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By Rztssell H. Bennett

A SURVEY of the Mesabi Range iron-ore industry demonstrates that a satisfactory degree of technical progress has been achieved in the last fifteen years. This advance has not been made over a uniform front, but, by salients and bays, according to the different initial positions, economic and technical, of the component producing companies. The producers endowed with rich ore have in general, because of natural advantages, progressed more slowly than their neighbors with leaner properties. The advance has been made step by step all along the line -blasting, loading, hauling and concentrating-and without the benefit of radical innovation. There have been, for instance, no such parallel discoveries in iron ore as leaching and flotation in copper. The much lower unit price of iron operates, of course, to bar the application of these and other practices developed for the higher-priced metals. There exists, however, a radically new concentration process whose technology is well worked out and which may become economically operative in the not too distant future; this, as so often happens in engineering advance, because of recent developments in an entirely unrelated field. First, however, let me devote a paragraph or two to the more important economies effected in mining.

Jan 1, 1932

Arizona Bureau of Mines and College of Mines and Engineering, University of Arizona, Tucson, Ariz. Lists of the maps and bulletins available will be sent upon request to G M Butler, Director, Arizona Bureau of Mines, Tucson, Ariz Bulletin 119, A resume of the geology of Arizona, by N. H Darton, is sold for $1, postpaid Fourteen other bulletins, any five of which may be had free of charge (additional numbers may be purchased for 10 cents each) are available as follows - 106, A geological reconnaissance of the Tucson and Amole Mountains, by Olaf P Jenkins and Eldred D Wilson, 114, Fluorspar, by M A Allen and G M Butler, 120, Gold and copper deposits near Payson, Arizona, by Carl Lausen and Eldred D Wilson, 122, Quicksilver resources of Arizona, by Carl Lausen and E D. Gardner, 123, Geology and ore deposits of the Courtland?Gleeson region, Arizona, by Eldred D Wilson; 125, The mineral industries of Arizona, by J B Tenney, 126, Asbestos deposits of

Jan 1, 1933

By H. Merton Ruth

THE northern anthracite fields, although facing the same economic problems as the southern fields, are confronted with the additional problem of fast dwindling reserves of anthracite which can be conventionally and economically won. The mining of coal which remains in thin beds and the recovery of pillars from areas heretofore considered unrecoverable, either by reason of cost or surface restriction, pose problems of in- creasing importance to the economic life of the northern field operations. Successful approaches have Been made to this problem and other work is being performed that holds promise. There exists in the northern anthracite field millions of tons of solid coal, the recovery of which has not been too successful due to the following inherent operating faults: inability to concentrate producing areas; the great quantities of rock to be cut and handled, lack of consideration of proper sequence and speed of second mining, and the continued use of the mine car as the basis of payment. A mining system that has successfully solved these major problems has been in operation in the Scranton area since 1944. Over 1% million tons have been produced from veins varying from 3 to 4% ft thick by this method, which is known as the multiple panel-triangular system-conveyor mining.

Jan 1, 1950

RECENTLY seven Boy Scouts were chosen from all the troops in the United States for their all-around ability, excellence in' scouting, and special ability as writers. Two of the members of this selected group were sons of In-stitute members. In the accompanying illustration, Ber-nard Queneau, son of A. L. J. Queneau, of the U. S. Steel. Corp., can be seen at the extreme left, just behind Theo-dore Roosevelt, who is shaking hands with the boy next to. Carl Zapffe (standing in the middle with his hands behind him), the son of Carl Zapffe, manager of the iron ore prop-erties of the Northern Pacific Railroad.

Jan 6, 1928

By Kidd, Robert L.

INASMUCH as galena and sphalerite flotation concentrates are being produced that contain over 95 per cent galena or sphalerite, it is not unreasonable to believe that a flotation concentrate assaying at least 80 per cent of native gold can he secured provided the gold is free. As yet, little if any fundamental information has been published concerning the flotation of free gold either alone or in the presence of sulfides. So far, we are unable to make as complete a separation of gold from sulfide minerals by flotation as is possible in, the separation of one sulfide from another by this method. An introductory study of the flotation of several free gold ores has been made in our Utah laboratories. For this study three general types of gold ores were chosen, namely: (1) Quartzite containing 0.26 oz. gold per ton in which no base metals are present; (2) sulfide ore (2.04 oz. Au; 1.31 oz. AO containing pyrite as the only sulfide present and the gold free; and (3) oxidized ore (1.10 oz. Au; 0.70 oz. Ag) from an oxidized zone of the vein from which sulfide ore (No. 2) was taken; little sulfide iron was present.

