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By J. A. Drozdiak

In this paper we examine, through pilot-scale testing, the possibility of operating a circuit comprised of two stages of HPGR comminution, followed by grinding through a horizontal stirred mill. In order to assess whether the novel circuit design could achieve the reduced energy requirements indicated in the literature, two more-established circuits, a cone crusher / ball mill and an HPGR / ball mill, were examined using a combination of testing and flowsheet simulation. The results showed that, based solely on the specific energy requirements for comminution, the HPGR / stirred mill circuit achieved a reduction of 9.2% and 16.7% over the HPGR / ball mill and cone crusher / ball mill circuits, respectively.

Jan 1, 2012

By W. Beres

Many non-commercial operators are currently faced with the need to operate fleets of legacy gas turbine engines. How long such engines can be kept in service safely, without replacing a significant portion of their aging structural components, has become a growing concern to engine life-cycle managers, due to uncertainties in the remaining lives of critical components of these engines. This paper describes the application of the damage tolerance methodology to predict safe inspection intervals (SII) for critical components of gas turbine engines, using components of a turboprop engine for a transport aircraft as examples. High temperature components of this engine are made of advanced aerospace alloys. The methods of predicting three-dimensional fatigue crack growth at the critical location of turbine discs and spacers using both commercial and in-house developed algorithms are presented. Sensitivity analyses are performed to determine the effect of variations in material parameters on the predicted safe inspection intervals. In addition, spin rig tests performed for both turbine discs and spacers are described. Finally, crack growth rates predicted by numerical analyses are compared to those obtained from spin rig testing and the damage tolerance of the turbine components is assessed.

Jan 1, 2015

By I. Haskett

ABSTRACT The Redwater pool, in north-central Alberta, was discovered in October, 1948, and at the end of 1950 was about 85 per cent developed. Production is from a Devonian coral limestone reef and is found at an average depth of 3,150 feet. The pool has an estimated .productive area of 37,000 acres, and a maximum oil zone thickness of slightly over 200 feet. Wide variations exist in both type and values of porosity, with the average value determined for the reservoir being 6.16 per cent. The total stock tank oil in place is about 1.24 billion barrels. The crude has a saturation pressure of 485 psig, com-pared to the original value of 1,050 psig for the reservoir, and the principal recovery mechanism so far has been the natural water drive. Cumulative production to January 1st, 1951, has been about 15.5 million barrels. These withdrawals resulted in a reservoir pressure decline of slightly over 100 psi. Predictions of future behaviour indicate that the water drive will not be of sufficient magnitude to maintain the reservoir above saturation pressure past 1953, and that, with natural depletion, the pool will be produced largely by a solution gas drive. Pressure maintenance by water injection has been considered and the possibility of raising the pressure to the original value of 1,050 psig is also discussed. INTRODUCTION WITH the discovery of the Lewduc pool, early in 1947, the Devonian limestone became the major geological formation contributing to Western Canada's rapidly growing petroleum reserves. Redwater, discovered in October, 1948, is now the largest known Devonian pool in Canada, with a recoverable reserve estimated at approximately 500 million barrels. Until recent years, the knowledge of the nature and characteristics of this type of limestone reservoir has been comparatively limited.

Jan 1, 1951

By D. O. DeGagné, J. F. Archibald

Spray-on polymer linings, as partial or full replacement for traditional forms of area rock support, have seen rapid growth in interest within Canada. A wide variety of lining types, ranging from polymer chemical to water-based cementitious materials, have been developed and commercially introduced to the underground mining market. Following approximately one decade of development work, several general conclusions have been reached regarding the performance of this class of area support media, these being that: health and safety assessment of the broad range of materials has shown the majority of them to be acceptable and safe for underground use; significant ground support potential can be offered by them, being equivalent to or better than capabilities offered by traditional support forms such as bolt-and-screen or shotcrete, and; their abilities to be rapidly deployed and provide near instantaneous support restraint offer considerable operational and cost benefits to underground operators. A review of currently available liner materials, their historic development within the Canadian mining industry, and overviews of support, health and safety, operational considerations and cost benefits are provided in this paper.

