Search Documents

By H. V. Makar

The Bureau of Mines successfully demonstrated, on a pilot plant scale, a sodium sulfate (Na2S04) refining process for removal of copper from molten ferrous scrap. Previous tests established technical feasibility on a bench scale for achieving low copper concentrations and determining the principal factors controlling process efficiency. The pilot plant study involved a 1-ton arc furnace for melting ferrous scrap and a 1-1/2-ton treatment ladle. Results of these tests showed that copper removals obtained in laboratory tests (approximately 40 relative percent copper removal with 10 weight-percent surface addition of sodium sulfate) could be achieved on the pilot plant scale when excess carbon was added with the sodium sulfate compacts. Data were obtained on temperature losses and carbon consumption related to Na2SO4 treatment.

Jan 1, 1974

By Victor Velho de Castro

O presente trabalho tem por objetivo avaliar a resistência ao desgaste de dois diferentes materiais (três amostras de aço SAE 52100 e três amostras de SAE 1045) em diferentes condições microestruturais, sem lubrificação e com lubrificação (utilizando biodiesel B100, óleo diesel/biodiesel B50 e óleo diesel comercial B7). Quatro amostras foram submetidas a tratamento térmico de têmpera (830°C resfriado em óleo nas amostras de aço SAE 52100 e 850°C resfriado em água nas amostras de SAE 1045) seguida de revenimento (100°C e 200°C nas amostras de aço SAE 52100 e 100°C e 300°C nas amostras de SAE 1045). Os ensaios foram realizados em um tribômetro do tipo pinosobre- disco, com rotação de 300 RPM, 1,5kg de carga e 4400m de distância percorrida, seguindo as recomendações da norma ASTM G99–95, medição da microdureza superficial da pista de desgaste e avaliação da superfície desgastada das amostras com um Microscópio Eletrônico de Varredura (MEV). Ao final dos ensaios, observou-se a variação da resistência ao desgaste entre os diferentes materiais, diferentes microestruturas e diferentes lubrificantes aplicados no ensaio. Os resultados indicaram uma tendência de diminuição de desgaste nos ensaios realizados com lubrificação de biodiesel B100.

Aug 16, 2017

By R. S. Shoemaker, F. W. McQuiston

The principle of cyanide recovery or regeneration of cyanide from solution, be it a weak wash or a foul solution, is by acidulation. All or part of the cyanogen is converted thereby into hydrogen cyanide which is fixed by an alkali (generally lime) and returned to the cyaniding system (the Mills- Crowe process). The solution is made acid by bringing it into contact with sulfur dioxide. The acidified solution is stripped in a packed tower by a counter flow stream of air in which the air and the solution are brought into intimate contact. The air leaving the tank charged with hydrogen cyanide is then passed to another tank in which it is mixed with an alkaline solution, the latter absorbing the HCN and leaving the accompanying air clean for reuse in removing more HCN from the acidified solutions. Of the recoverable cyanide, 9770 is usually recovered in a three-unit disperser tower system. When there is a considerable content of copper, the solution is made acid following standard procedure; however, the HCN is not removed until the copper is reduced by regeneration. Copper removal at this point may be as high as 90%. After reduction of the copper, the acidified solution then is contacted with a lime solution which brings down calcium sulfate as a second step. The silver, gold, and copper are recovered in the form of a precipitate which can be shipped to a smelter. At the Hudson Bay Mining and Smelting Company, Ltd., most of the precious metals have been recovered in the copper and lead concentrates, but it has been found profitable to submit the flotation tailings to a scavenger cyanide treatment for the recovery of exposed gold contained in pyrite rejected during flotation. Due to the fine particle size, maximum dissolution is achieved in a four-hour contact period. In the cyanidation plant, the thickened flotation tailing at 70% solids is given a four-hour agitation in weak calcium cyanide solution and filtered. Filter cake is repulped and discharged either to the tailing storage area or to the mine as backfill. Clarification, de-aeration and precipitation of pregnant solutions are carried out in a standard Merrill-Crowe unit. Typical analysis of precipitate is as follows: Au, 0.47%, Ag, 4. 88%; Cu, 46.0%; Zn, 13. 8%. Cyanicides in the form of copper minerals limit the reuse of barren solution. To recover the substantial quantities of cyanogen chemically combined with base metals in this solution, a cyanide regeneration plant is in operation. The barren solution is acidified with sulfuric acid in a spray tower where an upward current of air from a blower sweeps the liberated hydrocyanic acid gas to an adsorption unit where it meets a fine rain of lime suspended in plant solution with which it combines to form calcium cyanide. The lime--water solution carrying calcium cyanide is reused in the cyanide circuit. The return gases are drawn into the blower and back to the disperser tower thus