Jan 1, 1932

By C. P. Proctor, E. D. Hummer, J. A. Younkins

COMPARATIVELY little has been published relating practical experience with the control of the solid content of washery water systems that must be "closed." A "closed" system is one that for some reason may not be cleaned by the emission of any effluent of high solid content to be replaced by clean water in order to hold the solid content of the waste water within reasonable bounds for efficient washing. It is the purpose of this paper to describe practical experience with such a water system and to present data that it is hoped may be of some use to those who are faced with a similar problem. The washery clarification system described comprises three thickeners and three filters, and although it is admitted that each of the fields of water settling and thickening and filtration offers a wide scope for investigation, the authors are not attempting to present the answers to these problems but to set down the results of some experience along these lines, making no claims as to the ultimate solution of the problems in hand. The subject of water clarification has vast ramifications; for example, the science of flocculation as applied to water settling and filtering opens up fields of wide possibilities of which as yet little is known. A comparatively recent entry into the field of water treatment, which may offer possibilities of practical application, is the use of "wetting" agents such as "Aerosol." GENERAL ARRANGEMENT OF WATER SYSTEM So that the reader may obtain a comprehensive picture of the factors concerned in the control of a closed water system, it is necessary to give a brief description of the washing plant that is serviced by the water system to be described in the following pages. As we are not primarily concerned with the washing plant in detail, only a brief description of the flow of coal as it relates to the water system will be given here. For a more detailed account of this type of washer, the reader is referred to earlier papers.1.2 The washer, which handles 0 to 4-in. raw coal (750 tons per hour) is an American Rheolaveur plant of two complements; the coarse coal or sealed discharge plant and the fine coal or free discharge plant. The sealed discharge plant (Fig. r) consists of twin 48-in. primary launders, each having two washing boxes, supplemented by one 32-in. rewash launder, which also has two washing boxes. "Push" water for washing and conveying coal in these launders is supplied from a constant-head tank under pressure head of 10 ½ ft. at the point of introduction at the feed end of the launder. This tank also supplies water for upward or vertical currents used in the Rheolaveur or washing boxes. Each box discharges into a sealed boot elevator (Fig. I).

Jan 1, 1942

By AIME AIME

L H. DUSCHAK gave an interesting talk at a recent meeting of the. San Francisco Section, based -011 experimental work with a variety of ores at the laborator of the Treadwell-Yukon Co., in Berkeley, Cal. He was of the opinion that the customary "cut and try" method of flotation testing was still in evidence and that the flotation process was more of an art than an exact science. Propounded theories have enabled investigators to' classify reagents as well as to suggest new ideas both in the metallic and nonmetallic fields, but the application of theory has riot as yet been able to solve many of the unknown factors in operation.

Jan 1, 1932

AT THE beginning of February the British Government announced acceptance of the American terms for funding the war debt, the main features-being as follows: Interest rate of 3 per cent. for the first 10 years and then 3 ½ per cent. up to the final extinguishment of the debt at the end of 62 years. The interest from the time since the loans were made to the actual beginning of the funding arrangement to be recomputed at 4 1/4 per cent. instead of the 5 per cent. rate now in effect. Congress ratified the contract respecting the British debt. The ship subsidy bill was relegated to the rear, which is indeed an end to the project. The natural consequence of this will be a liquidation of the Govern-ment's shipping business. The joint committee on the reorganization of Govern-ment departments made its report to the President. In general this seems to offer a well-considered plan for the coordination of Government service. The pro-posals respecting the bureaus that are particularly associated with the mining industry may be subject to criticism. According to the committee's plan, the Geological Survey remains in the Department of the Interior, but the Bureau of Mines and the compilation of the statistics of mineral production are transferred to the Department of Commerce. The Lausanne conference among the European powers came to an end early in February, without any treaty having been signed, owing to the refusal of the Turks to accept the ultimatum of the Allies. Theoreti-cally this left conditions in a state-of war between the Allies and Turkey. However, no actual movements have resulted. The French and Belgians continued to hold the Ruhr district of Westphalia, extending their lines and administration of the region. The Germans adhered to the policy of passive resistance. In the Parliament of Great Britain the French and Belgian policy was plainly declared to be unwise and dangerous. French and Belgian industries, moreover, were troubled by strikes among their own coal miners for higher wages. The Poles and Lithuanians engaged in border war-fare;. the civil war in Ireland continued. The rest of the world was quiet insofar as military operations are concerned.