Jan 1, 2001

By Unknown

We are indebted to the Consolidated Mining and Smelting Company of Canada Limited for the several pictures in this issue showing the progress made in new construction at Tadanac and at Bonnington Falls, B.C. Plate No. 1 shows the down-stream side of the No. 1 hydro-electric station of the West Kootenay Power and Light Co. Limited at Bonnington Falls, on the Kootenay river, at which 17,500 k.v.a. will be developed to meet the requirements of the enlarged zinc and lead refineries at Tadanac. The generators, exciters and switch apparatus were supplied by the Canadian General Electric Company and the step-up transformers by the Canadian Westinghouse Company Limited. The plant throughout was designed, manufactured and erected by Canadians. Plate No. 2 shows the casting room of the new lead-plant and plate No. 3 the new tank-room of the electrolytic lead refinery at Tadanac. These additions will increase the capacity of the lead refinery by some 150 tons daily. The extensions also include practically doubling the zinc plant, and the gold and silver refineries. With the exception of the new stack, all designing construction and installation of the various plants is being done by the Company's own employees at Tadanac. Plate No. 4 shows the progress made in the construction of the new reinforced concrete stack at Tadanac. This stack, of 21 feet inside diameter at the top, will be brick lined and when completed will rise 400 feet above its foundations: the tallest of its kind in the world. The construction is in charge of the Rust Engineering Company of Pittsburgh, Pa. Plate No. 5 is a panorama of the Tadanac smelters and refineries. The photograph was taken from the roof of the lead smelter and gives an excellent idea of the size of the plant and its surroundings. Tadanac produces electrolytic lead, zinc and copper and fine gold and silver. During 1924 the company treated 1,333,523 tons of ore and produced: 161,400,- 769 pounds of lead; 92,197,824 pounds of zinc; l,382,750pounds of copper; 4,074,044 ounces of silver, and 26,976 ounces of gold.

Jan 1, 1925

By G. Morin, F. Chiesa, B. Duchesne

The Quality index is a useful tool allowing to assess the metallurgical quality of aluminium AlSiMg foundry alloys (A356/357). Its value is calculated from the ultimate tensile strength and elongation; it increases with the metallurgical quality of the casting, i.e. with finer dendrites, lower microporosity and inclusion count and lower iron content. This Quality index does not depend on the degree of temper applied during the heat treatment (aging time and temperature). After defining the Quality index, it will be explained how it can be predicted from the knowledge of the local solidification conditions obtained from modelling. The paper will then present four case studies of castings for which the prediction of the Quality index by modelling was made. These case studies from different authors will involve the three most common casting processes used for producing high integrity aluminium castings, namely: sand casting, permanent mould casting and investment casting (lost wax). In all cases, the accuracy of the prediction will be assessed by comparing the predicted Quality index distributions to the experimental values based on the published measured values of the local tensile properties.

Jan 1, 2006

By R. J. Mitchell

"High narrow backfills of significant strike length approach the stress conditions associated with a free standing wall but this lower bound design approach is not considered efficient for earth structures. This paper considers the design of high narrow backfills using upper bound limit analysis, retaining wall theory with backfill arching and centrifuge model tests. Complex toe support situations pertaining to operational factors at Falconbridge's Lockerby Mine are investigated. It is shown that the susceptibility of high narrow fills to toppling is much lower than anticipated from the retaining wall analysis and that toppling can be avoided by a suitable toe support. It is also concluded that adhesion between the fill and an adjacent pillar can markedly increase the stability of the fill. INTRODUCTIONFalconbridge's Lockerby Mine pillar projects will expose cemented tailings backfills to heights (H) up to 55 m. The exposed backfills may have strike lengths (L) greater than the heights but are relatively narrow, having widths (w) of up to 12 m. The bulk cement content of the fills varies from 8.3% (12.1 T:C) to 3.0% (33.1 T:C). Exposed faces may have up to a metre of overhang due to stope overblasting, as shown in Figure 1. These fills have a solids specific gravity of close to 3.0 and were placed at a pulp density of about 68%. The tailings have a fairly uniform grain size with 32% less than 0.075 mm (200 sieve) and about 12% slimes (less than 0.02 mm sizes). The fill exposed by early pillar removal will be backed by adjacent, near vertical, ore pillars. Future pillar removal may expose similar fills which are not supported by an adjacent ore pillar but which will be backed by either an open void or by uncemented backfill. Furthermore, these fills may be undercut when the pillar slot is driven, further reducing stability (case 3, Fig. I) or a wedge could be left to help support the fill (case 2, Fig . 1). This paper presents limit equilibrium analyses for these various backfill conditions in a 22: 1 T:C fill and com- pares the results to data from centrifuge model studies at a scale factor of A= 157 as indicated in Figure I. Prototype fills have a 6 m deep cone plug of 16:1 T:C which is accounted for in the comparison of centrifuge model test with analytical results."

