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By S. Tandanand, Powell. L. R.

The US Bureau of Mines assessed the geological effects on the ratio of maximum subsidence and extraction thickness, known as the subsidence factor, to develop a simple method for predicting the maximum subsidence. The study is restricted to the Northern Appalachian Coal Basin where data were collected from 13 coal mines. A model previously developed to relate the subsidence factor with the lithology and with the width and depth of working panels was used in the assessment. The result shows the possibility of determining maximum subsidence without resorting to a rigorous method.

Jan 1, 1985

By Frank Firmstone

It is proposed to consider here only comparisons made between results obtained when the materials employed are precisely the same, two furnaces at the same works for example, or the same furnace under different conditions as respects temperature of blast, method of charging, etc. With this limitation, the most important point in the comparison is the amount of fuel consumed per ton of iron. This, in fact, is the only item in the prime cost of the iron which can properly be said to vary in consequence of changes in the furnace process; for of the other principal items, the ore used per ton of iron, and also the amount of fluxing material, will depend on its richness, which, under the assumed conditions, will be practically invariable when taken over any considerable period; the labor cost also, from its nature, should be very nearly constant when estimated on the ton of iron. Moreover, as it consists principally in handling and moving material, it is greatly affected by circumstances quite independent of

By C. W. Donnelly, Louise B. Freeman, Coleman D. Hunter

It is with pride that the authors of this paper report. that during 1944 the production of petroleum in Kentucky passed its all-time peak, 9,496,985 bbl. being contributed. The delivery of natural gas, especially from eastern Kentucky, was maintained and possibly exceeded the highest deliveries of the past. To raise the production of oil and gas was not easy with inferior and worn-out equipment and the shortage of skilled labor. Reminiscent of the early days of the industry in the state, particularly in the hills of Elliott County, where mud roads often have gradients of 45º, was the replacement of worn-out trucks and tractors by oxen. Of the total 9,496,985 bbl., 17 per cent was contributed by the part of Kentucky that is east of the Cincinnati arch, as compared with the total of 7,010,776 bbl. contributed in 1943, of which 25 per cent was contributed by eastern Kentucky. This discrepancy is due to the discovery of more new fields in western Kentucky rather than to a decrease in production in the eastern fields, where the average has been maintained by reconditioning and secondary recovery of old fields and further development of new pools reported for the first time in 1943. In all, 1103 wells were known to have been drilled in the state during 1944, of which 659 were drilled in western Kentucky and 444 in eastern Kentucky. The exact number of wells drilled for any year is never known, as drilling permits are not required except 'when it is planned to drill through a workable coal seam. Of the total number of wells drilled, 459 were oil wells, 229 gas wells and 415 dry holes. The monthly production was as follows: January........................... 659,496 February.......................... 658,606 March...................... ...... 711,619 April............................. 649,926 May............................... 754,892 June.............................. 730,369 July.............................. 761,762 August............................ 878,053 September...................... 863,150 October........................... 968,732 November........................ 961,999 December........................ . 898,636 Western Kentucky In western Kentucky 352 wells were completed as oil wells with a total initial production of 47,922 bbl. of oil, the average being 136 bbl. per well. Two wells produced gas, the total initial production being 1,594,000 cu, ft. One hundred and eighteon wildcat wells were drilled, of which 107 were dry holes and 11 were oil wells. Three of the latter proved to be the discovery wells of commercial pools. Twelve new pools were discovered during 1944, none of major importance, as indicated in Table 2. The Utley pool, due to a stratigraphic trap on the flank of a closed structure, found oil in the Mcclosky at 2565 ft. in the discovery well, with initial productioil of 1000 bbl. The average daily production is now 227 bbl. from seven wells. The East Poole field, with the discovery well in the Tar Springs sandstone at 1811

