Search Documents

By M Battersby, F Parra, S Sanchez-Pino

Pneumatic flotation, as developed by Dr Rainer Imhof, has been applied in commercial beneficiation operations since 1987. Over 85 flotation cells have been installed in more than 30 operations, successfully beneficiating a wide range of minerals. Pneumatic flotation is however still considered as a non-conventional flotation process. Due to its highly selective nature in operation it has generally found its market in difficult applications where conventional flotation equipment has struggled. Compa±fa Minera Huasco (CMH) operates an iron ore pellet plant in central Chile. CMH desired to produce a low silica pellet for the export market where the SiO2 grade should not exceed 1.5 per cent. It is common practice in other countries to float quartz from hematite with relative ease. However, the silicates contained in the magnetite ore of CMH are only present as quartz in small quantities and the major undesirable silicates are difficult to float. CMH spent considerable time on developing a reverse flotation process and a reagent suite to try and reduce the silica content to acceptable levels. However, extensive laboratory and pilot plant tests with conventional mechanical tank type cells and with column cells failed to produce the required reduction in impurities. A series of tests with pneumatic flotation cells indicated that this type of flotation technology could achieve the required silica target. A 500 tph Imhoflot pneumatic flotation plant was subsequently constructed and successfully commissioned by CMH, making it one of the largest iron ore flotation plants in the world. This paper details the background and testing that was undertaken by CMH which led them to choose pneumatic flotation as the preferred process and compares the test results with that achieved in the industrial plant.

Jan 1, 2005

By D. Lawrence, H. Lawrence, G. Saevil

Plant operators in the mineral processing industry have been implementing expert systems for several years now, with the intent to increase production and reduce costs. Most systems have been installed on grinding circuits of some type and have proven that the technology is viable in that application. However, few applications on other processes have been attempted for various reasons, such as lack of reference sites, complexity of the process, lack of suitable measurements and lack of confidence in the technology. The successful application of an expert system on an Alumina Calcining kiln is the first such application in the West Indies Alumina Company (Windalco), and demonstrates how this technology can be successfully applied to other processes. The key objective of an expert system is to optimize the process. To achieve this the process must first be stabilized, followed by optimization of the operating parameters. The steps taken by this expert system to optimize the process are: • Check for abnormal conditions and deal with them using a prioritized set of special actions. • Use a selection of rule blocks to stabilize the process. • Once the process is stable, achieve additional benefits through further optimization of the process. The potential benefits that a user can expect to achieve through installing an expert system can be assessed on the basis that performance will improve from that of an "average operator" to that of the "best operator" performing at his optimum for 24 hours a day, every day. The quantifiable process benefits include: • Increased production • More consistent product quality • Reduced operating costs • Reduced standard deviation of key variables • Consistent plant operation The success of this project was measured in several ways: project implementation, realised economic benefits, and realised non-economic benefits. With a commitment to maintain the system over the long term, the longevity of the system is assured and will become an important part of the Alumina Calcining control strategy at Windalco. This paper discusses the various phases of implementing the ABB Expert Optimizer expert system, what the benefits were and the procedure that is in place to maintain the system to ensure continued long term success.

Jan 1, 2002

By G. Goodall, R. Jabbar, S. Muinonen, A. Cummings

"Glencore’s Sudbury Integrated Nickel Operations in Falconbridge, Ontario, has been in operation since 1930. Many new technologies and practices, now common in the industry, have been implemented at the site and contribute significantly to the industry’s understanding. Since the last update on the Smelter operation published at the NiCo 2005 conference, the Converter Aisle has undergone significant changes in the operation of vessels. The Aisle continues to consist of 4 Peirce-Smith converters, but the Slag-Make Converter, and the Slag Cleaning Vessel have been re-tooled or re-built as Hybrid Vessels (HV), each generating matte for the two Finishing Converters and clean slag for disposal. The complicated scheduling and the overall benefits to the Smelter are reviewed here as part of the on-going tradition of industry support.INTRODUCTION The Glencore Nickel Sudbury Smelter operates a Roast Reduction Smelting (RRS) flow sheet with a specialized Converter Aisle operation that maximizes metal recovery to Bessemer matte. The main process equipment consists of a pair of Fluid Bed Roasters, a submerged-arc Electric Furnace and four Peirce-Smith converting type vessels (Figure 1). In the RRS flow sheet, concentrate is roasted with air to remove sulphur as SO2. Calcine from the Roasters is mixed with petroleum coke upon addition to the six-in-line rectangular Electric Furnace. The coke back reduces the metallic oxides, producing a matte and a discard slag.Gases from the Roasters are sent to a single pass absorption Acid Plant where the SO2 is converted to acid. Under the previous operating targets, approximately 70–75% of the sulphur in the incoming concentrate was fixed as acid with the balance going to the Electric Furnace in calcine for Bessemer matte production. Off-gases from the Electric Furnace and Converters are combined in a balloon flue, treated for particulates in electrostatic precipitators and discharged to the Smelter stack without any SO2 fixation."

