Search Documents

By Bunting S. Crocker

PEBBLE grinding at Lake Shore is not a temporary wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct - 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 1 1/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to, $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change, back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1 1/4-in. steel ball charge. To change pebble size it is necessary only to change two sets-of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean. out the customary supply of grinding balls kept on hand. Also it has not always been ' possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant should be designed to use as many units-or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, .2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant. and an extremely efficient-one. In this plant the ratio of ball mills to tube mills was 1: 2. When the much cheaper pebble mills were substituted for the tube mills, this. ratio -was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was. proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on -.8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.1 This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without .grate discharges-both with 1 1/4-in. and 3/4-in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report' was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3 1/2-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are, lengthy and may be covered in a separate 'report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The -8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands .from the bowl sent to the primary pebble mills, see Fig: 1. In the pebble mill circuit the ore is ground to 90 pct -325 mesh (24 pct + 28 microns) In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO2, 80 pct insoluble. Its grindability at different. meshes has been shown near the top of 'the list in F. C. Bond's grindability tests' Lake Shore is not shown, but an adjacent-mine with identical ore, Wright-Hargreaves, is. Reasons for the-Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1952

By M. H. Gaskell, D. M. Kehn, G. T. Pyndus

A correlation of the bubble point pressure for black oil systems is developed using the standard physical-chemical equations of soiutions. The correlation is based on 158 experimentally measured bubble point pressures of 137 independent systems and is expressed in terins of the usually measured field pnrameters— flash separation gm-oil ratio, rank oil gravity, total gas gravity, and reservoir temperature. The data were obtained on systems produced in Canada, Western and Mid-Continent01 United States, and South America. The average error (algebraic) in the representation is 3.8 per cent, and the maximum error encountered is 14.7 per cent. INTRODUCTION In the absence of experimentally measured properties of reservoir fluids, it is often necessary for the field engineer to make estimates regarding the fluid properties based on the usually measured producing parameters. To aid in these estimations, various correlations have appeared in the literature in recent years. Among the pertinent properties of interest is the bubble point pressure. A correlation for this parameter has been reported by Standing'. However, this correlation was based essentially on California produced crudes and this limitation was pointed out with its presentation. The correlation presented in this paper utilized data on crude oil systems from Canada, Wes- tern and Mid-Continental United States, and South America. CORRELATION DEVELOPMENT The basic assumption used in this development is the same as employed by Standing', There is a wide variety of ways to combine these parameters; however, in this instance the combination was made on the basis of Henry's law2. Accordingly, the relationship proposed is Although Eq. 2 defines an individual system, it is of limited value since H' is a function of gas-phase composition and the system temperature. It was observed that for the systems where the bubble point was measured at several temperatures that the ratio of the bubble point pressures and the ratio of the corresponding absolute temperatures (R) were practically identical. Thus, for correlation purposes the bubble point pressure may be taken as a direct function of the absolute temperature. This relation is valid only for those systems that are not near the critical point. Accordingly, this correlation will be inadequate for systems in the region of the critical point. The solubility of the various hydrocarbons found in the gas phase increases with the molecular weight. Thus, the saturation pressure should be inversely related to the gas gravity. Applying these principles to Eq. 2 and rearranging terms gives: The variables 02 the left side of Eq. 3 were designated as the "hubble point pressure factor". The number of mols of tank oil per barrel is a function of the "mclecular weight" of the tank oil. Although the tank oil is a complex mixture, it was assumed for the purposes of this correlation that a unique molecular weight could be assigned to a given oil. This was designated as the "effective molecular weight", and was related to the oil gravity, This empirical relationship was developed simultaneously with the correlation by assuming values of M,. and working to obtain a smooth curve for both the correlation and the effective molecular weight. The relationship between the oil gravity and the effective molecular weight used in this correlation is shown in Fig. 1. The effective molecular weight is somewhat higher than the molecular weight of the C fraction. The difference between these values is largest for the low-gravity systems. It is noted that this effective molecular