Jan 1, 1945

The Lornex mine is located in the Highland Valley area of British Columbia, Canada, one of the most,promising districts of low-grade porphyry-type copper mineralization in the world; Lornex is operated by Rio Algom Mines, Limited, a member of the Rio Tinto Zinc Group of London. Exploration from 1965 to 1968 disclosed an ore body of 293 million tons with a grade of 0.43% copper and 0. 14% molybdenum; this led to the decision to construct a 38,000 tpd operation, and production commenced in the fall of 1972. The ore is open-pit mined with 15 cu yd (11 5 m3 shovels, which load 100- and 200 -ton trucks. These trucks dump directly into a 60 in. by 89 in. Allis -Chalmers gyratory from two sides of the 14 ft 6 in. (4.42 m) deep dump pocket and across splitters mounted on the pocket sides. The unlined pocket would hold 1,000 tons level full. The crushedore discharges into a 300-ton surge hopper equipped with an automatic sonic level indicator to prevent over -filling. A 96 in. by 23 ft (2.44 m by 7 m) apron feeder delivers the ore from the surge hopper to a 72 in. (1.83 m) transfer belt equipped with a tramp iron metal detector. The transfer belt discharges to a 60 in. (1. 52 m) conveyor which conveys the ore about 3,000 ft (914 m) to the 475,000 ton ore storage pile located near the concentrator. Four 48 in. (1.22 m) pan feeders transfer the ore to the mill feed conveyors, which in turn supply the two 32 ft by 15 ft (9. 75 m by 4. 57 m) semi-autogenous mills, which easily accept the 10 in. (254 mm) product of the crusher. Each semi- autogenous mill is followed by a secondary ball mill. The unusually large open side setting of the crusher accounts for its 6,300 tph maximum capacity. Concentrator capacity is now 48,000 tpd. Eccentric removal is eased by the design of the door to the crushed ore pocket, which has the cart rails mounted on its outer surface and is lowered to a horizontal position in the pocket by an electric winch. The top of the door rests on wear bar covered concrete piers in the pocket. Another feature is a chute through the floor at the crusher base that extends into the crushed ore pocket. Dust and spills of ore on that floor during crusher repairs are easily swept into the crushed ore pocket. This feature is also used at Bougainville and Palabora and has been added to others after operation has commenced. The design follows the concrete rectangular box concept with four main service and maintenance levels shown in the sectional drawings, although a third additional section has been added at the maintenance

Jan 1, 1978

By O. Katz, A. Spilners, M. Simnad

THE oxidation of titanium has been studied in recent years by a number of investigators.'-W ne of the most important aspects of the reaction of titanium with oxygen, which has been observed by several of the investigators, is the extremely high solubility of oxygen in the metal. A detailed study of the Ti-0 systeml% as shown that the solubility limit of oxygen in a-titanium is approximately 14.5 wt pct 0 from 800" to 1700°C. The unusual oxidation behavior of titanium may be ascribed to this phenomenon, since part of the oxygen uptake enters into solid solution in the metal and only a part is available for scale formation. A more complete study of the kinetics of oxidation of titanium would require information on the influence of oxygen in solution in the metal upon the oxidation rate. Such information would also be of interest in regard to the theory of oxidation of metals. The authors have measured the oxidation rates of oxygen-saturated and oxygen-free titanium between 800" and 1200°C and have found appreciable differences in their rates of oxygen uptake and in the activation energies. The experimental set-up was similar to that described in a previous paper.u The course of oxida-dation was followed continously by attaching the specimen to a sensitive spring balance and observing the extension of the spring with time of oxidation. High purity electrolytic oxygen was used and precautions were taken to purify the gases employed. The specimens were prepared from iodide titanium and were ground on metallographic polishing papers to a fineness of 4/0. The furnace was first evacuated and filled with argon; the specimen was lowered into the furnace, which was then evacuated and filled with oxygen, after the specimen had reached the oxidation temperature in argon. The samples were presaturated with oxygen by means of the following procedure: Each specimen was first heated in oxygen at 1200°C until its increase in weight indicated 13.8 pct by weight uptake of oxygen. The furnace was then immediately evacuated and filled with purified argon five times. The specimen was heated in the argon atmosphere for a period of 72 hr at 1200°C in order to saturate the metal with oxygen by complete absorption of the scale. The temperature was lowered to the oxidation temperature, the furnace was evacuated and filled with oxygen, and the oxidation rate was followed by noting the extension of the spring balance. No oxide film was visible on these specimens at the start of the oxidation experiments. Titanium samples in the oxygen-free as-received condition were also oxidized and their oxidation rates measured. The results are shown in Figs. 1, 2, and 3, which indicate that the initial rates of uptake of oxygen by both the oxygen-saturated and oxygen-free specimens are parabolic at these temperatures. The rate of uptake of oxygen is, however, appreciably lower

