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By Leo J. Vogel, E. D. Hummer

INTRODUCTION by E. D. HUMMER An efficient refuse-disposal system is a necessary part of the modem cleaning plant. The large-scale refuse system and disposal area, engineered for the lifetime of the plant, has become more important due to the greater tonnages of refuse produced from increased quantities of mechanically mined and prepared coal. Laws and increased government regulations have affected the design and operation of disposal systems. THE EFFECT OF MINE AND PLANT OPERATION Refuse disposal starts at the mining face. The mining system, the type of roof support, the machinery used, and the degree of discipline at the face contributes to the amount and type of refuse that must be removed through the disposal system. Economies gained through modification of the mining system or mining machinery may be partially offset by increased dilution of the mined coal with impurities from associated rock formations. In some cases the change in the mining system and equipment has reduced the dilution. In all cases the ratio of impurities that must be removed from the raw-coal feed to the plant not only affects the maintenance and operating costs of the plant, but greatly affects the design of the refuse disposal system. The design of a mine and plant should include the location and the estimated capacity of the disposal areas for the life of the property or plant. Complete evaluation of the disposal system and an estimate of its annual cost of operation over the life of the property or plant should be incorporated in the plant and mine evaluation. Plant design should be as simple as possible, consistent with the efficient production of clean coal. Material transfer points, crushing, screening and pumping operations should be kept to a minimum to reduce degradation and the subsequent increased difficulty and cost of refuse removal. The maintenance of good process control by operating management will result in a minimum of misplaced material in both refuse and clean coal. Reduction of misplaced float material in the plant refuse is an important factor in preventing combustion in the refuse storage pile. THE REFUSE DISPOSAL AREA Combustion in refuse piles occurs when air enters the pile in sufficient quantity to support oxidation. At certain rates of air in-filtration, the heat produced by the oxidation process is not dissipated and temperature rises.

Jan 1, 1968

By Herty, C. H.

The effect of temperature on the physical properties of structural materials has been the subject of many investigations during the past decade. The literature on the effect of elevated temperatures. is voluminous, and during the last four years a great deal of experimental work has been carried out on the effect of low temperatures on the impact properties of steel. For many engineering purposes it is important to know the shock resistance of steels at low atmospheric temperatures, and with the ever-developing field of refrigeration and separation of materials at low temperatures, the normal limits of atmospheric temperature have been exceeded in practice. Iron and its alloys in the unnotched condition have a high impact strength at low temperatures, but in practice variations in type of notches result either from design or accidentally through scratches and surface or internal defects. A study of notched bar impact tests is, therefore, highly important to the engineer. A survey of the literature on low temperature impact testing indicates that chemical composition, heat treatment and method of testing have been the variables which have been considered in most previous work. In only a few instances has any mention of the steel- making method been mentioned and in these the comparison was between steel-making processes and not between steel-making methods on the same process. In view of the wide variations in impact properties at room temperature, in quench and strain ageing and in grain growth of steels of identical chemical composition, (by this is meant the usual five element analysis reported), it was felt

Jan 1, 1957

By J. B. Austin

IN considering so complex a process as the smelting of iron in the blast furnace, there is obviously no single method of calculating efficiency that gives a complete appraisal of the performance of the furnace in all its several functions; such a comprehensive view as is required to evaluate producing ability and manufacturing costs for a given furnace is obtained only by considering a group of efficiencies, each of which measures the performance of the furnace from a specific point of view. For instance, there is the rate of consumption of coke per ton of iron produced, which is probably the most frequently considered, and is certainly one of the most important of the group; there is the efficiency of recovery of iron charged; and there is the efficiency of utilization of energy, both chemical and thermal, in some respects related to the former two yet in other ways independent of either. In each of these cases, the calculation can be made on the basis of the actual input of material or energy, or it may be based . upon a comparison with the minimum amount of material or energy that would be required by a "perfect" furnace; that is, an imaginary ideal furnace analogous to the perfect steam engine used in thermodynamic calculations, when this furnace is operating under optimum conditions. Moreover, the furnace itself can be considered either as an apparatus for smelting iron or, taking a somewhat less common point of view, as a gas producer yielding iron and slag as by-products. Some of these effi-ciencies are more useful than others, yet all are instructive; for a knowl-edge of them not only leads to good engineering practice from the standpoint of control but also discloses limits to possible improvement in performance and indicates the lines along which improvement is most likely to be profitable. Estimates of some of these efficiencies have appeared from time to time in the literature, but many of them are open to objection because they are based on data of uncertain accuracy or because they have been calculated without due regard to the thermo-dynamic principles involved. Moreover, so far as the author is aware, no general survey of the blast furnace comparing its performance in all the aspects mentioned has been made.