Jan 2, 1923

By T. J. Woodside

Prior to 1927 the lead blast furnace charge at El Paso consisted principally of direct-smelting carbonate ores, very low in zinc, and the resulting slag seldom carried more than 4.0 pct. With the exhaustion of the mines in Northern Mexico which supplied these ores, they were supplanted with a much smaller tonnage of sulphide concentrates containing appreciable amounts of zinc, and the zinc content of the slag rose to 10.0 pct. The Anaconda Company's slag fuming plant at East Helena, Montana, had then been in successful operation for several years, and it was immediately recognized that, while at that time the production of high zinc slag was small, if the grade held, a plant for the recovery of zinc from the accumulated slag might some day be practical. The grade not only held but became increasingly higher so that by 1946 suffcient slag had been accumulated to warrant the construction of a fuming plant regardless of future slag production from the blast furnaces, and with the removal of Government restrictions on construction at the end of the war the project was undertaken. The general design of the plant follows that at Bunker Hill, with waste heat boiler, unit coal pulverizers and even more extensive instrumentation and automatic control. The area around the lead blast furnaces is rather congested, being cut off from the nearest feasible site for the fuming plant by the copper plant, so that the most practical route for hot slag transfer is through the converter aisle. For this reason and also to make use of the existing converter cranes to serve the fuming furnace, an unused reverberatory at the north end of the converter building was dismantled and the fuming furnace and boiler located in its place. A flue at right angles to the furnace and boiler runs northward to the U-tube coolers and baghouse (Fig 1). From the bag-house a steel flue runs eastward to an existing and long unused 300 ft con- Crete stack. The coal handling system and de-leading plant lie north of the furnace and east of the cooling tubes and baghouse. The furnace is conventional, 8 X 21 X 33 ft high inside and completely water jacketed (Fig 2). There are 21

Jan 1, 1950

By P. T. Chiang

Chemical etching methods for the simultaneous revealing of the tellurium or selenium Phase and the chalcogenide grain boundaries of the alloy systems are given. A tellurium eutectic was found Present in zone-melted ingots. Similarly, a selenium monotectic was present in ingots. In general, the second phase (tellurium or seleniumn) occubies three different sites; viz., along the chalcogenide grain boundaries, as inclusions within the chalcogenide grain, and on the undersurface of the ingot. The detection limit for the tellurium phase is about 1 u in width. THERMOELECTRIC materials based on Group V (bismuth, antimony) and Group VI (selenium, tellurium) elements have aroused considerable interest in recent years in the practical application of thermoelectric cooling. In many cases, a small amount of excess tellurium (or selenium) was added to the material to optimize its thermoelectric properties. Then the question immediately arises as to the number of phases present in the resultant alloy. In the binary systems of Bi-Te, Sb-Te, and Bi-Se, the congruent melting compositions have been reported to be non-stoichiometric and are represented by Bi~Te respectively. It is to beexpected and known that Bi2Te3 and SbzTe3 crystallize from the melt with an excess of bismuth and antimony in the lattice and that tellurium forms a eutectic.~' The same could be assumed to take place in the pseudo binary systems of (Bi,Sb)zTe3 and Bi2(Se,Te)3 as well as in the system studiedby puotinen5 and other workers. Likewise, BiaSe3 crystallizes from the melt with an excess of bismuth in the lattice and selenium forms a monotectic.~ Therefore, in practice, alloys solidified from the melt often contain a second phase (tellurium or selenium) in one region or another of the solid mass even without the addition of excess tellurium (or selenium). ~u~~recht' studied the thermoelectric properties of (Bi,Sb)2Te3 alloys with excess tellurium and simultaneous additions of selenium. He mentioned that the materials show two phases because of the considerable excess of tellurium or selenium. However, he did not report as to how the tellurium or selenium phase was identified. It is generally believed that the presence of an excessive amount of tellurium or selenium phase in the alloy would adversely affect its thermoelectric properties and its uniformity. Consequently, there is a need for a simple method for the identification of the tellurium and selenium phase. The quantity of the second phase present is usually too small to be detected either by chemical analysis or by normal X-ray techniques. This investigation was therefore carried out, first, to devise a simple metallographic method for the identification of the tellurium or selenium phase coexisting with the chalcogenides and, second, to determine the distribution and specific location of the tellurium or selenium phase in the ingots. EXPERIMENTAL PROCEDURE The starting materials used for the alloy preparations were 99.999 pct pure bismuth, antimony, and tellurium and 99.997 pct pure selenium. The bismuth and antimony were obtained from Consolidated Mining and Smelting Co. of Canada Ltd., while the selenium and tellurium were obtained from Canadian Copper Refiners Ltd. The tellurium was purified further in the laboratory by zone refining. The elements were pulverized in a stainless-steel pestle and mortar. The amounts for the desired composition were weighed out each time on an analytical balance to make up a 100-g sample. Then the sample was introduced into a Vycor ampule (19 by 150 mm), pumped down to a vacuum of 10"5 Torr for 15 min, and sealed off. The ampule was then heated in a horizontal resistance furnace at 800" to 900°C for about 20 hr. During this period the assembly was rocked back and forth several times to ensure good mixing. At the end of the heating period, the ampule was quenched in cold water and then transferred to the zone-melting apparatus described in a previous publications to grow large-size aligned polycrystals. The background and ring-heater temperatures were adjusted to make the freezing solid-liquid interface slightly convex to the liquid. The recorded temperature gradient in the vicinity of the freezing solid-liquid interface was around 15°C per cm. The ampule was moved horizontally at a speed varying from 0.4 to 2 cm per hr so that the ring heater would cover the whole ingot length from end to end. A single zone-melting pass was used for the Bi-Te, Sb-Te, and Bi-Sb-Te ingots. Two passes in the forward and reverse directions were carried out for the Bi-Se and Bi-Se-Te ingots. Six passes in the forward and reverse directions were performed for the Bi-Sb-Se-Te ingot. The zone-melted ingots were found to contain several large crystals, with their basal planes (0001) approximately parallel to the growth axis. Samples of bismuth and antimony tellurides coated with a layer of tellurium, and bismuth selenide coated with a layer of selenium, were prepared for comparison in phase identification. These coatings were made by dropping a piece of the zone-melted ingot into some molten tellurium or selenium under argon atmosphere and allowing them to cool slowly to room temperature. The metallographic specimens were prepared by

