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By Edward B. Durham

At the Richmond mecting in February, 1901, I presented for discussion the proposed rail-specifications of Committee No. 1 of the American Section of the International Association for Testing Materials, with a review of their work, and other matters relating to the heat-treatment and chemical composition of steel (Trans., xxxi., 455). In order to bring out the views of our members on the important points referred to, I offered for discussion the following questions and comments (trans., xxxi., 972): " 1. The advantages to be gained by using more metal in the flanges and web of the heavier sections of rails.

Jan 1, 1903

By Philip W. Henry

There is a difference of opinion among engineers on the subject of depreciation in general, and still more on its application to any given case The committee which was appointed by the American Society of Civil Engineers to make a report on Valuation of Public Utilities, states "There is no subject connected with valuation about which there are more diverse views than those relating to depreciation;" and in the discussion which took place upon the presentation of the preliminary report of this committee in January, 1914, that part of the report dealing with depreciation called forth more discussion than any other. From this discussion it was evident that the word "depreciation" was not always used in the same sense, so it is quite important that some definition be adopted. As this paper relates to the proper method of treating depreciation by corporations or individuals operating oil properties, both of which must make a return of annual net income to the government in connection with the income tax, it will be in order to adopt the definition of the Internal Revenue Bureau. Under date of Mar. 29, 1910, the Commissioner of Internal Revenue of the Treasury Department stated "Deduction on account of depreciation of property must be based on lifetime of property, its cost, value and use." Again, on Form 1035, "Return of Annual Net Income" for miscellaneous corporations (Section 2, Act of Congress approved Oct. 3, 1913), a deduction from gross income is allowed for "Total amount of depreciation for the year." This item is defined more particularly as follows: "The amount claimed under Item No. 5 (b) for depreciation should be such an amount as measures the loss which the corporation actually sustains during the year in the value of buildings, machinery, and such other property as is subject to depreciation on account of wear and tear, exhaustion, or obsolescence. The amount taken credit for on this account in order to be allowable should be so entered on the books as to constitute a liability against the assets of the corporation. The amount claimed under this item should not cover losses in the value of stocks and bonds. Decrease in the book value of securities owned, so far as such decrease represents a decline in the actual value of such securities, should be deducted under item 5 (a) of the return. "Where depreciation of physical property is made good by renewals, replacements,

Jan 1, 1915

By Ellis H. Gates

BENEFICIATION of the manganese oxide ores at Three Kids Mine near Henderson, Nev., has evolved over a period of years. Commercial application of the process is on a secure basis, and an effective working hypothesis has been formulated to help solve the problems encountered in flotation. An outstanding difference between the Manganese Inc. operation and other flotation processes is the necessity for high recovery of slime mineral values. Part of the manganese minerals in the Three Kids ore deposit occurs as primary slime or as minerals that are easily slimed by grinding. High recovery of this slime portion is essential to an economic overall manganese recovery. The aim of much of the research at Manganese Inc. was to recover these slimed and finely ground minerals by flotation. Early Agglomeration Processes: Tunbridge' in 1880, followed by Everson2 in 1886 and Cattermole" in 1904, discovered that under certain pulp conditions precious metals and sulfide minerals could be agglomerated if enough soap or oil were mixed into the pulp slurry. These agglomerated particles, Which under ideal conditions ranged from 1/16 to 1/4 in. diam, were recovered by screening, tabling, and hydraulic classification. Even flotation by air occluded in the agglomkrate particles by vigorous stirring was suggested. Cattermole nearly succeeded in getting his process into commercial operation, but it was finally abandoned, presumably because he could nat compete with the low reagent requirements of the new agitation froth process. Since Cattermole's time Christensen in 1921, Borcherdt4 in 1926, and Chapman and Littleford- in 1934 have attempted to apply the principles of agglomeration. In the light of current research at Manganese Inc. on conditioning of pulp at high energy input levels, it is interesting to note the variety of equipment such as tumbling mills, mixing drums, and shakers described by Borcherdt in his patent to produce the agglomerated particles. Ore Characteristics: Manganese in the Three Kids ore occurs as a 'complex mixture of manganese oxides commonly Known as wad. The manganese minerals have Been identified as psilorflelane, hollandite, coronadite, and pyrolusite with minor amdunts of rhodonite and rhodochrbsite. The ppincipal gangue miherals are quaktz, opal, kaolinite, montmoril-lonite, calcite, gypsum, fcelesitite, and barite. Most of the mangandse ore minerals are either porous and sponge-like in structure or soft clay-like masses composed of fine grained mineral particles approaching the size range of colloids. Consequently the ore minerals from the Three Kids mine have relatively large surface areas and high void content per unit weight. The specific area of various manganese ore samples measured at the California Research Corp. laboratories ranged from 32 to 58 sq m per g. The great surface area and high void content explains in part the unusually high reagent consumption required in the treatment of this ore. Current Milling Practice: At Henderson6 he ore is crushed to —1 in. and ground with rod mills in a closed circuit with spiral rake classifiers and cyclones. Sulfur dioxide solution is added to the pulp as it leaves the grinding circuit, immediately followed by addition of a reagent mixture of soap, oil, and wetting agent in the form of an emulsion. The reagents are initially well mixed into the pulp during their passage through centrifugal pumps to the conditioning equipment. All reagents are added prior to conditioning. The intense conditioning or mechanical stirring of the pulp, in which large amounts of power are consumed, is the principal difference between the flowsheet used at Manganese Inc. and a standard flotation circuit. As much as 38 kw-hr of energy per dry short ton of ore in a pulp containing 18 to 23 pct solids is used for conditioning. After conditioning, the pulp passes to a flotation circuit that consists of one roughing stage, a scav-

