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By Edward C. J. Urban

ESSENTIAL requirements for ensuring safe working atmospheres in underground metal mines are planned systems of ventilation and provision for effective distribution of sufficient volumes of air by auxiliary equipment. Investigators1 in the past have suggested that quantities of air supply ranging from 300 to 800 cu. ft. of air per minute per man underground are necessary, to establish adequate general ventilation. These fundamentals are just as important as the use of water in dust suppression, control of blasting, the use of personal protective equipment, frequent inspection of working areas and education of employees in good dust-prevention practice. During the past two years the necessity for direct control of the underground working environment has become particularly urgent. The increase in production rate, with two-shift and three-shift operation replacing the single shift, has complicated the problem of dealing with excessive concentrations of blasting fumes and gases, carbon dioxide and monoxide, heat and oxygen deficiency. Although previous industrial hygiene programs had given consideration to these factors, they were generally regarded by mine operators as secondary to the control of dust hazards. Today, blasting smoke in a mining area means delayed production because the miner cannot return to his work place after a short interval. Smoke on the haulage level means a safety hazard and exposure of the underground employees to excessive concentrations of atmospheric contaminants. Exposure to the moderately excessive concentrations of carbon dioxide and monoxide found in metal mines has caused physiological disturbances and a decreased capacity for work; deficiency of oxygen, in some cases to as low as 3 per cent, has resulted in fatalities. During the past 10 years the Saranac Laboratory2 has been associated with the iron mines in the Upper Peninsula of Michigan and Wisconsin in developing and maintaining an industrial hygiene and silicosis-prevention program. At present, about 35 underground iron-ore mines are operated full time in this area. With two exceptions, all of them are ventilated by main fan installations installed either on surface or underground. ,The types of mining used in these mines may be classified either as sublevel cave, radial top slice or sublevel stope, and the procedures commonly employed in the ventilation of these three types are described in Figs. 1 and 2. A number of different methods of mine ventilation had been in use in this area when the study was started. Natural ventilation had proved unsatisfactory. The wide variation between low air volumes during the summer months and higher volumes during winter led to unpredictable conditions that were difficult to control. In the majority of the mines two separate shafts or openings to surface are utilized in

Jan 1, 1943

By E. C. Carris

The loading of bituminous coal mechanically was recognized about the year 1900, but during the following 28 years less than 5 per cent of the bituminous production in the United States was loaded by mechanical means. During the past 16 years, however, the usage of some form of mechanical loader has had a very rapid and accelerated growth, until at present production of bituminous coal by mechanical means has reached an all-time high of approximately 46 per cent, and is still increasing. This rapid growth in the use of the mechanical loader has been supplemented by an increase. in mechanical cleaning. The importance of mechanical cleaning of coal was recogniaed many years before the mechanical loader was introduced. because of the impossibility of cleaning manually the sizes below 2 in.. and the increasing competition from other fuels. Utilization of mechanical loaders without nleclianical cleaning 011 the surface for at least one or more sizes is the exception, rather than the rule. One exception is a mine that employs caterpillar-mounted mobile loaders exclusively, and mines the entire seam without any selectivity, and yet loads and ships six grades of coal of the very highest quality without mechanical cleaning. Such a fortunate situation can exist only where mining conditions are ideal, where top and bottom are exceptional and there are no impurity partings, and where there is loader height. The conversion from hand loading to mechanical loading has definitely introduced many problems in the preparation plant of yesterday and in the modern streamline plant as well. Many of these problenis have been solved by good, sound, conscientious initiative on the part of the plant personnel, with minor expenditures; the solution taking the pattern of rearranging 2nd improving the equipment at hand. Other problems have caused, and are causing, considerable concern and difficulty. Large Impurities The removal of large quantities of large impurities from the lump coal, and their handling and disposal, is definitely a troublesome and costly problem. The cost of manually removing the impurities from the picking table alone, based on the actual recovered lump coal, can reach a cost of 42 cents per ton or more. In the design of future plants it would be wise to give this phase of preparation considerably more time and thought, particularly if the seam of coal contains one or more heavy partings and the mining procedure is all seam. This same caution applies to seams that may not contain heavy partings but have poor and heavy draw-slate top, which falls with the coal during the mining cycle. One of the more recent of the modern preparation plants has answered this problem by installing a rock crusher under the picking table in such a manner that the impurities can be diverted by gravity directly to the crusher. This reduces the