Jan 1, 1969

By George W. Powel

IN the nation's effort to raise its magnesium metal supply to meet the ever increasing demand, the Government is relying not only on standard established practice but has extended its support to developments not so well proved-at least, not commercially proved. Within the past few months considerable progress has been made, not only in the definite output of metal but in the exploitation of new processes, the successful adaptation of which should, in a reasonably short time, assure an adequate supply of magnesium. Recent months have also witnessed a reshuffling of the favored methods as well as of the likely raw materials, with the mineral dolomite gaining in interest over its richer but less abundant competitor, magnesite, while the lowly sea water appears to be losing ground to the higher magnesian brines of Michigan and western Texas. Every month sees a new alignment and there are signs that still newer methods or raw materials will be in use before the industry finally stabilizes itself.

Jan 1, 1942

By Charles H. Jones

Jan 1, 1905

By L. W. Emery

Undergrorlnd combustion operations were initiated in a 60-acre Bartlesville sand "shoe-string" reservoir in Allen Connty, Kans., in 1956. Tests in separate patterns were conducted using various co~nbinations of air and recycle gas to propagate combustion fronts from the injection toward the producing wells. These patterns were made up of 6 injection and 20 prodrrcing wells Gas and liquid prorluctiorz from each pattern was measured on an individual-well basis, and comparisons were made between the three patterns to ascertain the relative effects of injected gas composition on production behavior. Breakthrough of the combustion front at a well was characterized by an increase in water production from the well followed by an increase in bottomhole temperatrrre to approximately 250" F. After burning fronts had broken through at five producing wells, operations were terminated in 1960. From the total project approximately 79,000 bbl of oil were produced during thermal operations at a cumulative produced GOR of 23 Mcf/bbl. No appreciable change in the character of the produced crude was observed. Combustion in the reservoir was maintained with injected gas compositions ranging fronz 6 per cent oxygen in recycle gas to 100 per cent air. lnjectiotz of large quantities of recycle gas resulted in higher producing GOR's from offset wells than were measured from a pattern into ~vhich straight air ~vas injected. The air required to move the combustion front through I acre-ft of reservoir was computed to be 20 MMscf. This valrre was found to be relatively independent of the quantities of recycle gas injected. The recovery efficiency from the swept area was esti~izated to be about 59 per cent. Areas swept were similar in shape to tlzose obtained with a laboratory potentiometric model. Samples of sund taken from behind the burning front by coring indicated almost total oil removal from the sand. Petrographic analysis of the core samples indicated that the sand had been heated to peuk temperature of rlbout 1,200" F. No rignificant difference in peak temperature was forrnd in two areas where compositions of injected gas were quite different. Compression costs for thermal recovery were estimated to be $1,20/bhl of produced oil. INTRODUCTION The use of the "forward combustion" process as an oil recovery method has received a great deal of attention. This method involves ignition of the formation in an injection well, followed by propagation of a combustion front through the reservoir. Combustion is maintained by the injection of an oxygen-containing gas to react with reservoir hydrocarbons. As the flame front progresses through the reservoir, oil and formation water are vaporized, driven forward in the gaseous phase and recondensed in the cooler part of the formation. In turn, the condensed fluids push oil into the producing wellbores. Completed field tests of the process were first reported by Kuhn and Koch,' and by Grant and Szasz.' Results from other tests have since been reported by Walter,3 by Moss, White and McNeil,' and by Gates and Ramey." ach of these tests essentially utilized a single injection well surrounded by four or more producing wells. Sinclair Research, Inc., elected to do field experimental work using a number of test patterns in a single field in order that comparisons between various operating schemes could be made. The site selected and purchased in 1955 for this experimental work was a 60-acre Bartlesville sand