Jan 1, 1958

By W. F. Kieschnick, and W. J. McGuire&apos, Eugene Harrison

This paper concerns the induction and extension of fractures into rock formations as involved in drilling, completing, and production stimulating operations on wells. Conclusions concerning formation breakdown are derived from (I) a review and extension of published analyses relating to mechanical theories of rock stress and the state of stress in the earth's crust and (2) a correlation of field data from fracturing operations. Conclusions concerning the mechanics of fracture extension, which indicate the relationship between fracture dimensions and rock properties, depth, and volume of injected fluid, are tentative and largely establish limits of relationships. These conclusions are derived from stress calculations, limited field data, and laboratory experimental studies. The experimental work involves the study of the stresses at the fracture boundaries and the geometry of pressurized fractures by means of photo-elastic modeling methods. Results of this investigation indicate that a large majority of pressure induced wellbore fractures are vertical, particularly in deeper wells; and variations in the pressures necessary to create and extend fractures can be explained largely on a basis of established rock properties. It is also shown that variations due to tectonic forces should usually be expected to be slight. Other results indicate that during that extension of fractures rather large fracture volumes are temporarily created by the parting of the formation. The purpose of this paper is to present the results of calculations and laboratory experiments concerning the mechanics of fracture induction and extension with a view to broadening existing knowledge relating to these phenomena. It is believed that continued progress in developing knowledge of this type is important to the further development of techniques for drilling and completing wells. When a wellbore is subjected to sufficiently high pressure, fracturing of the formation surrounding the well occurs. This phenomenon is a common cause of lost circulation during drilling operations, is a deliberately planned factor in most squeeze cementing operations, and is the controlling factor in many productivity stimulation operations. The exploitation of formation fracturing in recent years has been particularly impressive in sand-viscous fluid well stimulation treatments which were first introduced to the industry by the Stanolind research group1,2,3 several years ago. The borehole fracturing phenomenon has been analyzed mathematically by application of the principles defining the elastic and inelastic behavior of thick-walled cylinders of homogeneous, isotropic, and impermeable material.1,5,6 This basic argument is re-presented in this paper along with new considerations which strengthen it. In addition to these arguments relating to fracture induction some calculations and experimental results relating to the mechanics of fracture extension are presented. Since the nature of the pressure induced fracture at distances from the well is of great importance in considerations of well stimulation procedures there appears to be a need for an integrated approach which deals with the mechanics of fracture extension as well as with wellbore rupture.

Jan 1, 1955

By R. F. Janke, J. F. Bossler

Introduction The commercial use of ion exchange resins to recover uranium evolved in the decade following 1950 when significant efforts were made to recover this vital element economically and efficiently from low grade ores. It was discovered that in both acid and alkaline systems, uranium exists as an anionic complex. This led to the application of anion exchange resins to concentrate and purify the uranium product. While this is the simplest statement of what an ion exchange resin does in the recovery of uranium, it should not be forgotten that the industry is continuing to evolve new methods in mining and processing which affect ion exchange operation. The inception of solution mining, the use of expanded bed ion exchange columns and the development of continuous ion exchange systems represent several examples of this continuing evolution in the industry. Throughout all of the changes in equipment designs and variations in solution compositions, the purpose of the ion exchange resin remains the same - to concentrate and purify uranium. Ion Exchange Resins When evaluating ion exchange resins for a uranium recovery project, the task that faces the chemist or process metallurgist is an assessment of the efficiency and, therefore, the economics of the various candidate resins for a given project. This assessment should include consideration of both physical factors such as particle size, density, hydraulic expansion and pressure drop as well as chemical factors such as the uranium loading and elution characteristics. For any given application, the differences in the physical and chemical behavior of ion exchange resins are a function of the conditions used to manufacture those resins. Therefore, an understanding of the way in which manufacturing variables affect resin behavior is important. Since most uranium recovery projects are concerned with recovery of the anionic complexes of uranium using anion exchange resins, this discussion will be restricted solely to the family of anion resins. The majority of the ion exchange resins used in commercial recovery operations look similar to the resins shown in Figure #1. [ ] Most of these materials are based on a styrene- divinylbenzene copolymer. The liquid monomers, styrene and divinylbenzene, are charged to a co-polymer reactor containing water. The water insoluble monomers are then mixed with the water to suspend oil-like droplets of the monomer mixture in. the water. The entire reaction mixture is then heated causing the monomers to polymerize into sol id thermoset plastic spheres. The spherical nature of the particle, the particle size distribution and the primary crosslinking level are fixed in this step in the production process. Figure 2 is a simplified illustration of this step of the process. [ ]

