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By I D. Hair

A numerical groundwater model was used to predict the nature of interactions between heavy mineral sands mining operations and the groundwater regime at North Stradbroke Island, Southeast Queensland. North Stradbroke Island is formed mainly of massive vegetated Cainozoic aged dune sands. Since the 1950s, a number of operations have mined rutile, zircon and ilmenite on North Stradbroke Island. In such mining operations, a dredge and an associated processing plant are floated in a pond either above or below the prime groundwater table The numerical model was based on the North Stradbroke Island Groundwater Management Model, developed in 2001 by the Department of Natural Resources and Mines (NR&M). This model was constructed using the modular, cell-centred, finite-difference groundwater flow model, MODFLOW, developed by the United States Geological Survey (USGS). The NR&M model was refined to improve model resolution and detail for simulating dredge mine paths. Of particular interest were: water table fluctuations induced by mining for the identification of slope stability risk; the optimisation of sand spear arrays supplying mine make-up water requirements; inflow and outflow rates to the dredge pond for overall mine water balance; and interactions with Eighteen Mile Swamp and water supply bores controlled by the Redland Shire Council (RSC). Depending on the water level in the dredge pond compared with the surrounding groundwater table, water either flows out of the dredge pond into the groundwater regime or from the groundwater system into the dredge pond. The MODFLOW river package (RIV) was utilised for the modelling study. Dredge pond movement was simulated using specialised software developed by Watermark Numerical Computing (WNC) to interpolate pond positions and map the dredge path to the model grid both spatially and temporally. As tailings placement affects the groundwater system through increased recharge, the MODFLOW recharge package (RCH) was utilised to simulate different recharge efficiencies. Long-term model predictions were made for the period from July 2002 to May 2021. Results of the modelling showed that exchanges of flow between mining operations and the groundwater regime could be effectively managed. The study provided valuable information which was used to develop an appropriate mine water management plan.

Jan 1, 2003

By Harvey S

After considerable experience in connection with the magnetic iron-ores at the South, especially in the Cranberry district of western North Carolina and eastern Tennessee, the writer was led into a thorough investigation of the magnetic separation of iron-ores, and, from this investigation, has gradually been developed a separating machine which it is the purpose of this paper to explain. Numerous United States patents have been issued upon machines for magnetic separation. Since the original patents date back to 1849, it is not possible to-day to establish broad claims upon the principles of electro-magnets as applied in such machines. Among these inventions there are certain types of separators which have very marked practical advantages over the others; and these successful machines, while varying radically from each other in their proportions and arrangements, are all based upon a principle originally explained in a patent which was granted in 1870, and which has now expired by limitation. The problem of practically separating iron-ores, as found in the mines, is very different from the separation of a purely magnetic substance from a purely non-magnetic one, since the crystals of magnetite, scattered through the ore-bearing rock, and sometimes collected in masses, are of very different sizes. The average size of the particles may be definitely ascertained for each mine, and the crushing- and screening-machinery may be adapted thereto, but there will still be a large percentage of very fine crystals which, in the crushing for the average, will not be broken apart from the gangue. If these mixed particles are thrown into the heads they carry gangue with them, and if thrown into the tails they occasion a loss of ironore. There are two methods for preventing such loss. One is, to crush all the material to the size of the finest particles, which means, of course, a large crushing-plant, heavy wear and tear, and excessive cost for repairs, and the production of only a very fine-grained concentrate which our blast-furnace managers, to say the least, do not