Jan 1, 1975

By Anders Erlandsson

"The crude ores from the LKAB mines are sorted and refined into a number of products. As there are a number of different plants the flow of material is often complex.in connection with studies of a proposed extensive investment program it was necessary to get a good measure of the changes which would be caused in the total profit by the different possible investments. A linear programming model was set up to accomplish this, and to cover the consequential effects on the production system. The objective function of the model maximized the total profit, taking into account restrictions in materials balances, blending conditions, production capacities and market. The changes in total profit resulting from the different investment alternatives, together with investment and fixed costs, were used in a series of investment analyses. The application has proved to be very successful and the results come very near to what was expected."

Jan 1, 2014

By D Hart

In 1944 some ten or so pieces of legislation impacted on the mining industry. Today in excess of 80 in one way or another apply to the Northern Territory miner. This paper will look at the nature of the legislation that has been introduced and the reasons for its introduction. During the Second World War the mining of resources such as iron, tin etc for the war effort set the platform for the subsequent expansion of the industry in the post-war years. Thereafter, the change in attitudes over the years that have resulted from a more conscious public and more responsible governments and developers associated with the recognition of the legal rights of others and the threat of litigation, will be traced. With a major review of the Northern Territory Mining Act currently in progress, this paper will view ANZMEC's foreshadowed preference for a National Code and the NT's intent of drawing some commonalities with other states of Australia to achieve this aim. Finally, the questions of over-regulation and whether existing legislation is complete will be addressed. From this it is foreshadowed how legislation may develop during the lead into the 21st century.

Jan 1, 1994

By Felix Wolters, Michael Schütt

In over 40 years of operation, engineers at Aluminium Oxid Stade GmbH (AOS) have acquired a lot of experience whereby they have developed a series of optimizations regarding the specific primary energy consumption. The tube digester technology, invented by VAW in Germany and applied on an industrial scale for the first time in Stade, was brought to perfection by its innovative cleaning methods, which led to a constantly high heat transfer, low pressure loss and excellent availability of the digesters. Further improvements such as the heat recovery from the exhaust gas stream of the salt heater as well as the optimization of one of the fluid bed calciners and the installation of a combined heat and power generator (CHP) led AOS to where it is now—a leading alumina refinery with regard to the lowest energy consumption per ton of alumina.

Mar 1, 2017

By M. Mikhaylenko

The most common methods of processing of uranium ores is sulfuric or carbonate leaching with conversion of hexavalent uranium to uranyl carbonate or sulfate anionic complexes followed by sorption of the later by strong base anion (SBA) exchange resins. It is known that chloride anion has high affinity for SBA resins so that it depresses sorption of anionic complexes of uranyl. Application of the SBA resins becomes impractical for recovery of uranium when chloride concentration in pregnant leach solution exceeds 3-5 g/l [1-3].

Jan 1, 2012

By Leonard P. Vitt

This report details the work done to provide the U.S. Bureau of Mines with a customized Wireless Ground Monitor Wireless Ground Monitor was modified as follows: 1. A voltage regulator was added to prevent false trips with line voltage variations. 2. A Ground Wire Trap was designed to eliminate problems with parallel paths and provide spark suppression. 3. A three phase filter was designed to reduce 60 Hz leakage current. 4. A portable tester was designed to permit complete test of installed systems. 5. The design was modified to insure trip with an open ground wire in the event of a failure of any one semi-conductor component.

