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By W. O. Milligan, J. L. McAtee

THE constitution and structure of the hydrous forms of aluminum oxides and hydroxides have been studied in the laboratory of the Depart¬ment of Chemistry at Rice Institute for almost three decades. Progress reports have been made from time to time 13,17,18 summarizing and evaluating the existing state of knowledge concerning the complex inorganic system A12Oa-H2O. X-ray and electron diffraction analysis, together with dehydra¬tion isobars and isotherms, have been the most useful methods for increasing the understanding of the structure and interrelationships of the various forms of alumina and aluminum hydroxides. The purpose of this paper is to bring up to date the information now available, in the hope that a clearer understanding of the structures of these materials may prove useful in aiding mineralogists and geologists in studies of the genetic significance of the aluminum hydroxide minerals, and possibly of the closely related clay minerals. Early investigators found the system alumina-water to be confusing, since, prior to X-ray diffraction examination, it was not possible to clearly identify and characterize the various crystalline forms. One could not be certain as to whether or not different investigators were working with the same chemical individuals. In later studies, including some from the authors' laboratory, addi¬tional confusion resulted from difficulty in preparing pure samples of the various aluminum hydroxides and in obtaining relatively large crystals of dimensions above the colloidal range of particle size. Even at present, it is difficult to establish sets of interplanar spacings for the several forms of the aluminum hydroxides, as illustrated by the 29 different X-ray diffraction pat¬terns listed in a current publication. The nomenclature of the aluminum hydroxides is also extremely confus¬ing, inasmuch as two different systems are employed by large groups of in¬vestigators. Plans are being made to formulate a uniform system of nomen¬clature but in this report the designations employed will agree with previous papers from the same source.13,17,18 Four forms of crystalline aluminum hydroxides may be considered to be well defined and recognized, which, together with hydrous alumina "amor¬phous to X-rays," comprise an exceedingly complex inorganic system. Several

Jan 1, 1952

By James R. Stuart

AS the available remaining properties of iron ore reserves on the Mesabi Range are opened up for mining, the various properties located under lake beds are brought nearer an active status. The actual physical problems involved in stripping these properties do not act as a deterrent so much as the legal and political problems that are encountered. When it is proposed to destroy a natural lake that has been used by the public for many years, much local as well as state opposition may be encountered regarding its destruction. Public hearings must be held and some adverse publicity is likely to result. The ownership of the ore under the lake and the rights of the abutting property owners must be settled, and protection from damage caused by a disturbance in surface and subsurface drainage is likely to be demanded by property owners some distance from the proposed mine area. The Embarrass Mine, located near Biwabik, Minn., falls into this classification. A portion of the orebody lies under what was formerly Syracuse Lake, this body of water having been removed in the process of stripping the mine. An additional problem in the case of a meandered body of water is the establishment of a meander line that can be projected downward as mining progresses to form the basis for a satisfactory division between lake bed and upland ore shipments for royalty purposes. Fig. 1 illustrates the complications encountered in maintaining these divisions. A balance point was agreed upon in the center of the lake to make an equable division of lake bed ore to the abutting properties. The entire lake bed has since been adjudged the property of Minnesota. Lake Characteristics Lake bed stripping problems with which this paper is concerned necessarily are limited to a specific type of lake, namely the glacial lakes of the Lake Superior region. One characteristic common to these bodies of water is a deposit of fine black mud or silt on the bottom, frequently underlain by a layer of impervious blue clay. This is also true of the muskeg areas of the region, which present almost identical problems as lakes in stripping. The actual removal of the water and the lake bed material is a routine matter more or less standardized as to equipment, and the period of time required can be estimated easily on the basis of volume and capacity. More important than the foregoing is the execution of preliminary work, and above all, the timing involved. An account could be prepared based entirely on statistical and cost data which would give a very fair picture of the time required and cash outlay needed to effect the removal of a body of water preliminary to stripping the orebody. However, the real interest from the standpoint of the operator and the engineer who carry responsibility for completion of the job lies in the unexpected emergencies and the action of various materials involved in the stripping when the balance has been upset through diversion of water courses and the reduction of the lake level. Runoff and Drainage Lakes are located in natural basins that catch all the rain water and runoff water for a considerable area. Where a lake is involved having an inlet and outlet or a sizeable water course running through it, the drainage area may include a watershed covering many square miles. All available data then must be collected to supply a history extending over as many years for which information can be gathered on the flow of streams, annual rain and snowfall, and most important, the peak flows to be expected. Where the diversion of a stream around the stripping area is a part of the problem, this last factor is of great importance since it controls the cross-section to be selected for the diversion channel and the volume to be removed in its excavation, as well as affecting the hydraulic considerations to be met in the design of the completed channel. Characteristic material in the overburden found at the Embarrass Mine is illustrated in Fig. 2. Well Pumping Pumping from the well holes was started well in advance of the draining of the lake. Fig. 3 shows a gradual lowering of the water table with no noticeable fluctuations during the period in which the lake was being dewatered. Unfortunately, because of tight ground, a maximum flow to the wells was not maintained. This retarded the rate at which the water table was reduced so that in the course of stripping the excavation soon extended below the water table, and the great bulk of the pumping was handled from a system of sumps in the pit itself. Any dewatering program projected by prepumping from wells, a glorified well point system, would have to be started well in advance of the stripping to be of any great advantage. Preliminary drainage of the surface over the mine area is entirely apart from the actual elimination of the lake bed itself. Since the lake is what is called a perched water table because of the impervious character of the lake bottom, the adjoining surface may be dewatered below the surface of the existing lake and the flow will not be affected by the proximity of that body of water. This condition actually has been demonstrated through the establishment of a number of observation holes where a small churn drill was used to put down the holes and a 3-in. pipe was installed for taking water level