Jan 1, 2017

The suitability of char from Yallourn brown coal as a filler for carbon electrodes used in the aluminium smelting industry has been assessed on the basis of the specifications usually laid down by electrode manufacturers for petroleum coke fillers. The purity specifications can be met by a simple acid leaching of the coal before carbonization. The true density of char produced at 1200°C is within 5 per cent of the specification figure, but the bulk density of char obtained from as-mined coal is only half that specified for petroleum coke.Although this can be increased to three-quarters of the specified figure by briquetting the coal before carbonization, further increases would be very difficult to achieve, as the remaining shortfall in density is due to an extensive microporosity developed in the particles of char during carbonization.

Jan 1, 1970

By Fathi Habashi

During the recovery of copper, nickel, zinc, and lead from their sulfide ores, large amounts of sulfur dioxide are generated. A fraction of this gas is captured and transformed into sulfuric acid and liquid SOl and the remainder is emitted to the atmosphere. Numerous proposals were made to solve this problem. Hydrometallurgy offers the only valid solution by which elemental sulfur instead of sulfur dioxide can be produced. Sulfur can be easily stored, transported long distances, or transformed into S02 and H2S04 when desired.

Jan 1, 1996

By W. K. Gowans

The Sullivan Concentrator treats a complex sulphide ore consisting mainly of galena, marmatite and pyrrhotite. The ore carried minor values in silver, antimony, bismuth, indium and tin. The products of the concentrator are lead, zinc, iron and tin concentrates. The minor metal values other than tin are recovered from the concentrates by the refining processes in the smelter at Trail.The concentrator was completed in 1923 and first operated with a capacity of 2,500 tons per day. In the years 1926 to 1929 changes to equipment and buildings brought the operation up to 5,500 tons per day and then, in the years 1935 to 1945, the daily tonnage was raised to 8,500 tons. After 1945, the rising insoluble content of the ore made it necessary to consider the construction of a Sink-Float plant. At this time consideration was also given to the construction of a new crushing plant underground to replace the rolls crushers and cone crushers located at the concentrator. By 1949, construction of the Sink-Float plant was completed along with a new underground crushing plant, and an electric haulage way between the underground crushing chamber and the Sink-Float plant. The ore was hauled in forty 14-ton car trains which discharged into a 3,000 ton receiving bin by means of an air operated five car rotary dumper. With the start up of the,Sink-Float plant, a 1 ½ ft. by 12 ft. Hardinge rod mill was also brought into operation. The Sink-Float plant made it possible to treat 12,000 tons/day of mine run ore with up to 40% of this being rejected as float.

Jan 1, 1975

By R. M. Lauer

OVER THE YEARS concentration of the Sullivan ore has become progressively more difficult as the ore has become increasingly refractory. The introduction of new reagents and changes in the flow of the concentrator have been and continue to be necessary in order to maintain and improve the efficiency of metal extraction. Prior to the completion of the pilot mill in 1957 ail new ideas were developed in laboratory equipment. If these were found to be promising, there was no alternative but to introduce them directly to the main mill flow in one step. When a major change was involved there was always the possibility of serious losses while the mill was being adjusted to the changed conditions. It proved difficult to determine whether these losses were the result of the changed procedure or of failure to adjust the flow to a change in the condition of the ore during the test. The pilot mill has provided a valuable intermediate step between the laboratory and the mill. It avoids the losses which occurred when unsuccessful full-scale tests were attempted, and conversely it provides assurance that full-scale flow changes can be made without serious hazard.

Jan 1, 1960

The optimal production of the metals needed by society is best achieved by a close synergy between concentrate production and smelting. For over 75 years commencing in 1923, such a close relationship was enjoyed between the Sullivan Mine in Kimberley, British Columbia and the Trail zinc-lead complex. An effective program of continual change and improvement at the Sullivan concentrator was achieved through a strong entrepreneurial spirit. This spirit was in part maintained through a favourable location in south-eastern British Columbia that resulted in a dynamic local community that has survived the closure of the mine. Innovations to be reviewed include feed forward reagent control, the use of column flotation to zinc and lead cleaning and the application of ultrafine grinding coupled with concentrate dewatering using hyperbaric filter technology.