Jan 1, 1976

By R. V. Ramani

This report is volume 1 of an eleven-volume final report for a Grant completed by The Pennsylvania State University for the Bureau of Mines to develop a dynamic general purpose computer simulation model for underground coal mining. The simulator developed under the Grant has a wide range of possible applications in the design and evaluation of underground coal mining systems from the standpoints of health and safety, productivity, and mine environment. The Master Design Simulator consists of nine free-standing function-oriented subsystem models linked together into an integrated, comprehensive model of an entire mining system. The subsystems comprising the Master Design Simulator include geology and reserves, methane generation, roof support, subsidence, production, rail haulage, cost, water generation, and ventilation. This volume of the final report contains a statement of the purpose and scope of work of the Grant, and provides an overview of the Master Design Simulator with a brief description of each subsystem. Also included is a discussion of several areas of application for the simulator.

Jan 1, 1975

By Chikao Nishiwaki

INTRODUCTION The Japanese dependency for mineral raw materials from foreign sources increased rapidly within the 20 years since the Korean war. During this 20 year period Japan has attempted to industrialize and develop her economy while depending mostly on heavy import of nearly all kinds of mineral raw materials. She has fabricated them using energy generated mostly by mineral fuel that are also imported. Fabricated goods are either used domestically to accelerate more fabrication industry or exported after adding more refinement or fabrication. In this paper iron ore as a major raw material for the Iron and Steel industry, three major base metals, copper, zinc and lead, and petroleum as a major energy source have been selected as examples to represent the general features of various kinds of mineral raw materials that are imported. Among other important mineral commodities for which Japan dependa upon foreign sources are: Aluminium, nickel and cobalt, potash and phosphorous 100 percent dependent and tin, tungusten, chromium and molybdenum over 95 percent dependent. IRON ORE Statistics for the important iron and steel raw materials, namely iron ore, coal and scrap iron are shown in Tables 1, 2 and Fig. 1. Iron ore is by far the most important and basic among these three commodities. There has been practically no steel or iron imported as a metal into Japan during this period. Iron ore, including aggregate such as pellets, and metallurgical coal are mostly transported in large ore cargo boats to the port where very efficient port facilities unload them directly at the steel mill site. As a result