By Herbert D. Drechsler

The objective of this paper is to identify the components of income above that necessary to keep a mine in production and relate those components to the profits of a mining firm. This is a discussion of "Economic Rent". Economic rent is the payment to the factors of production -- capital, labour, entreprenuer and the owner of the mineral deposit -- above the minimum necessary to keep all of these factors cooperating in the production of minerals from a particular deposit. THE SOURCE OF ECONOMIC RENT IN THE MINING INDUSTRY In the mining business each mine has a unique cost of production unrelated to any other mine. The cost of production is developed from many factors such as the technical characteristics of the deposit: depth of deposit, depth of overburden, areal extent, and type of deposit. Furthermore, there are locational factors which also alter cost of production. For example, identical mineral deposits located in Zambia or in the southwest U.S.A. would have different capital and operating costs. These differences occur because of different qualities of labour, ease of acquisition of operating supplies or investment capital and other similar factors. To explain the source of rent, we shall use ideas developed by Alfred Marshall and Kenneth Boulding. Suppose a meteorite with special qualities exploded over earth and sent showers of small meteorites to the ground in various places. At some point in time, people learned of the special qualities of these meteorites and a demand arose and one mine was started. Suppose further that the cost of production to produce these meteorites was different for each deposit. Probably the lowest unit cost mine would commence production because if a higher unit cost mine started-up, eventually someone would discover a deposit that had a lower unit cost of production. The lower unit cost mine would take over the market and the high unit cost mine would cease production. The cost curve for the first mine would look like the following Figure 1. [ ] The output for the first mine is on the horizontal axis with the cost per ton on the vertical axis. The curve labelled AC I is the average total cost of production over the range of possible production that could occur at the property. The total costs are composed of variable costs, which are dependent on output and fixed costs, which are dependent on time. For example, the cost of operating supplies would be a variable cost and normal profit would be a fixed cost.1 The opportunity cost of the investment would also be part of the fixed costs. In our example, all factors of production, are receiving the minimum necessary payment that will keep the mine operating. The average total cost curve is U-shaped because at zero output, the meteorite mine has no variable costs, only fixed costs which, on a unit basis, obviously are very high. To produce a small output of meteorites, the mining firm adds

Jan 1, 1985

By T. T. Read

TWO papers on health and safety were given Thursday afternoon when a joint session of the Health and Safety Committee and the Mining Methods Committee was held. T. T. Read presided and the first paper, on ventilation at Climax, was presented by the author, L. H. Glanville. In the subsequent discussion by Cloyd M. Smith, P. B. Bucky, B. F. Tillson, D. F. Haley, J. L. G. Weysser, and others a number of interesting points not covered by the paper, which appears in full in the January Mining Technology, were brought out. Ventilation troubles in mines so often hinge on the air being too hot and damp that it was interesting to hear of one where it is frequently too cold and too dry.

Jan 1, 1943

By Walter C. Lawson

The primary objective in laying track in panel sections is to reduce the number of track laborers required. This is possible because the work is mechanized. Moreover, because the work is mechanized and each of the operations needs only a few men, and each is self-contained, the work can be carried on at night as well as in daylight; therefore, the method provides a means of preparing tracks on more than one shift. The laying of tracks in open pits and quarries in panel sections is not new but new methods have been made possible by the introduction of new types of equipment. The purpose of this paper is to describe the methods that are followed in the Morenci open pit. General Mining Operations Full-scale ore production at Morenci is about 50,000 tons daily. The normal yearly output of ore and waste is 32,000,000 tons, of which 30,000,000 tons is handled by rail haulage. This requires the building of approximately 47 miles of loading and other temporary tracks during a 12 months' period (Fig I). Broken ore and waste is loaded into 90-ton capacity dump cars with 5-cu-yd full-revolving electric shovels. Shovel banks are uniformly 50 ft high. They are blasted with churn drill holes and the broken material from a blast is loaded out with two shovel cuts, the first one being called (locally) the "splatter" cut and the second one the "clean-up" or "face-up" cut. When a shovel is making its clean-up cut, the loading track is about 70 ft from the toe of the bank and, under normal conditions, a bank is blasted against it without removing the track. It is thus evident that each track is used for two shovel cuts before relocation is necessary, inasmuch as a single position serves for both the clean-up cut prior to blasting and for the splatter cut following the blast. The normal advance of a solid bench by a single blast is 40 ft (Fig 2). Preparation OF Panel Grade In preparation of panel grades bulldozers and rooters are used for the bulk of the work and an auto-patrol road grader for the final stage. At Morenci, bulldozers have always been used for track-grade preparation, but before the introduction of the other two types of equipment grades were uneven and irregular at best, and always necessitated a big gang of track laborers to hand-block under the ties with rocks after a track was laid into place. High blocking also made poor track because of instability and caused frequent derailments. Under these conditions derailments were generally bad as re-railing became difficult. Much track was torn up and many tics broken in the process. The inadequacy of the hand-blocked track, together with the inefficiency of hand blocking, constantly pointed, to the desirability of reducing the amount of labor required, so various means were tried to improve conditions, such as wood