Jan 1, 2005

By Paul Ransom

Sullivan Production and Present Value ? Mined 150 million tonnes ? 6.0% Pb, 5.5% Zn, 2 opt Ag, + Sn, Cd, In ? May 2006 Value + US$43 Billion ? 50% of metal from first 37% (55MT) of tons ? True original size ~ 110-120 Mt @ 17%Pb+Zn ? 30MT unconsolidated fill + HW cave

May 1, 2006

It is shown that factors associated with conditions whIch are thermodynamically favourable for the formation of CuS04 in roasting lead to high copper recoveries on leaching. One of these factors is the ratio of the air to solids used for roasting. This can be varied widely without loss of temperature control in the sulphating roasting of many Australian copper concentrates.Thus, the prediction of the air to solids ratio corresponding to a maximum p(SO3) value, for a given concentrate and temperature, is of importance in fluidized bed roasting, since this could be associated with the formation of a maximum quantity of CuS04. Calculations are presented which predict this air to solids ratio, not only for a given concentrate composition, temperature, and pressure, but also for air which is enriched with sulphur dioxide or oxygen.

Jan 1, 1968

By Utley Wedge

(Cleveland :Meeting, October, 1012.) In general, the art of securing copper from sulphide ores or concentrates may be said to consist of : (1) separation, in the molten state, of copper sulphide with some iron sulphide, from the great bulk of gangue-material, followed by (2) elimination of sulphur and iron from the molten matte, by oxidation and fluxing, leaving metallic copper. The definition is made broad enough to cover the general practice, whether or not the ore or concentrate is given a preliminary roast for the removal of surplus sulphur. The above process, which I will designate as the smelting process, makes necessary the presence in the ore, or the addition thereto, of such silica, lime, and iron as may be necessary for the formation of a light and fusible slag, together with sufficient sulphur and iron for the formation of copper-matte, and also the addition to the matte of sufficient silica for the oxidation and fluxing-out of the iron. These items, together with the cost of fuel sufficient for melting the mass, constitute important items of cost in the copper-smelting process. Numerous cases where fluxing-materials are scarce or dear, or where fuel is expensive, have caused special attention to be given recently to the wet process of copper-extraction, which may be designated as consisting of forming soluble copper salts which are then washed or leached out of the ore or concentrates by dissolving in water or acid. Metallic copper may then be secured from the solution of copper salts by electrolysis; by cementation; or copper can be brought out of the solution as a * Presented also at the Joint Meeting of the Institute and Section IIIa, Metallurgy and Mining, of the Eighth International Congress or Applied Chemistry, New York, N. Y., September, 1912, and here published by mutual agreement.

Dec 1, 1912

The ore, as sent to the sulphide mill, is a mixture, in varying proportions, of oxidized ore from above the 450-ft. level and sulphide ore from above and below (to date mainly from above) that level. The oxidized ore consists essentially of quartz, with varying amounts of kaolin and clayey material which passes into the slime and, in excess, creates a serious problem in settling. Pyrite may be present in small amounts. The gold is free and extremely fine.The ore, classed as sulphide, consists of pyritic material in a quartz gangue. The sulphides have been exposed to weathering, leaching by descending solutions and attack by ascending gases and actual burning. The alteration products occur in widely varying proportions and this variation adds to the difficulties in settling and control of grinding and flotation.Copper Minerals.-Chalcopyrite is the principal copperbearing mineral with smaller and varying amounts of covellite, which is sometimes associated with chalcocite, both these minerals being apparently of secondary origin.Examination of tailing by Dr. Stillwell has revealed all stages between clean chalcopyrite, chalcopyrite tinged with covellite, covellite with residual cores of chalcopyrite and clean covellite.The chalcopyrite is free floating, very few particles of clean chalcopyrite being observed in the tailing from granular flotation. The chief tailing loss is probably as chalcopyrite occurring as small inclusions in pyrite...

Jan 1, 1939

By L. C. Graton

I. INTRODUCTION. The Relations of Scale in Geologic Work. MANY features of human accomplishment rest upon the possibility of representing natural objects on a scale of different magnitude from the actual one. Most drawings, for example, are of very different scale from that of the object depicted, and indeed are useful mainly on this account. More important than the question of mere convenience, however, is that of meeting the physical limitations of vision.1 Thus, a mural