Jan 1, 1952

By George G. Thomas

IN April 1949, Newfoundland became the tenth province of Canada. It had been a Dominion of the British Commonwealth, though actual Dominion status had been suspended for some fifteen years. During that fifteen-year period the Government had been by commission. Throughout the regime of the Commission Government, a limited program of investigation of the mineral resources of the country was in progress. It was during this period that the investigation of the iron resources of Labrador by the predecessors of the Iron Ore Company of Canada began. Since Confederation there has been a marked change in the attitude of mining capital to the new province and a notable acceleration in the tempo of prospecting and exploration. There is no private Newfoundland capital available for mineral exploration. In the past, as now, capital must be attracted to the province. Prior to Confederation, Canadian capital had to cross the barriers of customs and immigration. American capital hazily regarded Newfoundland as part of Canada, with one notable exception, the American Smelting and Refining Company at Buchans. Buchans The most significant mining development in Newfoundland today is the property of the Buchans Mining Co. Ltd., a subsidiary of the American Smelting and Refining Co., located at the town of Buchans. Their milling operation produces lead, zinc, and copper concentrates by selective flotation, and also produces a small amount of gravity concentrate. Presently the mill treats, about 320,000 tons of ore per year. Production started in 1928 at 500 tons per day and has been raised over the years to the present figure of 1200 tons per day. Conventional prospecting methods uncovered the original deposit. It was first examined as a possible source of sulphur from pyrite. Geophysical prospecting and diamond drilling resulted in discovery of enough additional ore to warrant a large scale operation. Mining is by open cut methods. In addition conventional underground cut and fill stoping and square set stoping methods are employed. Between 850 and 900 men are employed at Buchans. Concentrates are shipped by rail to Botwood and leave the country through the port. Buchans Mining is currently investigating a Crown concession covering an area of approximately 6250 square miles, adjacent to the area of present operations. Aerial photographs have been made of the area, and geologic parties, supplied by plane, are in the field during the summer months. The concession lies south of the Falconbridge concession. The geology of the Buchans area can be divided into three groups. The first category is the oldest and called the Buchans Series. It consists of volcanic lavas, breecias and tuffs and underlies most of the Red Indian Lake Basin. The second is a group of igneous rocks such as granite and quartz porphyry intrusive in the volcanics. It makes for rough country north and west of Buchans. The third group consists of sills and dykes penetrating the lavas near

Jan 1, 1952

By Robert Glass Cleland

[CAPITAL EXPENDITURES TWENTY YEARS, 1934-1950 Morenci Branch $ 67,261,000 New Cornelia Branch 24,677,000 Copper Queen Branch 7,766,000 United Verde Branch 11,037,000 Miscellaneous Total Mining Division Refining Division 5,584,000 Fabricating Division 17,032,000 Total Capital Expenditures $135,823300]