Jan 1, 1938

By Christian F. Baiz

INTRODUCTION Industrial minerals can have similar financing requirements to those needs of metaliferrous and coal mining projects. One of the outstanding differences is the understanding of the many different markets that these industrial minerals represent. There are in excess of 30 principal industrial minerals ranging in terms of annual production from the largest quantity such as sand and gravel, to possibly the smallest quantity of production of a legitimate industrial mineral which may be defined as industrial and rough gem diamond. (See Tables 1 and 2) Project finance techniques, analytical methods, and information requirements are covered elsewhere in this Chapter and Chapter 10. Some comments on equity finance aspects follow. EQUITY AMOUNT The amount of equity in a financing transaction provides a cushion to the lender in case the project runs into trouble. Generally speaking, the greater the amount of risk in a project, the more equity a lender will require. The amount of cash equity in a project financing transaction is between 25 and 40 percent. The amount of leverage (debt/equity) is negotiable based on such factors as the strength and extent of the sponsor's undertakings, of the project's economics, and the term of a strong sales contract. Naturally, a borrower seeks to obtain d loan with as little equity required as can be negotiated. Zero percent equity is usually associated with a take-or-pay contract (a hell-or-high- water unconditional commitment under which the purchaser must pay for the commodity as though it were shipped regardless of whether it is actually mined or shipped) with a financially strong corporate entity i.e., the parent company or an end user. As a take-or-pay contract is tantamount to a guarantee of payment, the banker views the credit basis of the loan as being other than the borrower's mining operation and therefore will analyze the creditworthiness of the institution purchasing the commodity on the basis of its corporate balance sheet. Do not confuse a take-or-pay contract with a take-if- tendered or take-and-pay contract. Another means in the United States of obtaining zero percent equity is the carved out production payment financing, typical of coal mine financing. This involves the sale, or conveyance, by the owner of the mineral property of a production payment to a nominally capitalized company. The nominally capitalized company is formed solely to buy the production payment and usually borrows 100 percent of the purchase price from a lender. Equity amounts of 10 to 20 percent generally are found in transactions which employ techniques such as an advance sale or a subordinated lease/loan. These agreements are viewed as quasi equity by most banks for several reasons. First, the proceeds represent a cash infusion early in the project's life - usually prior to the bank lending any funds. Thus, the repayment schedule is "behind" that of the bank debt. In the case of an advance sale, repayment is in the form of a long-term supply contract with favorable terms, usually priced close to the mining operation's costs. There is also usually a requirement that all costs of the project be paid for by the loan or equity, often making the sponsors financially responsible for any cost overruns. Cost overruns are a major risk in any project, large or small, and are often due to unforeseen factors such as wildcat strikes, a piece of faulty equipment, or any event which delays completion, e.g., remote location, weather, etc.. . The question of how overruns are covered can be negotiated. For example, in addition to the original debt/equity ratio of the project, there can also be a cost overrun ratio that will

Jan 1, 1985

By Charles F. Park

CORRECTLY speaking, iron ore is limited to any naturally occurring rock from which iron may be extracted at a profit, but in practice the term is frequently used to indicate borderline material or inaccessible material that may in the future become commercially profitable. This distinction is much more than quibbling about terminology. When a layman hears that there are about 45 billion tons of iron ore in the world he naturally thinks that sounds like quite a bit of rock, and what is the use of worrying about the depletion of so plentiful a material. He does not realize, nor does he stop to analyze. why iron ore in Europe and Japan may carry no more than 30 per cent iron while tremendous quantities of 50-60 per cent iron ore in Brazil are not exploited. To people in the iron industry the answer is obvious: the Brazilian material is inaccessible, and there is no suitable fuel near-by. In places closer to home, submarginal ore in considerable quantity contains undesired impurities, such as sulphur, phosphorus, or titanium.