Jan 1, 1956

By A. E. Perkins

fractured ingot will show a complete diagonal structure. If the temperature is lowered, the equi-axed zone is larger and the pine tree growth is less. A small ingot, though, is prone to this columnar structure, even though the casting temperature is kept down. The speed of cooling after solidification is slower in the hollow ingot than in the solid, therefore the granulation, which takes place after the crystals have formed from the liquid metal, is apt to produce larger grains than in the solid ingot. This is not a detriment. The original crystal size is of much more importance. The size of the grains due to this granulation constantly varies in response to mechanical work and subsequent heat treatment. Surface weakness, as shown by seaming and checking on the exterior when working or by cracking in hardening, is in some measure the result of the surface stresses of the ingot when cast. The temperature of the ingot mold or speed of freezing of the exterior of the ingot is of great importance and much seamy and cracked steel is caused by a too quickly cooled skin. Ingots that are turned or milled all over will not seam very much in working, in fact they are practically free from seams, whereas an ingot rolled without a preliminary machining operation is full of seams and cracks, some long and some short. It has been said that these short seams and cracks in the surface of the bar are the result of elongated depressions in the surface of the ingot and that if an ingot is cast very smooth it should give fewer surface defects. This is partly true, but an ingot cast very hot and having a very smooth skin will crack and seam very badly in the mill, as will the surface of an ingot poured very cold, which is very crinkly, because of the low temperature of the metal. The cause of the cracks and seams in one is not the cause of the cracks and seams in the other. The tube method of hollow drill steel manufacture is superior to the older method because of: (1) Greater freedom from external and internal straining; (2) the inherent small crystal size; (3) absence of harmful segregation, resulting in weakness of wall of hole. (4) less liability for steel to crack in inside of hole during forging or hardening; (5) toughening effect, arising from mild steel wall of hole limiting the intense hardening on quenching. Breakage and Heat Treatment or Rock-drill Steel By A. E. Perkins, Pittsburgh, Pa. Actual figures showing the superiority of one section or another are difficult to obtain. In a given size of steel, however, the cruciform sec -tion shows the most breakage; it is inherently a weak section. It contains less steel per foot than any other standard section of the same size and when rolled the wings tend to cool off setting up cooling stresses.

Jan 1, 1922

By Fred W. Gordon

It is my purpose to call attention to the close approximate correctness of certain coefficients and formulas concerning the flow of gases, and their applicability to the practical construction of flues and