Jan 1, 1949

By Frank C. Smith

THE following article covers the history of a metallurgical campaign, commenced in March, 1905, at the mines of the Socorro Gold Co., in the so-called desert region of Yuma county, Arizona. The results obtained were simply a function of necessity, as will be seen from the description of the local economic conditions and the character of the ores treated. The mines and mill of the Socorro Gold Co. are located near the small village of Harrisburg, about 60 miles from Congress Junction, Arizona, which, at the time this work was undertaken, was the nearest railroad station, and from which all supplies were brought by wagon through the desert, at a cost varying from $10 to $15 per ton in addition to the railroad freights. Wood, the only available fuel, cost $7 per cord. Water for all purposes was pumped to the plant from a well in the valley about 4 miles away. Under the Arizona law, a working-shift is 8 hr. Wages were $6 per day for foremen and mill-men, $4 per day for blacksmiths and engineers, $3.50 per day for miners, and $3 per day for shovelers and trammers. The ore consisted of quartz, in which, above the 250-ft. level, the iron-minerals were largely oxidized and some free gold was visible; below that level few traces of oxidation occurred, and pyrite constituted the principal mineralizer in the quartz, together with occasional pockets of galena and a few eccentric specks of covellite. The 20-stamp mill was equipped with plates for amalgamation, and three Standard tables for concentration. The tails from the tables were elevated by a centrifugal pump to a launder, by which they were conveyed to a 60-ton cyanide-mill for further treatment. The cyanide-mill was erected at so great a distance from the stamp-mill that a separate crew of workers and a separate power-plant were necessary. No arrangement had been made for the treatment of the concen-

Jan 1, 1907

By J. C. Donnell, O&apos

Sinking an inclined ventilation shaft at the Never Sweat mine of The Anaconda Co. has required a unique underground hoisting arrangement because the shaft changes direction from vertical to 71.5º. Access to the project is made through the High Ore shaft to the 2800-ft level, then by crosscut to the bottom of the vertical Never Sweat shaft. The 71.5º incline shaft starts at this point, which makes an unusual operation as the 60-cu ft sinking skip coming up the incline changes into a vertical position above the 2800-ft level before going into the dump scrolls. This requires both rollers and guide shoes on the skip. Coming up the incline to the collar of the shaft the loaded skip enters into a guard before reaching the vertical section. The tendency of the surface- mounted Never Sweat hoist is to lift the front wheels of the skip off the track, so the guard keeps the skip on the rails during the transition.