Jan 5, 1913

By William Phillips

THE earliest mention of the sulphur deposits in what is now Culberson County, Texas, seems to be contained in "Explorations and Surveys for a Railroad Route from the Mississippi River to the Pacific Ocean, United States War Department, 1853-54." Captain John Pope (he was not commissioned Captain until 1856) had charge of one of the divisions of the Survey. In his report on the geology of the route, William P. Blake noted the analysis of a greenish earthy powder from the bed of Delaware Creek, at the springs, which contained 18.28 per cent. of free sulphur. The report states: "From the chemical examination of the waters of the spring, by Dr. Booth, it appears that chloride of sodium is the principal solid ingredient, but considerable quantities of sulphide of sodium are present. * * * * * The waters from the springs examined by Dr. Booth were alkaline, but it is possible that some of the springs are acid; and in this case the precipitation of sulphur would ensue on the mixture of the acid waters with those charged with alkaline sulphurets and the origin of the deposits of sulphur would he thus explained." This is the first reference to the existence of sulphur in that region, the first mention of the character of the water found there and the first suggestion as to the' origin of the deposits. The possibility of the existence of acid waters was confirmed by E. M. Skeats, nearly 50 years later.1 In this publication, Skeats says that the water from the sulphur deposits in Section 11, Block 114,, Township 5, contains 79.08 grains of free sulphuric acid per gallon. This locality is about 13 miles (21 km.) south of Delaware Creek and about 35-miles (56 km.) east of the Delaware Springs. The next published reference to the existence of sulphur in Trans-Pecos Texas is in the First. Annual Report of the Geological Survey of Texas, 1889, page 226, where W. H. von Streeruwitz mentions that sulphur is found in that region, but gives no further information. In the

Jan 9, 1917

By Neill A. O, Worth I, O&apos

The formation of Kalgoorlie Consolidated Gold \lincs 1't% Ltd hati pla,xd all the major the gold producing areas adjacent to the city of Kalgoorlie/Boulder under one management, enabling the development of the 'Super Pit'. This has facilitated the economic optimisation of the resource. Optimisation has been achieved by the timely collection and presentation of data and the use of these data to reduce costs and improve planning procedures.In particular the following issues have been addressed:ò the reduction of the impact of the mining activity on the adjacent urban population centre,ò implementation of safety procedures for mining through underground voids,ò the determination of the optimum bench height and mining method for ore production benches and blast design. This system is fundamentally based on maintaining a high degree of selectivity while achieving a balance of relatively low unit operating costs,ò assessment of the comparable benefits of contractor versus owner operated earth moving equipment, andò the development of effective ore body models, production reconciliation and the long term mine planning procedures to facilitate effective strategic planning.

Jan 1, 1993

By H. D. B. Wilson

THIS PAPER has resulted from a preliminary draft of a background paper prepared anonymously for the Science Council's Special Study of the Solid-Earth Sciences in Canada. The draft was circulated to eighty-four Canadian geologists; forty-eight of these responded with criticisms of the draft. All critical re-marks were considered carefully and most have been incorporated in the final paper in an attempt to balance the author's bias, although this bias undoubtedly still appears. Indications of support for a particular statement or section were also helpful, because many of these contradicted criticisms and enabled the author to judge the consensus. Thus, all criticism could not be incorporated because the referees did not agree. It is interesting that some remarks were con-sidered overstatements by some and too bland by others. The most perplexing critical comment was one which suggested a new author was needed, and went on to name the existing author as the most likely candidate.

Jan 1, 1970

By T Berryman, C Adams, L Griesel, C Buxton

The Kalgoorlie Consolidated Gold Mines (KCGM) schedule model is used for short- to medium-term planning and scheduling at the Fimiston Open Pit Mine (the Superpit). The model bridges the gap between grade control models, which are generated one to seven days prior to mining, and the annual Resource and Reserve model update that is used for budget and life-of-mine plans. The schedule model uses close-spaced grade control data (up to 20 000 m drilled each month) as well as resource definition drilling to estimate production volumes three to six months ahead of current mining. From this, a rolling 12 week mining and processing schedule is generated on a monthly basis. As the KCGM operation is mill and sulfur constrained, the schedule model needs to be an accurate predictor of ounces as well as tonnes, gold and sulfur grades. The process must also be time efficient in order to meet monthly planning and reporting deadlines. Due to the large amount of new data generated each month, model updates using conventional interpretation and estimation techniques have proven to be impractical and too time consuming. This paper will examine the evolution of the Superpit schedule model process over time, recent improvements and the effect of different model outcomes on short- and medium-term mine planning.CITATION:Griesel, L, Buxton, C, Adams, C and Berryman, T, 2014. The superpit schedule model – A time-efficient data-driven estimation for short- and medium-term planning, in Proceedings Ninth International Mining Geology Conference 2014 , pp 397–410 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By J. Ivon Graham, T. D. Jones