Jan 1, 1976

By Keith R. Bock, Norman Parlee, Albert M. Barloga

Coils of pure iron and iron-carbon alloy wire (0.05 to 0.80 pct C) and sufficient sulfur to saturate the solid phase were equilibrated in evacuated or argon filled tubes. After rapid cooling, and removal of the outside nonmetallic layer, the wires were analyzed for carbon and sulfur and the data used to construct an Fe-S binary and isotherms of the Fe-C-S ternary in the range 1149° to 1427°C. THE solid solubility of sulfur in steel is of interest in connection with such phenomena as hot shortness, burning, and so forth. "Burning," the more or less permanent damage that some steels suffer when heated for forging or rolling, has been shown to be related closely to the behavior of sulfur and less closely to carbon and oxygen.' Attempts to interpret burning phenomena in steels fail because of lack of data on the Fe-C-S and the more complex systems in this family. Rosenqvist and Dunicz 2 and Turkdogan, Ignatowicz, and pearson3 have largely elucidated the Fe-S diagram in the region of interest but no information on the Fe-C-S diagram in this region appears to be available in the literature. This paper deals with the elucidation of the Fe-C-S diagram in these interesting ranges. The method employed is different from those used by previous workers2'3 on the Fe-S system. EXPERIMENTAL METHOD Pure iron wires (Ferrovac E) or iron carbon alloy wires of 1 mm in diam were cleaned with acid and acetone, coiled, and placed in silica tubes (7 mm OD and 5 mm ID) previously closed at one end. Enough sulfur was added to assure saturation of the solid iron phase. The filled tubes were either simply evacuated and sealed, or filled with argon at a reduced pressure and sealed. The argon was required at the higher temperatures to prevent collapse of the tubes. The filled and sealed tubes were placed in the middle uniform temperature zone of a Globar tube furnace and equilibrated at temperatures ranging from 2100°F (1149°C) to 2720°F (1493°C). After equilibration the tubes were removed from the furnace and quenched in air or water, the form of quenching being found to have no effect on the results. The tubes were broken open and the coils were placed in a 1 : 1 HC1 solution to remove the sulfide-rich layer. The coils were then cut into small pieces and analyzed for sulfur and carbon. In the early stages of the investigation different equilibration times ranging up to 17 hr were tried and the cores of the wires were analyzed to test for saturation. One hour appeared to be sufficient to reach maximum sulfur content at 1454°C and 2 hr sufficient at 1149°C. The practice adopted was to use at least 3 hr at the higher temperatures and at least 5 hr at the lower temperatures. The iron-carbon alloy wires used were made by carburizing pure iron wire with carbon monoxide gas in a one inch diameter ceramic tube at about 1204°C. Differing carbon contents were obtained by allowing fairly large coils to react with the gas for varying lengths of time at a flow rate of 800 cc per min. The reacting times ranged from 15 min to 4 hr depending upon the amount of carbon desired. The furnaces used were controlled by means of Leeds and Northrup Speedomax Type H instruments with D.A.T. Control attachment. Thermocouples were calibrated against the melting points of gold and copper. The temperatures recorded appear to be accurate within i2.7OC. Starting with run No. 85 and continuing with the

Jan 1, 1962

By William E. Lamont

The Mineral Resources Institute (MRI) of the School of Mines and Energy Development (SOMED), University of Alabama conducted studies to determine the potential for upgrading the Eastern Devonian shales by flotation. Results of those studies indicate that flotation concentrates containing 150 liters of oil/t (36 gal/t) can be produced from shale containing 50 liters of oil/t (12 gal/t) as determined by Fischer Assay with good recovery of the kerogen. Organic carbon and trace elements such as molybdenum, vanadium and uranium also are concentrated with the kerogen. Kerogen concentration was found to be a function of the size of grind.

Jan 1, 1983

By Christopher Boys

"Abstract -Many water inflows in Saskatchewan potash mines are linked to salt anomalies. Potash ore and three anomalies in the PCS Cory potash mine in the Patience Lake Member of the Prairie Evaporite Formation were studied to understand the processes involved in their formation. Another goal was to find geological indicators of the proximity of anomalies. A geological model that relates the potash to the salt anomalies has been developed.Up to five post-burial facies may be recognized . These range from carbonate-mud collapse breccias to bedded sylvinite, and probably result from post-burial leaching. ln this model, the same fluids that caused widespread recrystallization and possibly enrichment of the potash salts, started by dis­ solving huge cavities in the evaporites . The effects of the leaching were highly variable, indicated by selectively preserved delicate laminae and chevron textures, to deformed and totally dissolved salt beds . ln at least two major events, Late Devonian and Cretaceous, collapse structures may have acted as conduits for overlying formation waters to cause pervasive leaching and recrystallization of the Prairie Evaporite Formation .Geological indicators of the proximity of an anomaly may include: a change in Insoluble seam color from greenish-gray to mottled brown and greenish-gray, unusual local increases in ore grade, large patches (e.g., >200 cm2) of sylvite-poor potash crosscutting units near the top of the ore zone, and drops in topography >10 m of the marker seams. Anhydrite was more common in an anomaly than in ore. The local abundance of anhydrite in ore may indicate the proximity of some salt anomalies.Once an anomaly has been encountered, indicators of a major collapse feature may include: stretched clay seams, folded beds, small collapse features (1 m to 20 m), numerous washouts, or split clay seams with injected salt. Any number, or none, of these indicators may be found close to a major collapse feature."