Jan 1, 1947

By E. C. Carris

The loading of bituminous coal mechanically was recognized about the year 1900, but during the following 28 years less than 5 per cent of the bituminous production in the United States was loaded by mechanical means. During the past 16 years, however, the usage of some form of mechanical loader has had a very rapid and accelerated growth, until at present production of bituminous coal by mechanical means has reached an all-time high of approximately 46 per cent, and is still increasing. This rapid growth in the use of the mechanical loader has been supplemented by an increase. in mechanical cleaning. The importance of mechanical cleaning of coal was recogniaed many years before the mechanical loader was introduced. because of the impossibility of cleaning manually the sizes below 2 in.. and the increasing competition from other fuels. Utilization of mechanical loaders without nleclianical cleaning 011 the surface for at least one or more sizes is the exception, rather than the rule. One exception is a mine that employs caterpillar-mounted mobile loaders exclusively, and mines the entire seam without any selectivity, and yet loads and ships six grades of coal of the very highest quality without mechanical cleaning. Such a fortunate situation can exist only where mining conditions are ideal, where top and bottom are exceptional and there are no impurity partings, and where there is loader height. The conversion from hand loading to mechanical loading has definitely introduced many problems in the preparation plant of yesterday and in the modern streamline plant as well. Many of these problenis have been solved by good, sound, conscientious initiative on the part of the plant personnel, with minor expenditures; the solution taking the pattern of rearranging 2nd improving the equipment at hand. Other problems have caused, and are causing, considerable concern and difficulty. Large Impurities The removal of large quantities of large impurities from the lump coal, and their handling and disposal, is definitely a troublesome and costly problem. The cost of manually removing the impurities from the picking table alone, based on the actual recovered lump coal, can reach a cost of 42 cents per ton or more. In the design of future plants it would be wise to give this phase of preparation considerably more time and thought, particularly if the seam of coal contains one or more heavy partings and the mining procedure is all seam. This same caution applies to seams that may not contain heavy partings but have poor and heavy draw-slate top, which falls with the coal during the mining cycle. One of the more recent of the modern preparation plants has answered this problem by installing a rock crusher under the picking table in such a manner that the impurities can be diverted by gravity directly to the crusher. This reduces the

Jan 1, 1947

By W. I. Garms, J. L. Stevens

Because of the many conflicting opinions regarding the functioning of the ball load and the characteristics of ball wear in ball-mill grinding, the following résumé of reports covering recent tests conducted at the Hayden plant of the Kennecott Copper Corporation is submitted with the view of pointing out certain salient features that should throw additional light on the subject. Mills Used Cylindrical drum mills with trunnion feed and discharge were used. The discharge trunnion was 13½ in. in diameter. Dimensions inside the lining were 6 ft. 6 in. in diameter and 9 ft. 6 in. long. The mills were Britannia-lined and nine lifter bars projected 2 in. from the lining. The mill speed was 162/3 r.p.m. Feed to the mill was 9 per cent on 10-mesh, 85 per cent on 100-mesh, 77 per cent solids. New balls measured 2 in. in diameter and the approximate weight of the ball load was 40,000 pounds. Initial Test During a test that was being conducted to determine the relative economic value of standard cast balls vs. forged steel balls in the local 'grinding operation, certain observations were made in regard to the functioning of the various ball sizes and the characteristics of ball wear, which warranted a supplementary test for the purpose of confirming such observations. For the initial test, the necessity for developing a graduated load of the new forged balls before the test could be started was realized, but the time required to obtain a graduated load and the functioning of the various ball sizes was not clear. No advantage could be seen for starting out with a full load of forged balls, as this initial load, at the least, would have to be completely ground out and replaced by daily additions of balls before the graduation could be complete and a normal static condition of ball sizes reached. For a considerable period prior to exhaustion of this initial charge there would be an unusual quantity of worn balls of a certain size, which would not be typical of a normal charge and could influence the characteristics and functioning of the entire load. Therefore, it was decided to start by daily addition of the forged balls to a mill that contained a normal operating load of cast balls. After 88 days of operation the mill was shut down for inspection. By that time the load consisted of a mass of forged balls functioning around a large core of the residual material of the former cast-ball load. This core of cast slugs, locally referred to as "the kidney," was well defined from the forged-ball load.