Jan 1, 1989

By E. J. Sides

In this paper, the argument is put forth that the codes used for public reporting of mineral resources and mineral reserves have some significant limitations with regard to the information provided to investors on the short- to medium-term value of mining properties. Such limitations include a) failure to explicitly incorporate a time element and b) failure to state explicitly the quantity and quality of mineral product that will be produced for sale. A time-value framework for reporting mineral assets is proposed in order to overcome these limitations. The proposed system is based on two classification axes: a) an X-axis representing time in years during which the value of the asset is likely to be realized and b) a Y-axis representing the value identified by a mining or exploration company\u2019s activities. The potential benefits of this system are discussed, including the ability to incorporate factors, such as permitting and construction times, directly into the magnitude of the reported mineral assets. The proposed framework would assist in the standardization of mineral assets reporting in such a manner that they could eventually be more readily included in financial statements. Year of Publication: Not explicitly mentioned in the document.

Jan 1, 2010

By Dalton Fowler

"Pressure vessels used in pressure leaching and pressure oxidation processes in the mining industry are often lined with acid resistant brick and membrane systems. Although the options available when it comes to selecting materials and linings designs for these vessels appear somewhat limited there are, nevertheless? several choices which must be made in order to successfully design, manufacture and construct these linings. These choices require careful evaluation of materials properties which include, among others, thermal conductivity, thermal expansion, chemical swell, and modulus of elasticity, as well as a variety of construction methods.Some of the options available, and some of the means of evaluating them objective of this discussion."

Jan 1, 1999

By D. A. Temple, G. O. John

Until recently, the approach to the problem of treating complex ores bas been to produce, insofar as is possible, mono-metal concentrates which are treated in traditional zinc, lead and copper smelting circuits, with the residues and by-products from some of these circuits being retreated for the recovery of contained values either by a separate residue treatment process or along with the primary feed to the appropriate circuit. The continued development of the Imperial Smelting Process, with its ability to treat poly-metal concentrates, calls for a re-examination of the traditional approach to this problem, and this the paper attempts to do. A survey of recent progress in the technical development of the process is made and a summary of current operational performance is included.

Jan 1, 1971

By J. Cayouette, C. Bolduc, P. Blatter

"The LaRonde mill was commissioned in 1988 and began treating ore from the Penna shaft in 2000. The mill has been expanded to accommodate 7,500 t/day, but ramped down to 6,000 t/day since the deep extension began. The process recovers gold, silver, and base metals. In 2008, the Goldex plant began trucking gold-pyrite concentrate to a treatment plant feeding pulp to the LaRonde mill, increasing gold output by 40,000 oz. In 2009, the high-grade Lapa ore processing facilities were annexed to the LaRonde plant, increasing gold output by 120,000 oz. This paper outlines the metallurgical considerations related to the treatment of different ore types.RÉSUMÉ Le concentrateur de la mine LaRonde a été mis en service en 1988 et a commencé à traiter le minerai du puits Penna en 2000. La capacité du concentrateur a augmenté afin de permettre le traitement de 7 500 tonnes de minerai par jour, capacité qui est retombée à 6 000 tonnes par jour (tpj) depuis le début de l’extension en profondeur de la mine. Ce procédé permet de récupérer de l’or, de l’argent et des métaux communs. En 2008, l’usine de traitement Goldex a commencé à transporter par camion du concentré de pyrite et d’or vers une usine de traitement alimentant en pulpe le concentrateur de LaRonde, augmentant ainsi la production d’or de 40 000 onces. En 2009, les installations de traitement du minerai à haute teneur de Lapa ont été annexées à l’usine de traitement de LaRonde, augmentant la production d’or de 120 000 onces. Cet article passe en revue les considérations métallurgiques liées au traitement de différents types de minerai.INTRODUCTIONThe LaRonde metallurgical plant in the Abitibi region of Quebec, Canada, evolved into a mature milling site following expansions in 2000–2003, thanks to improved processes, manpower, and water treatment capabilities that better accommodate the complex constituents of the LaRonde ore. The ore requires a sequence of grinding, copper- lead and zinc flotation, and precious metals leaching, ending with a carbon-in-pulp (CIP) gold recovery. A portion (35%) of the tailings is sent back underground as paste backfill after being treated with the SO2-air cyanide destruction process. The tailings area has a dedicated cyanide destruction and metal precipitation area through which the used water passes before returning as recycled water. A biological water treatment plant was built in 2004 to address water toxicity in the final effluent. The water toxicity was the result of the high sulfide content of the LaRonde ore (from thiocyanate or stable cyanide species buildup) and the use of 90% recirculated water in the mill. The water treatment plant uses bacteria to transform thiocyanate into nitrates and precipitate phosphorus from the water before it is discharged into the environment."