Jan 1, 1886

By D. S. Sanders, E. W. Witcomb

FLOTATION tests, from which the flowsheet for the concentrator was determined, were made in a pilot mill at Chuquicamata and later comparative testing of full size flotation cells was done at the Andes Copper Mining Co. concentrator in Potrerillos. The principal sulphide copper mineral is chalcocite with varying amounts of covellite, enargite, and chalcopyrite. The oxidized portion is principally antlerite, a green basic copper sulphate. The gangue minerals include quartz, sericite, pyrite, feldspars and kaolin. Chuquicamata pit ores have proved to be hard grinding. The results of the flotation and metallurgical work prove that good extraction and concentrate grade can be obtained under a rather wide range of conditions. Metallurgy In general the metallurgy of the ore requires a fairly close control of the grind to the primary rougher flotation and of the regrind of the primary rougher concentrate before cleaning. Regrinding is necessary to reduce the insoluble content of the final concentrate since a large part of the middling combinations are between the copper and silica minerals. The alkalinity of the pulps can vary between wide limits and control of the percent solids gives the desired grind. Alkalinity Control Arroyo Salado water is supplied to the concentrator head tanks. It contains about 5000 ppm of salts but test work shows no deleterious effect. Actually this water permits a decrease of as much as 1 lb of lime per ton of ore as compared to San Pedro water, which is used for leaching oxide ore. About one quarter of the water flow through the concentrator will be new water from the reservoir. The remainder will be reclaimed from the thickeners. For primary grinding and rougher flotation, new and reclaimed water are combined in the head tanks. To avoid erratic results, only new water is used in the regrind system. Milk of lime is added to the rod mill feed chute to obtain the maximum benefit of mixing with the ore. This protects the equipment from corrosion due to entrained acids and soluble copper salts. Lime consumption is 3 to 5 lb per ton of ore. Grinding and Flotation The six bin gates over each belt feeder are opened in controlled combination and sequence to keep a live bed of ore in the bin to prevent arching and to obtain mixing. The milk of lime added ahead of the rod mill is regulated to suit rougher flotation requirements. The primary promoter reagent, Aerofloat 238, is added to the 10-mesh discharge of the rod mill as it is laundered to the ball mill scoop box. The classifier overflow is regulated to produce a grind of about 12 to 15 pct +65 mesh and 60 pct -200 mesh. A secondary promoter Z-4 (aml zanthate) and the pine oil frother are added in the overflow launder where a pH of 10 to 11 is maintained. Comparatively heavy beds of froth on the

Jan 1, 1952

Jan 1, 1961

By W. P. Blake

Wm. P. Blake, New Haven, Conn. (Communication to the Secretary, June, 1889): It is gratifying to find that Prof. Carpenter in his paper sustains the accuracy of my original determination of

Jan 1, 1889

By John K. Hammes

INTRODUCTION From the point of view of the consumer, the cost of mineral commodities might be viewed as the total price industry pays for mine output. Similarly, the mining company engaged in the ongoing business of finding, developing and operating new mines is concerned very much today with the distribution of all the revenue paid for mine output and not just with its direct or controllable cash costs. In other words, what will the labor force, the supplier of energy, explosives, and reagents, and the supplier of government services, management, capital, and the ore reserves receive for their contribution to production? A significant amount might be categorized as what economic theory classifies as economic rent. Mining companies are finding it increasingly difficult to negotiate concession or work contracts to develop ore reserves in lesser developed countries, just as they find themselves accused of taking excess profits in their own countries. The principal difficulty in reaching mutually satisfactory agreements in many cases is the difficulty in establishing an acceptable distribution of income, in particular the distribution of economic rent. (See Chapter 7). Further, it is logical that the decisions and policies of the interested parties be they government or industry should be and will be influenced by their competitive position relative to determining the distribution of economic rent. The raising of finance for a mining project today may be inhibited by too biased an extraction of economic rent by entities other than the risk takers, the mining companies. ECONOMIC RENT A British economist (Ricardo D. McGraw-Hill, 1955) first proposed the concept of economic rent (See Drechsler, H.D. earlier in this Chapter.) Ricardo pointed out that no rent would be paid until land became scarce. Ricardo also stressed the distinction that the price of farm produce was not higher because of the rent paid on farm land but that the rent or price of more fertile farm land was higher because the price of produce is higher. This led to the concept that economic rent is not a cost and that it is therefore not price determining, but is itself price determined. Economic rent may be viewed as a cost if cost is equated with the concept of payment or distribution of income to any factor used in production. However, if it is a criterion that a payment must be made to the supplier of the factor of production, then economic rent is not truly a cost because it is not necessarily distributed or "paid" to the supplier. Consider the commodity A which occurs in three grades: Al, A2, and A3. Any grade may be used as a substitute for another grade in all of the industrial uses of A. However, in industries X and Y the different grades have different productivities. Table 1 shows the relative productivities of each grade in industries X and Y, and in all other industries [~]. [ ]