Jan 1, 1947

By W. I. Garms, J. L. Stevens

Because of the many conflicting opinions regarding the functioning of the ball load and the characteristics of ball wear in ball-mill grinding, the following résumé of reports covering recent tests conducted at the Hayden plant of the Kennecott Copper Corporation is submitted with the view of pointing out certain salient features that should throw additional light on the subject. Mills Used Cylindrical drum mills with trunnion feed and discharge were used. The discharge trunnion was 13½ in. in diameter. Dimensions inside the lining were 6 ft. 6 in. in diameter and 9 ft. 6 in. long. The mills were Britannia-lined and nine lifter bars projected 2 in. from the lining. The mill speed was 162/3 r.p.m. Feed to the mill was 9 per cent on 10-mesh, 85 per cent on 100-mesh, 77 per cent solids. New balls measured 2 in. in diameter and the approximate weight of the ball load was 40,000 pounds. Initial Test During a test that was being conducted to determine the relative economic value of standard cast balls vs. forged steel balls in the local 'grinding operation, certain observations were made in regard to the functioning of the various ball sizes and the characteristics of ball wear, which warranted a supplementary test for the purpose of confirming such observations. For the initial test, the necessity for developing a graduated load of the new forged balls before the test could be started was realized, but the time required to obtain a graduated load and the functioning of the various ball sizes was not clear. No advantage could be seen for starting out with a full load of forged balls, as this initial load, at the least, would have to be completely ground out and replaced by daily additions of balls before the graduation could be complete and a normal static condition of ball sizes reached. For a considerable period prior to exhaustion of this initial charge there would be an unusual quantity of worn balls of a certain size, which would not be typical of a normal charge and could influence the characteristics and functioning of the entire load. Therefore, it was decided to start by daily addition of the forged balls to a mill that contained a normal operating load of cast balls. After 88 days of operation the mill was shut down for inspection. By that time the load consisted of a mass of forged balls functioning around a large core of the residual material of the former cast-ball load. This core of cast slugs, locally referred to as "the kidney," was well defined from the forged-ball load.

Jan 1, 1947

By C. Richard Tinsley

INTRODUCTION Project financing has been defined by various authors in this book. It is founded on reliance upon the project's own future cash flows and secondarily on the collateral value of the asset base developed. It is chiefly provided by the banking community who, after an assessment of the risks involved (See Chapter 9) analyse the company's presentation (See Chapter 10) in order to structure a financing package often at a leverage of up to 80% debt. This paper describes the various support and structuring techniques used to delimit and abate the risk components perceived from the banks' experience on project development especially for the period before cash flow is generated (pre completion). ADVANTAGES The advantages of project financing extend beyond the offloading of risks to banks. Among the advantages are the following: 1. Longer loan maturity than conventional, unsecured financing; 2. Loan disbursement and repayment tailored to specific mine; 3. Ability of a joint venture without a "track record" to raise financing; 4. Obtain financing where further direct borrowing (corporate credit) limited by existing indentures; 5. Possibility of arranging finance off-balance sheets; 6. Leverage return on equity; 7. Obtain tax savings with certain project financing structures; and 8. Obtain flexible loan repayment conditions. It can also be said that from the point of view of the company, project financing has certain disadvantages which can be summarised as follows: 1. Perceived higher cost (risk premium) compared with straight corporate credit; 2. More documentation; and 3. Need to negotiate risk-sharing aspects. The main attraction of project financing is the ability to pass certain risks on to the lenders within a flexibly funded transaction. (For a more detailed description of the specific risks see the author's paper at the beginning of Chapter 9. The same risk system is followed here.) The typical project financing has corporate support prior to completion and will usually carefully control the resultant cashflow after completion to ensure the bank does not unnecessarily assume further risks of an equity nature. WITHIN THE COMPANY'S CONTROL Operating Risk This is broken into three interrelated segments for ease of description. Technical Component: This must be specifically addressed if new technology or unproven adaptations are to be used. The banks will have scrutinized the sampling and bulk testing work, perhaps grilling the geological, mineral processing, metallurgical, and mining engineering staff or consultants directly. A number of supports are available such as: TECHNOLOGY GUARANTEE: Either the company itself or the process developer will guarantee that the technology will work for a period of, say, six months to two years after startup. If not, then financial penalties will apply although these are rarely sufficient to repay the loan. Many process developers are weak financially and may have little substance to back such a guarantee.