Jan 1, 1893

By Rudolf Kudlich

"The realization that fuel costs, high as compared with those enjoyed prior to the war, are apparently permanent, has caused steam power plant operators to apply their best efforts towards improving the fuel economy of their plants.Though they have given much attention to the reduction in combustion losses by improving methods of firing and improved equipment, and to diminishing the heat losses by more care in maintaining clean heating surfaces and tight and adequate baffling, they have almost universally neglected the waste that occurs from discarding unconsumed combustible with the boiler refuse.This loss varies between extremely wide limits, depending on the quality of the coal, the type of plant and the firing practice in general, the higher the ash content of the coal the greater will be the amount of combustible appearing in the refuse, and the losses will be greater as the boilers are driven at higher ratings. The refuse from stoker fired boilers is apt to have a lower aver-eve percentage of combustible than that from hand fired boilers if operated at the same rating, but stoker fired boilers are usually pushed to higher ratings and the combustible in the refuse will run higher.In actual operating practice the amount of unconsumed combustible in the boiler refuse may range from a few percent to as high as forty or fifty percent. In an informal statement one well known combustion engineer estimates that the average amount of combustible in the refuse from a large stoker fired central station boiler plant is about 15 or 20%. Assuming that the coal burned contains 12% of ash, and allowing for flue dust losses, approximately two percent of the coal will be lost through this channel. The operating engineer of one large modern central power plant reports that when burning coal of about the same ash content, the average amount of combustible in the refuse was in the neighborhood of 33%, en tailing a loss of approximately 4% of the heating value of the fuel fired. Where the coal used daily runs up into hundreds of tons, an appreciable economy could be effected by recovering this percentage."

Sep 1, 1924

Discussion of the paper of H. R. Barnhurst, presented at the New York Meeting, October, 1913, and printed in Bulletin No. 82, October, 1913, pp. 2523 to 2532. H. R. BARNHURST :-I would say that in that experimental furnace we were trying to adopt the proper form for burning low-volatile coal. I was enabled on one occasion to burn coke having 15 per cent. ash. Tailings, or washings from the preparation of coal for by-product coke ovens, which came to me 52 per cent. ash, were burned without any trouble. So that ash really might be regarded as not burning, but simply diluting the fuel that was there, without doing any particular harm except that we had to put in more of the material to get the same quantity of heat. With anthracite, I burned coal right along with from 24 to 27 per cent. ash, but it took a furnace of peculiar form in which the products of combustion had to be diverted to the point of entrance of the fuel. With an ordinary furnace the fuel is deflagrated by its surroundings as soon as it enters the furnace. In a cement furnace this action is secured by the hot clinker passing clown, and the walls of the kiln are heated upon contact with it, revolving as they do', so that deflagration occurs very quickly in a cement kiln. Deflagration would occur very quickly in an open-hearth furnace also. There is no trouble in burning anthracite coal under those conditions for three or four days successively. It should be very well dried and thoroughly pulverized. Aside from that, there is no trouble. WILLIAM A. EVANS,* Boston, Mass. :-In a 52 per cent. ash, or any high-ash coal, what disposition is made of the ash ? MR. BARNHURST :-Most of it went out through the flues. Some of it remained in the boiler and a quantity remained in the furnace. We raked it out and carried it away. MR. EVANS :-Do you experience the forming of stags in flues with such coal as that?

Jan 12, 1913

By Fred Wise

THE Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed, The oxide ore is not segregated. GENERAL METALLURGY The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxi-dized, The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices, On this basis metallurgical results are viewed in the light of an optimum tailing rather than a per-centage of extraction.

Jan 1, 1940

By Arthur F. Wendt

In the manufacture of charcoal-blooms, washing or cleansing of the ore from adhering gangue has been practiced for many years. A sluice-box is even to-day used for that purpose in the Southern Appalachian Mountains. For the Lake Champlain bloomaries hard magnetic ores are crushed with stamps and then sluiced or jigged in hand-jigs. The stamps used in the Lake Champlain region are of the most primitive kind. They consist of a trip-hammer, dropping on a perforated cast-iron plate, through which the crushed ore escapes. Only friable rock can thus be successfully treated. To render the raw ore friable, a calcining or burning in piles on a bed of cordwood is resorted to. An expense of 35 cents per ton of crude rock treated is thus incurred. The crushed ore is shovelled or otherwise fed into jigs without previous sizing, except such as is incidental to crushing through the perforated bottom of the stamping-machine. In jigging this ore, it is allowed to pass through the bedding into the "hutch." The tailings are scraped off and wasted. They contain a large percentage of fine ore. In many instances, this is regained by subsequent sluicing. Such was the condition of the industry, when the writer's attention was first called to it in 1882. No attempt had been made to introduce into thiS branch of ore-dressing the notable improvements adopted in concentrating the ores of other metals. The result of the writer's investigations was the building in 1882 of the "Crown Point Separator." At the Crown Point Mines, N. Y., lean ore unfit for the blastfurnace (containing less than 331/3 per cent. of iron), was concentrated and used in the bloomary connected with the mines. The ore was burned here as elsewhere, and treated as described. The plant designed by the writer for Crown Point, is nearly a counterpart of that since erected by Messrs. Copeland and Bacon, contractors, for the Theal mine, Putnam County, New York, which is illustrated by the accompanying plate. The only difference is in

