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By J. Lee, R. G. Rosehart

"The basic problem of arsenic contamination from mine water is discussed. Effective methods of removal involving chemical precipitation, activated carbon adsorption, ion exchange and reverse osmosis are discussed. As well as an economic analysis, detail is also paid to other ions in solution which are removed simultaneously.INTRODUCTIONThe object of this paper is to review arsenic and its removal from gold extraction plant effluents. The first method considered is chemical precipitation. A second method is sorption processes, such as ion exchange resins and activated carbon adsorption. A third method is the possible application of reverse osmosis for the treatment of mine waste waters containing arsenic.Arsenic pollution in the gold mining communities of Canada and Ghana have been extensively studies (l, 2). Lynch(3) Shabunin, Gutman and Kogan( 4), Hollo, Troth, Vennes and Illinigi (5) and Hau(6) have reported their removal methods of arsenic by chemical precipitation, ion exchange and biological methods. Their conclusions were that the precipitation with Fe2+ and Fe3+ ions was most effective.Scope of the ProblemInitial aspects of this research program involved a detailed analysis of the operation of a gold extraction plant located in Canada. The name and location of the mine is proprietary information. Figure 1 gives a basic process schematic with the main tailing streams identified.Table 1 gives information as to the various mine streams and their physical and chemical composition. In this study, primary attention is paid to the removal of dissolved arsenic from waste waters. Solids from this type of operation can always be back-filled. As could be anticipated, the dissolved As content of the primary flotation tailing stream is very low. The dissolved As level in the secondary processing stream and the roasting quench water is quite high. The reason being that at this stage of the process, the arsenic has been freed from the arsenopyrite forms."

Jan 1, 1972

By Norman O. Lotter, André R. Laplante

"The pursuit of concentrator surveying for purposes of metal accounting or process optimization has not yet connected with the well-published realm of geostatistics. The reality of lognormality in time has seldom been discussed. One publication by Krige, 1962, examined such effects for a gold concentrator in South Africa, towards an information system that would provide feedback to the mining and surveying teams on grade control of ore mined.During present studies which are designing and testing a new survey model, called statistical benchmark surveying, it was found that the distributions of feed, concentrate and tailings bear significant residual signs of the parent lognormality in space. The multiple hosting of paymetals in different mineral species complicates the understanding of these distributions. These features have to be recognized and treated in order to perform appropriate quality control on the diagnostic sampling performed in the pursuit of statistical benchmark surveying.It is shown in this paper that the above two features can actually be used to advantage in the diagnosis of continuous periods of operations data ahead of such surveying, including the estimation of recoverable paymetals in flotation tailings from a graphical method.INTRODUCTIONAssessment of Concentrator PerformanceMilling and flotation, sometimes accompanied by gravity separation, are well-established unit processes, which have traditionally been used to treat ore that has been mined. This is because these processes liberate, then separate, the valuable minerals, which contain paymetals, from host gangue. This is often done by a combination of crushing and grinding processes, or autogenous grinding, to liberate the paymetal minerals, followed by gravity separation and/or flotation, to separate these valuable minerals into a concentrate, or several concentrates. These concentrate(s) have a lower bulk than the parent ore from which they are derived. Further, they contain a high percentage of the valuable minerals. Accordingly, they are more economically treated in subsequent beneficiation processes than the ore from which they were produced."

Jan 1, 2005

By L. Lui, A. Cauchon, A. R. Laplante

"The gravity circuit at Les Mines Camchib Inc. consists of two pinched sluices to bleed a feed to one of two 76 cm (30"") Knelson concentrator. The Knelson concentrate is upgraded to smeltable grade with a smaller Knelson (18/30 cm) and a riffleless table. Their tails are column cyanided and washed prior to recycle to the grinding circuit. This paper details size-by-size circuit performance. The Knelsons recover around 90% of free gold (as determined by Mozley Laboratory Separator), even below 37 µm. The sluices are ineffective, with upgrading ratios around 2, but approaching unity in the fine sizes. Much of the effectiveness of the circuit lies in the high gold circulating loads in the grinding circuit, which partly atone for the poor sluice performance, and the ability of the Knelsons to recover free gold. Secondary upgrading recovers slightly above 99% of the primary concentrate gold, 90% of it in a 65% Au concentrate and the balance in a heavily loaded (10 to 20 oz/st) pregnant solution. INTRODUCTIONBackgroundThe operating performance of the mill of Les Mines Camchib Inc. has been described earlier (Laplante et al., 1989). The presence of significant chalcopyrite (1 % and up) makes flotation necessary prior to cyanidation, to limit cyanide consumption and minimize environmental impact. In 1982, the presence of very high gold circulating loads in the grinding circuit lead to the commissioning of a gravity circuit, schematically shown in Figure 1. Gravity recovery of gold has been found beneficial because it increases metallurgical recovery by 1-3% (Laplante et al, 1987), as only part of the flotation tail is cyanided, and economic recovery, because of the lower net smelter return of gold in the flotation concentrate (as compared with the gravity concentrate). The latter increase is estimated, with a net smelter return of 95%, at 1.5% when gold recovery by gravity is equal to 30%. In early 1988, a sampling program was initiated to characterize the behaviour of gold in the grinding/gravity circuit. This paper reports on the results of this study."