Jan 1, 1985

By George A. Roberts, Arthur H. Grobe

Tensile, hardness and density properties are presented for a new 18-8 stainless steel powder for the —50, —100, and —140 mesh cuts and also for a prepared blend containing 62 pct —325 mesh powder. The data were obtained for sintering temperatures of 2100° to 2350°F and for compacting pressures of 30 to 50 tons per sq in. A NEED has existed for a stainless steel powder of uniform composition from particle to particle, which could be readily molded and sintered, and which would yield moderate to high strength and ductility when processed into parts by standard powder metallurgy techniques. The literature reveals that previously available powders have been lacking in one or several of these desirable characteristics. Originally attempts to produce stainless steel powder by blending and diffusing elemental metal powders were unsuccessfu1.l A truly alloyed stainless steel powder reported by Dale in 1947 yielded reasonable mechanical properties but required an exceptionally high sintering temperature of 2400°F in pure dry hydrogen.' A later development reported in 1950 by Stern provided a stainless type powder of excellent molding properties' but this powder likewise required sintering at temperatures over 2350 °F and did not yield properties as high as those previously reported. The powder was not truly alloyed prior to sintering. A prealloyed stainless steel powder has been developed that is readily moldable at 30 to 50 tons per sq in., can be sintered at low temperatures (2100" to 2300°F), and which yields compacts with both high strength and ductility. The powder is prepared by the liquid disintegration method," direct from molten metal. Early attempts to make 18 pct Cr, 8 pct Ni stainless steel powder by this method did not produce a usable powder for pressing and sintering because of a preponderance of spherical particles with relatively high, surface oxide content. It was learned that silicon additions to iron alloys, powdered by this process, gave bright powder of low-oxide content and such additions were made to 18-8 stainless with similar results. An attendant change in particle shape occurred reducing the number of spheroids materially. The optimum composition which has been studied and is reported here contains approximately 2.5 pct Si. Silicon additions to stainless steel of the 18-8 type improve the resistance to high-temperature oxidation and promote the formation of ferrite. Corrosion resistance is, however, controlled by chromium content and even though ferrite may exist the corrosion resistance is not generally impaired.4 Experimental Procedure A detailed study of the 18-8, 2.5 pct Si powder has been made in accordance with the following outline: 1. Four blends a. — 50 mesh b. —100 mesh

Jan 1, 1952

By P. W. Bakarian

PREVIOUS determinations1,2,3 of the texture of magnesium and its alloys have shown only slight variations in the principal features of the structure. This investigation presents pole figures for magnesium and two of its alloys, in which differences of some magnitude will be shown. Schmid and Wassermann have considered the texture of magnesium in analogy with that of zinc-also a hexagonal lattice. Whereas the zinc texture is said to be derived primarily through basal slip and secondarily through {102}† twinning, the {102} twinning is considered impossible at the later stages of the rolling in magnesium because of its smaller axial ratio. This state of affairs appears reasonable for pure magnesium, but for alloy textures the situation is more complex. A double orientation similar to that found in zinc has been reported by J. D. Hanawalt,4 and the addition of small amounts of calcium to magnesium and its alloys to produce a double texture was noted by J. C. McDonald.5 The pole figures of this investigation, using an improved X-ray technique, show these various textures. The dependence of directional properties of rolled sheet on modifications of texture is well recognized. Correlation of physical properties with existing texture should be even more significant if quantitative valuations are attached to the various areas on the pole figures. Such a quantitative technique is developed herein. EXPERIMENTAL METHOD Two alloys were used, besides high-purity magnesium. In each case, the metal was extruded to bars 32 in. thick and hot-rolled to a thickness of about 0.075 in. It had been established previously that such a reduction was more than sufficient to eliminate the extruded structure. The rolling proceeded in the same direction as that of the extrusion, without any cross rolling. The temperatures corresponded to normal commercial practice, averaging around 600°F. A unique method of preparing the sample was developed in this work. Instead of using a flat specimen, a stem of the metal was cut out of the rolled sheet in the desired direction (rolling direction) as shown in Fig. 2. From a sheet of 0.075 to 0.065-in. gauge, the stem was cut to a square cross section with a jeweler's saw, rounded with a file and polishing paper to a diameter of 0.060 in, and evenly etched in dilute hydrochloric acid to a diameter of 0.038 ± 0.001 in. In this way, a cylinder of the original sheet with its axis in the desired direction was produced. By using this type of specimen for the pin-point pictures, any variance in the absorption of the diffracted beam at the various angular settings of the specimen, was completely eliminated.*

Jan 1, 1941

By L. H. Levenson, Charles S. Barrett

Numerous determinations1-' of the texture of cold-rolled polycrystalline iron, steel, and ferritic alloys have been made with good agreement among the various observers as to the principal features of the texture common to all, but there has been no adequate explanation of the texture in terms of, the behavior of individual grains. This study presents an explanation of the texture based on the observed rotation of single crystals when deformed in a manner similar to a grain in the aggregate, a method we have found useful in leading to a better understanding of compression and tension textures;8,9,10 in particular, the aim has been to determine experimentally the end orientations of rolled single crystals of iron. These are discussed in connection with the point of view of Weverl and of Boas and Schmid2 that the rolling texture is a simple compression texture (compression normal to the rolling plane) superimposed on a simple tension texture (tension in the rolling direction), and the point of view of Kurdjumow and Sachs3 that three ideal end orientations exist. Experimental Method The single crystals used were in the form of disks cut from a stock of crystalst pre- pared in this laboratory by a strain-anneal method from mild steel decarburized and purified in hydrogen, the details of which have been described elsewhere.11 To ensure maximum elongation in the rolling direction and to simulate approximately the rolling of a grain in an aggregate, lateral flow was constrained by placing each crystal in a close-fitting hole in a plate of copper or iron of equal thickness and rolling the plate and crystal together. About one-third of the crystals were rolled in a "sandwich" pack by adding top and bottom cover plates. The two-high mill used had 8-in. diameter rolls and was operated at 15 r.p.m. The initial diameter of the crystals was between 1/4 and 1/2 in., while the starting thickness varied from 0.031 to 0.376 in. With a single exception, which was rolled 63 per cent, the crystals were rolled to a final reduction in thickness of 82 to 97 per cent in 20 to 40 passes, alternately changing the sense of the rolling direction. The initial orientations of the crystals were determined either by X-ray back reflection12 or by optical reflectionl3 from etch pits; the mean final orientations were determined by X-ray Laue photograms using molybdenum, silver, or tungsten radiations, by optical reflection, and in many cases by both methods. Results Twenty-six crystals were studied. It was found that there were three typical be-