Jan 1, 1885

By N. Nehoroshev, P. Solozhenkin, N. Voyloshnikova, A. Pshennikov

"Molecular modeling of clusters of major bismuth minerals, Bi2S3, was carried out whereby molecular models of minerals and reagents clusters were created. Their HOMO and LUMO energy levels were calculated and using known modelling equations, the absolute hardness ?, eV and electronegativity were calculated. The charges of separate atoms of clusters and the change of properties of the minerals were determined which allowed predictions of the flotation properties of raw ores containing rare bismuth minerals. Modeled complexes containing the clusters of minerals and various collectors were constructed, consisting of atoms and bonds within clusters of minerals. Major physical and chemical parameters were determined for the complexes created, including the total and electronic energy. Based on the energy calculations, predictions of the collector efficiency were proposed. The efficiency of collector was established for various thiolic reagents, the compounds of phosphorus (III) and mixtures of reagents. Theoretical results were checked and confirmed against laboratory ore flotation tests. A gold-bismuth-magnetite ore type was studied where gold, bismuth and iron (magnetite) are of industrial interest. The ore contains 2.8 g/t Au, 1.0 g/t Ag, 0.26% Bi, 45% Fe and 1.25% As. Sulphide mineralization can be mainly found in the form of pyrrhotite, pyrite and arsenopyrite. Bismuth can be mainly found in the form of bismuthite and native bismuth. Gold is fine. A flow sheet for Zhelezny Kryazh ore treatment was proposed. This flow sheet consists of grinding to 95-98% -74 µm, sulfide flotation with two cleaner with xanthate addition, cyanidation of flotation concentrate and flotation tailings included treatment with with CaO (6 kg, 6 hours), cyanidation (8 hours) and CIL (8+24 hours). Saleable products are gold doré, bismuth and iron concentrate."

Jan 1, 2014

By Clare B. Carlson, C. P. Proctor

THE history of the froth flotation of coal is relatively short. The flotation process was applied to fine-coal cleaning about the time of the end of World War I. Coal flotation finds more widespread use in Europe than in the United States and is used on a commercial scale in nearly all major coal-producing countries in the world. In Fig. 1 the general arrangement of a froth-flotation plant at a British colliery is shown. GENERAL FEATIJIEES OF THE PROCESS Application The flotation process is applicable to the cleaning of the fine sizes, particularly minus 48-mesh coal, although it may be used on sizes as coarse as 1/8 in. Since most cleaning processes usually are not effective on sizes smaller than 48-mesh and certainly not effective on minus 100- mesh, the froth-flotation process offers a solution to the problem of cleaning these sizes. This process consists-of agitating the feed pulp with a small amount of reagents, usually 1.0 to 2 ½ lb. per ton of feed in the presence of water and air; the reagents help to form small bubbles that collect on the particles having a greater affinity for the bubble surface than for the water surface, in this case the fine coal, and the bubbles buoy these particles to the surface to form the frothor concentrate. Slate and shale particles remain wetted and suspended in the water to be drawn off as tailing. The frothers commonly used in coal flotation are cresylic acid, pine oil, and alcohols. More expensive frothers, used in ore flotation, work satisfactorily but do not compare with these cheaper reagents for over-all cost. Collectors- used are low-grade petroleum fractions, such as cheap, partly refined fuel oil, or lubricating oil, sometimes mixed with small amounts of kerosene. Other collectors used are the fatty acids and their