Jan 1, 2017

By Jean-Luc Collins

"This new tool can be used as a quick guide f or the assessment of the potential economic wealth of any risky capital project. It does provide an unambiguous recommendation of the type ""go or no go"" to the decision-maker simply by comparing two different ratios profit/investment. One ratio determines the minimum value required to at least break even while the other measures a probable value based on the asymptotic theory of statistical extremes.Its application is briefly illustrated by means of an underground gold mine extension project, an oil sands mine project and a polymetallic open pit project.Once a favourable recommendation has been obtained then a detailed ""cash flow after taxes"" analysis can be carried out for the project.IntroductionMany tools are used in financial decision-making such as the payback period, the discounted cash flow, the internal rate of return and the wealth growth rate (Berry , 1971).These tools are useful and meaningful in evaluating normal projects closely related to regular business where the number of assumptions made is usually very small and where the cost of capita! is usually selected as a discount rate or a hurdle value. In situations where a significant number of assumptions have to be made,two problems can emerge, namely:I. the risk involved is underestimated leading to a potential financial disaster; or2. a good opportunity is lost because the approach used in evaluating the project is too conservative.Each assumption has its own degree of uncertainty and it is well known that combining uncertainties together can lead to a total uncertainty of critical proportion. The easiest way to account for risk is to increase the discount rate or the hurdle value, but this approach alone is quite subjective and can easily lead to the problems described above."

Jan 1, 1993

By Descarreaux. Jean

"The Abitibi volcanic belt was sampled along a cross section extending from Matagami to a point east of Rouyn. A few samples were also collected in Louvicourt township, Quebec. The major elements were determined for 344 samples, and rock names were defined using normative minerals. The analytical data were studied by means of variation diagrams. The Abitibi volcanic belt is characterized by rocks belonging to both the tholeiitic and calc-alkaline magmatic series, but they show some differences as compared to some typical reference series, particularly those from Japan. Basalt is the predominant rock type and accounts for 44 per cent of the sampled outcrops. The volcanic rocks are relatively poor in K2O and TiO2. Regionally, the volcanogenic copper-zinc mineralization is associated with the calc-alkaline volcanic sequences. Locally, within such sequences, some types of abnormal deviations from the averages are recognized only in association with the copper-zinc mineralization. The anomalies are ascribed to hydrothermal alteration consisting of the introduction of MgO and removal of N a20; enrichment in K20 is also observed where zinc mineralization predominates. Examples of such anomalous zones show that, in places, this alteration phenomenon can be quite extensive. Petrochemical characteristics can be used as geochemical ore guides in exploration programs."