reservoir located in Allen County, Kans. Combustion operations were initiated in mid-1956. Between that time and termination of the project in mid-1960, combustion fronts were propagated from injection wells to producers in three separate well patterns, using different mixtures of air and recycle gas. The test was terminated before sweep of the three patterns was complete so that information about the effect of combustion on the swept areas could be obtained by coring. Results from the test in the form of injection and producing well performance have been carefully recorded, and these form the general basis for this paper. DESCRIPTION OF RESERVOIR The reservoir in which the combustion tests were conducted is a Bartlesville sand "shoe-string", typical of a number of small reservoirs in Southeastern Kansas. Average reservoir characteristics are shown in Table 1. Fig. 1 is an isopachous map of the producing sand showing the reservoir to be approximately 400-ft wide and 2,500-ft long. Maximum net productive sand thickness is 21 ft. Fig. 2 shows a typical core analysis obtained by coring with water-base mud. The reservoir has no appreciable dip and is closed on the sides by degradation of sand into shale. The main body of sand is heavily laminated with shale stringers, which are not continuous between wells. The main reservoir is overlain by 30 to 40 ft of laminated low-permeability sand and shale streaks. No information is available on the original properties of

By D. C. Shelton

This paper discusses the metallurgical testing procedures and their applications at the laboratories and mills of Pima Mining Co. The author shows, using discussions and case histories, that routine sampling and evaluation of daily and monthly data can be as valuable as thorough laboratory investigations and mill test work. A great deal of valuable information can be made available this way. This paper discusses laboratory and mill metal-lurgical testing procedures and their applications at Pima Mining Co., located about 25 miles southwest of Tuscon, Ariz. The Company, incorporated in 1951 (the year the orebody was found) in California to operate the Pima mine, is jointly owned by Cyprus Mines Corp., Utah Contruction and Mining Co., and Union Oil Co. Milling started in December 1956. The main Pima orebody is a contact metamorphic vein of copper-bearing hornfels with an average dip of 45° to the south with an east-west strike. The hanging and footwalls are quartzitic pyroclastics (arkosite) containing some sulfides. Barren igneous intrusions are also present. Conglomerates and alluvial wash cover the deposit to a depth averaging about 200 ft. The principal copper mineral is chal-copyrite. Native copper, chalcocite, and minor amounts of chrysocolla and other copper minerals are also present, together with sphalerite and some molybdenite. Generally, it has been noted that the ratio of chalcocite to chalcopyrite is higher in the lower grade ores, while as the grade increases, there are only minor amounts of chalcocite. Open pit mining is employed at Pima, using both trucks and an inclined skiphoist system. A detailed description of Pima Mining Co. operations was presented by Spaulding et all. More recently another orebody to the north and east of the present pit has been discovered and will be placed in operation within a few years. This ore is of a porphyritic rather than a lode nature since values are disseminated in both igneous and metamor- phosed sedimentary rocks. The new orebody is expected to be harder and cleaner than the present deposit. The views expressed in this article are those of the author and do not necessarily reflect those of Pima Mining Co. nor of the profession in general. PlMA MILL The following outline of milling practice is presented to furnish a background for later chapters of this report wherein specific problems will be detailed and analyzed. Run of mine ore is crushed in three stages, using conventional equipment to 85% minus 1/2-in. for feed to two 10 x l6-ft rod mills operating in open circuit. Each rod mill discharge is split between two 10% x 13-ft ball mills in closed circuit with three 20-in. cyclones. In addition, a 7 x 12-ft ball mill with one 20-in. cyclone receives part of the discharge from one of the rod mills. This equipment was manufactured by Allis Chalmers and Equipment Engineers. The cyclone overflows are combined, sampled and, after joining the regrind cyclone overflow, are distributed into 12 10-cell Fagergren 66-in. flotation machines. The primary rougher concentrate produced