Jan 1, 1993

By M Rafty, C Lane

Managing water is a key aspect of mining, whether a site is in water surplus or water deficit. Water in tailings storage facilities (TSFs) needs to be managed carefully to maximise the density of the solids, maximise the re-use of available water and to minimise the risk of water overtopping the spillway in the event of heavy rainfall. Developing a water balance models is a standard method for monitoring the major flows of water around a site. However, models created by specialist consultants are often overly complex and are not suited to being used by process plant personnel after they have been handed over to the mine operator. Models typically offer information on water management at the time that they are created, but then ‘sit on the shelf’ and offer no further insight or value for operating mines. Advisian has developed a customised water balance model for a mine in Sub-Saharan Africa, which can be adapted to other locations, to assist in managing water at the site and in particular to predict the current and future water level and volume in the TSF under various rainfall scenarios. The model was developed using Microsoft Excel and requires no specialist skills to operate. Site personnel enter in daily production, water transfer and rainfall data and then view the output dashboards to observe the current water level and volume, decant rates, licensed water abstraction and water supply versus demand. The output dashboards were developed in close consultation with the client to ensure that all of the relevant key performance indicators were included, from the perspectives of all the key stakeholders. This modelling approach focusses on the needs of the people who will actually use the model to actively manage TSFs with the need to involve the model developer as part of ongoing operations eliminated. The simplicity of this approach is now being embraced by other mine sites in Africa and Australia. CITATION:Rafty, M and Lane, C, 2018. Water balance management 101, in Proceedings Mine Waste and Tailings Stewardship Conference 2018, pp 485–490 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 23, 2018

By Martin A. Mayers

It has long been recognized that it would be highly desirable to be able to predict the temperatures at various points in a burning fuel bed and their variations with changes of the properties of the fuel and the rate of air supply. Experimenta1 investigations of fuel-bed temperature1-4 have not been lacking, but the absence of any theoretical background has made it impossible to use the results of these tests as a basis for prediction, except in a very general may. A mathematical analysis", based on a somewhat simplified theory of the distribution of lieat in a fuel bed, lias recently been completed by the author. Even though numerical values of some of the coefficients are not known, the calculations result in some conclusions of general significance. Thus, for example, it appears that when the temperatures throughout the fuel bed can be calculated, the rate at which ignition of the fuel by underfeed action can be maintained can also be calcu1ated. The uncertainty concerning the values of the coefficients referred to above is serious limitation on the use of that method. The following paper presents a discussion of these coefficients and includes estimates of their values based on the data now available. A brief description of the theory on which the analysis is based, and some results of the calculations showing their degree of agreement with experimental results, are presented by way of introduction to the discussion of the coefficients. For a more. extended discussion of the theory, see the A.S.M.E. paper5. Theory and Comparison with Experiment The theoretical analysis is similar to that used in the calculation of the stationary temperatures in a conducting solid in the classical theory of heat concluction. The fuel bed is pictured as a uniform, continuous, porous solid through the interstices of which air can be blown and ill which combustion can take place. The solid lias a finite thermal conductivity, so that heat may be transported in any direction, against the flow of air as well as in the same direction. While it is true that nothing