Jan 1, 1985

[Allen-Sherman-Hoff Pump Co. Third Cover Dee Carlton Brown, Advertising Allis-Chalmers Mfg. Co. 822, 831, 840 Compton Advertising, Inc. American Cyanamid Co. * James J. McMahon, Inc. American Manganese Steel Div., American Broke Shoe Co. 839 Fuller & Smith & Ross, Inc. Anaconda Copper Mining Co. * Kenyon & Eckhardt Inc. Carton Products Corp. * Penn and Hamaker, Inc. Chester Cable Corp. * George Homer Martin Associates]

Jan 1, 1952

We are glad to print the following letter from Herbert C. Hoover, who has recently been appointed Food Administrator, by President Wilson. Mr. Hoover has a tremendous task before him, and we are sure that mining engineers will give him their hearty support by responding to his appeal to stop the waste of food.

Jan 7, 1917

By R. G. Clay, C. F. Seaman

THE mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that time, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered, the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. ORIGINAL EQUIPMENT AND INSTALLATION A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots and the up position would connect it to the line from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1942

By R. S. Sandhu, S. H. Advani, J. K. Lee, B. L. Aboustit

A comprehensive finite element formulation for thermo-elastic consolidation based on a general variational principle is presented Sample one dimensional problems associated with isothermal consolidation, heat conduct ion, and thermo-elastic consolidation are presented. In addition, evaluations of thermo-elastic consolidation for an underground coal conversion field site are conducted.

Jan 1, 1982

By M. C. Leverett, S. E. Buckley

The production of oil is accomplished as a result of its displacement from the reservoir by either gas or water, and the amount of oil recovery is limited by the extent to which the displacing gas or water accumulates. This paper describes the mechanism by which the displacement is effected and the advantages of water over gas as a displacing agent. In the light of the results of experimental observations of the flow of mixtures of oil and/or gas and/or water through sands, certain conclusions are drawn relative to the changing character of the displacement as depletion proceeds, and on the effects of the properties of the fluids and of producing conditions on the ultimate oil recovery. Introduction Crude oil has no inherent ability to expel itself from the pores of the reservoir rocks in which it is found; rather, it must be forcibly ejected or displaced by the accumulation of other fluids. Accordingly, a knowledge of the mechanism by which one fluid is displaced by another is essential to an understanding of the fundamental process by which oil is recovered. The displacing fluids normally available are gas and water, either or both of which may exist originally associated with the oil in a potentially usable form or may be supplied to the reservoir from external sources. Gas is present in most oil reservoirs. If the quantity is relatively small, it may exist originally completely dissolved in the oil, but if the quantity exceeds that which may be held in solution by the oil at the prevailing pressure, the excess will be found in the free state. Most reservoir sands appear to contain some water,l and in the majority of fields the oil is trapped in the structure and held over water. In certain conditions the entrapping water may advance into the oil zone and displace the oil therefrom. This paper describes in a qualitative manner some of the characteristics of the displacement of oil by either gas or water, with an attempt to elucidate somewhat the mechanism by which such displacement is effected. Role of Dissolved Gas When the pressure on an oil containing dissolved gas is reduced to the saturation or bubble-point pressure, gas escapes from solution and expands, imparting to the oil-gas mixture a high degree of expansibility and causing it to flow in the direction of reduced pressure. This expansibility has frequently been attributed to the oil; closer analysis, however, reveals that as gas escapes from solution the oil phase actually. contracts. Hence, it is the accumulation and expansion of the liberated gas that supplies the propelling force for the oil. It has been shown experimentally2-5 that within certain limits a sand containing oil, gas and water will permit the simultaneous flow of all three; the readiness with which they may flow depends upon the characteristics of the sand, of each of the fluids, and upon the relative amounts of the fluids in the sand. Thus, in the initial stages of production by the mechanism of dissolved gas drive, the liberated gas is present in the sand in low concentration and is unable to