Jan 1, 1941

By Charles S. Barrett, L. H. Levenson

Numerous determinations1-' of the texture of cold-rolled polycrystalline iron, steel, and ferritic alloys have been made with good agreement among the various observers as to the principal features of the texture common to all, but there has been no adequate explanation of the texture in terms of, the behavior of individual grains. This study presents an explanation of the texture based on the observed rotation of single crystals when deformed in a manner similar to a grain in the aggregate, a method we have found useful in leading to a better understanding of compression and tension textures;8,9,10 in particular, the aim has been to determine experimentally the end orientations of rolled single crystals of iron. These are discussed in connection with the point of view of Weverl and of Boas and Schmid2 that the rolling texture is a simple compression texture (compression normal to the rolling plane) superimposed on a simple tension texture (tension in the rolling direction), and the point of view of Kurdjumow and Sachs3 that three ideal end orientations exist. Experimental Method The single crystals used were in the form of disks cut from a stock of crystalst pre- pared in this laboratory by a strain-anneal method from mild steel decarburized and purified in hydrogen, the details of which have been described elsewhere.11 To ensure maximum elongation in the rolling direction and to simulate approximately the rolling of a grain in an aggregate, lateral flow was constrained by placing each crystal in a close-fitting hole in a plate of copper or iron of equal thickness and rolling the plate and crystal together. About one-third of the crystals were rolled in a "sandwich" pack by adding top and bottom cover plates. The two-high mill used had 8-in. diameter rolls and was operated at 15 r.p.m. The initial diameter of the crystals was between 1/4 and 1/2 in., while the starting thickness varied from 0.031 to 0.376 in. With a single exception, which was rolled 63 per cent, the crystals were rolled to a final reduction in thickness of 82 to 97 per cent in 20 to 40 passes, alternately changing the sense of the rolling direction. The initial orientations of the crystals were determined either by X-ray back reflection12 or by optical reflectionl3 from etch pits; the mean final orientations were determined by X-ray Laue photograms using molybdenum, silver, or tungsten radiations, by optical reflection, and in many cases by both methods. Results Twenty-six crystals were studied. It was found that there were three typical be-