Jan 1, 1947

By A. E. Back, K. E. Tame, W. F. Wyman, S. F. Ravitz

IN 1940, when it appeared that the United States soon might be cut off from foreign sources of high-grade manganese ore, the Bureau of Mines began an extensive series of investigations on the production of ferrgrade products from low-grade domestic ores by ore-dressing, pyrometallurgical, and hydrometallurgical methods. Among the hydrometallurgical processes for oxide ores investigated was a sulphur dioxide process in which the manganese was recovered from solution by precipitation with lime;1 this process is essentially as follows: The ground ore is suspended in water and leached with SO2 gas to extract the manganese as manganese sulphate. Calcium chloride is then added to the slurry to precipitate calcium sulphate and form manganous chloride solution, and the resultant slurry is filtered to remove the calcium sulphate together with the in- soluble portion of the ore. Calcium hydroxide is added to the filtrate and the mixture is agitated with air to precipitate hydrated oxides of manganese and regenerate calcium chloride. The precipitate is calcined, and the calcium chloride is recycled. To determine manganese extraction, reagent consumption, grade of product, and similar data for this process, a laboratory investigation was started. Two minor modifications in the process were soon made: (I) the ore was suspended in calcium chloride solution instead of in water in order to extract the manganese and precipitate the calcium sulphate in one step, and (2) the agitation with air after the addition of slaked lime to the filtrate was omitted because manganese hydroxide is readily precipitated without air agitation and can be nodulized or sintered to a satisfactory product. When the process was carried through a series of cyclic tests, it was noted that the concentration of dithionate ion (which is one of the products of oxidation of SO2 by MnOz) in the cycled solution gradually increased; at the same time the chloride content of the solution gradually decreased, owing to soluble losses in the leach residue and in the precipitate. After a number of cycles, the calcium chloride became virtually completely replaced by calcium dithionate, the quantity of dithionate ion formed during the leach being more than enough to make up for the losses of dithionate in filtering. The calcium chloride process thus developed into a calcium dithionate process. Since the latter eliminated the need for make-up calcium chloride, and the dithionate solutions appeared much less corrosive than the chloride solutions, work on the calcium chloride process was dropped in favor of the calcium dithionate process. ACKNOWLEDGMENTS The dithionate process was developed by the Bureau of Mines as part of a

Jan 1, 1946

By Charles W. Eichrodt

So much has already been wimitten on this vast subject of ground movement and subsidence, and so many data collected and commented upon, that in this paper the author proposes to confine himself to the submission of some plans showing what has actually occurred in the mining of several of the orebodies which have come under his direction during the last 20 years or so at Rio Tinto, Spain. Some notes are added for the purpose of describing the plans. The author does not attempt to give his opinion as to the reasons why the subsidences have occurred in the manner in which they are found, but confines himself to giving as true a picture as possible of the movements that have taken place. Effects of Excavation on Ground Movement Broadly, the effects of excavations on the overhead or adjacent ground can be classified under three headings, as follows: 1. Ground movements due to excavations made from surface only, such as those made in opencasts where no excavations underlie them. The stability and strength of the ground to resist movement must determine the slopes. 2. Ground movements due to underground excavations only; i. e., where the surface is not removed. 3. Ground movements due to a combination of both; i. e., from excavations taking place from the surface for the purpose of stripping an orebody and recovery of the ore to a certain horizon combined with excavations being carried out below, at the same time, for the winning of the ore below the horizon at which mining by stripping or opencasting under the specific conditions is no longer economical. Nos. 1 and 2 represent independent and distinct problems and No. 3 is a combination, which presents results much more difficult of explanation. The accompanying plans and sections give examples of what has occurred under the three conditions. Unfortunately, the class of ground is not exactly the same in them all, but the difference is not greater than is generally encountered over districts of equal area. Excavations from Surface To illustrate No. 1 three cross-sections are shown, of excavations made for the stripping and excavation from three different orebodies

Jan 1, 1930

By H. W. Graham

The young metallurgist of today who thinks casually of the technical literature of the steel industry might conclude that little has been published concerning the Bessemer process. This conclusion is true as relating to the last decade or two, but there is a fairly voluminous Bessemer literature in the period of 1860 to 1910. Since that time, however, an entirely new level of metallurgical knowledge of steelmaking has been reached; and the present paper is simply an effort to look broadly from this new level and to bring the recorded history of the process up to date. It is an effort to appraise the acid Bessemer method of steelmaking in the light of today's technical knowledge thereof. In this paper matters of history will be dealt with only insofar as they contribute to an understanding of today's situation. Attention will be directed to the apparent causes of past developments in seeking to perceive the true direction of trends now in existence and effective as bearing upon future developments. The paper will discuss the economics of the Bessemer process, but only to the extent that will bear upon the metallurgist's understanding of his responsibility. The problem of the quality of Bessemer steel will be reviewed and examples will be given of the technical data upon which the reasoning of the paper is based. Matters of engineering design, mechanical maintenance, operation and metallurgical investigation will be surveyed. Comparisons will be made between the Bessemer and open-hearth processes for the sake of clear and ready understanding. Production First of all it is considered logical and informative to show how much Bessemer steel has been and is being produced, and the relation of that production to the tonllage made by the open-hearth process. Table 1 shows the tonnages made by the two processes in representative years. It is to be noted that these figures are in net tons of 2000, lb., whereas previous to 1940 the Institute figures were compiled in gross tons. id open-hearth production and steel going into castings are included, since separated data are not available for the early years of the period covered. The ratio shown in Table I is merely the Bessemer tonnage for a given year divided by the open-hearth tonnage of the same