Jan 1, 1942

By S. E. Buckley, M. C. Leverett

The production of oil is accomplished as a result of its displacement from the reservoir by either gas or water, and the amount of oil recovery is limited by the extent to which the displacing gas or water accumulates. This paper describes the mechanism by which the displacement is effected and the advantages of water over gas as a displacing agent. In the light of the results of experimental observations of the flow of mixtures of oil and/or gas and/or water through sands, certain conclusions are drawn relative to the changing character of the displacement as depletion proceeds, and on the effects of the properties of the fluids and of producing conditions on the ultimate oil recovery. Introduction Crude oil has no inherent ability to expel itself from the pores of the reservoir rocks in which it is found; rather, it must be forcibly ejected or displaced by the accumulation of other fluids. Accordingly, a knowledge of the mechanism by which one fluid is displaced by another is essential to an understanding of the fundamental process by which oil is recovered. The displacing fluids normally available are gas and water, either or both of which may exist originally associated with the oil in a potentially usable form or may be supplied to the reservoir from external sources. Gas is present in most oil reservoirs. If the quantity is relatively small, it may exist originally completely dissolved in the oil, but if the quantity exceeds that which may be held in solution by the oil at the prevailing pressure, the excess will be found in the free state. Most reservoir sands appear to contain some water,l and in the majority of fields the oil is trapped in the structure and held over water. In certain conditions the entrapping water may advance into the oil zone and displace the oil therefrom. This paper describes in a qualitative manner some of the characteristics of the displacement of oil by either gas or water, with an attempt to elucidate somewhat the mechanism by which such displacement is effected. Role of Dissolved Gas When the pressure on an oil containing dissolved gas is reduced to the saturation or bubble-point pressure, gas escapes from solution and expands, imparting to the oil-gas mixture a high degree of expansibility and causing it to flow in the direction of reduced pressure. This expansibility has frequently been attributed to the oil; closer analysis, however, reveals that as gas escapes from solution the oil phase actually. contracts. Hence, it is the accumulation and expansion of the liberated gas that supplies the propelling force for the oil. It has been shown experimentally2-5 that within certain limits a sand containing oil, gas and water will permit the simultaneous flow of all three; the readiness with which they may flow depends upon the characteristics of the sand, of each of the fluids, and upon the relative amounts of the fluids in the sand. Thus, in the initial stages of production by the mechanism of dissolved gas drive, the liberated gas is present in the sand in low concentration and is unable to