Jan 1, 1943

By C. Bazin

"Discrete transfer functions which use the grinding circuit feed tonnage to predict mass flowrates at reagent addition points in the flotation circuit were calibrated using a dynamic grinding/flotation simulator. Predicted values are multiplied by g/t ratios to calibrate reagent flow. The transfer functions are currently integrated into the distributed control system at Les Mines Selbaie.INTRODUCTIONThe development of adequate tools for the control of flotation processes is a continuous challenge. Since the early 1970's there have been numerous publications dealing with practical applications of flotation process control. A review of that literature shows that basic feedback and feedforward controls are the two preferred strategies for flotation processes. Some advanced control schemes started to emerge (e.g. Zaragoza and Herbst, 1988). Feedforward and feedback aim at reducing the effect of the following flotation process disturbances (e.g. McKee, 1991):0 Changes in throughput;0 Changes in feed grade;0 Changes in particle floatability (size, liberation, surface oxidation ... );0 Deliberate changes of set-point;Feedforward strategy acts on manipulated variables, mainly reagent addition, to attenuate the effect of disturbances such as changes of feed rate and grade. The feedforward strategy is driven by measurements of disturbances. Feedback control acts to compensate unmeasured disturbances, such as ore floatability, and to bring process variables to targeted set-points of concentrate or tails grades. The feedback control action is driven by set-point error, caused by either a deliberate change of set-point or unaccounted disturbances.This paper proposes a way to improve feedforward control schemes by addition of a simple model of solids transportation. The feedforward strategy considered in this paper is of ratio type, i.e. the manipulated variable is changed according to a predefined ratio between this variable and the measured disturbances. A typical example is the ratioed reagent addition to mass flow of ore."

Jan 1, 1993

By Clare B. Carlson, C. P. Proctor

THE history of the froth flotation of coal is relatively short. The flotation process was applied to fine-coal cleaning about the time of the end of World War I. Coal flotation finds more widespread use in Europe than in the United States and is used on a commercial scale in nearly all major coal-producing countries in the world. In Fig. 1 the general arrangement of a froth-flotation plant at a British colliery is shown. GENERAL FEATURES OF THE PROCESS Application The flotation process is applicable to the cleaning of the fine sizes, particularly minus 48-mesh coal, although it may be used on sizes as coarse as % in. Since most cleaning processes usually are not effective on sizes smaller than 48-mesh and certainly not effective on minus 100¬mesh, the froth-flotation process offers a solution to the problem of cleaning these sizes. This process consists of agitating the feed pulp with a small amount of reagents, usually 1.0 to 91/ lb. per ton of feed in the presence of water and air; the reagents help to form small bubbles that collect on the particles having a greater affinity for the bubble surface than for the water surface, in this case the fine coal, and the bubbles buoy these particles to the surface to form the froth or concentrate. Slate and shale particles remain wetted and suspended in the water to be drawn off as tailing. The frothers commonly used in coal flotation are cresylic acid, pine oil, and alcohols. More expensive frothers, used in ore flotation, work satisfactorily but do not compare with these cheaper reagents for over-all cost. Collectors used are low-grade petroleum fractions, such as cheap, partly refined fuel oil, or lubricating oil, sometimes mixed with small amounts of kerosene. Other collectors used are the fatty acids and their

Jan 1, 1943

By Zhao-shu Sun, Mao-lin Li, Rui Cui

"Fankou Mine produced pyrite concentrate (PC) from the tailings by rougher and scavenger flotation before 2010, with the averaged S grade of 40.54% at recovery of 79.34%. The economic value of PC increases with its purity. When the S grade of PC feed to sulfuric acid plant is over 47%, the pyrite cinder can reach a high Fe grade of 62% which can be directly sold as iron concentrate. On this background, engineers in Fankou Mine set the PC with S grade of 47% as target.The direct way to reach that goal was to clean the rough PC Lots of cleaner flotation tests were conducted and the balanced strategy was to produce two ranks PC with the S grade of 47% and 38% at the recovery of 44.20% and 36.17%, respectively. The dual-product flowsheet was successfully industrialized in 2010. In order to further improve the quality of PC, gravity separation was introduced into the flowsheet. The numerical simulation and tests were performed to evaluate the beneficiation and classification performance of hydrocyclone as it treated the rough PC. It was shown that the underflow of hydrocyclone could addressed the S grade of 47%, and the finer size overflow could also be enriched to PC with S grade of 47% by flotation, due to the contained pyrite with well liberation. That is to say, with the ingenious addition of hydrocyclone to treat the rough PC, single PC with S grade of 47% could be achieved. The new flotation-gravity separation flowsheet was commissioned to produce the high quality PC at the end of 2013. The averaged S grade is 47.20% at recovery of 82.26% in 2014. The two steps flowsheet improvement maximizes the enterprise benefits and leads to simple utilization of Fe in pyrite in downstream industry."