Jan 1, 1941

By E. Douglas Sethness

The uranium industry is booming. In Texas alone, there are about 22 different companies with active exploration programs. Twelve solution mines have been permitted; three surface mines have been authorized; and two mills are currently in operation. However, the industry also has a problem, and that is the disposal of radioactive wastes. Over the past several years, stories concerning nuclear wastes have appeared frequently in the news. One of the most frequently cited cases occurred in Grand Junction, Colorado. In 1966, after ten years of investigations, the U. S. Public Health Service (PHS) discovered that tailings from a uranium mill were being used as fill material and aggregate for local construction purposes. It was estimated that between 150,000 and 200,000 tons of material had been removed and used under streets, driveways, swimming pools, and sewer lines. In addition, tailings had been used under concrete slabs and around foundations of occupiable structures. Further studies prompted the Surgeon General to warn that the risk of leukemia and lung cancer could be doubled at the measured radiation levels. More recently, the L. B. Foster Company discovered that its building site in Washington, West Virginia, was radioactive. While digging a foundation, the ground erupted and a ball of fire 30 feet high shot out. Evidently, the dirt was laced with radioactive thorium and zirconium, a potentially explosive mixture contained in a Nigerian sand which had been used by the previous site owners in the manufacture of nuclear fuel rods. Just this month we have read about legal suits to stop exploration for a nuclear waste disposal site in Randall County, Texas. The U. S. Department of Energy is trying to locate a deep underground nuclear waste depository for final burial of over 76 million gallons of high-level wastes. The problem is acute, the wastes are accumulating at a rate of about 300,000 gallons per year. Nor do these numbers include the spent fuel elements from nuclear power plants that are in temporary storage facilities. Fortunately, public awareness of these and other related issues is high. Unfortunately, the differences in the waste products from the nuclear fuel cycle are not always apparent to the general public. There are two distinct types of radioactive wastes: "high-level", which consist of spent fuel or wastes from the reprocessing of spent fuel; and "low-level", which, in general, are by-product wastes. There are numerous non-technical definitions that can be applied to help the layman differentiate between high-level and low-level wastes. For this latter purpose, it is best to think of them in terms of what we can see and feel. In general, high-level wastes are physically hot and can cause acute radiation sickness in a short period of time. Low-level wastes are not hot, but may cause chronic health effects after long exposure. The wastes which we are concerned with in the uranium mining and milling industry are low-level wastes. As recently as ten years ago, there were very few controls or regulations governing tailings disposal methods. At the same time, mine reclamation was not enforced through either state or Federal laws and the long-term viability of abandoned tailings ponds was not assured. The regulatory climate has changed significantly in the last decade, however. The low-level radioactive wastes generated by uranium mining and milling are generally contained in a tailings pond. Approximately 85-97% of the total radioactivity contained in uranium ore is present in the mill waste that goes to such tailings ponds. The isotope Radium-226 is probably the most potentially harmful radioactive parameter in the ponds. Radium emits gamma radiation and is also an alpha particle emitter. Because gamma radiation is very penetrating, it presents a potential health problem when a source is located external to the body. Gamma radiation will go through the body, causing damage to each cell encountered on the way. Although alpha particles have very little penetration capability, they can cause extensive cell damage. For this reason, alpha particles are a problem after inhalation or ingestion. Radium creates a health hazard by both of these mechanisms. Radium decays to radon gas which can be inhaled and serve as an alpha particle emitter. Additionally, radium is very soluble and readily enters the natural hydrologic cycle if allowed to leach from a tailings pond. With a half-life of 1620 years, radium has plenty of time to be taken into the food chain and end up in our bodies, emitting alpha particles. Because the potential health problems are better understood today than ten years ago, and because the Nuclear Regulatory Commission (NRC) has developed increasingly stringent government regulations, the uranium mining industry applies a high level of technology to the disposal of nuclear wastes. In most cases, low-level radioactive wastes are disposed of at or near the site where they are produced. There are six commercial burial grounds for low-level wastes, but it would not be economical to ship all mine or milling wastes to these sites. The on-site disposal methods most often used are ponding

Jan 1, 1979

By Walter Crafts, C. M. Offenhauer

IN a previous study' of the carbide phase of chromium steels, it was shown that chromium carbide (Cr7C3) is a more stable carbide than cementite (Fe3C) at tempering temperatures above about 500°C. in quenched steels containing from about r1to 7.5 per cent chromium. Below this temperature the carbide phase was found to consist of cementite in all of the steels examined. In the present investigation it was found that a similar change in the carbide phase takes place in molybdenum and in chromium-molybdenum steels at approximately the same tempering temperatures. The approximate ranges of occurrence of the alloy carbides Cr7C3, Cr4C, and Mo2C have been determined in steels containing up to 5 per cent chromium, 1 per cent molybdenum and 0.60 per cent carbon. The experimental steels were made in small high-frequency furnaces from Armco ingot iron, low-carbon ferrochromium and calcium molybdate and were poured into 2-in. ingots, which were forged to 1-in. round bars. Sections of the bars were heated to the carbide-solution temperature and cooled rapidly to induce austenite transformation at relatively low temperatures. The steels were subject to some segregation, therefore comparatively high temperatures were necessary to dissolve all of the carbides. The microstructure of the quenched specimens was examined to verify the belief that the carbides had been dissolved, and also to determine the nature of the nonmartensitic structure of incompletely hardened steels. Ferrite and pseudomartensite were present in some of the specimens, but all specimens were quenched at such a rate that pearlite was not formed, in order to ensure the formation of the carbides at temperatures lower than about 500°C. It had been found previously' that in chromium steels the chromium carbide could be produced by transformation of austenite in the pearlite range above 500°C., and, in order to start the tempering treatment with the carbide present as cementite, the absence of pearlite was considered to be essential. The specimens were tempered for 64 hr. to permit the formation of the stable carbide. After tempering, the specimens were submitted to microscopic and X-ray examination. The carbides were separated either electrolytically in a 5 per cent HCl electrolyte or with an acidified copper-potassium chloride solution. The crystal structure of the carbide phases was determined by the Debye-Scherrer type of X-ray patterns. The X-ray patterns obtained from specimens near the boundaries between two types of carbides were generally weak and diffuse as compared with the other patterns. For this reason, and because the more soluble components may be masked, the boundaries between the phases could not be determined precisely. However, the X-ray analysis of electrolytically separated carbides has given consistently reproducible results