Jan 1, 1990

By G. H. Wigton

THE oxidized ores of the Chief Consolidated Mining Co., in the Tintic mining district, have never yielded to metallurgical treatment by any standard method except smelting. These ores occur in comparatively small, irregular, siliceous replacements in limestone, and carry silver, lead, and gold as the principal valuable metals. But smelting methods. are unsatisfactory, as the lead smelters heavily penalize the large amount of silica in the ore, and the copper smelters do not pay for the lead in the ore; therefore, an alternative treatment method was sought. In 1916, a metallurgical research department was organized under; the direction of the writer. Many standard metallurgical methods were investigated but, on oxidized ore, none yielded extraction great enough to make its use feasible. During some chloridizing-roasting experiments in 1916, made in an assay muffle, the writer noted that large proportions of the silver, lead, and gold were volatilized from certain oxidized ores. The possibility of volatilizing the metals and collecting them in the form of a concentrated fume was then studied. But, as it was shown that satisfactory chloride-volatilization would be limited to a furnace feed with a high fusing temperature of the non-volatilizable constituents, and as the results obtained in a rotary kiln with the admixture of common salt with the ground ore were inferior to muffle tests, further volatilization experiments were discontinued until 1920. During this interval, an experimental sulfidizing and flotation plant was operated. After the cessation of attempts to float the oxidized minerals, chloride-volatilization experiments were resumed. The highly infusible flotation. tailing from the operation of the flotation mill was used for furnace feed. The first tests made on flotation tailing mixed with 10 per cent. common salt showed -nearly complete lead and gold volatilization but considerably less volatilization of silver than in muffle tests. In attempts to increase the silver extraction, the percentage salt in the furnace feed was varied, and it soon became apparent that the use of smaller percentages of salt raised the fusion temperature of the furnace charge and thus,

Jan 8, 1925

Possible consequences of continuing internal political unrest and the imposition of foreign sanctions raise concerns about the traditional reliability of the Republic of South Africa as a major world supplier of important minerals. A recent Bureau of Mines' technical report (Open File Report 76-86) examined the impacts of temporary peacetime disruption scenarios for six key materials and mitigating options. Because of South Africa's dominant producer position for these minerals (except for cobalt), expansion of primary production capacity outside of South Africa is not a short term mitigative option. However, prospects of a longer term disruption would likely spur investment in capacity expansion in alternative Market Economy Country supply sources. This report identifies the resources, production costs, and burdens that would be placed on alternative MEC producers of chromium, cobalt, gold, manganese, and platinum-group minerals if South Africa's production were subject to a long term; disruption. When long term supplies from South African deposits are discounted, more than a few years reserves were found only for cobalt and manganese. Even for these two minerals, higher sustained prices would be required to increase manganese production capacity and to support means of cobalt transportation not through South Africa.

Jan 1, 1986

By R. W. Nice

"The Similkameen Mining Company property is situated about 9 1/2 miles south and west of Princeton B.C. Owned and operated by Newmont Mining Corporation, mining is presently taking place in the Ingerbelle pit. At 15,000 tpd of 0.50% Cu ore the Ingerbelle pit is sufficient for about 8 years. Another 8 years ore supply is available from the old Granby workings of the Copper Mountain property which operated from 1927 to 1957.The pit operations commenced in late 1970 with the concentrator starting in March 1972. Designed and built by Bechtel Corporation under the supervision of Newmont the concentrator's design was based on single-stage crushing and single-stage modified autogenous grinding. The modification was installation of a crusher to crush the critical size build-up. The description of the operation follows.CRUSHERThe primary crusher is an open pocket Allis-Chalmers. 54"" x 74"" gyratory. The basic closed side setting is 8"". The crusher is fed from one side with 100 ton Electra-haul trucks. The crusher pocket discharges to a 96"" wide conveyor belt, then to a 48"" conveyor system including a stacker belt. The crushed product is piled on a stockpile with a capacity of about 30,000 live tons. This stockpile feeds the concentrator."