Jan 1, 1979

By H. Macpherson

Durhm has a well earned reputation for supplying some of the finest coking caals in the world. The caals, in general, vary in rank from 301 to 501/2. Durham has traditionally produced foundry coke for the major proportion of the foundries in Great Britah. Durham coals can be described as clean, soft coals, which yielded, with the old hand methods of coal getting, low ash small coal because the extraneous shale was normally harder than the coal and occurred in larger pieces. Under these conditions, with a plentiful supply of cheap labor, it used to be sufficient, at most collieries, to hand clean the large sizes of coal which could then be remixed with the untreated small coal to produce carbonization coals, ready for the coke ovens, at under 5 pct ash. With the introduction of mechanical methods of coal mining, the coal gradually required more cleaning. The first method adopted to resolve this problem was the introduction of pneumatic dry cleaners for the small coal. Although such machines had little effect on the fine coal (say below 1/8 in.), they could clean the intermediate sizes of coal. This, coupled with hand cleaning the larger sizes above 1-1/2or 2 in., resulted in a combined run-of-mine mixture below 6 pct ash, capable of maintaining the quality and reputation of the cokes produced. In more recent years, the intensification of mechanization and power loading, coupled with gradual exhaustion of the cleaner seams, has created the need for a more complete method of coal cleaning. This particularly applies in the fine sizes (say below 1/50 in.) which normally vary, under present day conditions, between 15 and 35 pct ash and are much too dirty to be included in the raw state in a carbonization mixture. This pronounced change has been accelerated because legislation controlling the dust conditions of mine airs for the prevention of pneumoconiosis has resulted in tk~e extensive use of water underground and a consequent increase in moisture content of the run-of-milie output. The presence of damp fine coal decreast the efficiency of prescreening and dry cleaners, so that this type of preparation for low ash coking coals is decreasing, although it is still used satisfactorily for industrial coals in the medium ash range. Table I shows the gradual increase in mechanization, the reduction in manpower, the increased use of explosives per ton of coal extraction, and the increase in the proportion of coal cleaned by mechanical means in Great Britain. Although similar figures are not available for Durham Div. until after the date of nationalization of the coal industry, it is probable that the increase in mechanical cleaning, particularly by dry cleaners, was more marked in the Durham collieries than elsewhere in the country. As dry clealiers were replaced by coal washeries in the Gritish ccal industry, no special attempts were made to recover the slurry, with the result that large outflows of dirty water were allowed, deposits of slurry came in to lagoons and neighboring streams, and a proportion of fine material was lost from the coking coal. This position, coupled with the higher moisture of the washed coking coal, resulted in adverse effects on coke oven throughputs and coke quality. It is now realized that the natural coal fines are an essential ingredient of coking coals in obtaining the correct coke structure in metallurgical cokes. This, together with economic pressure, led to the introduction of flocculation and filtration plants for the recovery of slurries, and later, when the ash contelits of the filter cakes were too high, to the introduction of froth flotation equipment. After this position had been reached, the tailings from the froth flotation pIants were, in many cases, still allowed to constitute an undesirable effluent. Recent legislation on river pollution has changed this picture; it is now necessary to provide a circuit which is completely closed so far as solids are concerned. The gradual increase in the coal cleaned by wet methods and froth flotation in Durham Div, is shown in Table 11. It is now an accepted feature of new Washeries that. froth flotation should be an integral part of the washing process from the initial installation of the plant.