Jan 1, 1941

By H. W. Graham

The young metallurgist of today who thinks casually of the technical literature of the steel industry might conclude that little has been published concerning the Bessemer process. This conclusion is true as relating to the last decade or two, but there is a fairly voluminous Bessemer literature in the period of 1860 to 1910. Since that time, however, an entirely new level of metallurgical knowledge of steelmaking has been reached; and the present paper is simply an effort to look broadly from this new level and to bring the recorded history of the process up to date. It is an effort to appraise the acid Bessemer method of steelmaking in the light of today's technical knowledge thereof. In this paper matters of history will be dealt with only insofar as they contribute to an understanding of today's situation. Attention will be directed to the apparent causes of past developments in seeking to perceive the true direction of trends now in existence and effective as bearing upon future developments. The paper will discuss the economics of the Bessemer process, but only to the extent that will bear upon the metallurgist's understanding of his responsibility. The problem of the quality of Bessemer steel will be reviewed and examples will be given of the technical data upon which the reasoning of the paper is based. Matters of engineering design, mechanical maintenance, operation and metallurgical investigation will be surveyed. Comparisons will be made between the Bessemer and open-hearth processes for the sake of clear and ready understanding. Production First of all it is considered logical and informative to show how much Bessemer steel has been and is being produced, and the relation of that production to the tonllage made by the open-hearth process. Table 1 shows the tonnages made by the two processes in representative years. It is to be noted that these figures are in net tons of 2000, lb., whereas previous to 1940 the Institute figures were compiled in gross tons. id open-hearth production and steel going into castings are included, since separated data are not available for the early years of the period covered. The ratio shown in Table I is merely the Bessemer tonnage for a given year divided by the open-hearth tonnage of the same

Jan 1, 1941

By Rudolf Gahl

Page Tests Conducted in Small Test Mill................... 577 Demonstration Tests Conducted by Minerals Separation Co....... 577 Flotation Tcsts Conducted by Inspiration Co............. 577 Sampling of Orebody........................ 578 Other Results Obtained in Small Test Mill.............. 580 Tests Conducted in Large Test Mill................... 580 Flow Sheet............................. 580 Discussion of the Value of Preliminary Table Treatment......582 Callow Flotation Installation..................... 585 Flinn-Towne ~nstallation...................... 587 Cole-Bergman Installation...................... 588 Methods Followed in Competitive Tests................ 590 Advantage of Pneumatic Flotation Machines.............. 591 Pneumatic Machines of Minerals Separation Co............. 591 Float Skimming Device......................; 594 Experimental Inspiration Flotation Machine.............. 596 Arrangement of Callow Cells in Series ................ 598 Variations of Flow Sheet....................... 599 Coal Tar as Flotation Agent..................... 600 Experience with Primary Slime.................... 602 Influence of Iron on Flotation..................... 605 Influence of Chemicals on Primary Slime............... 607 Fineness of Crushing Desired.................... 607 Operation of Large Concentrator..................... 608 Flow Sheet of Callow Sections.................... 608 Flow Sheet of Inspiration Sections.................• 609 Inspiration Flotation Machine of Steel Construction.......... 610 Minerals Separation Section..................... 613 Filter Plant............................ 613 Results Obtained in Commercial Concentrator............... 616 Tonnage............................... 616 Air and Power Consumption...................... 616 Floor Space............................ 617 Water Consumption......................... 617 Recovery............................. 618 Oil Consumption.......................... 618 Operating Costs for Flotation..................... 620 Discussion of Results Obtained.................... 620 Prospects for Futurc Development of the Flotation Process......... 621 Porous-Bottom Experiments..................... 622 Raising the Grade of Concentrates.................. 624 Recovering Carbonates by Flotation................. 624 Recovery of Silicates by Flotation.................. 627 General Theory......................... 628 Conclusion............................... 630

