under the same conditions as lanthanum arsenate were amorphous.

Jan 1, 1998

By Gary Jobson

The Mt Todd Gold mine is situated near Katherine in the Northern Territory of Australia. Commissioning of an 8 Mtpa processing facility incorporating a four stage crushing circuit, ball mills, flotation, concentrate regrinding, high rate thickening and CIL processing was commenced during late 1996. A 4.5 M tonne /annum crushing and heap leach facility was commissioned during late 1994, and was subsequently upgraded to a nominal 6 M tonne / annum circuit during 1995. Processing equipment used included primary gyratory, secondary cone and tertiary Barmac VSI crushers. Some of the features of that crushing circuit and heap leach facility were presented at the Randol Perth 95 conference by myself. The location is described as sub tropical with average climatic conditions including 1100 millimetres annual rainfall, predominantly during the December. to March wet season, and 3000 millimetres evaporation and maximum temperatures ranging from around 26 to 46 °C. The main gold occurrence at the Mt Todd site is the Batman deposit, with current published Reserves of some 60 Mt@ 1.20 g Au / tonne ( 0.0425 oz Au / ton ). The primary ore from the Batman deposit is an extremely competent greywacke and is very hard at an average Bond Ball Mill Work Index of26.5 kWhrs/ tonne. Other indicators include a Bond Impact Crush Work Index of 20 kWhrs I tonne, an Abrasion Index of 0.15 and Ultimate Compressive Strength values of 100 - 200 MPa. The main gold occurrence is very finely disseminated free gold (at an average grain size of around 5 micron) within a combination of quartz and mixed sulphide. This gold bearing component is confined to largely parallel veins ranging from 0.5 to 25mm wide within a greywacke matrix. Although the gold is essentially non­refractory, there is a strong recovery versus grindsize relationship due mainly to the liberation characteristics and the relatively low permeability of the host greywacke rock.

Jan 1, 1997

By James H. Nulty, Joseph S. Roti

Wetting agents used for heap leach gold applications would appear to have technical merit in aiding in recovering higher levels of precious metal. Many questions, however, remain as to whether they will function in practice, on what type of ore are they effective and if there are any negative affects associated with their use. A series of experiments are conducted at Bateman Labs, Sparks, Nevada, to determine under what conditions reagents of this type are useful. Carbon kinetics test work was also performed to determine if there are negative aspects associated with carbon loading with the use of these treatment chemicals. This paper outlines the results of these research evaluations, the general chemistry associated with action of wetting agents on leach ores and future work which is planned to further determine the viability of these products in the mining industry.

Jan 1, 1998

By C. F. Bonney

Spherical or oil agglomeration has been recognised for many years as a possible process for the recovery of mineraVmetal values. BP has investigated the potential of spherical agglomeration for a number of minerals as have other workers. Testwork has also been carried out by both BP and others on the selective agglomeration of fine coal. The spherical agglomeration of both minerals and coal depends on the use of an immiscible liquid e.g. a hydrocarbon which is added to a suspension of the solids in water. On agitation the oil is distributed over oleophilic/hydrophobic surfaces and particle impact allows inter-particle liquid bridges to form causing agglomeration. It is also possible to enhance the oleophilicity of the material to be recovered by the use of collectors similar to those used in froth flotation.

Jan 1, 1998

By Terry M. Hertel

The chemistry of waters processed by Acid Rock Drainage water treatment plants can affect the Total Suspended Solids quality of the plant effluents. The sulfate ions in the ARD water will combine with the calcium ions from the lime that is used to adjust pH to form calcium sulfate. When the sulfate content of the waters to be treated is greater than about 1400 parts per mil­lion, the resulting calcium sulfate concentration will be approx­imately 2100 parts per million. This concentration is above the saturation point for calcium sulfate, and the salt will drop out of solution until the concentration falls below the saturation point. This "salting" phenomenon affects the total suspended solids of the treatment plant effluents by virtue of the fact that the salts will report to the filtration apparatus used for the analy­sis. This scenario is exacerbated as an effluent sample ages. Zortman Mining's water treatment plant effluent was being affected by "salting" during the spring and summer of 1995, and it seemed as though the effluents were exceeding the TSS lim­its; for no obvious reason. A testing program was initiated to verify an observation that "salting' was taking place. Treatment plant samples were collected and allowed to age, with TSS analyses being performed every day that the samples aged. Synthetic solutions were also prepared to study the "salting" phenomenon.