Jan 1, 1975

The Kaiser Aluminum and Chemical Corp. announced plans to raise June ingot prices by 4 cents to 57 cents per pound. Producers? April value for 99.5 percent unalloyed aluminum ingot was at 53 cents per pound. Domestic merchants last month were quoting the metal at 48-49 cents per pound. Primary aluminum producers began restarting production facilities following the settlement of the coal strike. Noranda Aluminum Inc. recalled workers to prepare potlines for production at its plant at New Madrid, Mo., which had closed February 14, 1978 due to the shortage of coal. Eastalco Aluminum Co. resumed full capacity production at its Frederick, Md. smelter where production had been cut by 30 percent. Anaconda Aluminum Co. resumed full capacity production at its Sebree, Ky. smelter.

Jan 1, 1978

By F. du Breuil

BCR National Laboratory (formerly Bituminous Coal Research, Inc.), under a USBM con­tract (originally with the Department of Energy), monitored the installation and operation of a computer-based belt-monitoring and control system at the Barnes & Tucker Coal Co. Mine No. 20, Ebensburg, PA. The system installed was the National Coal Board's Mine Operating System (MINOS) coal transport monitoring and control system. The program objec­tive was to document (1) the installation requirements for a conveyor control system, (2) the operating experience with the system, and (3) the impact of monitoring and control on mine operation. Analysis of the system operation indicates that monitoring and control did improve belt availability significantly, and reduced downtime associated with belt delays. Economic analysis showed the increased belt availability would potentially result in significant production increase, and, therefore, excellent payback on the initial investment. The system performed satisfactorily, was well accepted by the work force, and did not increase the maintenance workload.

Jan 1, 1987

By Concentrator Staff

"IntroductionA brief literature search will reveal that the information available concerning rod mills and ball mills is quite extensive. This type of equipment is used to treat every element from antimony to zinc. The size and capacity of individual mills, the number of units employed, and the configuration of the equipment within a concentrator add another dimension to the possible scope of this discussion. It would be impossible to treat this volume of material with any degree of thoroughness.It is the intention in this paper to describe the simplest and generally the most typical circuit wherein a single rod mill and ball mill operate in parallel. Some of the subject matter will not be entirely applicable to every installation. However the principles should be generally relevant to most applications.Typical Circuit DescriptionFigure 1 illustrates the arrangement of the equipment associated with a rod-ball mill circuit.Ore is extracted from a fine ore storage bin (1) the capacity of which is large enough to allow .continuous mill operation independently of the mine and crushing plant production schedules. A number of feeders (2) may be employed to control the flow of ore to one or more feeder conveyor belts (3). In larger tonnage operations several such conveyors are employed at least one of which will be equipped with a variable speed control to allow adjusting of the amount of feed reporting to the rod mill. A rod mill feed conveyor (4) accepts all of the ore from these feeder belts for ultimate transfer to the rod mill. This conveyor is normally equipped with a weightometer (5) to monitor the flow rate of ore to the rod mill and ultimately to allow control of this flow. Typically these conveyor belts would be equipped with feed deficiency switches which would give a remote alarm of the fact that no ore is present on the conveyor. The conveyor belts are protected by underspeed monitors which will provide a remote alarm and automatic shutdown of the conveyor in the C'.'~nt that belt slippage occurs. Finally the belts are protected by monitors which will alarm and shut down a belt if it is not running straight in the idlers."