Jan 1, 1965

By B. D. Thomas

When the origin and early plans, of Battelle Memorial Institute were described in this journal in October 1929 by R. W. Gillett the first director, the doors of the laboratory had just been opened and the hope of the founder that the institution serve the world by research for industry still existed only as an ideal in the minds of the thirty or so employees who then constituted the staff. Since that time the country has experienced troublous times. A devastating depression plunged the entire world into an unprecedented period of financial and industrial inactivity; and the slow recovery of the late thirties was soon to be overtaken and overwhelmed in the violence of war. Through this time of stress Battelle has continued to grow. The staff has increased to almost 600; buildings and equipment have been added, until now it stands in a place of pre-eminence in the metallurgical and engineering world.

Jan 1, 1944

By Westwater, J. S.

The Mather mine of The Negaunee Mine Co. embraces nearly all of Sections 1 and 2, T47N, R27W. within the limits of the cities of Negaunee and Ishpeming on the Marquette iron range of Michigan's Upper Peninsula. Operations of the Negaunee Mine Co., the stock of which is jointly owned by Bethlehem Steel Co. and The Cleveland-Cliffs Iron Co., are under the direction of the latter company. The opening of the Mather mine was described by C. W. A1len. Mather "A" shaft was completed to an initial depth of 2352 ft in early 1943. The shaft capacity of over one million tons yearly is in prospect of being realized by orderly development and mining, and sinking of the second or "B" shaft then started in accordance with previous planning. Mather " B, " also a vertical shaft, was designed as a twin to Mather "A" shaft and the sinking technique developed at Mather "A" was utilized in "B" shaft operation.

Jan 1, 1949

By J. Kostuik, M. J. de Bastiani

The method of mining the uranium orebody at Denison mine in the Elliot Lake district of Ontario is discussed. Details of the development of haulage-ways and rooms are giuen, as well as methods of ore haulage. The milling and beneficiation stages are noted briefly. Located in the Elliot Lake mining district of Canada, the Denison mine lies between Sudbury to the east and Sault Ste. Marie to the west, Fig. 1. This area is overlain by Huronian sediments filling a basin in the Keewatin basement complex. Subsequent folding of these sediments developed a synclinal structure, known locally as the Quirke Lake Trough, which forms the northern portion of the now famous big Z structure, the geological trademark of this region. Plunging to the west and extending nine miles from the granite outcropping in the north to the Keewatin lava contact to the south, the syncline has dips of 10" to 45" on the north limb and 10" to 20" on the south limb, Fig. 2. The Denison orebody, located on the north limb of the syncline, lies between 800 and 3000 ft below the surface. The main ore zone is within 50 ft of the Missisagi-Keewatin contact (see Fig. 3), and it varies in thickness from 6 to 32 ft. This ore zone has a composite dip of 19" to the south and is composed primarily of two pyritized conglomerate beds separated by a band of pebbly quartzite stratum. The upper conglomerate bed of the main zone varies from 2 to 15 ft in thickness, while the lower conglomerate varies from 2 to 12 ft. Over extended areas, mining is carried out over the entire section comprising the two conglomerate layers and intercalated quartzite; in other areas, the conglomerates are selectively mined, depending on the thickness of the quartz-ite. The ore zone contains approximately 6 pct sulfides and 80 pct silica. The predominant uranium-bear ing minerals are brannerite, uraninite, and an unidentified mineral (probably thucolite). Some of the uneconomic metallic minerals identified to date are pyrite, pyrrhotite, and chalcopyrite. MINE LAYOUT Shafts: Underground exploration and development (Fig. 4) were performed through two shafts 1/2 mile apart. The No. 1 shaft, sunk to a depth of 1856 ft with the main station established at the 1600-ft ore level, is a 250-sq ft, 5-compartment shaft capable of handling 4500 tpd. Two 5x5-1/2-ft compartments serviced by a 12-ft diam double-drum hoist, are used for hoisting 190-cu-ft bottom-dump skips in balance. A 30-man cage, offset by a counterbalance, operates within a 5-1/2x11-1/2-ft compartment. The fifth compartment is used as a manway, pipe, and cable-way. This shaft is also provided with an intermediate pumping station used during the initial mining phase as well as loading-pocket and spill-pocket installations and pumping facilities. The 480-sq ft No. 2 shaft, located down-dip from No. 1 shaft, has a 2776-ft depth. Serving as the main hoisting shaft with a maximum capacity of 9000 tpd, it has eight compartments, four of which are used for hoisting ore in 10-ton capacity bottom-dump skips. These four compartments, measuring 5x6-ft in section, are serviced by two 12-ft diam double-drum semi-automatic hoists. A 50-man cage is operated in a 15x8-ft compartment, and it is counterbalanced by a 12-ton weight. Two compartments are used for ventilation, manway, electric cables, and pipes. A station was constructed at the 1325-ft horizon of the No. 2 shaft for relay pumping. The main sta-