Jan 1, 1938

By Martin A. Mayers

It has long been recognized that it would be highly desirable to be able to predict the temperatures at various points in a burning fuel bed and their variations with changes of the properties of the fuel and the rate of air supply. Experimenta1 investigations of fuel-bed temperature1-4 have not been lacking, but the absence of any theoretical background has made it impossible to use the results of these tests as a basis for prediction, except in a very general may. A mathematical analysis", based on a somewhat simplified theory of the distribution of lieat in a fuel bed, lias recently been completed by the author. Even though numerical values of some of the coefficients are not known, the calculations result in some conclusions of general significance. Thus, for example, it appears that when the temperatures throughout the fuel bed can be calculated, the rate at which ignition of the fuel by underfeed action can be maintained can also be calcu1ated. The uncertainty concerning the values of the coefficients referred to above is serious limitation on the use of that method. The following paper presents a discussion of these coefficients and includes estimates of their values based on the data now available. A brief description of the theory on which the analysis is based, and some results of the calculations showing their degree of agreement with experimental results, are presented by way of introduction to the discussion of the coefficients. For a more. extended discussion of the theory, see the A.S.M.E. paper5. Theory and Comparison with Experiment The theoretical analysis is similar to that used in the calculation of the stationary temperatures in a conducting solid in the classical theory of heat concluction. The fuel bed is pictured as a uniform, continuous, porous solid through the interstices of which air can be blown and ill which combustion can take place. The solid lias a finite thermal conductivity, so that heat may be transported in any direction, against the flow of air as well as in the same direction. While it is true that nothing

Jan 1, 1938

By Benjamin F. Tillson

Although portable fire extinguishers are valuable for fires in an incipient stage, some medium that will dissipate a large amount of heat is needed to fight a fire in mine timbers, because of the size to which the fire will probably have grown before it is discovered. Scaling off, or otherwise smothering such a fire with gases that will not support combustion, is often ineffective for the residual heat will cause reignition when the area is opened and fresh air reaches the hot charred timber. The most effective ways of fighting mine fires, therefore, are to drown and cool the hot zone with water, and to "load out" the partly burned material. Either of these methods requires a direct attack of water ; upon the inflamed material; therefore, every mine containing a large ' quantity of timber or other combustible should be as completely equipped with a fire-hydrant system as it is with compressed-air mains. Mine managers may object to the expenditure, especially if they have not suffered from mine fires. The cost of such preventative measures, however, is only a small portion of the possible losses, and the water pipes can be used for other purposes, which will pay the carrying charges. They can supply water to the rock drills, which will encourage more wet drilling with hollow-drill steel, for the trouble and expense of packing water to individual tanks at the drills will be removed. The carbon-tetrachloride hand extinguishers seem to have no place in mining except for fires that start in electrical equipment. They give off noxious gas, which may prove poisonous in confined quarters; and as it is heavier than air it is of no value for smothering fires overhead or in side timbers and would be dissipated by a good ventilating draft. The old soda-acid extinguisher or the more modern "foam" types ' are preferable. The former probably does not rely for its efficacy on the carbon dioxide generated but on the smothering and cooling effect of the water and the trajectory of stream that may be obtained because of the gas pressure developed. The production of a foam having a tough film to smother a fire, even on overhead surfaces, with the added

Jan 1, 1923

By Benjamin F. Tillson

Although portable fire extinguishers are valuable for fires in an incipient stage, some medium that will dissipate a large amount of heat is needed to fight a fire in mine timbers, because of the size to which the fire will probably have grown before it is discovered. Scaling off, or otherwise smothering such a fire with gases that will not support combustion, is often ineffective for the residual heat will cause reignition when the area is opened and fresh air reaches the hot charred timber. The most effective ways of fighting mine fires, therefore, are to drown and cool the hot zone with water, and to "load out" the partly burned material. Either of these methods requires a direct attack of water ; upon the inflamed material; therefore, every mine containing a large ' quantity of timber or other combustible should be as completely equipped with a fire-hydrant system as it is with compressed-air mains. Mine managers may object to the expenditure, especially if they have not suffered from mine fires. The cost of such preventative measures, however, is only a small portion of the possible losses, and the water pipes can be used for other purposes, which will pay the carrying charges. They can supply water to the rock drills, which will encourage more wet drilling with hollow-drill steel, for the trouble and expense of packing water to individual tanks at the drills will be removed. The carbon-tetrachloride hand extinguishers seem to have no place in mining except for fires that start in electrical equipment. They give off noxious gas, which may prove poisonous in confined quarters; and as it is heavier than air it is of no value for smothering fires overhead or in side timbers and would be dissipated by a good ventilating draft. The old soda-acid extinguisher or the more modern "foam" types ' are preferable. The former probably does not rely for its efficacy on the carbon dioxide generated but on the smothering and cooling effect of the water and the trajectory of stream that may be obtained because of the gas pressure developed. The production of a foam having a tough film to smother a fire, even on overhead surfaces, with the added