Jan 1, 1962

By O. N. Carlson, H. A. Wilhelm, J. M. Dickinson

On the basis of microscopic studies, melting-point observations, and X-ray analyses, a phase diagram is proposed for the Cb-V system. A complete series of solid solutions is formed with a minimum in the solidus at 1810°C near 35 wt pct Cb. No compounds or intermediate phases were found in the system above 650°C. THERE is an ever increasing need for better structural metals and alloys for use in nuclear reactors. In addition to the normal properties of engineering structural materials, such as high temperature strength, resistance to corrosion, and fabric-abil~ty, the nuclear properties of the material must be considered. In a nuclear reactor it is important to conserve neutrons, so a material which removes these neutrons from the reaction excessively is considered to have unfavorable nuclear properties. In nuclear-reactor design the engineer must have nuclear as well as other data available on alloys in order to make a wise selection of materials. Due to the fact that many of the common structural materials have undesirable nuclear properties, it is vital that new alloys of metals having more favorable nuclear properties be investigated. Columbium and vanadium are both high melting metals, both exhibit resistance to chemical attack, and no great difficulty is encountered in fabricating them into desired shapes. With proper treatment both metals can be cold rolled extensively without failure. In addition they have desirable nuclear properties for certain types of reactors. Therefore, the alloys of columbium and vanadium should be of interest in the atomic energy program. Since an alloy-development program is enhanced by a knowledge of the phase equilibria of the components, this investigation was undertaken to establish the phase diagram for the Cb-V system. According to the Hume-Rothery rules for alloying,' the chemical similarity, crystal structure, and atomic-size factor are favorable for a complete series of solid solutions for this system. Both elements are in the same family of group V of the periodic table and thus are quite similar chemically. The crystal structures of columbium and vanadium are compatible for extensive solid solubility, since both have body-centered-cubic structures. The atomic diameters of columbium and vanadium are 2.85 and 2.62Å, respectively. This difference of slightly more than 8 pct is well within the 15 pct maximum difference allowed for extensive solid solubility. Experimental Procedures Source of Materials: Columbium powder and sheet trimmings were obtained from the Fansteel Metallurgical Corp. According to the manufacturer the metal contains less than 1 pct impurity. An analysis of the metal showed approximately 1800 ppm C in the powder while the sheet trimmings contained less than 500 ppm C. Spectrographic analysis showed minor amounts of Ca, Cr, Fe, Mn, Si, Ti, V, and Zr in both forms of the columbium. No commercial source of vanadium having the ductility and purity desired was available to the authors at the beginning of this investigation. As a result, all of the vanadium used in this study was prepared by the bomb reduction of vanadium pen-toxide with calcium employing the method reported by Long.' Yields of massive vanadium normally were about 80 pct. Chemical analysis of the vanadium prepared in this manner showed the presence of 200 to 500 ppm N and 800 to 1000 ppm C. Minor amounts of Ca, Fe, Mn, Si, Zr, Cr, and Cb were detected by spectrographic analysis. This vanadium metal was ductile and was cold rolled into 5 mil sheet. Annealing was not necessary during this rolling and the metal retained its cold-rolling characteristic after are-melting. Preparation of Alloys: The Cb-V alloys were prepared by melting pieces of vanadium sheet togethel-with columbium in the form of sheet or pellets of powder. The melting was carried out under argon in conventional arc-melting equipment employing a tungsten electrode and a water-cooled copper crucible. Each alloy was remelted three or four times, inverting the alloy after each melting in order to assure complete mixing. Alloys normally were obtained as round flat disks, weighed approximately 70 grams, and had roughly the shape of a disk 1 1/2 in. in diameter and 1/4 in. thick. Half of each alloy

Jan 1, 1955

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By William P. Jones

THE consumption of sulphuric acid, one of the most important commodities in our modern industrial world, is often used as a barometer for industrial activity. The economics of acid manufacture are largely dependent upon the location of the place of consumption and the availability of raw materials in that locality. Sulphuric acid is made from SO,, oxygen from the air and water. Therefore the sulphur dioxide is the only raw material to be considered in an economic study. SO, can be obtained from almost any material containing inorganic sulphur, such as elemental sulphur, pyrites, coal, sour gas and oil, metallurgical gases, waste gases, or gypsum and anhydrite. Many tons of acid can also be reclaimed by the recovery and concentration of spent acids. The aim of this paper is to present a guide to the economic aspects to be considered when the installation of an acid plant is contemplated. It must be remembered that 1 ton of elemental sulphur produces 3 tons of sulphuric acid and that the shipping of sulphuric acid by tank car is very costly. The size of the plant must also be given careful consideration. For instance, operation of a plant producing 5 tons of acid per day might be warranted in Brazil or Pakistan, whereas economics usually favor buying quantities up to 50 tons per day for use within the United States. Elemental sulphur, when available at the low price of 1M4 per lb delivered at an acid plant, has always been the raw material most frequently used for sulphuric acid. All conditions favor its use at this price. The so-called sulphur shortage has been the subject of so many technical papers, magazine articles, and newspaper items during the past y6ar that it hardly seems necessary to mention it again, but a very brief review of the matter will serve as a foundation for the discussion that follows. There is no shortage of sulphur. Only a shortage of low-cost Frasch-mined brimstone exists today. Other more expensive sulphur-bearing materials are plentiful, both in the United States and abroad. The low cost of Frasch-mined brimstone has discouraged the development of higher cost sources. However, the approaching depletion of Gulf Coast dome deposits and the greatly increased demand for sulphur here and abroad have made it necessary for industry to prepare for conversion to utilize sulphur in other forms. For future planning this situation must be considered permanent and not temporary. This conclusion is based on the fact that although sulphur demand will continue to rise, the production of Frasch-mined sulphur probably will not increase greatly beyond its present level of about 5,000,000 long tons per year. The International Materials Conference in Washington estimates 1952 requirements of the free world at nearly 7 million long tons; and the Defense Production Administration has recently set a new goal for 8,400,000 long tons annual domestic production by 1955. The total sulphur equivalent produced in this country in 1950 was 6 million tons. What, then, are the alternatives for the manufacture of the vital chemical, sulphuric acid? Today about 85 pct of this country's sulphur, and nearly 50 pct of the world supply, comes from our Gulf Coast salt domes and is extracted from the earth by Frasch's hot water process. The Gulf Coast salt dome deposits have been the most important known natural deposits in the world, producing 90 million tons of sulphur during the past 50 years. However, at the present rate of extraction these deposits cannot be expected to last indefinitely. Pyrites Pyrites are, and have been for many years, the source of more than 50 pct of the world's sulphur requirements. The principal use, of course, is in the manufacture of sulphuric acid. The use of pyrites in the United States has diminished greatly because of the availability of low cost native sulphur, but pyrites have continued a major source of sulphur in many other countries. The most available pyrites for use in this country are in the form of lump pyritic ore and in mill tailings from flotation of other minerals such as lead, zinc, copper, gold, and silver. An important factor, when the use of pyrites for acid manufacture is being considered, is the disposal of calcine. A ton of sulphuric acid requires approximately ton of high-grade pyrite and results in 1/2 ton of calcine. If the calcine is a fairly pure oxide, free of harmful impurities, it can be used, after sintering, in an iron blast furnace burden. Its value might be as high as 15d per unit of Fe at the blast furnace; or possibly $10.00 per ton of sinter, if it assays 65 pct Fe. This might result in a credit of $4.00 per ton of acid if the sintering plant and blast furnace are both located adjacent to the acid plant. On the other hand, several factors must be considered before this credit can be realized, i.e., freight to blast furnace, availability of sintering facilities, methods of eliminating impurities, and the removal of valuable metal values. In some locations it would be most economical to dump the calcines.