Jan 11, 1965

By Leon W. Dupuy

A brief elemental discussion of the fundamental factors involved in diamond drilling often fills a need, particularly when a mine operator is contemplating for the first time an exploratory drilling campaign. Mr. James A. Barr, of the Mining Methods Committee of the Institute, has suggested that many engineers must undertake a drilling program of one kid or another and, in the case of diamond drilling, sometimes even are at a loss as to how to start correspondence. The Bureau of Mines has acquired much experience in managing exploratory diamond-drilling activities. Its activities in diamond drilling included exploring for potash in Texas, Utah and New Mexico1-' during the 1920's. The Strategic Minerals Exploration program was started in 1939 and was materially expanded during the war to include critical and essential minerals. At the peak of its war-time activity more than 200 engineers were employed, a majority of whom supervised diamond drilling. Many of them had little experience with such work prior to their employment with the Bureau, and it was necessary to instruct them in their duties. The purpose of this paper, therefore, is to present the fundamentals of exploratory diamond drilling from the viewpoint of management in such a way that they will Serve as a basis on which those beginning exploratory diamond drilling may build an experience and acquire ability to direct exploration as they gain that experience. Selection of Method of Drilling Selection of the method of drilling is governed by the type of material to be penetrated and by the type of sample needed to determine the quality and importance of a deposit. Types of drills that might be used for a given job include diamond drills, churn drills, hand auger drills, power auger drills, bucket drills, and rotary drills (sometimes called "seismograph rigs"). Each type has a specific field in which it is most useful. While the Bureau of Mines utilized all the types of drilling mentioned, diamond drilling was used on a majority of its exploratory projects. Factors Governing Selection of Diamond Drilling The prime consideration in selecting a method of drilling is first to determine what results are desired and then to choose a method that will furnish reliable data in the ground to be penetrated. Use of diamond drills permits recovery of a core as well as sludge at a comparatively low cost. The core provides an excellent means of studying the formation being drilled and furnishes much information that would be difficult to obtain otherwise. Diamond drills will recover a satis-

Jan 1, 1948

By Leon W. Dupuy

A brief elemental discussion of the fundamental factors involved in diamond drilling often fills a need, particularly when a mine operator is contemplating for the first time an exploratory drilling campaign. Mr. James A. Barr, of the Mining Methods Committee of the Institute, has suggested that many engineers must undertake a drilling program of one kid or another and, in the case of diamond drilling, sometimes even are at a loss as to how to start correspondence. The Bureau of Mines has acquired much experience in managing exploratory diamond-drilling activities. Its activities in diamond drilling included exploring for potash in Texas, Utah and New Mexico1-' during the 1920's. The Strategic Minerals Exploration program was started in 1939 and was materially expanded during the war to include critical and essential minerals. At the peak of its war-time activity more than 200 engineers were employed, a majority of whom supervised diamond drilling. Many of them had little experience with such work prior to their employment with the Bureau, and it was necessary to instruct them in their duties. The purpose of this paper, therefore, is to present the fundamentals of exploratory diamond drilling from the viewpoint of management in such a way that they will Serve as a basis on which those beginning exploratory diamond drilling may build an experience and acquire ability to direct exploration as they gain that experience. Selection of Method of Drilling Selection of the method of drilling is governed by the type of material to be penetrated and by the type of sample needed to determine the quality and importance of a deposit. Types of drills that might be used for a given job include diamond drills, churn drills, hand auger drills, power auger drills, bucket drills, and rotary drills (sometimes called "seismograph rigs"). Each type has a specific field in which it is most useful. While the Bureau of Mines utilized all the types of drilling mentioned, diamond drilling was used on a majority of its exploratory projects. Factors Governing Selection of Diamond Drilling The prime consideration in selecting a method of drilling is first to determine what results are desired and then to choose a method that will furnish reliable data in the ground to be penetrated. Use of diamond drills permits recovery of a core as well as sludge at a comparatively low cost. The core provides an excellent means of studying the formation being drilled and furnishes much information that would be difficult to obtain otherwise. Diamond drills will recover a satis-

Jan 1, 1948

By Glenn Southward

This describes a highly concentrated mining system consisting of a series of short faces arranged to get complete recovery of coal with the maximum protection for the men and equipment. A series of conveyors collect the coal from eight faces and discharge into mine cars at a central loading point on a main haulageway. The conveyor is sectional and, by a multiple-drive arrangement, the same standard parts are used to construct a unit of any length. THE. V system of mining used at the Norton mine of West Virginia Coal & Coke Co. was designed as a modified longwall with face conveyors, FIG. 1.-GENERAL DESIGN AND ARRANGEMENT OF FACES IN V SYSTEM. for use under roof conditions that would not permit regular longwall methods. This mine is in the Lower Kittanning seam, which averages about 6 ft. in height. There is from 20 to 30 ft. of slate and shale immedi-