This paper presents methods in use in British coal-mines for the suppression of dust. While the effect of present-day treatment will not be apparent for several years, in pits in South Wales, concentrations as high as 50 gm./1,000 cu. ft. were some years ago fairly common in various operations at or near the coal-face, with an average concentration of 20 gm./1,000 cu. ft. Now the average concentration in what were previously dusty pits is probably not greater than 3 gm./1,000 cu. ft.Nevertheless, every effort must be made to improve conditions still further. In South Wales it is claimed that 99% of the coal-faces requiring treatment have been dealt with either by wet cutting, infusion, or hand-spraying. The extent of the treatment by these methods can be gauged by the fact that the total length of coal-faces now being treated by water infusion in Cardiff and Swansea areas is about 33,000 yds., by wet cutting 25,000 yds., and by hand-spraying nearly 14,000 yds. The influence of methods of mining upon dust production is only mentioned very briefly in this review, and there is no doubt that much more is to be learned concerning this aspect. It can be stated, however, that good mining practice is essential to prevent unnecessary dust production, and the problem of dust in mines will be solved only by the closest co-operation between mine managers, workmen, and scientists.

Jan 1, 1947

By L Reynolds

Oceana Gold (New Zealand) Limited (OceanaGold) has commenced mining operations at the Globe-Progress mine in the Reefton Goldfield, an area that historically produced 2.1 million ounces of gold from underground quartz lode deposits. During the construction phase of the operation, advanced exploration programmes targeting near-mine prospects commenced. The focus of this ongoing exploration is the discovery of near surface, refractory gold deposits that are amenable to open-pit mining methods and treatment through the Globe-Progress process plant.   One such discovery is the Supreme deposit, located along strike approximately 3 kilometres to the south of the Globe-Progress mine. The Supreme area was tested extensively with diamond drilling in 2006, defining a deposit with similar mineralisation styles and grade to the Globe-Progress deposit. The mineralisation comprises a central core of gold-bearing quartz veins, locally containing visible gold, surrounded by a more extensive halo of refractory, disseminated gold associated with an arsenopyrite-pyrite sulphide mineral assemblage. Three subparallel mineralised structures host the mineralisation. These structures are highly oblique to the country rock orientations, extending approximately 250 metres along a northeast-southwest strike and at least 220 metres down-dip, with an average thickness of 12 metres and moderate dip to the southeast. Records of historical workings indicate that mineralisation extends further down-dip and remains open at depth. 5175 ounces of gold were mined historically at the Supreme deposit. OceanaGold have defined an Inferred Resource for the Supreme deposit of 0.84 million tonnes at 1.46 grams per tonne for a total of 40 000 ounces (at a 0.8 grams per tonne cut-off). Higher-grade zones of mineralisation exist within the deposit and further drilling is planned to test the extent, continuity and geological controls of the mineralised structures.

Jan 1, 2007

By R. Xie, J. Liu, Y. Zhu, Y. Li

The surface chemistry features of spodumene and feldspar were investigated by XPS, zeta potential measurements and first principle calculation. The results indicated that the zeta potential of feldspar was more negative than that of spodumene. The most possible active sites on spodumene and feldspar were Al sites and O sites. However, the number, type and activity of these sites were different. The flotation behaviour of five new collectors (cationic collector decyloxy-propyl-amine (DPA), anionic collector α-bromododecanoic acid (α-BDDA), N-Dodecyliminodiacetic acid (DIDA), and mixed collector DPA-BDDA and DPA-DIDA) were investigated through micro flotation experiments and analysed based on the surface chemistry features of spodumene and feldspar. The results revealed that cationic collector DPA processed a high collecting capacity for both minerals. The highest recovery of spodumene and feldspar was 91.60% and 95.30% respectively. The anionic collector α-BDDA and DIDA showed poor flotation behaviour. The recovery of spodumene and feldspar when using α-BDDA as collector was less than 5% and when using DIDA as collector was less than 1%. For the mixed collector DPA-BDDA, the recovery of spodumene and feldspar was 82.27% and 32.61% respectively at pH 4.5. For mixed collector DPA-DIDA, the recovery of spodumene and feldspar was 35.69% and 9.08% respectively at pH 5.8. The results indicated that the cationic collector had high collecting capacity, but low selectivity. The collecting capacity of anionic collector for spodumene and feldspar was poor at the same concentration compared with cationic collectors. Mixed collector DPA-BDDA showed good flotation behaviour and selectivity on spodumene. The ―COO- in mixed collector might improve the selectivity. However, too much ―COO- would increase the electrostatic repulsion, which might cause low selectively and poor recovery. Keywords: Spodumene, feldspar, surface chemistry features, mixed collector, flotation

Jan 1, 2020