Jan 1, 1917

By K. C. Mills, E. T. Turkdogan

The results on the rates of vaporization of Fe-Ni alloys, levitated by an electromagnetic field in a stagnant atmosphere of helium, are shown to be in close agreement with those predicted theoretically. AS shown in Part I of this series of papers, the rates of vaporization of metals enhanced by convection currents are increased further by the process of condensation of metal vapor when there is a steep temperature gradient in the boundary layer. At the time this theoretical work was completed one of the authors (K. C. Mills) had already carried out some experimental work on the rates of vaporization of molten Fe-Ni alloys. The experimental work was done under conditions which make it possible to compare the experimental rate data with those predicted on theoretical grounds. EXPERIMENTAL The object of the experiments was to evaluate the activities in a number of binary metal systems. The details of the experimental technique are given in another paper by Mills, Kinoshita, and Grieveson,2 and, for the present purpose, a brief mention of the relevant features of the experimental method will be adequate. Using an induction coil, the metal weighing about 1 to 1.5 g was levitated by the electromagnetic field in a stagnant atmosphere of pure helium. The helium was freed from oxygen and water vapor in the usual manner, and its pressure in the vaporization chamber was maintained at 1 atm through a bleeder device. The vaporization chamber consisted of a pyrex tube, 2.8 cm diam, with a side arm connected to the helium bleeder. The bottom of the tube, shaped conical, was surrounded by the induction coil, and the top was sealed with an optical flat for the purpose of measuring the melt temperature with a two-color optical pyrometer. By suitable adjustment of the high-frequency power input, the metal was brought to the desired temperature in about 30 sec of levitation. After a specified vaporization time, e.g., 3 to 4 min, the alloy was quenched by switching off the power. The amount of metal vaporized was determined from the difference between the initial and final weights of the sample. During the vaporization time, temperature fluctuations of about ±10°C were observed. COMPARISON OF EXPERIMENTAL DATA WITH THEORETICAL VALUES The experimental results are given in Table I. As already mentioned, the attainment of the vaporization temperature took about 30 sec from the moment of levitation. Therefore, the times recorded in Table I are 30 sec less than the total time; this is considered to be a reasonable estimate of actual vaporization times.

Jan 1, 1964

By N. Plint, E. T. Pinkney

Copper segregation developments within the Anglo American Group are described with particular reference to the work of the TORCO Project Team. Some radical differences have been adopted in the mechan-ical approach as compared with methods used by previous workers. Most important among these are the following: (1) The heating unit is a fluidized-bed roaster. (2) A moving bed segregation chamber is used in which a packed bed of heated ore, to which reagents have been added, moves slowly downwards under gravity and is discharged at a controlled rate by means of a suitable valve mechanism. (3) The quantities of reagents used (sodium chloride and some form of carbon) are only a small fraction of those recommended by other investigators. (4) Retention times of only 2 min have been satisfactory for the complete segregation of many types of ore. Project TORCO is the code name given by Anglo American Corp. of South Africa, Ltd., to the investigation undertaken in Zambia to develop a plant capable of treating large tonnages of refractory copper-bearing ores by the segregation process. The process has now been under investigation for six years (1960-66) and during that period work has progressed from bench-scale laboratory tests through small-scale pilot-plant work to a 500-tpd commercial pilot unit. Test work has been carried out mainly on the ores of the Zambian Copperbelt and results have been encouraging. Recoveries on the order of 90% have been obtained on all Copperbelt ores tested with high concentrate grades. PRELIMINARY INVESTIGATION Laboratory scale segregation test work on Nchanga banded sandstone ore and oxidized ore from Kansan-shi was originally conducted by Professor Maurice Rey. SociCtC Miniere et Mktallurgique de Pefiarroya, in 1957. Both ores responded favorably to the process. In November, 1960, the consulting engineers of Anglo American Corp. of South Africa, Ltd., initiated preliminary investigations on the application of the segregation process to Copperbelt ores. In early laboratory work, samples were heated in a closed, externally heated tube furnace for one hour at temperatures ranging from 800 to 830°C. The free moisture content of the ore feed was 4% and reagent additions were 0.5% salt and 1.0% coal. Both salt and coal were pre-crushed to -48 mesh Tyler.* All samples except an ore from the Bancroft mine responded well to the process and recoveries of over 90% were obtained. The presence of dolomite was responsible for the relatively poor recoveries from the Bancroft ore, as tests on one of the other ores with additions of calcite or dolomite confirmed. Recoveries dropped from 94 to 45% Grades of all the rougher flotation concentrates produced were low at 10-15% copper. The laboratory investigation was then followed up on a pilot-plant scale. An attempt was made to carry out the segregation process in a coal-fired 5-ft-diam fluidized-bed roaster. The feed consisted of a banded sandstone from Nchanga to which 0.5% sodium chloride had been added. Reducing conditions were maintained in the roaster by the addition of coal in excess of that required for heating the feed. The excess coal and added salt constituted the segregation reagents and the temperature of the bed was held at approximately 750°C. Recoveries of copper and grades of concentrates were low. The reasons for lack of success were: (1) The reaction is a gaseous one and a fluidized-bed roaster is therefore unsuitable as a segregation unit. (2) Coal in excess of that required for heating was introduced into the fluidized bed and under these conditions some of the copper would be reduced in situ. The term "reduced in situ" is used here to describe any copper mineral reduced to the metallic state in its original site within the ore particles. If this copper is to be recovered by subsequent froth flotation it must first be liberated from the gangue by grinding or some other method which, in practice, is usually not feasible on ores of this type.