Jan 1, 1961

By A. P. Svenningsen

IN the early stages of design it was not considered necessary that separate crushing plants be built for the new sulphide concentrator and smelter until sometime in the future. The plan was to use the existing crushing facilities for both oxide and sulphide ore. A few additions were contemplated for the existing plants, such as increased bin capacity, and possibly two new secondary crushing units. The more the problem was studied and discussed with the plant operators, the more it became evident that it was complex. It involved the classification of different kinds of ore from the open pit mine -sulphide, oxide and mixed-and how best to separate them so that each kind of ore was given the proper processing and treatment. It also involved the problem of keeping the different ores from being contaminated in bins, hoppers and chutes. Added to these, transportation became complicated and would involve additional handling and loading of ore from crushing plants to conveyors, to bins, and finally to railroad cars which were to be hauled to the concentrator and dumped into the fine ore bin. General In the early part of 1951 it was decided that the concentrator be constructed with ten grinding units instead of seven as originally authorized. The smelter was to be increased proportionally and naturally also the overall tonnages of ore to be handled by the new sulphide plant. Due to this increase in plant capacity and the larger tonnages involved, the difficulties which would arise by using the existing crushing plants were increased to a point where it became evident that the building of new crushing plants for sulphide ore exclusively was technically, as well as economically, advantageous. Authorization was, therefore, given by the company to construct new crushing plants to handle 30,000 tons of ore per day, and capable of reducing the run-of-open-pit ore to the proper size feed for the 10x14-ft rod mills in the concentrator. The ore, mined in the open pit, sometimes comes in pieces as large as 6 to 7 ft diam. The rod mills may call for ore crushed to 3/4 in. The large .size of ore delivered from the open pit determined that a 60-in. gyratory crusher be used as primary breaker. Such a crusher will have a capacity considerably in excess of 30,000 tons per day. The crusher will be a single discharge unit driven by a 500-hp electric motor through a tear coupling and a floating shaft. This type of drive has proven successful at a number of other crusher installations which our company has operating in the United States, Mexico and South America. The tear coupling will protect both the crusher and motor against damage in case of overload. No new features are incorporated in the design of the crusher itself, except that the, discharge chute is made the full width of the crusher with parallel sides instead of the usual converging sides. This change in detail should eliminate, a feature which has been a bottleneck in some of the operating plants and has caused loss of production due to ore hanging up and blocking the chute. The secondary crushing plants will have three 7-ft standard Symons cone crushers and six 7-ft short head Symons crushers. Between the primary and secondary crushing plants a coarse ore bin will be constructed with a nominal draw-off capacity of 30,000 tons of ore. The standard Symons and the short head Symons will be in separate buildings. All the crushing plants and the coarse ore bin are interconnected with conveyor belts for transporting the ore to the crushers at the tonnage rate desired. The final product of the new crushing plants is produced by the short head crushers. It will be delivered onto a conveyor belt leading to the top of the fines ore bin in the concentrator. A separate conveyor belt running the full length of the fines ore bin and provided with a movable tripper of rugged design will discharge the sulphide ore into the bin. The concentrator bin is planned and designed so that the installation of this additional conveyor will not interfere with the operation of the two railroad tracks on which crushed ore is brought from the existing oxide plant. Thus when completed the bin can be filled simultaneously by ore from the new crushing plant and by ore from the existing leaching plant.