Jan 1, 1943

By Alfred Boyles

The experiments described below are a continuation of work on the graphitization of cast iron conducted as part of the program of fundamental research at Battelle Memorial Institute. In previous work it was found that melting in hydrogen increased the amount of eutectic carbide appearing in certain alloys and in others produced marked changes in the size of the graphite flakes. These effects seemed to be closely related to the amount of sulphur and manganese present in the material studied. Iron-carbon-silicon alloys did not respond to hydrogen in the same manner as cast iron and there was no evidence to show that hydrogen was able to stabilize the carbide in such alloys unless additions of sulphur were made. The two alloys examined contained about 2 per cent of silicon. The present paper is concerned with alloys of lower silicon content and describes the structural changes produced by melting in various atmospheres at pressurcs ranging from a partial vacuum up to 300 lb. per sq. inch. Experimental Method The experimental technique was similar to that previously employed, so that direct comparison might be made with earlier work. Alloys were

Jan 1, 1939

By Alfred Boyles

The experiments described below are a continuation of work on the graphitization of cast iron conducted as part of the program of fundamental research at Battelle Memorial Institute. In previous work it was found that melting in hydrogen increased the amount of eutectic carbide appearing in certain alloys and in others produced marked changes in the size of the graphite flakes. These effects seemed to be closely related to the amount of sulphur and manganese present in the material studied. Iron-carbon-silicon alloys did not respond to hydrogen in the same manner as cast iron and there was no evidence to show that hydrogen was able to stabilize the carbide in such alloys unless additions of sulphur were made. The two alloys examined contained about 2 per cent of silicon. The present paper is concerned with alloys of lower silicon content and describes the structural changes produced by melting in various atmospheres at pressurcs ranging from a partial vacuum up to 300 lb. per sq. inch. Experimental Method The experimental technique was similar to that previously employed, so that direct comparison might be made with earlier work. Alloys were

Jan 1, 1939

Large equipment for quarrying and for the various stages of crushing and grinding are the trend for new and existing operations. Included are large size haulage units, fast drilling equipment, impact crushers and large mills capable of high tonnage rates for metallic and nonmetallic minerals. In grinding, large-size mills continue as the choice in industries where grinding constitutes a major portion of the processing. Typical is a 10 x 34 ft two-compartment mill in closed circuit with an 18-ft diam air separator for cement. Other cement mill sizes are in the order of 12 x 36 ft, 2500 hp; 10 x 37 ft ball and rod (wet), 1400 hp; 13 x 36 ft, 3000 hp; 13 x 40 ft (wet), 3000 hp. Several installations of combination ball and rod mills have been made, and one of the largest is 12 x 45 ft. In connection with the wet grinding of raw materials, the DSM 270-degree screen has been installed in several cement plants with the intention of producing a material finer than 50 mesh. In ore processing, large autogenous mills (121/2 x 20 ft) are being installed.

Jan 2, 1964

By W. E. Hall

The Darwin mining district of California, 160 miles north of Los Angeles, has yielded an estimated $45 million in lead, silver, zinc, and copper since 1875. The deposits are in silicated limestone of Penn-sylvanian and Permian age close to a stock of horn-blende-biotite-quartz monzonite. Steep left-lateral strike-slip faults that strike N 50o 70o E served as feeder channels for the ore solutions, and minor folds, favorable beds, and faults close to the feeder channels localized the orebodies. The orebodies are irregular, steep, pipe-like replacements that cut across bedding, bedding replacements, and fissure fillings in a calc-silicate country rock consisting mainly of wollastonite, grossularite-andradite garnet, idocrase, and diop-side. Near surface the ore consists of a soft, crumbly mass of cerussite, hemimorphite, jarosite, limonite, and some secondary copper minerals. The primary ore consists of galena, sphalerite, pyrite, pyrrhotite, and chalcopyrite with smaller amounts of andorite, matildite, scheelite, tetrahedrite, and a high silver-lead-bismuth-selenium-bearing galena. Preliminary X-ray fluorescence study of the primary ore has indicated unusually high concentrations of bismuth, selenium, and silver in galena from certain parts of the mine. One phase of this study concerns the distribution of these elements. Twenty-two samples of galena and thirty of sphal- erite have been purified and analyzed or are being analyzed. The selenium in 21 samples of galena ranges from 0.0045 to 2.11 pct; four contain more than 1 pct. These four samples were collected from the Essex orebody, where most of the selenium is concentrated. Three are from high-grade, near-surface ore, and one is from the 700 level. Four of the seven analyzed galena samples from the Thompson workings immediately north of the Essex orebody contain 0.29 to 0.74 pct Se; the others show less than 0.05 pct. Galena from the Defiance workings about 2000 ft south of the Essex has less than 0.03 pct Se. The selenium-bearing galena also contains concentrations of bismuth and silver. Preliminary spectographic analyses indicate as much as 3 pct Bi and more than 1 pct Ag in galena with high selenium content. Antimony is the only other minor element present in galena in amounts greater than 0.02 pct; most galena contains between 0.1 and 0.2 pct and the maximum is 0.72 pct. A second phase of the study is an attempt to determine the temperature range of the formation of the ore. The iron content of sphalerite deposited in equilibrium with pyrrhotite is being used as a geologic thermometer as it is a function of the temperature of deposition of the ore.' Preliminary results indicate temperature range from 430°C in replacement orebodies in the Defiance workings to 180°C in shallow fissure fillings.