Jan 2, 1924

By W. M. Aubrey, P. L. Steffensen

AUTOMATIC grinding control for ball mills in closed circuit with cyclones was developed for the concentrator of the Marmoraton Mining Co., a subsidiary company of Bethlehem Steel Corp. The concentrator, located at Marmora, Ont., Canada, treats an ore that analyzes 35 to 40 pct magnetic iron. Magnetite, the recoverable iron mineral, is concentrated by fine grinding and magnetic separation. The mesh of grind obtained in the plant is not determined by the point at which liberation is sufficient to yield a satisfactory grade of concentrate, but by the physical requirements of the downstream pelletizing operation. This is considerably finer than is metallurgically necessary. The concentrator has two identical units, each consisting of a rod mill in open circuit, followed by a ball mill in closed circuit with a cyclone and two rougher magnetic separators. The cyclone was chosen as a classifier because of its high efficiency on heavy gravity, magnetically flocculated material.

Jan 1, 1957

By Y. S. Bonillas

IRA B. JORALEMON, Warren, Ariz.-This paper has covered the situation so well that I have few suggestions to offer. Two points, however, are shown more clearly in the lower mines of the Calumet Arizona Mining Co. than in Copper Queen workings, and may be worth mentioning. The first is the usual association of high-grade primary sulphide orebodies with much larger masses of pyrite and silica, averaging less than 1 per cent. copper. The richer areas occur sometimes around the borders of the lean pyrite bodies, and sometimes entirely within them. While as yet we have not been able to find any certain reason for the location of high-grade sulphides in low-grade masses, it seems likely that the age of fracturing is one controlling feature. The first stage of mineralization, with pyrite and silica, had the effect of sealing up the rock. The later solutions, richer in copper, could pene-

Jan 12, 1916

By R. L. Dietrich

[ ] We at The Dow Chemical Co. have recently done some work on the deformation of magnesium alloys and although it is far from complete there have been a few observations which may be of interest in connection with this paper. In the first place, I should like to point out that the deformation of polycrystalline magnesium alloys has not been completely explained by the work on single crystals. Bakarian and Mathewson found slip on the ( 101) and {002) planes, and (102) twinning as the deformation mechanisms in magnesium single crystals, but in compression when the (002} planes were within ii° of parallel t0 the compression surfaces, fracture occurred at all temperatures up to 340°C (644°F). In commercial working processes, particularly in rolling, an orientation is quickly produced which should be unfavorable to further deformation by the three established mechanisms. The first few reductions in rolling quickly produce an orientation having

Jan 1, 1947

By C. J. Barber

This paper Donald describes how New Jersey Zinc Co. made surveys for a connection between the Ivanhoe and Van Mater shafts at Austinville, Va. Except to say that the two faces had to meet "accurately" on line and grade, Donald does not indicate the required precision. Assuming that there were 24 angles in the 1 ½ -mile traverse and 15 in the one-mile traverse, it can be shown that if the average error in plumbing each shaft was ±30" and the average error in measuring each traverse angle was ±10", then the average error at the point of connection would have been about ±1.9 ft normal to the line between the shafts. This calculation assumes that any errors in the triangulation would be negligible compared with the errors in the plumbings and traverses, and it also neglects taping errors. With no constant errors or blunders, the latter would be important only in lines normal to the line between the shafts.

Jan 7, 1959

By John P. Bingham

GRACE mine and mill, owned by Bethlehem-Cuba Iron Mines Co. and operated by Bethlehem Cornwall Corp., are being developed to produce 9600 tons of 42 ½ pct Fe magnetite ore, to yield 5760 tons per day of agglomerated iron concentrates averaging 65 pct Fe. Located about 14 miles south of Reading, the property is ideally situated in relation to good highways and railroad. Three major steel plants, at Bethlehem, Steelton, and Sparrows Point, Md., are about equidistant from the mine. Discovery of the orebody resulted from an airborne magnetometer survey, completed in 1948, con¬sisting of 7000 lineal miles of flight lines spaced one-fourth of a mile apart and extending from the Delaware River on the east to Gettysburg, Pa., on the west. After examining the magnetic maps, Bethlehem's geologists made limited surface traverses over the Grace mine and Warwick area to check the location, and realtors were engaged to purchase the desired properties over and adjacent to the area showing favorable indication of ore.