Jan 1, 1973

By Dale A. Sketchley

The objective of Quality Assurance and Quality Control programs on advanced exploration projects is to ensure that data from sampling, assaying, and recording of geological observations are of high integrity for the purpose of obtaining reliable mineral resource and reserve estimates. National Instrument 43-101 Standards of Disclosure for Mineral Projects states that all scientific and technical reports must be prepared and certified by a Qualified Person in accordance with professional and industry standards following Mineral Exploration Best Practices Guidelines and Definitions, and Guidelines of the CIM Standards on Mineral Resources and Reserves. Additional information is given in Companion Policy 43-101CP and Technical Report Form 43-101F1. Although the Qualified Person must discuss and describe the Sampling Method and Approach; Sample Preparation, Analyses, and Security; and Data Verification; there is little specific information in the guidelines on Quality Assurance and Quality Control programs. Examples from the Oyu Tolgoi Cu-Au Project are given for optimizing sampling protocols, following proper sample preparation procedures, safeguarding sample shipments through chain of custody, ensuring quality of assaying with Round Robin monitoring programs, verifying transcribed data, and validating data with check assays to ensure high integrity for resource and reserve estimation. Oyu Tolgoi, also known as Turquoise Hill, is in the South Gobi desert, Mongolia, about 550 km south of the capital city of Ulaanbaatar, and 875 km northwest of Beijing, China. The project covers a major copper-gold system, which consists of the South West, South, Central, and Far North deposits, comprising high sulphidation and porphyry mineralization with oxide, supergene, and hypogene ore types. Exploration is done by multiple diamond drill rigs, and samples are prepared on-site and shipped to laboratories in Mongolia, Canada, and Australia for further preparation and analysis.

May 1, 2003

By S. A. Scott

In 1913, potash mineralization was recognized at a depth of over 7,000 feet in an oil-well hole (Norcanol's Radville No. 1 well) drilled near the town of Raddlle, Saskatchewan. Although, at this depth, potash is not considered mineable by conventional underground mining methods, this discovery shed the first light on the potash deposits of Saskatchewan in much the same manner as oil-well drilling was responsible for the first discovery of potash in the Carlsbad Basin of New Mexico in 19:25. The discovery near Radville alerted the oil-well geologists and, as a result potash mineralization was recognized in the Verbata No. 2 well, near Unity, Saskatchewan, in 1946. The mineralization, which was at a depth ?of 3,466 feet, graded more than 20 per cent K20 over a thickness of eleven feet. This discovery, as its depth, thickness and grade were all amenable to mining,

Jan 1, 1963

By Paul W. Shelley, Daniel J. Hodges

The mining industry is undergoing a period of rapid transformative change.  This is the result of the social and financial accountabilities that the industry is faced with because of the world’s desire to cap average temperature increase to 1.50C above pre-industrial levels (UNFCC, 2016).  To limit global warming to 1.50C, greenhouse gas emissions must peak before 2025 at the latest and decline 43% by 2030.  In hard rock mining, comminution is identified as the major energy consumer, being responsible for 40-60% of the electricity use.  It follows that comminution is the major greenhouse gas emission process of the mine.  All effort is being made to radically reduce the carbon emissions of comminution. Grinding media plays a large role in the energy consumption and carbon emission story.  In recent years industry practitioners have attempted to quantify the grinding media carbon footprint.  This paper makes a definitive contribution to that discussion.  The grinding media carbon footprint is quantified in Scope 1, 2 and 3 emission terms in the context of a grinding media supplier.  Major emission contributors, steel, natural gas and electricity are discussed in detail.  The steel industry transition is a key factor. The paper will explore the steel industry transition and its effect on grinding media manufacturing.  The carbon footprint of heritage steel process routes are analysed and compared with the process routes of the future.  Emerging steel making technologies are presented and discussed.  The future of the steel manufacturing and procurement route for grinding media is proposed. This narrative allows for a comparison between the heritage carbon footprint and the future carbon footprint of manufactured grinding media.  Of equal importance is the opportunity of reducing the grinding media carbon footprint in the field.  The paper explores possibilities of reducing the grinding media carbon footprint at the customer site.  Media in use options are presented that take account of the mineral characteristics, the mill dynamics, and the product metallurgy to optimise both carbon footprint and media performance. It is intended that this is a landmark paper in the discussion of the carbon footprint of grinding media.

May 4, 2025

By Jing Li, Weiqiang Lai, Joe Zhou, Jiezhen Ma, Shuibo Chen, Ashele Copper Company, Feng Gao, Guofu Feng