Jan 1, 1960

By Albert F. J. Bordeaux

INTRODUCTION. THE following general survey of the character and present condition of the silver-mines of Mexico, though not offered as a detailed treatise, has been prepared with care, is believed to he generally up-to-date, and is offered as a summary which may be found useful by mining engineers. Mexico has ranked for several centuries first among the silver-producing countries. It is now second as copper-producer, and seventh as gold-producer, with a fast-increasing output every year. The total amount of capital invested in its mining- and smelting-works is more than $250,000,000, while the yearly metallic production is $160,000,000. There are more than 20,000 recorded mining properties. The silver-mines of Mexico represent deposits of three principal classes: (1) veins and stock-works in eruptive rocks; (2) fissure-veins crossing the crystalline schists, granite, diorite, etc. ; and (3) veins and replacement-deposits in limestone. The true fissure-veins in some districts contain (in value) more gold than silver. In other deposits, the silver is decreasing in depth, leaving galena and zinc-blende as the principal ores.

Jan 9, 1908

By Lennart Wellenius

LAISVALL - ONE OF THE BOLIDEN MINES Until the Aitik Mine (2,000,000 mtpy) started in 1968 the Laisvall Mine (1,200,000 mtpy) was the biggest unit of about 20 producing mines, belonging to the Boliden Company Ltd. As shown in Fig. 1 the mines are spread over a wide area. Most of them are small complex ore mines (50,000 - 300,000 mtpy), for which reason the total mine production does not exceed 5,500,000 mtpy. When our company was invited to present a paper, Laisvall was chosen because it is the biggest European galena-mine in production today. The Laisvall Mine is situated in the county of Norrbotten, Sweden, at the eastern border of the Caledonian mountain-chain. The road distance to Slagnäs, the nearest railway station, is 60 miles. The railway between Slagnäs and BOLIDEN's smelting plant, the Rönnskär Works, is 140 miles long. PRODUCTION STARTED IN 1943 In the Autumn of 1938 a boulder was discovered, consisting of sandstone with galena, on the shore of the Lake Stor-Laisan. The boulder came from an Eocambrian sandstone formation, which is overlain by Cambrian-Silurian rock.

Jan 1, 1970

By O. B. Bucklen, J. P. Matoney, P. G. Meikle, R. L. Terry

INTRODUCTION The trend in the coal industry for many years has been to make every- thing "bigger and faster." This also holds true for loading and storage facilities. Where once a 1 million st per yr (910,000 mt/yr) mine was rare, 3-5 million st per yr (2,730,000-4,550,000 mt/yr) mines are not uncommon today in the bituminous industry. This great increase in production, made possible through mechanization, requires greatly improved loading facilities just to get the coal away from the mine. Also, there is a trend to- ward combining production from more than one mine and cleaning it in a single preparation plant. In addition to increased loading capacity, this frequently necessitates raw-coal storage and blending facilities. The development of the unit train with its attendant economics has contributed more than any other factor to the widespread construction of storage and high-speed loading facilities. In the export market, the construction of ever larger coal colliers has added to the need for faster and larger loading facilities. In this climate of rapidly changing technology, any written ac-

Jan 1, 1979

By W. E. Foreman, James A. Redding

INTRODUCTION Screening is a mechanical process that separates particles by size. Modern coal preparation is made possible by efficient sizing. Sizing of coal is defined in terms of the opening through which an individual particle will or will not pass, or as two openings, the larger of which passes the particle and the smaller of which retains the particle. Screening accomplishes sizing by passing coal of a wide range of sizes through a series of screens, each of a decreasing size. The individual screen discharges are then directed into different screen products for subsequent sale or processing. Screening in modern preparation plants is done for the following reasons: 1 ) Removal of fine coal before a size reduction step. 2) Size separation of the coal into a coarse and fine fraction for sized products or processing in coarse and fine coal circuits. 3) Recovery of heavy media from heavy media coal cleaning circuits. 4) Dewatering. 5) Rock and trash removal by handpicking on a screen. Fig. 8-1 is a generalized flowsheet of a plant showing the use of screens in the coal preparation process. Table 8-1 shows the vibrating screen equipment utilized in Fig. 8-1

Jan 1, 1979