Jan 1, 1989

By G F. Brent, R D. Hart, C Carr, F Felipe

In underground operations, where copper, gold and other metals are mined, the open stopes can be backfilled to stabilise surrounding workings. This fill may consist of tailings from the processing plant, usually mixed in slurry with a binder such as ordinary Portland cement (OPC). Extensive use is made of tailings in OPC-based paste fill, cemented hydraulic fill or cemented aggregate fill applications. A large underground operation such as the Mount Isa Copper Operations utilises over three million tonnes of backfill per annum. Geopolymers offer an alternative binder system with significant environmental advantages over OPC. Geopolymers are formed by polymerisation of aluminosilicate precursors in an alkali medium. They generally form products with superior strength characteristics as well as increased resistance to degradation, particularly under acidic conditions. Geopolymers have lower carbon dioxide emissions associated with their manufacture than OPC and offer further potential advantages, such as the use of less water and more waste materials. The application of geopolymers to weak material systems such as mine backfill has not been widely studied. The aims of this investigation were to assess a range of Australian mine process wastes for suitability as feedstocks in fill applications and to construct a range of these cemented minefills using a geopolymer binder. The physicochemical characteristics of each of the wastes are discussed as well as how they ultimately contribute to the geopolymer cement through direct replacement of OPC. Results from various tests on the candidate geopolymer fill formulations are presented and the potential of such fill formulations to replace OPC in mine backfilling operations is highlighted. The significant potential contribution to reducing greenhouse gas emissions from the use of geopolymer-based mine backfill across the industry is noted. Conversion of a single mine, such as the Mount Isa Copper Operations, is shown to match the reductions in greenhouse gas emissions due to the current total installed photovoltaic electricity generation capacity of Australia. A similar examination of wastes available to a gold mine could potentially yield new binders for use in established or new fill systems.

Jan 1, 2007

By E. Ozberk

The copper and nickel smelting industries are important contributors to the technical and economic well-being of Canada while some of the smelters are among the largest sources of sulphur dioxide emissions in Canada. A review of the operations in six smelters - Falconbridge Limited, Falconbridge, Ontario; Hudson Bay Mining and Smelting Co. Limited, Flin Flon, Manitoba; Inco Limited, Copper Cliff, Ontario and Thompson, Manitoba; Noranda Mines Limited, Horne Division, Noranda, Quebec and Gaspe Mines Division, Murdochville, Quebec - have shown that about 80% of the total sulphur input to these smelters was emitted to the atmosphere in 1980. A simultaneous review of the existing and emerging technologies which can directly or indirectly reduce the sulphur dioxide emissions indicated that it is possible to reduce the emissions to below 20% of the sulphur input to these smelters. This can be achieved by varying degrees of modifications to the smelting operations and installing sulphur fixation facilities at significant costs which may increase the metal production costs. This would consequently make the competitiveness of the Canadian copper and nickel producers more difficult in the already tough world metals market.

Jan 1, 1983

By Ben C. McLellan, Philip J. Bangerter, Andrew Murphy, Glen D. Corder, Damien Giurco

Sustainability now forms a critical part of metallurgical plant design. In addition to compliance with requirements and corporate principles, design engineers must also consider how community and societal responses to the metallurgical industry affect the way they design and build plants. Three broad areas to consider are: the design engineer’s contribution to sustainability through the design process and the relationship to the social license to operate (SLO)—the unwritten license communities give companies to safely and responsibly operate the plant and equitably share benefits; the realization of sustainability goals through methodologies emerging in the design activities; and the integration of design frameworks into a contemporary project design context and its investment objectives. This chapter explores sustainability in metallurgical plant design in the context of SLO, design tools and frameworks, and community and societal influences.

Jan 1, 2015

By Willem P. C. Duyvesteyn

"Scandium International Mining Corp (SCY) has been involved in the commercial development of its Nyngan Scandium Project for about ten years. Various studies at a variety of global R&D organizations have been carried out in an attempt to define the deportment of scandium in both the limonite as well as the saprolite phases of Nyngan ore to assist with process development. Technologies such as sulfuric acid baking, high pressure acid leaching, sulfuric and hydrochloric acid atmospheric leaching and heap leaching have been extensively investigated both at the bench and pilot scale level to arrive at the most economic process flow sheet that will be discussed in some detail in this paper. Solvent extraction, ion exchange and other refining technologies have been tested and an optimum configuration has been established for the Nyngan project which is expected to be the first primary scandium mine and metallurgical plant in the world.Because of a lack of product availability the current scandium market is limited to less than 20 tons per year. Technologies have been identified to produce product with different specifications for aluminum alloys and solid oxide fuel cell applications. Preliminary economics have identified that Nyngan will be a robust and low capital cost project that can readily be expanded as the size of the scandium market increases."