Jan 1, 1996

By J. Libaude

The Changement mine in French Guyana has, since 1983, processed gold ore by classical gravity methods. A large part of the gold is in the form of very small, free or enclosed grains. Recovery by gravity methods is thus limited (25-30%) and the present plant produces about 40,000 t annually of auriferous waste assaying about 6 git. The company therefore decided to retreat the waste by heap cyanidation. After laboratory tests which showed the feasibility of the method, a pilot test was set up on site. An experimental 3000 t heap, composed of waste of 0-8 mm grain size, was laid out by bucket loader on an impermeable pad covered by PVC and geotextile cloth. The upper layer of the heap was reworked by mechanical shovel to break up the area that had been compacted by the passage of the loader so as to restore adequate permeability.

Jan 1, 1998

By Harrie van Dijk

Feeding laterite nickel slurry into high pressure acid­leach (HPAL) autoclaves in an aggressive environment with slurries heated up to 200°C requires a special pump design. For decades the GEHO piston diaphragm pumps have been used for autoclave feeding at gold ore pressure oxidation plants with slurries up to l 77°C. Lessons learned from these installations have been instrumental in designing second generation feed pumps for processing laterite nickel with operating temperatures of 200°C and a flow rate of 400 m3/h. The Australian nickel industry is using momentarily 8 GEHO positive displacement piston diaphragm pumps. For this application six GEHO ZPM 900 "dropleg" pumps have been supplied to the Murrin Murrin project and two to Bulong.

Jan 1, 2000

By Gary P. Marshall, Jeffery S. Thompson, Robert E. Jenkins

The effectiveness of the new DuPont Passivation Technology in eliminating acid generation from iron sulfides is presented. Passivation is achieved by treatment with a dilute basic perman­ganate solution. This treatment produces a thin impervious layer of iron/manganese oxide on the sulfide surface, providing a bar­rier to the environment so that the sulfide is not available for undesirable reactions such as the production of acid.

Jan 1, 1998

By P. Brooks, W. LaRochelle, J. M. Tranquilla

EMR's recently commissioned Pilot Plant in Fredericton, N.B. Canada, the latest and largest in the company's sequence of prototype and demonstration systems, is being used as a process development facility for mining clients worldwide. Fitted with two microwave reactors and 75 kW generators, this plant is designed as a two-stage continuous processor with a throughput in the 5-10 tpd range depending upon the material feedstream. This facility has been used principally to calcine pyrite and arsenopyrite gold concentrates although it can readily accommo­date carbonaceous materials as well as a wide range of other ore types. Engineering details of the plant and the results of the first nine months of operation are presented. The plant has proven very successful as an advanced process development facility as well as a basic production unit module.