Jan 1, 1957

C. H. WILSON,* New York, N. Y. (written discussion?).-On page 2680, under the heading Junction Box and Zone Box, the authors say, with reference to the zone-box principle of connections between primary couple, compensator couple, and copper leads, which method in its appli-cation to a single couple is shown in Fig. 34, that if this principle is adopted for a multiple installation, so as to save compensating leads; a selective switch may be (and the inference is that it must be) inserted between the zone box and the different couples, putting the switch at an inconvenient point, or putting up with some complicated interlacing of connections.. Wilson-Maeulen Co. has employed the zone-box method extensively for several years and in none of the many multiple installa-tions has either of those alternatives been resorted to. Instead there has been used a method of wiring with the switch at the indicator and involv-ing no more complications than the junction-box method shown III

Jan 11, 1919

By W. L. Price

Since they were first introduced in the United States in 1964, rail mounted bucket wheel stacker-reclaimers have gained wide acceptance by the coal burning power plant group. Other installations include those in phosphate and steel mill ore storage areas. First Installation in the U.S.A. Three basic types of bucketwheels for handling stored bulk materials are now being used in the U.S. Bucketwheel stacker-reclaimers came into use in the U.S.A. in 1964 at the Chalk Point power station built by the Potomac Electric Power Co. It was decided, after serious study, to install the first slewing-boom, rail mounted bucketwheel stacker-reclaimer at this installation. It was designed with a separately counter- weighted wheel boom because the fixed counterweight of the European designed bucketwheel made it possible for the machine to overturn backwards if the wheel rested on the ground due to a possible failure of the luffing system.

Jan 1, 1970

By Matsubara, A.

THE problem of the equilibrium between iron, carbon, and oxygen was first carefully investigated by E. Baur and A. Glaessner,1 who determined the equilibrium conditions of the two reactions Fe304 + CO ? 3FeO + C02 (1) FeO + CO ? Fe + C02 (2) within the range of temperature of 350 to 900° C. The next series of investigations was made by R. Schenck and his co-workers,2 who sought to check the results of Baur and Glaessner's work within the range of 550° to 700° C. and to elucidate the conditions of formation of iron carbide. The equilibrium conditions at the higher temperatures, however, are generally understood to be much more complicated, owing chiefly to the formation of solid solutions between iron, its oxides, and carbide; nevertheless their determination is of much more practical importance than those at the lower temperatures. It is well known that each of the reactions is reversible; equilibrium is established when the partial pressure of oxygen in the gaseous phase has become equal to the dissociation pressure of magnetic oxide in equation 1 and that of ferrous oxide in equation 2. These equilibria are subsequently designated by "the Fe304-FeO equilibrium" and "the FeO-Fe equilibrium," respectively. As the Fe304-FeO equilibrium has three phases with three components, it must be bivariant. But the reaction involves no pressure change in the gaseous phase, hence the final composition of the latter is practically unaffected by pressure. To know, the equilibrium conditions, the compositions of the gaseous phase at different temperatures must be determined. At the higher temperature, however, iron and its oxides dissolve one another in various proportions forming solid solutions; in many cases the content of oxygen in the solid phase becomes a new independent variable.

Jan 2, 1921

By D. S. Sanders

IN the four mills of the Cerro de Pasco Copper Corp. in Peru, some 3000 tons of complex sulphide ores are treated daily, with four kinds of concentrates produced: copper, lead, zinc, and pyrite, each containing values in silver, gold, bismuth, and some other metals. Copper occurs in the various ores as chalcopyrite, enargite, tetrahedrite, tenantite, bournonite, and chalcocite; lead as galena; and zinc as sphalerite and marmatite. Small amounts of silver and tungsten minerals are also present. The gangue consist mainly of shale, quartzite, calcite, rhodochrosite, pyrite, and pyrrhotite. In general the values are freed with a relatively coarse grind, about 70 prr cent minus 200 mesh being the finest to which the pulps are ground. In Morococha and Casapalca however. sphalerite particles much finer than this contain inclusions of chalcopyrit r. At Cerro de Pasco the silver and gold are present in all sizes of the ground pyrite.