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Rannome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 31/4x41/4-in. tool joints; 8-in. bits, 23/4x 33/4-in. tool joints; 6-in. bits, 2Y4x3Y4-in. tool joints; 4-in. bits, 15/ix25/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 23/8-in. drill pipe with tool joints, 31h-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 55/8 to 3 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Ranaome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 3Y4x41/4-in. tool joints; 8-in. bits, 23/4x 3 3/4-in. tool joints; 6-in. bits, 21/4x31/4-in. tool joints; 4-in. bits, 15/ix25/s-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 2-in. drill pipe with tool joints, 31/2-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 5 to 37/8 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By E. F. Reed

CONSIDERABLE deep hole prospect drilling has been done in the last few years in the Globe-Miami mining district about 70 miles east of Phoenix, Arizona, and in the San Manuel-Tiger area about 50 miles south of the Globe-Miami region. More than 205,000 ft of churn drilling have been completed by the San Manuel Copper Corp. at their property in the Old Hat Mining District in southern Pinal County. The deepest hole on this property is 2850 ft; there are 49 holes deeper than 2000 ft. At the adjoining Houghton property of the Anaconda Copper Mining Co., where only one hole reached 2000-ft depth, there were 27,472 ft of churn drilling and 3436 ft of diamond drilling. Three churn drill holes were deepened by diamond drilling methods. Near Miami in the Globe-Miami district the Amico Mining Corp. drilled four holes by combined churn and rotary drilling methods, the total amounting to 13,879 ft, of which 2256 ft were drilled with a portable rotary rig. In the same district, besides doing a large amount of shallow prospect drilling, the Miami Copper Co. drilled two holes of 2560 and 3787 ft, respectively, which were completed by churn drilling methods. The rocks encountered in drilling at San Manuel and Tiger are described by Steele and Rubly in their paper on the San Manuel Prospect' and by Chapman in a report on the San Manuel Copper Deposit.' The rocks are well-consolidated Gila conglomerate, quartz monzonite, and monzonite porphyry. In some places these formations stand very well while being drilled, and three holes were drilled without casing, the deepest of which was 2200 ft. In other holes faulted and fractured ground made drilling difficult. In the Globe-Miami district the deep drilling was done in the down-faulted block of Gila conglomerate east of the Miami fault and in the underlying Pinal schist. The geology of this area is described by Rannome. In the Amico holes the conglomerate varied from material consisting entirely of granite boulders and fragments to a rock made up of schist fragments in a sandy matrix; in the Miami Copper Co. holes there were more granite boulders and the material was poorly consolidated. Drilling was much more difficult and expensive in the Miami area than in the San Manuel district, mainly because of the depth of the holes and the formations drilled. All the deep hole prospecting described in this paper was done with portable rigs. The churn drill rigs were of several types, of which the Bucyrus-Erie were the most popular. Bucyrus-Erie 28L, 29W, and 36L rigs were used on some of the deeper holes on the San Manuel property. A few Fort Worth spudder types were tried, and the deepest hole at San Manuel was drilled with a Fort Worth Jumbo H. The spudder type is considerably larger than most other rigs used on this work and required a larger location site. The spudders were belt-driven machines with separate power units, and time required for setting up and moving was much longer than with the more portable drills. All the churn drilling was done by contractors or with machinery leased from them. A few of the contractors had complete equipment, including most of the necessary fishing tools. Unusual and special fishing tools were obtainable from the supply companies in the oil fields of New Mexico or in the Los Angeles area. Most of the contractors used equipment with standard API tool joints, so that much of it was interchangeable. Failure of tool joints is one of the principal causes of fishing jobs. It can be minimized if the joints are kept to the API specifications and the proper sized joints are used in the various holes. The minimum sizes that should be used with various bits are as follows: 12-in. and larger bits, 4x5-in. tool joints; 10-in. bits, 31/4x41/4-in. tool joints; 8-in. bits, 23/4x 33/4-in. tool joints; 6-in. bits, 2Y4x3Y4-in. tool joints; 4-in. bits, 15/ix25/8-in. tool joints. Two rotary drill rigs were tried at San Manuel on the same hole, and a portable rotary drill rig was used on the Amico drilling for test coring the formation and for drilling in holes 3 and 4. Rotary drilling differs from churn drilling or cable tool drilling in that the bit is revolved by a string of drill pipe and the cuttings are removed from the hole by a thin solution of mud pumped through the drill pipe. The principal parts of a rotary rig are the power unit, a rotating table to revolve the drill pipe, hoists to raise and lower the pipe and to handle casing, and a pumping system to circulate the drilling liquid. The rig used on the Amico property at Miami was mounted on a truck. The larger rig used on the San Manuel property was hauled by several trucks and had separate turntable and pumping units. Diamond drill coring equipment was used successfully with the rotary rig in the holes on the Amico property. To allow for 23/8-in. drill pipe with tool joints, 31h-in. core barrels and bits were used. With the standard 31h-in. core barrel there was considerable difficulty in maintaining circulation with mud, so a barrel was designed with a smaller inner tube and a broad-faced bit. This allowed coarser material to circulate between the barrels. Rock bits of 55/8 to 3 in. were used with the rotary rig for drilling between core runs. Diamond drill equipment is much lighter than churn drill tools, so that fishing tools can usually be obtained from supply houses by air express when needed. Three churn drill holes on the Houghton property at Tiger were deepened by diamond drilling with Longyear UG Straitline gasoline-driven machines. The open churn drill hole was cased with 21h-in. black pipe. In deep hole churn drilling, casing is one of the most important items, especially in drilling in un-consolidated material like the formations drilled by