"Copper-zinc ores often contain gold and silver, which can be recovered into copper concentrate by flotation as by-product. The challenges for operations that produce copper, zinc, gold and silver are the separation of copper and zinc and the low grade of precious metals. Ashele copper-zinc ore in north China contains 2.38% Cu, 0.69% Zn, 0.21 g/t Au and 18.10 g/t Ag and the plant produces copper and zinc concentrates, with gold and silver being recovered into copper concentrate. Copper occurs mainly as chalcopyrite and Zn occurs as sphalerite. Copper is well recovered with a recovery of over 94%. However, the recovery of Zn is substantially low (around 50%) due to fine grain size of sphalerite. In addition, gold recovery is much lower than expected and only 33% of gold in the flotation feed was recovered. The grade of gold in copper concentrate is also lower than the required payable grade (1.0 g/t). To determine the cause for lower Zn and Au recoveries, a plant survey program was conducted on the operation. The information acquired shows that the loss of Zn was mainly in the copper concentrates (accounting for approximately 38% of the head assay); of which, 24% was liberated, 8.5% was associated with chalcopyrite and 4.6% was associated with pyrite, indicating that the separation of copper and zinc was the main cause for the Zn loss, with liberation being the second cause for Zn loss. Future effort in improving Zn recovery should be focused on the separation of zinc from copper. Gold in the ore occurs as both microscopic gold and submicroscopic gold, accounting for 36% and 64% of the head assay, respectively. Microscopic gold occurs mainly as native gold and electrum; it is relatively coarse with a grain size of 10 to 240 µm and a large portion can be recovered by gravity. Submicroscopic gold was mainly carried by pyrite (accounting for 38%), and to a less extent, by tennantite (15%) and chalcopyrite (10%). Gold in sphalerite is nominal (1%). It is suggested that more coarse gold and tennantite be recovered and a gold recovery increase can then be achieved to obtain a copper concentrate with over 1.0 g/t Au through flowsheet optimization. Gravity concentration should be investigated and may be worth added into the existing flowsheet to maximum the gold recovery and grade.INTRODUCTIONCopper-zinc ores containing gold and silver is a subtype of polymetallic sulphide ores. Unlike porphyry copper sulphide ores and iron-oxide copper-gold ores, this type of ores often contains substantial amounts of zinc and lead. Polymetallic sulphide ores are mainly distributed in Canada, such as at the LaRonde and Horne deposits. Other examples include Mt. Morgan in Australia and Boliden in Sweden. Globally, production and reserves from gold-rich VMS deposits account for 1370 t gold. This is equivalent to 1% of the world production and reserves, whereas the Canadian total for gold-rich VMS deposits is 870 t, representing close to 10 % of Canadian production and reserves (Gosselin & Dube, 2005a, 2005b). Flotation is the principal process for processing the copper-zinc sulphides ores, for subsequent smelting, roasting, or hydrometallurgical treatment. Oxidized copper-zinc ores can also be processed using leaching technology. For example, at LaRonde operation (a Cu-Zn-Pb ore containing 3-4 g/t Au and 60-100 g/t Ag) the flowsheet consists mainly of a copper flotation circuit and a zinc flotation circuit followed by a cyanidation process of flotation tails. The precious metals are recovered with copper concentrate, zinc concentrate and cyanidation of zinc tails via Merrill Crowe precipitation (Zhou et al., 2005)."

Jan 1, 2016

By H. C. Horwood

INTRODUCTION THE Argosy Gold Mines, Limited, which joined the ranks of active Ontario gold producers during the summer of 1936, holds a group of 21 claims on Casummit lake, in the Patricia portion, district of Kenora, northwestern Ontario. The Company is developing and mining goldbearing quartz veins that occur in a series of greywackes and slates of probable Keewatin age. The writer made a geological examination of the property and the surrounding area for the Ontario Department of Mines during the early part of the field season of 1936. This paper is a report on the general geology, structure, and ore deposits at the mine and is based almost entirely on data compiled during the examination. HISTORY Mineralization was first discovered on the property in May, 1929, by A. F. Hewitt, who, with J. A. Borthwick, decided that the country had prospecting possibilities and commenced work in April. Mr. Hewitt staked, and with his partner, J. Ziyone, still holds the group of claims lying immediately east of the present Argosy holdings. Mr. Borthwick staked six claims, K.R.L. 9,681 to 9,686, inclusive, adjoining Hewitt's on the west and later added claims K.R.L. 9,733 to 9,740, inclusive, to his holdings. F. Fiskness, who was also prospecting in the region, discoveted No. 1 vein on Borthwick's claims, and with his two partners, L. Parth and H. Jones, staked claims K.R.L. 9,758 to 9,763, inclusive, to the northwest of Borthwick's and Hewitt's ground. In July, 1929, the Casey Mountain Operating Syndicate, which held ground to the northwest, secured an option on Borthwick's claims. The Fiskness-Parth-Jones property was added in January, 1930, to protect the west dip of the veins. During the winter of 1930-31, the veins were explored by diamond drilling, No. 2 for a length ?of 900 feet and a vertical depth of from 51 to. 375 feet? and No. 3 .t~ a slope depth of 175 feet. I~ August, 1.931, after the msrallat10n of a miming plant, the Casey Summit Gold Mmes, limited, was incorporated to take over the property from the Syndicate. The Maple leaf Mines optioned a controlling interest in the stock in August, 1933, and on September 17th, 1934, brought a 50-ton amalgamation mill into production. From early in October to December 27th, 1934, when the mill was closed down, 4,094 tons of ore were treated to yield 887.89 ounces of gold and 87 .26 ounces of silver( ! ).