Jan 1, 2016

By F. N. Lathe

In 1882, The Arizona Copper Co., Ltd., acquired producing copper mines at Metcalf and Morenci (locally called Longfellow). Metcalf is situated a distance of 7 miles, and Morenci a distance of 6 miles from the general office of the company at Clifton, Ariz. A reduction works with five blast furnaces was built in the center of the town of Clifton in 1884 and operated until January, 1914. Concentration of oxidized ores was begun in 1890-1891, in a mill built adjoining the smelter, and id October, 1893, a leaching process was started for treating tailings from the oxide mill. A sulphide concentrator with 300 tons capacity was built at Clifton and started treating chalcocite ores in July, 1896. This was the first mill to treat low-grade copper-siilphide ores successfully. Another concentration mill for sulphide ores was erected at Morenci in 1900. The sulphide mills are now being remodeled to treat ores by concentration and flotation, with a capacity of 4,000 tons at Morenci and 500 tons at Clifton. The leaching plant has not been running since September, 1914. The average yearly production from this company's mines and reduction works from 1885 to date has been: Pounds 10 years ended 1894................................... 6,864,902 10 years ended 1904................................... 20,439,614 10 years ended 1914................................... 32,665,905 Part of year 1915, to Sept. 12........................... 30,206,106 4 months of 1916, to June.............................. 15,964,840 June, 1916........................................... 4,900,000 Construction on a new smelting plant was started in February, 1912, and finished in February, 1914. In the Transactions, vol. 49, E. Horton Jones published "Unit Construction Costs from the New Smelter of the Arizona Copper Co., Ltd.," which fully describes the construction of this plant, with details of costs.

Jan 1, 1917

By E. V. Daveler

The milling plant of the Alaska-Gastineau Mining Co. is located at the town of Thane, Alaska, on Gastineau Channel, 4 mi. south of Juneau and directly across the channel from the Ready Bullion mine of the Treadwell mines on Douglas Island. Previous to the organization of the Alaska-Gastineau Mining Co., in 1912, the milling was done at the stamp mill of the Alaska Perseverance Co. located at the mine in Silver Bow Basin. But the contemplated increase in production made it necessary to locate the mill on Gastineau Channel, as not enough water was available at the mine during the winter months, nor was there room for taking care of the large tonnage of tailings that would ultimately be discarded from the mill. Character of Ore The gold and silver values occur in quartz lenses associated with pyrrhotite, galena, arsenical pyrites, and zinc blende. These lenses occur in large bodies of slate, schist, and metagabbro, which also carry some valuable metallic contents. With the essential low cost of mining, it was necessary to mine large bodies of low-grade material containing the high-grade streaks. Of the minerals, pyrrhotite, the magnetic sulfide of iron, is predominant, and galena is the next sulfide present in quantity. The gold occurs free and coarse and associated with the iron sulfide in its finer state. The silver occurs with the lead or is alloyed with the gold. A chemical analysis of the ore shows Silver............ 0.10 oz. Magnesia......... 0.4 per cent. Silica............. 59.7percent. Sulfur............ 1.2 per cent. Iron.............. 5.2 per cent. Lead............. 0.1 per cent. Alumina.......... 21.4percent. Zinc............. 0.2percent.. Lime............. 5.0 per cent. Experimental Work Investigations at the old Perseverance stamp mill had developed the following points: Roughing concentration following stamping gives a fair recovery.