Jan 1, 1945

By W. O. Philbrook, R. S. Kaplan

Rates of oxidation were studied for levitation-melted samples of Fe-C, Fe-Si, Fe-C-Si, Fe-C-P, Fe-C-Mn, and Fe-C-S alloys oxidized by mixtures of either 5 or 10 pct O2, in helium at several temperatures in the range 1850o to 2050°C. The concentrations of the oxidized elements decreased linearly with time. Mass transport in the gas phase was found to be the rate-controlling step. Ilowever, a mulliple -film model was necessary to account for the flux of iron and manganese vapor counter to the oxygen flow. THE advent of the basic oxygen and spray-refining processes of steelmaking has generated considerable interest in the kinetics of reactions between oxygen-bearing gases and liquid iron alloys. The object of this research was the determination of the role of mass transport in the gas and metal phases and of surface chemical reactions in the overall oxidation process. Of the reactions for elimination of elements dissolved in liquid iron, the oxidation of carbon has been investigated most extensively. Studies by Fuji and Ura,1 Fuji,' and Gunji et a1.,3 done with pilot-scale basic oxygen furnaces, indicated that mass transport of gaseous oxygen was rate controlling when the metal carbon content was greater than 0.15 to 0.30 pct. When the carbon content was less than this critical amount, diffusion of carbon dissolved in the iron was thought to be rate limiting. L. A. Baker et a1.4,5 levitated samples of iron-carbon alloys and oxidized them in CO-CO2, CO-C02-He, and O2-He atmospheres at 1660°C. These investigators found that the metal carbon concentration decreased linearly with time and their samples exploded when the carbon content fell below about 0.8 pct. They concluded that the rate-limiting step was the transfer of oxidant to the gas-metal interface. A model based on this rate-limiting step predicted decarburization rates within a factor of two of those experimentally observed. The explosion of the samples was attributed to a shift in rate-controlling step from gas-phase mass transport to transport of carbon dissolved in the metal. L. A. Baker and R. G. ward6 and R. Baker7 used freely falling droplets to investigate oxidation kinetics. In both papers, the authors speculated that flame fronts may have been partially responsible for the lack of better agreement between experiment and theory. Swisher and Turkdogan8 oxidized liquid iron having an initial carbon content of 0.97 pct with CO-CO2 and CO-C02-Ar mixtures at 1580°C. They concluded that a surface reaction was the rate controlling step. Iron-silicon alloys have been oxidized by R. Baker7 and Filipov and Martynov.9 R. Baker 7 also studied iron-carbon-silicon alloys falling as liquid droplets through oxygen. He found that the rate of both carbon and silicon oxidation varied inversely with the drop radius. When the initial silicon concentration was greater than 6 pct, the samples cooled during their fall. Manganese oxidation has been studied by R. Baker,7 and by Filipov and Martynov.9 ward10 investigated the loss of manganese from liquid iron under vacuum. He found that mass transport of manganese vapor was rate limiting in argon atmospheres of 0.02 torr to 700 torr pressure. Iron-sulfur alloys were desulfurized with H2-He and H2-Ar mixtures by Peters et al.11 A model based on the transport of H2S from the gas-metal interface into the bulk gas gave fair agreement between the predictions of the model and the experimental results. EXPERIMENTAL TECHNIQUE Essentially, the technique consisted of using helium-oxygen gas mixtures to oxidize levitated samples of liquid binary and ternary iron-base alloys. The fused silica reaction tube is shown in Fig. 1. The coil design is similar to that of L. A. Baker et al.4'5 A Tef-lon-sheathed, ferritic stainless steel lifter (B) and an alumina pedestal (C) were used to support the sample in the coil prior to levitation and for the recovery of the sample. Samples were inserted into

Jan 1, 1970