Jan 1, 2016

By R. D. Wunder

The Chino Mines Company, a Kennecott-Mitsubishi partnership, located in Hurley, New Mexico, recently completed a $405 million modernization of its mine, concentrator, and smelter facilities. The mine/ concentrator modernization commenced in 1980 and the smelter modernization began during 1982. Modernization of the mine and concentrator facilities involved increasing the mining rate from 98 kt/d (108,000 stpd) to 152.2 kt/d (167,800 stpd). Ore mining was designed to increase production from 20.2 kt/d (22,300 stpd) to 34.0 kt/d (37,500 stpd). A new concentrator was constructed near the mine using semi-autogenous grinding. The new smelter uses the INCO flash smelting process, increasing capacity by 40.8 kt/a (45,000 stpy) to 99.8 kt/a (110,000 stpy) of salable copper.

Jan 1, 1985

By T. Grewar

"South Africa is a water-scarce country, rated as one of the 30 driest countries in the world. Worryingly, government expects water demand to outstrip supply as early as 2025. The South African government believes that substantial volumes of water can be made available through the reuse of treated mining-impacted water. Currently, however, large volumes of treated mining-impacted water are produced, often with no end use in mind other than discharge or disposal, when these treated effluents may have further uses in applications which are currently and unnecessarily using high-quality water. Treating the wastewaters to levels that meet ’fitness-for-use’ guidelines for alternative applications such as agriculture, sanitation, or industrial processes will reduce the treatment cost burden compared to potable water production. The primary conclusion of this review is that the most suitable option for re-using treated mine-impacted water is in agriculture, for irrigation of crops (food, forage, or energy crops), which currently and unnecessarily makes use of the majority of South Africa’s high-quality resources or potable water. Other findings include the realization that although guidelines and legislation for water re-use in South Africa exist and are readily available, they tend to be contradictory and confusing in many cases. It is imperative that these ambiguities and incongruities in the available guidelines and legislation for water re-use be clarified and updated swiftly IntroductionSouth Africa is a water-scarce country and is currently rated as one of the 30 driest countries in the world, with an average rainfall of 490 mm/a, approximately half of the global average. As an indicator of the degree of regional variability in South Africa’s water supply, it has been shown that 70% of all runoff is from approximately 20% of the land area. Regardless of the water scarcity faced in South Africa, our water conservation track record is poor, with an average consumption of 280 L/d per person, almost 60% more than the global average of 175 L/d per person. Around 40% of this allocation is utilized in watering lawns and gardens (Zhuwakinyu, 2017).The South African government predicts water demand to outstrip supply as early as 2025. On an international scale, the situation appears just as dire, with the United Nations High Level Panel on Water (HLPW) expecting a 40% water shortfall by 2030, which may affect up to 1.8 billion people based on current water demand trajectories (Zhuwakinyu, 2017)."

Mar 1, 2019

By Margaret R. Lickes

Mercury was produced originally at the Idaho Almaden mine from 1939 to 1942. When the ;.nine was shut down, more than 100,000 tons of ore, containing from 2.5 to 3 pounds of mercury per ton, was known to be adjacent to old workings but had been considered too low grade for commercial operation. Later, detailed, systematic exploration by Rare Metals Corporation of America increased and delimited enough minable reserves for further operation to be foreseen for several years. The average grade of the ore is estimated at slightly more than 3.5 pounds of mercury per ton. Open-pit mining and installation of the largest rotary kiln in the United States furnacing mercury ore have made possible profitable exploitation of these reserves (fig. 1). The bulk of the exploration was by percussion drilling, although some diamond drilling was done to check the results. When enough reserves were indicated to assure a successful operation, stripping and plant construction were started simultaneously. Figure 2 is a general view of the plant. The ore bodies, although irregular in size, shape, and grade and covered by essentially barren cap rock, are amenable to low-cost, trackless, open-pit mining. Development consists of stripping the overburden and removing, barren cap rock. Ore is being mined at approximately [5],000 tons per month and treated in a 175-ton-per-day furnacing plant.