Jan 1, 1953

By W. J. Slatosky

W. 0. Philbrook (Cairiegie Institute of Technologyogv—Mr. Slatosky has presented an interesting and constructive paper that represents another step along the way of converting steelmaking from an art to a science. I am confident that his computer will be practical and successful and that with a very few months of experience it will provide a significantly better degree of control than his record of 65 pct of heats within range obtained with the slide-rule calculator . A paper such as this, with a lot of symbols and condensed mathematics, is difficult to comprehend quickly. Since I have had an opportunity to study it carefully, perhaps my evaluation of its validity and accomplishments will save time for others. Mr. Slatosky has correctly used standard principles of stoichiometry and heat balances, which are available to anybody, but he has also brought to them two original contributions: 1) He has developed from operating data some empirical relations for predicting the final FeO content of the slag (at 0.5 pct C end-point) as a function of slag basicity, lance height, and scrap, ore, and scale in the charge. This improves the accuracy of prediction of temperature or scrap requirement compared with assuming an arbitrary, constant FeO content at the end of each heat. There is no assurance yet that exactly the same relations will hold for other furnaces or practices, but similar correlations can be expected. 2) He has combined calculations that are ordinarily carried out laboriously as a number of individual steps into a single, simple linear equation that can readily be fed into a machine. This involved a tremendous amount of painstaking detail work as well as the imagination to see the possibility and work out the steps. While his particular Eqs. [3] and [6] are valid only for the furnace design, charge weight, and blowing time used at Aliquippa Works, only a few numerical values have to be changed to adapt it for other conditions. In order to arrive at a useable solution, Mr. Slatosky had to make some basic assumptions about the process that are similar to those used by others. He neglected variation in some process variables and assumed an arbitrary average value for waste gas analysis and temperature for want of more exact information at the present time. All of these judgments are clearly stated. In addition, some thermody-namic data presently available are not adequate for the job, notably in relation to heats of formation and sensible heat in slag, and some expedient has to be adopted to get around the difficulty. Other people might prefer slightly different judgments about these details and hence obtain somewhat different numerical solutions. This is not of serious importance, however, because the errors accumulate in the "heat loss" term and are largely self-compensating for a constant heat time. Although the extended Eq. l(a) in Appendix I was set up as a rate equation originally, for convenience in using an analogue computer as stated in the paper, the time dependence was removed by later mathematical manipulations and assumptions about the process. The final result is an integration of element and energy balances from initial to final states; this procedure is as legitimate here as in any other form of heat-balance calculation. The formal handling of stoichiometry and thermochemistry appears to be correct, and it is assumed that any arithmetical errors would have come to light in applying the calculations to furnace practice. Mr. Slatosky's approach is not necessarily unique, in that other people might start with apparently different equations or prefer another form of final equation for another type of computer. However, he has presented an accomplished result that appears to be a theoretically sound and practically useful way of applying scientific principles and rapid computation for better control of steelmaking. His success will undoubtedly encourage himself and others to improve on the mathematical model and its use as better informatioq becomes available. John F. Elliott (Massachusetts Institute of Teck-t2ology)-The last comment by Mr. Richards that a calculator is quite unnecessary for an L-D operation ?-equi??es a rebuttal. The L-D furnace is a very high capacity process which places a premium on close control. When one is making steel at rates between 100 and 200 tons per hr, one cannot afford the luxury of an extra 5 or 10 min at the end of a heat correcting for an error that should never have been made in the first place. Mr. Slatosky's paper is a very sound application of the simple principles of stoichiometry and the energy balance. It is a satisfactory and valuable start, but only the start of the development of methods of control for this process. An analysis of the process shows that it should be very suitable to control by a computer. This is especially the case when various grades of steel are to be made. In fact, it would seem that the organizations who are planning new and bigger installations of L-D vessels should consider carefully the advantages that would stem from computer control of a vessel with the operator present to do little more