Jan 1, 1920

By B N. Morrison

In all metallurgical plants metallurgists require analyses of the various intermediate products, eg flotation feeds, flotation concentrates, tailings, mattes, slags, etc to monitor the efficiency of the plant processes and for plant control purposes. Traditionally these samples have gone to off-site laboratories and there can be some considerable delay before analytical data are received. On-line monitors have to be used for plant control purposes and these are difficult to maintain and can give poor quality data. Recent technological advances in automation in analytical laboratories now allow a far quicker turn-around time for analytical data so they can be of more use to the metallurgist. Automation, which involves the use of robots and fewer people in laboratories, has the added benefit of providing higher quality data as the possibility of human error and bias is minimised. Differences between conventional manual laboratories and automated laboratories are described. Automated analytical laboratories are now becoming the laboratories of choice for mining companies which require large numbers of samples to be analysed on a 24/7 basis on similar sample types using the same sample preparation methods and analytical procedures. Automated laboratories have now been installed in platinum, iron ore, copper and aluminium producing mining companies. A further advantage to metallurgical plants is now the possibility of siting automated analytical laboratories in-plant and designing the plant around the laboratory. Concentrator streams, slags, etc can be sampled by automated samplers, the samples transferred by automation to the in-plant laboratories and all sample preparation and analyses carried out by automated processes. These laboratories require a minimum of staff. Such automated in-plant laboratories have been installed on a platinum mine flotation plant, a platinum mine smelter and steel plants. One is in process of being installed on an iron-ore mine with the sampling being carried out by an automated sampling tower. The procedures followed in planning an automated analytical laboratory for a metallurgical plant are described in detail in the paper. The platinum plant smelter automated system which includes rapid transport of samples by air-tube to the laboratory for the analysis of the samples by XRFS is also described in the paper. It is concluded that automated analytical laboratories can be retro-fitted to most metallurgical plants but they are best suited to greenfield sites.

Jan 1, 2008

The Ashanti mine is located in the Gold Coast region of West Africa in Ghana, where gold is found over a wide area of about 105,000 sq km (40,000 sq miles). Ashanti has been the most productive of all the lode mines and has been one of the world's notable gold producers for over 50 years. The mine is located at Obuasi, 113 km (70 miles) north and east of Tarkwa, which is 80 km (50 miles) inland from the coast. The Ashanti lode is worked on 300 m (1,000 ft) of strike length over a width of 12 m (40 ft), and by 1934 it had attained a depth of 700 m (2, 300 ft). The lode is in pre-Cambrian granite intruding a system- of shists, phyllites, etc., containing fragments of the wall rock with quartz, carbonates, sericite and small amounts of pyrite, arsenopyrite, sphalerite, galena, and gold. The shattered quartz fractures carry sulfides and gold, most of which are invisible. (William Harvey Emmons, "Gold Deposits of the World," McGraw -Hill, 1937.) Because the mill treats an ore composed of refractory sulfides - chiefly arsenopyrite and fine free gold - the flowsheet reflects the complex ore characteristics. Plane tables are used in the grinding, circuit for free gold recovery. After roasting the sulfide flotation concentrates, additional free gold is recovered from the calcines with a rotary belt concentrator and the two concentrates are combined for feed to the amalgamation pans, with the resulting gold amalgam being retorted. Following four-stage crushing to minus 9. 5 mm (318 in.) with two stages of grinding, the cyclone overflow is floated for sulfide recovery. The flotation concentrate is roasted with the calcine reground and roughly cyanided to scalp out the gold readily dissolved in one of two cyanidation circuits. The cyanide residue joins the thickened flotation tailing for cyanidation in a complete CCD circuit using filters. The gold precipitates are calcined before melting in pot-type furnaces. The tonnage to the mill of 3,175 mtpd (3,500 s tpd) with a content of over 14.7 g per mt (0.43 oz per st) places the Ashanti production as one of the top producers in the world; and with its past long life, the total production rates it as one of the world's great mines.