Jan 1, 1952

By H. L. Gilbert, W. W. Stephens

Production of anhydrous zirconium tetrachloride by direct chlor- ination of a zirconium oxide carbon mixture in a silica-brick-lined chlorinator is described. Theory and thermodynamics of reactions are discussed. A pilot-model chlorinator and full-scale production equipment are described and operating data are included. ANHYDROUS zirconium tetrachloride required as a starting material in the Kroll process for production of ductile zirconium has been produced in this country principally by chlorination of the carbide or "carbonitride" made by reduction of zircon sand concentrates with carbon in the arc furnace. This process and the equipment used have been fully described.'.' It is well known that zircon or badde-leyite ores may be chlorinated directly with chlorine in the presence of carbon, and this one-step approach would seem at first glance to be preferable to the two steps involved in the carbide-chloride operation. As previously discussed1 considerations leading to adoption of the longer process were: 1—Chlorination of the carbide proceeds rapidly at temperatures below 500°C, whereas temperatures above 900°C were considered necessary for chlorination of zircon-carbon mixtures. 2—The highly exothermic nature of the carbide-chlorine reaction makes it self-sustaining, while heat must be supplied continuously in direct chlorination of the ore. 3—Silicon was thought to be chlorinated along with zirconium in the direct chlorination of zircon, leading to high chlorine consumption. In production of carbide in the arc furnace, silicon is driven off as silicon monoxide and does not enter the chlorinator. 4—For efficient direct chlorination, the ore must be finely ground and intimately mixed with carbon, and the mixture briquetted. 5—Direct chlorination requires a much larger chlorinator for a given production capacity than chlorination of carbide. Recently it has become necessary to produce large quantities of zirconium metal from a chemically purified zirconium oxide. Since the cost of the latter is high, the relatively high losses encountered in production of carbide in the arc furnace could not be tolerated, and a graphite resistor furnace5 was developed for production of carbide, which was then chlorinated in equipment previously used for chlorinating arc-furnace carbide. This method of operation was quite satisfactory, and losses in the carbid-ing step were minimized. However, operating costs were relatively high, and the process did not lend itself particularly well to large-scale operation because of the multiplicity of small units required and the hand labor needed to load, unload, and maintain the furnaces. To chlorinate the carbide, a vertical-shaft chlorinator was used in which the charge was heated with a central split graphite-rod resistor.' Operation of this chlorinator was somewhat less satisfactory with the resistor furnace carbide than it had been with the arc-furnace carbide due, principally, to the differences in physical properties of the carbide. The arc-furnace carbide is obtained as a fused metallic-appearing mass which can be crushed to -1/4 in. with production of a minimum of fines, whereas the resistor-furnace product is lightly sintered and produces a large proportion of fines in crushing and handling. These fines tend to pack in the chlorinator and promote channeling. which results in poor chlorine efficiency and low capacity. Data based on production of 22,000 lb of chloride from resistor-furnace carbide in this equipment are shown in Table I. Necessity for increasing production of chloride from about 2,000 to 15,000 lb per week led to further investigation of possible methods for direct chlorination of the oxide or oxide-carbon mixtures. Direct chlorination of the pure oxide presents much less difficulty than chlorination of zircon sand or oxide ores. The oxide is easily ground to —200 mesh, and no silica is present to cause excessive consumption of chlorine or contamination of the product. Theoretical Considerations The principal chemical reactions involved in chlorination of mixtures of zirconium oxide and carbon may be written as follows: ½ ZrO2 (c) + C(c) + Cl2(g) = ½ ZrCl,(g) + CO(g) [I] 1/2 ZrO, (c) + CO(g) + Cl2(g) = ½ ZrCl,(g) + CO2(g) 121 ½ ZrO2 (c) + ½ C(C) + Cl2(g) = 1/2 ZrCl4(g) + ½ CO2(g) [3]

Jan 1, 1953

By H. N. Wolcott

AMONG the many cases of replacement of one mineral by another, that of quartz or silicates by pyrite, or even other sulphides, is not uncommon, but the reverse of this process does not appear to have been recorded, even if it has been observed. It, therefore, seemed to the writer that the case described below was worthy of record. FIG. I.-GROUND MASS OF CHALCOPYRITE (C) CONTAINING ROUNDED GRAINS OF PYRITE (P) SHOWING DIFFERENT DEGREES OF REPLACEMENT BY QUARTZ (Q), AND ALSO QUARTZ VEINLETS CUTTING THROUGH THE PYRITE. (PHOTOGRAPHED BY REFLECTED LIGHT.) The specimens described in this paper were obtained by R. B. Morton of Idaho Springs, Colo., from the Old Town Mine, in Russell Gulch, and are in the collection of Cornell University, where the writer studied them. In the rough, the specimens appeared to consist largely of chalcopy- rite, with some grains of tetrahedrite. The polished surface of some of

Jan 6, 1917