Jan 1, 1962

By B. W. Gandrud

WET-PROCESS coal-washing tables as we know them today have been in use in this country for approximately 25 years. The literature records only a few table installations worthy of note prior to adoption of the present-day differential-motion table. According to Phillip,1 six Campbell bumping tables were being operated by the St. Bernard Coal Co., Earlington, Ky., in 1893. As late as 1923, according to Richardson, 2 such tables were in operation in the Rosedale washery of the Cambria Steel Co., at Johnstown, Pa. Bumping tables probably were not used to any considerable extent except in the plant at Earlington, and at two or three plants in Pennsylvania, but they may be considered as. forerunners of the present-day coal-washing table. The transition from bumping tables to the type of tables now in use signalized a revolutionary change in table design. The advantages of the new design over the old were so outstanding as to make its adoption inevitable. BUMPING TABLES Bumping tables, as the name implies, utilized a bumping action for conveying the refuse and discharging it over the refuse end. The tables were hung from rods attached to overhead supporting beams. A shaft and cam arrangement at one end of the table served as the actuating mechanism and imparted to the table an endwise, back-and-forth, swinging motion. A bumping block was fixed in the proper position at the shaft end, so that the table would strike against it at the end of each backward swing. The hanger rods or stirrups were so adjusted as to give the table a slight uphill slope toward the refuse end; that is, the end next to the driving mechanism. The feed box was near the upper end of the table and designed to spread the mixture of coal and water over the width of the table. Wash water was added a short distance below the feed box. With the table in motion and coal mixed kith water being fed to it through the feed box, a certain amount of stratification took place, depositing heavy refuse at the bottom of the bed in contact with the surface of the table. Constant bumping and a certain amount

Jan 1, 1950

By C. D. Woodward

PrioR to 1902, the tramming of ore from the stopes to the shafts, in the Butte mines, was done by man or animal power, but the demand for greater tonnage and the need for more improved methods of tramming the ore resulted in the installation of the compressed-air locomotive in a few mines. Some of the objections to animal haulage were thus eliminated, but this locomotive was not generally adopted, because of a number of limiting features. It was superseded by the electric trolley locomotive. The early types of trolley locomotives were not satisfactory from the operators' viewpoint, but they had many economic advantages and after a number of years, in which improvements were constantly made in the structural details, they became dependable, so that the use of animals for main haulage in the mines has now been practically abandoned. There are now operating in the Butte mines of the Anaconda Copper Mining Co., 185 trolley-type locomotives, each of which has replaced five animals. It cost approximately $50 per month to keep an animal, hence the economies that resulted fully justified the expenditure of electrifying the haulage systems. The haulage conditions in this company's mines possess many advantageous features that arc not always found in underground mine workings. The drifts and crosscuts are large, approximately 5 ft. (1.5 m.) wide by 71/2 ft. (2.3 m.) high, and have curves of no less than 18 ft. (5.5 m.) radius; and the roadbeds of the drifts and crosscuts have a uniform grade of 0.5 per cent. toward the shaft, or in favor of the loaded trips. In all of the mines, except two, the track gage is 18 in. (45.7 cm.): and all tracks in the main haulageways are laid with 25-lb. (11.3 kg.) rails. The stations at the shaft are large, well-lighted, clean, and cool. The average temperature of the stations, however, is about 75" F. (24" C.), while the average temperature in some of the drifts and crosscuts is 92" I?. (33.5" C.). As the orebodies lie at varying distances from the hoisting shafts,