Jan 1, 1923

By E. W. Engle

For several years the Fansteel Products Co. has been engaged in e production and development of various of the rarer metals. It is at present engaged in the commercial production of tantalum, tungsten, and molybdenum in a very high state of purity. The information contained in this paper is partly a general compilation from various sources and partly information obtained as a result of recent years of work in the laboratory of this company. Tantalum, tungsten, and molybdenum form a group peculiar to themselves among the commercially available metals. Their very high melting points (tantalum 2770' C., tungsten 3350' C., and molybdenum 2550" C.) set them apart so far as the metallurgical processes for their production are concerned. Further, when properly prepared and treated they will stand a somewhat unusual degree of cold work, that is work below the temperature at which marked crystal growth takes place. Such cold-worked metal may show a high tensile strength; tantalum to 130,000 lb. per sq. in., tungsten to 490,000 lb., and molybdenum to 260,000 lb., in the form of fine wire. In physical appearance, these metals are not unlike some steels; they may be better described by saying that polished tantalum somewhat resembles lead; polished tungsten, iron; and polished molybdenum, nickel. The polished surfaces of all three remain untarnished in ordinary air. Air containing corrosion-producing impurities will, in general, show molybdenum the least resistant to corrosion and tantalum the most resistant. Molybdenum has, however, a high degree of resistance to corrosion by hydrochloric acid, a property that should render it of commercial interest. On the other hand, it is markedly attacked by some of the weaker organic acids. The three metals oxidize rapidly when heated at or above a dull red in air. This feature limits their use in many applications where their high melting point and strength make them appear usable. Tungsten and molybdenum, at red heat, are quite inert to hydrogen and nitrogen; while tantalum has a rather remarkable affinity for all the ordinary gases, which results in its becoming embrittled in their presence. This

Jan 1, 1925

By George W. Sheary

Situated 45 miles west of Philadelphia and 50 miles southwest of Bethlehem Steel Company's Bethlehem plant, the integrated Grace mine facilities produce 4000 tpd of high-quality iron ore pellets. Startup of the Grace mill was in November 1958, and iron concentrates were shipped until March 1960, when the pellet plant went into operation. The underground mine, when fully developed, will hoist about 9000 tpd, including development rock. The crude ore, which has an iron content of approximately 42%, is magnetically cobbed and crushed in closed circuit with 3/4-in. slotted vibrating screens. The ore is washed and screened in the mill and the plus 1/4 in. screen product is again cobbed prior to feeding to the rod mills. Grinding is done in rod mills, and in ball mills that are in closed circuit with 30 in, cyclones, to 1% on 100 mesh and 65% minus 325 mesh. The ground ore is magnetically concentrated by three-drum permanent type finisher magnets to 67% Fe. The magnetite concentrate is pumped from the mill to a 50 ft thickener in the pellet plant. The thickened magnetite concentrate is relayed to the filters through a two-stage pump arrangement. The filter cake is fed by belt from filter concentrate bins to mixer-muller belts where it is intimately mixed with bentonite and a slight amount of water for moisture control. The fluffed concentrate then drops into balling cones. The green pellets from the balling cones are fed in a prescribed feed pattern to shaft furnaces through the use of shuttle car mechanisms. The green pellets are dried, heated to 2500°F and cooled to 1200°F as they descend through the furnace and into pellet coolers where the pellets are further reduced in temperature to 250°F. Heat is recovered in the cooling process through the use of air-to-air heat exchangers. The cooled pellets discharge on a rubber conveyor belt and are transferred to 6000-ton pellet railroad loading bin. The final pellet product consists of 90% plus 1/4 in. material (after tumble testing1) and contains 66% Fe plus 3% silica.

Jan 5, 1962