Jan 1, 1942

By W. J. Reagan

All of the material described in the following paper is within the following specifications: carbon, 0.50 to 0.85 per cent; phosphorus and sulfur, 0.04 per cent max.; manganese, 0.60 to 0.85 per cent; silicon, 0.20 to 0.30 per cent. Steels with much lower carbon contents can be handled in exactly the same manner. One of the first things necessary to basic open-hearth slag control is uniformity of charge, particularly as to the silicon content and the percentage of lime charged. Our past work has shown that the silicon content of the charge is more important than the carbon content in determining the carbon at which a heat will melt. We have found that by keeping the silicon content of the charge about 0.60 to 0.65 per cent, and with a lime charge of about 7 per cent (limestone equivalent), our heats melt sufficiently high in carbon to allow a moderate feed of ore and the slag will be of low iron oxide content. In making a heat of 0.75 carbon steel we endeavor to have the heat melt at about 1.20 per cent carbon. The iron oxide content of the finishing slag depends upon the final phosphorus desired; for a given ratio of lime-silica in the slag, the lower the phosphorus desired in the final steel, the higher the iron oxide content of the finishing slag must be. By reducing the lime charge to a minimum, the heats melt faster, and by keeping the silicon content at the desired figure the heats melt with a high SiO2 slag, which has been shown in previous work to give a low iron oxide slag. In most of our work, using a 180,000-lb. charge and with basic pig containing about 1.20 per cent silicon, we use a lime charge of 8500 lb. of limestone and 2000 lb. of burnt lime. The pig charge is about 48 per cent of cold pig, and the charge contains a large percentage of works scrap. This causes a heat to melt with about 1.20 to 1.30 per cent carbon, for a heat that will finish about 0.75 per cent carbon. This gives a heat that will need a feed of about 2000 lb. of ore, and also a slag with an FeO content of about 4 to 5 per cent at melting. The additional FeO obtained from the ore addition will bring the FeO in the slag at tapping to about 6 to 7 per cent. It is necessary to carry quite a fluid slag to prevent pickup of FeO, as a too heavy slag causes FeO to build up in it, giving a finishing slag with a greater FeO than is desired. By adjusting the lime charge of a basic open-hearth heat, a very sensitive adjustment of the slag may be obtained. The figures in Table 1

Jan 1, 1935

By W. J. Reagan

All of the material described in the following paper is within the following specifications: carbon, 0.50 to 0.85 per cent; phosphorus and sulfur, 0.04 per cent max.; manganese, 0.60 to 0.85 per cent; silicon, 0.20 to 0.30 per cent. Steels with much lower carbon contents can be handled in exactly the same manner. One of the first things necessary to basic open-hearth slag control is uniformity of charge, particularly as to the silicon content and the percentage of lime charged. Our past work has shown that the silicon content of the charge is more important than the carbon content in determining the carbon at which a heat will melt. We have found that by keeping the silicon content of the charge about 0.60 to 0.65 per cent, and with a lime charge of about 7 per cent (limestone equivalent), our heats melt sufficiently high in carbon to allow a moderate feed of ore and the slag will be of low iron oxide content. In making a heat of 0.75 carbon steel we endeavor to have the heat melt at about 1.20 per cent carbon. The iron oxide content of the finishing slag depends upon the final phosphorus desired; for a given ratio of lime-silica in the slag, the lower the phosphorus desired in the final steel, the higher the iron oxide content of the finishing slag must be. By reducing the lime charge to a minimum, the heats melt faster, and by keeping the silicon content at the desired figure the heats melt with a high SiO2 slag, which has been shown in previous work to give a low iron oxide slag. In most of our work, using a 180,000-lb. charge and with basic pig containing about 1.20 per cent silicon, we use a lime charge of 8500 lb. of limestone and 2000 lb. of burnt lime. The pig charge is about 48 per cent of cold pig, and the charge contains a large percentage of works scrap. This causes a heat to melt with about 1.20 to 1.30 per cent carbon, for a heat that will finish about 0.75 per cent carbon. This gives a heat that will need a feed of about 2000 lb. of ore, and also a slag with an FeO content of about 4 to 5 per cent at melting. The additional FeO obtained from the ore addition will bring the FeO in the slag at tapping to about 6 to 7 per cent. It is necessary to carry quite a fluid slag to prevent pickup of FeO, as a too heavy slag causes FeO to build up in it, giving a finishing slag with a greater FeO than is desired. By adjusting the lime charge of a basic open-hearth heat, a very sensitive adjustment of the slag may be obtained. The figures in Table 1

Jan 1, 1935

By Freeman Bishop

Nothing has disturbed the copper- lead-zinc producers in recent years more than smelter tolerances set by some states. What hurt most, copper spokesmen told the Government, was the Government's own erroneous estimates of smelter pollution controls. Government charts indicated that 23 firms smelting copper, lead and zinc may need to invest about $108 million by fiscal 1976 to control stack emissions. The copper industry's independent study, however, indicates that the required controls could cost as much as $550 million to install and $200 million a year to operate.

Jan 1, 1971