Jan 1, 1969

By J. H. Clemes

The process of leaching silver ores with sodium hyposulphite is comparatively new, and published accounts of the details and losses of the process are as yet very few. The following account of a careful series of experiments to determine the actual loss in this process may therefore be acceptable to members of the Institute. The ore treated was highly calcareous, and the principal sulphide present was iron pyrites. Following is an analysis by Professor Price, of San Francisco, of the ore treated : Per cent. Silica,..........15.13 Sulphur,.........13.31 Arsenic,.......... 9.82 Iron............ 17.33 Alumina,..........1.35 Zinc,...........4.92 Lead,..........1.78 Carbonate of lime,..33.78 Magnesia,.......... 2.58 100.00 Ten tons were crushed dry with a 30 X 30 screen = 900 holes per square inch. In order to feel our way with this ore, two experimental or "pilot" charges were roasted. Pilot Charge No. 1. Treated 1000 pounds in a four-hearth reverberatory. 8.30 p.m.—Entered fourth hearth, remaining there one and onehalf hours; sulphur began to burn. 10 p.m.—Changed to third hearth; ore became red-hot half an hour after entering; the fireplace was kept almost destitute of fuel. 2.30 A.M.—Changed to second hearth. Charge continued, giving off a large quantity of sulphurous fume; the fireplace was kept dark.

Jan 1, 1884

By E. B. Hughes, F. G. Norris

Oxidation is characteristic of all processes for making steel from pig iron. This thought has been aptly expressed by H. W. Graham13 in the recent Howe Memorial Lecture, "The process of steel-making consists of using iron oxide to reduce the carbon content of the pig iron to the desired level." Direct oxidation refers to the reaction that results from introducing oxygen either alone or mixed with other gases directly into the steel bath. In working a heat, oxygen in some form is added to the steel to remove the undesirable constituents as oxides. There are two general methods of supplying this oxygen to the open hearth bath. The method in which the oxygen is added as a compound such as Con or iron ore is designated an indirect method because the steel bath must first decompose the compound to attain oxygen in a usable form. The other method in which oxygen is added to the bath in the elemental or gaseous form, more or less diluted by other gases such as nitrogen, is a direct method because no decomposition is necessary before the oxygen reacts with iron to form FeO which is dissolved in the bath in a usable form. Bubble formation (at least of a macroscopic size) presents no problem in case of direct oxidation because the reacting agent (oxygen) is in the form of bubbles when entering the metal. Regardless of whether oxygen is intro- duced into the bath directly or indirectly, several reactions proceed concurrently. In the presence of the bath of molten iron, FeO is formed either by oxidation of iron by oxygen or reduction of oxides by iron. This FeO dissolved in the molten iron is the source of oxygen for the oxidation of manganese, carbon, silicon, and phosphorus if these are present. The oxidation of carbon from steel is the controlling steelmaking reaction. The introduction of oxygen into the bath, its reaction in the elimination of impurities, and its control by the use of de-oxidizing additions have been subjects of importance since steel was first made by the bessemer process. Direct oxidation is, in a sense, an application of the bessemer principle to the open hearth process. Graham13 states: "The use of oxygen or air jets introduced into the metal bath is essentially an effort to approach in the open hearth the turbulence, rapid interface reaction, and high production rate per hour that exists in the bessemer process." The general subject, "Oxygen in Steel-making," was given new life by the studies of Herty and associates starting about1925. The current interest in elemental oxygen is intimately related to the large scale production of oxygen developed during the war period. Some of the first applications were made during the early summer of 1946 at the Steel Co. of Canada using oxygen with the fuel. Such a use is of considerable interest and importance in contributing to faster heats but is not the subject of the present discussion which deals entirely with the introduction of elemental oxygen directly into the open hearth bath.