Jan 1, 1952

By J. C. Allen, E. V. Herring

CHILE EXPLORATION CO's. oxide plant at Chuquicamata, the largest of its kind, started in 1915 with initial operations at the rate of 10,000 tons of ore per day. Anaconda Copper Mining Co. acquired the controlling interest in 1923 and in the next few years modernized the plant and mine, extending operations to a nominal capacity of 490 million lb of electrolytic copper yearly by 1942. Plant production averaging 495.7 million lb of copper per year was maintained for the eight-year period extending from 1941 to 1948 inclusive. During the World War II production reached its highest level when in 1944 the output was 531.8 million lb. The maximum production for any one month was 49.6 million lb, achieved in May 1947. E. A. Cappelen Smith, consulting metallurgist for M. Guggenheim's Sons, worked out the first process for the treatment of Chuquicamata copper oxide ore about 1913, and directed a staff of 20 young engineers operating a pilot plant at Perth Amboy, N. J., on three shifts for an entire year. The copper is extracted from the oxide ore by a hydro-metallurgical treatment described in detail by T. A. Campbell in AIME Transactions, Vol. 106, Copper Metallurgy, 1933. The Oxide Plant is divided into five main operations as follows: (1) Crushing, (2) Leaching, including dechloridizing and sulphur dioxide treatment, (3) Tailing disposal, (4) Electrolytic precipitation, (5) Smelting and melting. Crushing The present Primary Crushing Plant has been in operation 25 years, during which time 4,342,112 cars of 70 tons net capacity have passed through the dumpers. The ore as received from the mine is frequently blocky and up to 5 feet in section. It is reduced to 9-in. size by two 60-in. gyratory crushers, and transported to the east and west ore bins by 60-in. conveyors. In the secondary crushing operation the ore is first reduced to 2 to 3-in. size by five standard 7-ft cone crushers and two No. 10 gyratories. This product is then reduced to 3/8-in. size by fifty-two 48-in. vertical Symons disc crushers and is transported by belt conveyors to the loading bridges from which it is charged° into the leaching vats. Conveyor belt No. 14 which serves the eight vats in the east section of the leaching plant has transported over 125 million tons of ore, which is a record for any type of conveyor. This 60-in. belt travels at a speed of 600 fpm and carries loads up to 5500 tons per hr. Leaching The ore is leached in 14 vats, averaging 13,500 short tons net capacity each. These vats are lined with a natural mastic which adequately resists the sulphuric acid leaching solutions. The leaching process as presently conducted comprises soaking the ore with successive volumes of leaching solution and spent electrolyte and withdrawing successive volumes of enriched solution from the vat to a total of approximately 3.2 million gal. The first 1.6 million gal of this volume withdrawn is called strong solution and contains about 35 parts of copper per 1000. The balance of 1.6 million gal subsequently withdrawn is stored in a

Jan 1, 1952

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915; when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps: Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible: Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn-drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blasthole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated' by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Such shots were fired from either end .by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1952