Jan 1, 1937

By M. B. Shongwe, R. K. K. Mbaya, C. K. Thubakgale, P. Mendonidis

"A low-grade nickel-bearing ore was pre-concentrated on a SB40 laboratory model Falcon Concentrator. Three size fractions, -212+125, -125+75 and -75 µm were tested with the intent to upgrade the nickel content by reducing their content of fine silica, which becomes amorphous during leaching. Separation of silica from nickel-bearing magnesium silicate minerals; hornblende, clinochlore and talc, was more efficient in the finer particle sizes, -125+75 and -75 µm, and at high G-forces. For these two size fractions, nickel was upgraded from 0.06 to 0.07%, with a reduction in silica content from 52.12 to 50.71%. Leaching of feed samples prior to pre-concentration resulted in the formation of amorphous silica, indicative of the presence of fine silica in the samples. Leaching of nickel from the -125+75 µm fraction increased with increasing G-force and feed pulp percent solids, while that of -75 µm was not influenced. The optimum nickel recovery was 15.68% to the Falcon underflow, using the feed size fraction of -75 µm, 30 G and a feed pulp solids content of 10% solids, after 120 minutes of treatment.INTRODUCTION A low-grade South African nickel-bearing ore with about 0.06% Ni was previously studied for atmospheric pressure acid leaching using sulphuric acid. The study was motivated by the continuing search for methods to treat of low-grade ores, which appear to be an increasingly important source of valuable metals as a result of a declining production from rich sulphide ores. Nickel is a base metal that is used as an alloy in steelmaking and chemical industries because of its resistance to corrosion (Haudet et al., 2015). Nickel laterite ores with grades lower than 1.50% Ni are considered to be complex ores, accounting for 70% of the global nickel reserves (MacCarthy et al., 2016). Their complex nature may be attributed to the indefinite mineralogy of these ores, in which nickel may occur in more than one mineral host. In some cases, nickel may be finely disseminated in the ore. The grade of such an ore and complex mineralogy do not justify the application of pyrometallurgical methods for extraction of nickel (Zhu et al., 2012). This suggests that there is a need for less expensive alternative methods of extraction. This study investigates the application of centrifugal gravity concentration, using the Falcon concentrator, followed by leaching to determine the effect of pre-concentration of the ore prior to extraction by atmospheric pressure leaching with sulphuric acid."

Jan 1, 2017

By J. Forster, C. A. Pickles, Mackowiak. K.

The future development of nickel laterite deposits is crucial with regards to satisfying impending nickel demands. The commercial implementation of new hydro/pyrometallurgical processes in recent years has led to significant improvements in the extraction of nickel from the laterite ores. One potential alternative pyrometallurgical processing route would be to selectively reduce the nickel and produce a concentrate with nickel contents similar to those attainable in the conventional flotation of sulphide ores. However, even bench scale studies have not yet been able to produce a high grade (>15% nickel) concentrate from a low grade laterite ore. The present research involved the microwave vacuum reduction roasting of a nickel laterite ore to yield a high grade concentrate, rich in nickel. Firstly, the effects of temperature and pressure on the nickel grade and recovery were investigated using a thermodynamic model. Secondly, briquettes containing a mixture of charcoal and a low grade silicate nickel laterite ore were reacted in a microwave vacuum system. The experimental variables studied were the input microwave energy, pressure and absorbed microwave power. Thirdly, the ferronickel in the reacted briquette was concentrated by magnetic separation and nickel grades of up to 21% could be achieved.

Jan 1, 2015