Jan 1, 1949

By E. B. Hughes, F. G. Norris

Oxidation is characteristic of all processes for making steel from pig iron. This thought has been aptly expressed by H. W. Graham13 in the recent Howe Memorial Lecture, "The process of steel-making consists of using iron oxide to reduce the carbon content of the pig iron to the desired level." Direct oxidation refers to the reaction that results from introducing oxygen either alone or mixed with other gases directly into the steel bath. In working a heat, oxygen in some form is added to the steel to remove the undesirable constituents as oxides. There are two general methods of supplying this oxygen to the open hearth bath. The method in which the oxygen is added as a compound such as Con or iron ore is designated an indirect method because the steel bath must first decompose the compound to attain oxygen in a usable form. The other method in which oxygen is added to the bath in the elemental or gaseous form, more or less diluted by other gases such as nitrogen, is a direct method because no decomposition is necessary before the oxygen reacts with iron to form FeO which is dissolved in the bath in a usable form. Bubble formation (at least of a macroscopic size) presents no problem in case of direct oxidation because the reacting agent (oxygen) is in the form of bubbles when entering the metal. Regardless of whether oxygen is intro- duced into the bath directly or indirectly, several reactions proceed concurrently. In the presence of the bath of molten iron, FeO is formed either by oxidation of iron by oxygen or reduction of oxides by iron. This FeO dissolved in the molten iron is the source of oxygen for the oxidation of manganese, carbon, silicon, and phosphorus if these are present. The oxidation of carbon from steel is the controlling steelmaking reaction. The introduction of oxygen into the bath, its reaction in the elimination of impurities, and its control by the use of de-oxidizing additions have been subjects of importance since steel was first made by the bessemer process. Direct oxidation is, in a sense, an application of the bessemer principle to the open hearth process. Graham13 states: "The use of oxygen or air jets introduced into the metal bath is essentially an effort to approach in the open hearth the turbulence, rapid interface reaction, and high production rate per hour that exists in the bessemer process." The general subject, "Oxygen in Steel-making," was given new life by the studies of Herty and associates starting about1925. The current interest in elemental oxygen is intimately related to the large scale production of oxygen developed during the war period. Some of the first applications were made during the early summer of 1946 at the Steel Co. of Canada using oxygen with the fuel. Such a use is of considerable interest and importance in contributing to faster heats but is not the subject of the present discussion which deals entirely with the introduction of elemental oxygen directly into the open hearth bath.

Jan 1, 1949

Jan 1, 1885

By Jean A. Andre

In 1969 "Société des Mines et Fonderies de Zinc de la Vieille-Montagne" produced 221,000 tons zinc ingots and 22,000 tons zinc dust, thus rating highest on the world's zinc producer list. The Company's plant at Balen, Belgium, with an output of 135,000 tons electrolytic zinc is the most important one of this group; it is an integrated metallurgical zinc and lead unit with recovery of most by-products. The present paper describes the zinc production facilities. The roasting section includes four fluid bed roasters (total concentrates charged: 740 tpd) and two sulfuric acid having a corresponding capacity. Calcine is leached in a fully automated plant by a two-step continuous process. Leaching residues, mixed with lead furnace slag, are treated in a shaft furnace. The recovered zinc fume is leached separately. In the near future, the leaching residues are to be treated by a wet process, developed in the Company's own laboratories. Purification of solutions is accomplished in two steps: antimony and zinc dust are added in the first one, only zinc dust in the second one for removal of cadmium. Electrolysis cells are partly of a conventional type. A new type of cell configuration was put onstream in October, 1969, planned output 60,000 tpy. Main features of the new cell rooms are the fully automatic pulling, stripping and replacing of the jumbo-size cathodes (immersed area 2.6 m2). Zinc cathodes are remelted in induction furnaces followed by in-line automatic casting and stacking machines.