By Shiou-Chuan Sun

THIS paper presents the design and operation of a frothmeter capable of measuring the frothing characteristics of pine oils and other frothing reagents. The experimental data show that the frothability of pine oil is governed by: 1-rate of aeration, 2-time of aeration, 3-height of liquid column, 4-chemical composition of pine oil, 5-pH value of solution, 6-temperature of, solution, and 7-concentration of pine oil in solution. The effect of mineral particles on the behavior of froth also was studied, and the results can be found in a separate paper.1 The results also show that the relative frothabilities of pine oils in the frothmeter generally correlate with those in actual flotation, provided that other factors are kept constant. In addition to pine oils, the other well-established flotation frothers were tested, and the results are included. In this paper, compressed air frothing is the frothing process performed by means of purified compressed air, whereas sucked air frothing is the frothing process accomplished by purified air sucked into the glass cylinder by a vacuum system. The term vacuum frothing denotes that froth was formed by degassing of the air-saturated liquid under a closed vacuum system. Apparatus The frothmeter, shown in Fig. 1, is capable of reproducibly measuring the volume and persistence of froth as well as the volume of air bubbles entrapped in the liquid and is capable of being used for compressed air frothing, sucked air frothing, and vacuum frothing. Fig. la shows that for compressed air frothing, the apparatus consists of an airflow regulating system, 1-3; a purifying and drying system, 4-8; a standardized flowmeter to measure the rate of airflow from zero to 500 cc per sec, 9; and a graduated glass cylinder, 13; equipped with an air regulating stopcock, 10; an air chamber, 11; and a fritted glass disk to produce froth, 12. The fritted glass disk, 5 cm in diam and 0.3 cm thick, has an average pore diameter of 85 to 145 microns. The pyrex glass cylinder has a uniform ID of 5.588 cm and an effective height of 63 cm. The inside cross-sectional area of the glass cylinder was calculated to be 24.53 sq cm, or 3.8 sq in. For sucked air frothing, Fig. lb shows that the apparatus for compressed air frothing is used again, with the following modifications: 1-compressed air and its regulating system, 1-3, are eliminated; and 2-a vacuum system, 16, equipped with a vapor trap, 15, and a vacuum manometer, 17, is added. The vacuum system can be .either a water aspirator or a laboratory vacuum pump. Any desired rate of airflow can be drawn into the glass cylinder, 13, by adjusting the opening of the air regulating stopcock, 10. The sucked air stream is cleaned by the purifing and drying system, 4-8, before entering the glass cylinder, 13. When this setup is used for vacuum frothing, the air regulating stopcock is closed. The frothmeter has been used for almost 3 years and has proved to give reproducible results, as illustrated in Table I. With a magnifying glass and suitable illumination, the frothmeter also can be used to study the attachment of air bubbles to coarse mineral particles.2 Experimental Procedures Except where otherwise stated, the data presented were established by means of the compressed air method. The volume and persistence of froth were recorded respectively at the end of 4 and 6 min of aeration at a constant rate of airflow of 29.3 cc per sec which is equivalent to 71.6 cc per sq cm per min, or 462.6 cc per sq in. per min. The aqueous solution for each test, containing 1000 cc of distilled water and 19.2 ± 0.5 mg frothing reagent, was adjusted to a pH of 6.9 ± 0.2. The volume of froth is expressed as cubic centimeter per square centimeter and is equivalent to the height of the froth column (the distance between the bottom and the meniscus of the froth). The volume of froth was obtained by multiplying the height of froth by the cross-sectional area of the glass cylinder, 24.53 sq cm. Before each test, the glass cylinder, 13, was cleaned thoroughly with jets of tap water, ethyl alcohol, tap water, cleaning solution, tap water, and finally distilled water. The cylinder with stopcock,

Jan 1, 1952

By Adam L. Wesner, A. C. Richardson

DURING World War II the demand for electrode carbon was greater than could be met by the supply of petroleum coke available for this use. It was believed that coke made from an extremely low-ash coal might be a suitable substitute for petroleum coke. Requirements for electrode carbon to be used for the production of aluminum are that the silicon plus iron content shall be not more than 0.14 pct of the coke. However, during the period when petroleum coke was in short supply, it appeared that relaxations of the specifications might be allowed to permit 0.4 pet iron plus silicon in the coke. For this work, the specifications were that the coal should contain less than 1 pet of ash, and the combined iron and silicon content of the coal should not exceed 0.28 pct. Heavy-liquid separations made on a number of coals showed that if certain size fractions of some coals were separated at low specific gravities the resulting float products would meet the required specifications. The objective of this work was to develop a continuous, commercially feasible process for producing low-ash coal. After consideration of various processes, the method chosen was a float-and-sink separation using a solution of calcium chloride as the separating medium. Eagle Seam coal was used for the experiments. The -10 +35 mesh fraction was found to be the most promising feed for producing the low-ash [ ] content coal. A batch of 30 tons of -3/8-in. coal was screened at 10 and 35 mesh; the yield of -10 +35 mesh was 20 pct of the feed. This fraction comprised the feed for the separation tests. In spite of the fact that this portion was screened a second time at 35 mesh, the coal still contained 16 pct that was finer than 35 mesh. Three-fourths of the undersize was in the -35 +48 mesh range. A representative portion of the -10 +35 mesh fraction was subjected to batch float-and-sink separations to determine the best gravity to be used in the continuous tests to obtain a high recovery of low-ash content coal. The heavy liquid used was a mixture of carbon tetrachloride and benzene which assured complete wetting of all of the particles. As usual in batch separations, adequate time was allowed for each separation so that even the near gravity particles had ample time to separate. The separations were made at increments of 0.01 sp gr from 1.25 to 1.28 and also at 1.32 and 1.59. Each specific-gravity increment was assayed for ash content. From these data it was concluded that 1.27 sp gr was best for the continuous separation of the coal. The products from the batch separation were composited into float 1.27 and sink 1.27 sp-gr fractions; these were assayed for ash, iron, and silicon. Table I shows the results of the batch separations. The composite float 1.27 fraction contained 66.04 pct of the total weight, and had an analysis of 0.79 pet ash, 0.073 pct iron, and 0.112 pct silicon, or 0.185 pct iron and silicon combined. These data are indicative of the results possible if a perfect separation is made at 1.27 sp gr under static conditions. Inasmuch as the method under investigation is a dynamic system these results could not be obtained. It remained to be determined, however, just how close to the theoretical results the actual continuous separation would be.