Jan 1, 1923

By C. D. Woodward

PrioR to 1902, the tramming of ore from the stopes to the shafts, in the Butte mines, was done by man or animal power, but the demand for greater tonnage and the need for more improved methods of tramming the ore resulted in the installation of the compressed-air locomotive in a few mines. Some of the objections to animal haulage were thus eliminated, but this locomotive was not generally adopted, because of a number of limiting features. It was superseded by the electric trolley locomotive. The early types of trolley locomotives were not satisfactory from the operators' viewpoint, but they had many economic advantages and after a number of years, in which improvements were constantly made in the structural details, they became dependable, so that the use of animals for main haulage in the mines has now been practically abandoned. There are now operating in the Butte mines of the Anaconda Copper Mining Co., 185 trolley-type locomotives, each of which has replaced five animals. It cost approximately $50 per month to keep an animal, hence the economies that resulted fully justified the expenditure of electrifying the haulage systems. The haulage conditions in this company's mines possess many advantageous features that arc not always found in underground mine workings. The drifts and crosscuts are large, approximately 5 ft. (1.5 m.) wide by 71/2 ft. (2.3 m.) high, and have curves of no less than 18 ft. (5.5 m.) radius; and the roadbeds of the drifts and crosscuts have a uniform grade of 0.5 per cent. toward the shaft, or in favor of the loaded trips. In all of the mines, except two, the track gage is 18 in. (45.7 cm.): and all tracks in the main haulageways are laid with 25-lb. (11.3 kg.) rails. The stations at the shaft are large, well-lighted, clean, and cool. The average temperature of the stations, however, is about 75" F. (24" C.), while the average temperature in some of the drifts and crosscuts is 92" I?. (33.5" C.). As the orebodies lie at varying distances from the hoisting shafts,

Jan 1, 1923

By Linn Bradley, W. W. Strong, H. D. Egbert

A hot-dry method of cleaning the gas from blast furnaces has been shown1 to conserve the sensible heat energy of the gas, and in general it thus permits of a higher flame temperature. The electrical method of cleaning gases removes from 98 to 99 per cent. of the suspended matter in one operation when operating properly, thus making possible the use of a very clean gas for burning in the stoves. The use of such clean gas thus makes feasible a different type of hot-blast stove construction from that being used at present. It therefore seems advisable to construct the stoves from the viewpoint of getting the maximum amount of heat energy into the products of combustion. By preheating the air added to the top gas and obtaining a better mixture of the air and top gas, less excess air will be required and a higher flame temperature obtained from the combustion of the top gas. The interior of the stove should be provided with a kind of brick that will withstand a high flame temperature, will conduct the heat energy along certain directions, insulate the flow of heat along other directions and at the same time store heat energy so as to make the flow of such energy more uniform. It is quite likely that these results can be best obtained by using several kinds of bricks: one type which is a good insulator of heat energy to be used for lining the stove and for making partitions through the stove to restrain the flow of heat; another type of brick or material with high refractory qualities; and a third type which will be a good conductor of heat energy. The surface of the stove exposed to the atmosphere should be a minimum in comparison to the mass of checkerwork which it is possible to obtain, due regard being paid to other conditions. With clean and hot furnace top gas it will be advisable to consider a rearrangement of the checker work in the stoves. One season for the present size and shape of checker openings is to provide facility for

Jan 1, 1917

By Warren E. Wilson

The transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,1 of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer3 concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. Energy Requirements Carrying out the idea contained in ' Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: Solids load in term of mass per unit time by W Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L slope of energy grade line by s Mass per unit volume of solids by m. Mass Per unit volume of liquid by mi Mass per unit volume of mixture by mm We then have as an expression for the solids load: W = pQmm The work done per unit mass, P, on the mixture to move it along the pipe the

Jan 1, 1948

By Warren E. Wilson

The transportation of solids in pipe lines is a matter of deep concern in many fields of engineering. Much experimental and theoretical work has been done in an effort to devise means of designing pipe lines that will satisfactorily transport given solid loads, but there remain many unanswered questions in this field. The present discussion will be confined to a rather narrow portion of this field and will endeavor to point out certain factors that may well be considered when comparing the relative economy of various installations. The data have been presented previously but a new interpretation was prompted by a suggestion in a paper by Professor O'Brien,1 of the University of California, that the friction work per pound of dry material should be considered in determining the relative efficiency of various installations. This method of considering relative economy in various pipe lines leads to some rather unexpected conclusions, especially with reference to the economy of large pipe sizes. A previous suggestion by the writer3 concerning the economy of a rather peculiar type of installation is strengthened. It is not believed that the considerations set forth herein apply directly in the case of dredge lines, where economy is largely a matter of a high rate of pumping in order to reduce the unit overhead charge. However, it is believed that some value may be gained from these considerations in cases involving the disposal of a fixed amount of solids per unit time with a minimum expenditure of energy and at a low cost for replacement of worn pipes. Such situations are not uncommon in the disposal of mill tailings and the disposal of solids deposited from irrigation waters. It was found that data leading to general conclusions for various pipe sizes were not available but certain definite trends are indicated and methods of analyzing available data are presented in the hope that they may be of assistance in preparing engineering estimates on designs of pipe lines of the type described above. Energy Requirements Carrying out the idea contained in ' Professor O'Brien's paper concerning the method of computing energy required to transport solids, the following analysis leads to some rather interesting conclusions. Let us represent: Solids load in term of mass per unit time by W Ratio of mass of suspended solids to total mass of mixture flowing by p Volume rate of flow of mixture by Q Length of pipe by L slope of energy grade line by s Mass per unit volume of solids by m. Mass Per unit volume of liquid by mi Mass per unit volume of mixture by mm We then have as an expression for the solids load: W = pQmm The work done per unit mass, P, on the mixture to move it along the pipe the

Jan 1, 1948