Jan 1, 1998

By Andy Cole, Stanley Bunk, Steve Dunn, Thom McCord

In 1994, Goldstrike initiated a research program to investigate processing options for treating what was then the small amount of carbonaceous ore identified in the reserves. The focus of the research was on developing a process which incorporated the autoclave technology already utilized at Goldstrike. Over the next two years, Barrick pursued the development of an ammonium thiosulfate leaching (ATS) process. During the period in which the ATS research was being conducted the amount of carbonaceous ore in Goldstrike's reserves increased substantially and continued growth was projected. By 1996, significant advancement was made on the development of the ATS process. The work culminated in a series of integrated pilot plant runs incorporating autoclaving, ATS leaching, resin in leach, stripping and cementation. Several questions about the process chemistry remained unanswered, preventing commercial development. Barrick was committed to producing gold from the carbonaceous ore reserves by 2000. A due diligence review of carbonaceous ore processing options was initiated during the second half of 1996. Processes considered in addition to ATS were fine grinding/autoclaving, as practiced then by Santa Fe, and whole ore roasting as practiced by Independence Mining and Newmont. The objective by year-end was to recommend the most robust and economical process for Goldstrike's carbonaceous resources. In the fall of 1996, a detailed review of roasting as a processing alternative for Goldstrike was initiated. Throughout the ATS development period, whole ore roasting was used as the benchmark. All the ATS metallurgical and financial results were compared to roasting bench top test results. It was clear this data was too imprecise for use in due diligence study, a more definitive recovery value was needed to complete the carbonaceous ore review. The primary goal of the roasting review was to conduct pilot plant roast tests to gain a more thorough understanding of Goldstrike's carbonaceous ore's response to roasting and its effects on gold recovery. Using the test facility at Hazen Research, four pilot plant campaigns were conducted to determine optimum roast conditions and further define metallurgical design parameters. Conventional two-stage fluid bed technology based on Freeport­McMoRan's roasting patent was the specific process basis for testing. The ore's response to roasting exceeded expectations. In January 1997, a preliminary economic evaluation comparing ATS and roasting was presented, and the outcome was approval to proceed with a roasting prefeasibility study. In July 1997, approval for Goldstrike's Roaster facility was received and basic engineering started immediately thereafter. Detailed engineering _began in January 1998 and plant construction commenced in October 1998,t, Goldstrike's Roaster facility is designed to treat 12,000 tpd of a refractory gold ore, which contains both pyrite and organic carbon. Circuits in the facility include primary and secondary crushing, dry grinding, roasting, gas handling, neutralization, thickening and standard CIL. Gold desorbtion and recovery is carried out at Goldstrike's existing refinery. Plant start-up of the first dry grinding and roasting train is scheduled for March 2000, with the second train coming online in May 2000.

Jan 1, 1999

By Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By Daniel W. Kappes

In late 1995 at the Randol symposium in Squaw Valley, I gave a paper which described the Itos project flowsheet. At that time, project construction was nearly completed and stmtup was close. This paper describes the ltos operation since then. To summa­rize things, startup occurred partially in late 1995 with full start­up of all components in about mid 1996. From then until now, April 1998, the plant has been running at partial capacity and has yet to achieve full production rates.

Jan 1, 1998

By Dan Mackie

Since the production prototype continuous vat leaching pilot plant was installed at Ariel Resources in Costa Rica in 1995, the test results and lessons learned have been invaluable in plotting the future of this method. Vigorous test programs in base and precious metals as well as design developments for studies have shown the advantages, economically, technically, and environmentally that favor this method over current practices in heap leaching and stirred tank leaching. Much larger individual vats have been developed with throughputs as high as 5000 t/d with larger vats envisaged, and there has been a trend towards using vats with ground ore. Five projects are slated to begin detailed engineering this year, three in precious metals and two in copper. Of particular interest to future producers is the ability to deposit d,y, detoxified tailings at a fraction of the cost of conventional methods. Proprieta,y work is ongoing which promises to eliminate the need for tailings dams when using other methods of processing such as flotation. This is of particular importance in light of dam failures around the world in recent years.

Jan 1, 2000

By Frederick W. DeVries

The following charts were presented at the Randol Gold & Silver Forum 2000, held at the Hotel Vancouver (Vancouver, B.C.) from April 18-21, 2000. It must be borne in mind that the "facts" stated and con­clusions drawn by the author were not as the result of a rig­orous technical investigation, but rather as a member of a journalistic team seeking a rapid view for presentation as news to a TV audience. Since the author's visit, some retrenchment on the part of the operators on estimating the cyanide content of the spill seems to have occurred. This does NOT alter the conclusions that (a) the spill was an intolerable embarrassment for the world mining industry, nor (b) the potential damage to future mining investments, should prompt action not be taken: 1. to understand the reasons for the dam failure, and 2. to draw up firm concepts of how to avoid future fail­ures of dams at mining properties containing materials that are toxic, environmentally unfriendly, or with as highly dis­comfiting a profile as cyanide.

Jan 1, 2000

By R. Bruce Thorndycraft

Heap leaching should not be chosen for development of a property solely on the basis of paper studies. Management should be made to understand where the risks are in heap leaching, and management should be alerted to the variability of costs in heap leaching, many of which are based upon company philosophy. These concepts are complex, but easily illustrated with guided tours of existing heap leach facilities.