Jan 1, 1970

By W. J. Carlson, I. lwasaki

Manganese and silica in nonmagnetic tailings from the direct reduction-magnetic separation step were so closely associated that no physical concentration methods were effective. Dilute sulfuric acid dissolved manganese and iron rapidly; manganese in the form of silicates was virtually insoluble in sulfuric acid solutions, thereby governing the maximum extraction attained by leaching. Autoclave leaching at elevated temperatures and at high oxygen pressures reduced the iron extraction to less than one tenth without affecting the manganese extraction. There was an indication that autoclaving solubilized the manganese silicates appreciably. Elimination of iron from leach solutions was carried out by aerating the solution in the pH range of 6.5-7; a platinum-calomel electrode pair provided a reliable indicator for the end point of iron removal. The purified leach solutions were free of interfering impurities for the recovery of manganese as metal, dioxide, or carbonate. Manganiferous iron ores of the Cuyuna Range, Minn., are too low in iron content to be directly useful as iron ores, and too low in manganese content to be termed manganese ores. Common methods of physical beneficiation such as gravity concentration and flotation have not been successful in upgrading this material. The R-N process, when applied to manganiferous iron ores,1'2 produces sponge iron by direct reduction of iron oxides, while leaving manganese as manganous oxide. From fine grinding and magnetic separation of the reduced ore, a tailing is produced which is relatively rich in manganese and low in iron content. The nonmagnetic tailing can readily be leached with acids and then further processed to recover manganese in various forms. In this way, two processes, neither of which would be an economic solution to the Cuyuna ore problem by itself, may be combined to produce two marketable products, thereby rendering the over-all process economically more attractive. This article reports on the results of a leaching study on the nonmagnetic tailings, both in a batch agitation vessel and in an autoclave; on the problems involved in the purification of leach liquors by aeration; and on the recovery techniques used to produce manganese metal by electrolysis, manganese dioxide either by electrolysis or by autoclaving, and

Jan 1, 1968

By T. M. Chance

THE necessity of adopting improved methods of bituminous-coal cleaning at. many collieries is recognized by all familiar with the requirements of the various bituminous markets, the approaching exhaustion of the relatively small areas of inherently high-grade-coal, and the difficulties encountered in maintaining a uniform grade of shipped coal. Many coals lose much of their .market value by degradation in size, and, in any event, if the cleaning process is to contemplate wetting the coal the minimum amount of fines should be produced to prevent sludge losses and the difficulties resulting from the moisture content of the wet fines. Hence, it is desirable to by-pass the finer coal dry, if this can be done without too-great increase in impurities in the final product. This procedure has given excellent results at washeries at Steelton and in the Birmingham district. If the raw fines are too impure to permit admixture with the washed coal, it may be desirable to use dry methods, such as the pneumatic table, rather than to wet an excessive amount of such fine material. The bituminous-coal cleaning plant should therefore: (1) Treat the raw coal in as large lump as possible and without sizing; (2) eliminate degradation of pure coal -in the washing process; (3) avoid wetting the fines, if possible; (4) produce washed coal practically free from mechanically included impurities; (5) crush the refuse and retreat, if much intergrown coal, and refuse are present, rather than precrush the entire feed. Ordinary gravimetric methods of concentration, whether hydraulic or pneumatic, will not meet all of these requirements, as the large lump includes those sizes. now sent to the picking tables, if it is desired to treat material of such size. I do not know of any prior commercial method, based on the falling velocity, coefficient of friction or resistance to fracture, that permits the production of clean coal varying in size from, say, 10-in. lump to 1/16 -in. mesh.

Jan 6, 1924

By Henry Knox

AMIDST the tumult and the shouting of the times, there are perceptible certain tendencies, quiet though serious, to take stock of the situation with a view to putting our house in order, and of these none is more encouraging than the awakening, interest in the higher education, and especially in technical education. Various committees of the professional organizations have cast much light on the matter and have pointed the way to progress and reform, but their scope is necessarily restricted by the requirements of precision and any excursion beyond these bounds must be undertaken by the unfettered individual who is free to rush in where responsible committees fear to tread. A discussion of the bearing of the higher education on character scarcely lends itself to objective treatment and yet, fraught as it is with imponderables and abstractions, the theme cannot be escaped in any com-prehensive scheme of education. The question is being pressed with growing insist-ence, why these United States, relatively to Europe, are deficient in intellectual output, and the ready reply, "Because America is a new country and Americans are too intent on developing their natural resources," is too facile and too false to carry conviction.

Jan 12, 1922