Jan 1, 1957

By Maurice A. Roche

Introduction Most of the Flin Flon ore-body, like many others in northern Canada, was covered by water. The original discovery was made where the ore-body outcropped on a strip of land jutting out into Flin Flon lake. The outcrop could be traced for a distance of 600 feet across the strip of land, but disappeared under the lake on both sides of the peninsula. Following the discovery or location of the property, in 1915, and subsequent to its examination by the H.P. Whitney interests in 1925, an extensive diamond-drilling programme was completed, which proved the existence of an ore-body at least 2,600 feet long and 900 feet deep. In plan, the central part of the ore-body at surface attains an overall width of about 450 feet; including waste horses, and gradually becoming narrower to the south. To the north, it stretches out as a much narrower tongue. The main body is a solid sulphide ore containing gold, silver, copper, and zinc. Adjacent to, and on the footwall side of, the solid sulphide ore-body, there are considerable widths of talcy disseminated copper ore, particularly in the upper 300 feet of the deposit, where the maximum widths occur. The ore-body occurs in a greenstone area, the rock being quite hard on the hanging-wall side, with the footwall schisted, talcy, and much softer. The greater part of the ore-body was overlain by 12 to 15 feet of water and from 15 to 90 feet of mud and clay, this material requiring eventual removal, regardless of the system of mining adopted. The total of mud and clay overburden was estimated at 1,000,000 tons. After considerable study, it was decided to mine the upper 300 feet of the ore-body by open-pit methods; the balance of the ore to be recovered by underground mining . The present paper gives an account of the methods adopted in both operations, and of the problems encountered.

Jan 1, 1935

By Corby G. Anderson

Since 1985, the Sunshine Mining Company has used a patented nitric-sulfuric acid pressure leaching system for production of silver and copper from a refractory sulfide ore body. To better understand this novel process, a program of lab scale testing was undertaken to quantify the effect of system variables on silver recovery. Some of the factors studied include temperature, agitation, acidity, vessel pressure, particle size, percent solids and oxidizer amount. In turn, this knowledge will be applied in actual plant practice to upgrade reaction time, increase concentrate throughput, raise silver recovery and lower reagent costs. A discussion of plant process changes resulting from lab scale testing is included.

Jan 1, 1991

By R. L. Rock

The Bureau of Mines conducted radiation-ventilation studies in six large underground uranium mines to investigate the extent of radon-daughter exposure experienced by the miners and to analyze the ventilation systems to set what changes were required to achieve recommended radiation standards. The studies entailed tracing and measuring the mines' ventilating air, quantitatively and qualitaitively, from its entrance into the mines, through production areas, and lack to the surface. Pressures consumed in creating airflow through mine circuits were determined by altimeter surveys. Air quantities were measured with anemometers or by chemical smoke cloud methods. Air quantities delivered through tubing were determined using U-tube water gages to magnahelic gages to measure velocity pressures. Radon-daughters were filtered from mine air to measure alpha activity; activity levels were converted to "working levels," the established unit for reporting miner exposure. Attention was given to locating unnecessary contamination influx sources and recirculating air. The radiation-ventilation surveys showed that where basic principles of food ventilation were not followed, high alpha radiation levels invariably resulted. Some mines required more primary air for controlling radiation levels throughout active areas, but more efficient use of available air was generally possible. Premature contamination of intake air was prevalent where development Openings intersected ore or passed by worked-out stopes. Several auxiliary systems were inefficient in secondary distribution, recirculating air already too highly contaminated for ventilation. Many got these common problems can be traced try the lack of dine planning with radiation control in mind.

Jan 1, 1969

By W. E. Davies

"This paper is the sequel to that on Developments in Sintering Efficiency delivered by Davies and Mitchell at the International Ore Dressing Congress, Stockholm, 1957. Attention is drawn to the fundamental concept of sintering as a heat transfer process, in which the ignition heat is continually transferred onward through the bed as a heat wave. By its passage, this heat initiates endothermic 'changes, such as carbonate decomposition, whose energy requirements are supplied by the combustion of the fuel contained in the bed.Ideally, the heat wave should be generated by the ignition hood followed by the pre-heating of the incoming air. The fuel content of the bed should be no more than the theoretical quantity needed to balance the endothermic changes. In practice, the heat wave is seriously short of energy and this is offset by excess fuel in the bed. As a result, sinter quality suffers, more return .fines are produced, and fuel is wasted, a loss of efficiency which could be avoided if the fundamentals of correct ignition were more widely appreciated.Sintering speed, otherwise heat wave velocity, is a function of the rate at which the incoming air acquires .heat from the cooling sinter and imparts it to the unburned feed below the combustion zone. It is thus a function of rate of air flow and so depends on bed permeability>. Ways of• improving and maintaining permeability at the maximum consistent with thorough sintering are discussed, particular reference being made to taconite concentrates. The possibilities of recrystallizing refractory ores into forms more amenable to physical concentration techniques by sintering are visualised, very little expenditure on heat energy being required when the process is operated near the ideal conditions."

Jan 1, 1960