Jan 1, 1981

By Buks Kruger

"After the milling and flotation of the platinum bearing ore, the flotation concentrate is electrically smelted in 6-in-line electric arc furnaces and the resulting furnace matte treated in Peirce Smith converters to remove the iron sulphide. The iron free (less than 1% Fe) converter matte is then granulated and sent to the Base Metals- and Precious Metals Refineries for further processing.During the smelting and converting processes two distinct gas streams are produced.The Peirce Smith converter off gas contains 4% - 9% SO2 which is treated in a conventional Sulphuric Acid Plant to produce sulphuric acid (H2SO4).The electric furnace off gas contains 0.9% SO2 and was (until September 2002) vented through a 91 meter high stack to atmosphere.Off gasses which contain less than 3% SO2 is difficult to handle and quite a few processes had been developed over the years to cope with this gas. A broad spectrum of process suppliers and technology approaches is available and the different technologies can be divided into a few main approaches i.e.Scrubbing of the off gas with an alkaline solution to produce an environmentally friendly and stable waste product which can be discharged onto a tailings dam or permanent landfill site.Physical absorption of the SO2 into a chemical solution to produce high quality chemically pure process chemicals.Catalytic reaction of SO2 with O2 to produce SO3 which is absorbed in water to produce sulphuric acid of less than 90% strength.In October 2002 Impala Platinum commissioned a Sulfacid Plant at its Smelter in Rustenburg which recovers the SO2 from the electric furnace off gasses into a weak sulphuric acid solution.The Sulfacid Plant technology is widely used in the paint pigment industry. This is, however, the first application of the technology on electric furnace off gas.The furnace off gasses contains flue dust which is removed in electrostatic precipitators. The gas exiting the precipitators normally still contains between 40 mg/Nm3 and 60 mg/Nm3 dust and passes through a quench-scrubber before entering the reactors. The quench-scrubber cools the electric furnace offgas down from ± 200 °C to 60 °C and act as a final dust removal stage."

Jan 1, 2003

El Indio, eighty percent -owned by Compania Minera an Jose Inc., a subsidiary of St. Joe Minerals, and twenty percent-owned by private Chilean investors, will open their new flotation and Carbon -In -Pulp plant near La Serena, 480 km (300 miles) north of Santiago, Chile, in early 1981. The underground mine, at an altitude of 4,270 m (14,000 ft), will produce about 1,270 mt per day (1,400 st per day) of ore containing about 2. 2570 Cu, and gold and silver to the extent of 16 and 254 g per mt (0.47 and 7.4 oz per st) respectively. While construction of the plant was underway, development ore, containing some 1,400 kg (45,000 oz) of gold, was shipped to the United States for recovery in a copper smelter. At El Indio the underground ore is rail-hauled to the mill a short distance from the portal where it is crushed, ground, and floated to produce a copper concentrate containing both gold and silver. The crushing plant also crushes waste for use as mine backfill. Approximately 92% of the copper and 60% of the gold and silver are recovered in the flotation concentrate. Flotation tailings, at a grade of 19 g Au and 116 g Ag per mt (3.4 oz Au and 0.55 oz Ag per st), are cyanided for 72 hours in six Pachucas with auxiliary air lifts for pulp transfer. Lime, cyanide, lead nitrate, and a phosphate are also fed to the cyanidation Pachucas. After cyanidation, the pulp is treated in four air lift Carbon-In-Pulp tanks equipped with external vibrating screens for carbon advance. Loaded carbon is stripped in two pressure vessels in parallel. Pressure stripping must be used because of the high altitude. Stripped carbon is dewatered in a spiral classifier before being reactivated in a rotary kiln and returned to the C. I. P. circuit. Pregnant solution is then sent to four banks of four Zadra cells for gold and silver recovery on steel wool. Zadra cell cathodes.are washed with the slimes being captured in a plate and frame filter. The filter cake and the washed cathodes are then dried on an infrared dryer and smelted in two stages to bullion bars. Published reports indicate that the total investment for this unique operation was in the neighborhood of $80 million.