Jan 1, 1998

By Steven Schurman

Acid Rock Drainage (ARD) is one of the most serious and persistent pollution problems facing today's mining industry. ARD is formed when pyrite (FeS) is exposed to natural weathering conditions. Pyrite weathering, formally the oxidation and hydrolysis of pyrite, leads to the generation of soluble, hydrous iron sulfa-salts and the production of acid in the form of H+. The resulting effluent is water that is characterized by high levels of iron (and any other metal contained in the rock undergoing weathering) as well as high concentrations of sulfate and a low pH (acid). Rocky Mountain Remediation Services, L.L.C., has developed a patents pending Envirobond'™ product, Envirobond™ ARD, to treat acid rock drainage. Treating pyrite with Envirobond™ ARD inhibits the oxidation and hydrolysis of pyrite thus curtailing the pyrite oxidation cycle.

Jan 1, 1999

By Arif Emre Dursun, Ali Aras, Hakan Özşen

Specific energy (SE) is defined as the amount of work required to cut a unit volume of rock and is used to estimate the performance of excavation machines. Generally, the SE value is produced from full-scale or small-scale rock cutting tests. However, these tests require expensive equipment and experienced personnel, and the testing procedure is time consuming and impractical. Therefore, for estimation of SE, this study aimed to find a solution to estimate SE using rock mechanics test results and Bond work index values (BWi), which can be produced in a more practical, less time-consuming manner and with inexpensive test equipment. Small-scale rock cutting, rock mechanics and Bond work index tests were carried out on seven rock samples, which can be grouped into two different type of rocks: marble and travertine. In addition, the instantaneous cutting rates (ICR) of a selected roadheader were calculated using SE values. Rock mechanics tests performed in this study were uniaxial compressive strength (UCS), Brazilian tensile strength (BTS), ultrasonic velocity (Vp), Schmidt rebound hardness (RL), corrected point load index (Is(50)) and density (ρ). The simple regression method was used to estimate SE from rock properties and BWi. As a result of this simple regression, a significant and highest correlation was found between SE vs. UCS and Is(50). A good correlation was found between SE vs. BTS, RL and BWi values. A moderate correlation was found between SE vs. Vp and ρ.

Jul 28, 2021

By Amila Sandaruwan Ratnayake, T. Dilmi. U. Wijewardhana, H. C. S. Subasinghe

The value addition potential of ilmenite was examined using several characterization techniques. Raw ilmenite is composed of 93.65 wt.% of FeOTiO2, 3.55 wt.% of SiO2, 1.22 wt.% of Al2O3, and the remainder being other minor oxides. Ilmenite and powdered carbonized coconut shells were mixed in the weight ratio of 4:1. A separate fraction of the same mixture was added with powdered seashells in the weight ratio of 4:1:0.5. Six fractions of each mixture were distinctly milled from 1 to 6 h in a planetary ball mill. XRD spectra (broadened and diminished reflections of ilmenite) and FTIR observations (Ti―O―Cbonding) of milled samples indicate the possible incorporation of carbon into the ilmenite structure. Any clues of an occurrence of new rutile peaks were not observed in XRD spectra of milled samples. Consequently, the carbothermic reduction has not been initiated during mechanical activation. In this case, samples powdered for 6 h in a mill were isothermally annealed for 2 h under normal airflow at temperatures of 800 °C, 1000 °C, and 1200 °C, respectively. According to the X-ray diffractograms, the annealing temperatures of 1000 °C and 1200 °C exhibited almost similar trends with rutile (R), pseudobrookite (PB) and elemental iron (F), and a very few ilmenite (I) peaks. Consequently, 6 h of milling and 1000 °C annealing were concluded as the optimum conditions for the carbothermic reduction. Moreover, this study indicated seashells as a potential rate raiser for the carbothermic reduction of ilmenite at 800 °C. Therefore, this process is applicable to upgrade ilmenite into a mixture of synthetic rutile and elemental iron.

Mar 29, 2021