Jan 1, 1981

By J. E. Tobey

UNQUESTIONABLY, manufactured gas will stage a comeback of such huge proportions as to dwarf its previous history. Timing will depend on two things: the diminishing supply of natural gas and the perfecting of processes for large-scale production of high-Btu gas from, coal. What the timing will be, and how soon the new era will begin for manufactured gas, no one knows. It may be just around the corner and it may be several years, but it is certain to come. It will depend on economic changes, and any major world upheaval might accelerate it tremendously. Supplemental gas for pipelines will come no doubt from two sources at least: (1) synthetic-fuel plants of the coal-chemical group, which produce chemicals and liquid fuels with a tail gas high in methane as a by-product; and (2) large-scale gasification plants, which will produce high-Btu gas entirely or as a major product with chemicals and liquid fuels as by-products. There is important activity at the present time, in the development of new processes and the improvement of those now in existence. The first large-scale gasification plants of the second category probably will be strategically located in the coal fields on or near the long-distance transmission gas lines, and these plants probably will operate jointly with large natural gas storage fields. Inasmuch as virtually all the major gas-transmission lines cross one or more coal fields, such locations will not be difficult to find. It will be a practical necessity to produce a high-Btu gas -that is, 750 to 1000 Btu-in order to maintain the capacity of the main transmission lines and local city distribution systems and further to eliminate the problem of appliance changeovers in the cities. If gasification plants on a medium or smaller scale prove to be economical, local gas utilities no doubt will consider .the construction of such plants to supplement their supply of natural gas. If low-temperature distillation processes like those now being developed by the Pittsburgh Consolidation Coal Co. are successful, and there is reason to believe that they will be, utilities that supply both gas and electricity to communities would find this method particularly attractive, as they could burn the char product directly under power boilers, thereby minimizing thermal

Jan 1, 1953

By T. M. Morris, F. M. Lewis, W. C. Lay, Gordon M. Bell

Part I of this chapter is concerned with the description of and metallurgical data pertaining to the flotation circuits used in the London mill of the Tennessee Copper Co., at Copperhill, Tenn. The ores treated at the London mill are massive complex sulfides, in which pyrrhotite and pyrite are the predominant minerals. Other sulfides are chalcopyrite and sphalerite; magnetite is also of economic value. The approximate mineralogical analysis of the ore is shown in Table 1. [ ] The chalcopyrite and sphalerite are closely associated with pyrrhotite as shown by the tracings of micrographs (Fig. I), which are representative of the mill feed. Fig. la illustrates the type of locking between pyrrhotite and chalcopyrite; Fig. lb illustrates the way in which sphalerite is attached. Thesephotographs indicate that it is not practical to break all of these attachments. Hence, in order to produce concentrates with satisfactory grades and recoveries, most of the locked chalcopyrite particles must be recovered into the copper concentrate, and the locked sphalerite-pyrrhotite particles must be separated from the iron concentrate. A copper concentrate, a zinc concentrate, and an iron concentrate are made. Typical metallurgical results are tabulated in Table 2. HISTORICAL In 1922 the London mill started as a copper concentrator, with a capacity of 300 tpd, and in 1926 an additional flotation circuit was added to recover the iron sulfide remaining in the copper tailing. In 1927 a zinc flotation circuit was installed between the copper and the iron flotation circuits. This procedure was not satisfactory because the iron recovery was low, due to the

Jan 1, 1962

By C. W. Lawr

THE ores mined by the Santa Gertrudis Co. at Pachuca, Mexico, are mainly silver-bearing; they also yield some gold and carry a little copper. Strong cyanide solutions are used to dissolve the silver and gold, and the mechanical loss of cyanide at the tailings filter has been considerable. Prior to the year 1923, the pulp, after agitation, was given a 35 to 37-min. wash with barren solution in a Butters-type leaf filter; the discharged slime cake was repulped and pumped into eight Merrill self-sluicing slime presses in which the slime was given a further wash with barren solution for 25 to 35 min., followed by a 10-min. water wash. At that time the barren solution titrated around 0.350 per cent. free KCN and 0.500 per cent. total KCN. (Aero brand cyanide is used.) It can readily be seen that a 10-min. water wash would leave considerable cyanide in the filter cake, which of course was a total loss. This averaged 0.522 kg. (1.15 lb.) KCN per ton of solution. As the daily discharge to. the dam was 1600 tons1 of slime of a 2: 1 ratio of liquid to solid, the value of this mechanical loss of cyanide was an important item and its recovery held alluring possibilities.

Jan 1, 1929