Jan 1, 1952

By W. D. Keller

THE high-alumina clays of Missouri are interesting for several reasons. They are the only sedimentary deposits in North America of first grade (over 70 pct A1,0,) diaspore and boehmite sufficiently large to be mined commercially. Mineralogically, they are composed of diaspore and boehmite in intimate coexistence, but gibbsite, the common hydrated alumina mineral in bauxite and laterite, is absent, or is exceedingly scanty if present.* Therefore, the Missouri high-alumina clays do not represent typical laterization. The physical chemistry leading to their mineral genesis, and the geolbgical setting for their formation, present important problems for scientific study. The origin of the Missouri high-alumina clays has been discussed in earlier reports by Allen1 (1935) and McQueen 8, 9 (1929, 1943). These investigations were made under difficult conditions, however, because no diaspore deposit had been opened where a Paleozoic (Pennsylvanian) rock cover lay undisturbed over the high-alumina clay. The relationship of the clay to only the underlying rocks could actually be observed. Lacking the guidance of critical postdepositional geologic evidence, undue and erroneous emphasis was placed on the postdepositional leaching of the deposits in the theories on the origin of diaspore and boehmite from clay (hydrous aluminosilicates) . With new geologic evidence at hand, a clearer picture can be made for the origin of the Missouri high-alumina deposits. In this paper the geology of a diaspore deposit whose original Paleozoic cover had not been eroded will be described and interpreted genetically. The close relationship of boehmite and diaspore (also with kaolinite and a chloritic mineral), will be established. A theory of origin for the high-alumina clays, consistent with the new data, will be postulated.

Jan 1, 1952

PAST PRESIDENT WILLIS MCGERALD PEIRCE New York City PRESIDENT-ELECT AND TREASURER ANDREW FLETCHER New York City VICE-PRESIDENTS O B J FRASER, New York City J B MORROW,' Pittsburgh, Pa. J L GILLSON, Wilmington, Del L F REINARTZ, Middletown, O PHILIP KRAFT, New York City W THOMAS,' Tucson, Ariz.

Jan 1, 1952

Without having asked for it, the writer has apparently been drawn into the discussion regarding relative hoisting capacities which was started by Woodward Iron, egged on by Joe Haller of Cleveland-Cliffs [Mining Engineering, August p. 642] and Axel Kjelgaard [November p. 919], who took exception to Mr. Haller's remarks of tonnage hoisted and avers that they do better in Sweden. It does not appear to the writer that Mr. Haller was inclined to refer to the tonnage hoisted at Cleveland-Cliffs in a manner to imply that they claimed any hoisting record, but rather as a statement that they had hoisted a certain amount of ore during the year 1950. Mr. Haller has been in the mining game long enough to appreciate what Mr. Kjelgaard did not apparently take into consideration, that length of haul, shifts hoisted, nature of ore, method of loading and the many other factors which influence the hoisting cycle make a comparison of tonnages hoisted of two given mines almost an impossibility. Mr. Cheyette sent you .a copy of his letter to Bob Manegold relative to the capacities of the Orient #2 mine. There are several mines around the Country which constantly set up records about which little or nothing at all is ever heard. We cite the Snow Hill Coal Co. mine at Terre Haute, Ind., which has an automatic Nordberg hoist and which hoists 12 tons per trip or 700 tons per hr from 700 ft. This is a counter-weighted installation and a complete cycle is made every 62 seconds or 58 trips per hr. Another mine, The Potash Company of America mine at Carlsbad, N. Mex., has a Nordberg automatic hoist, hoisting eight tons per trip from 1150 ft on a one minute cycle. The materials hoisted in these two mines are ideal for automatic hoisting-they are free flowing so that no hang up is anticipated and hoisting is continuous. Multiplying out the above figures you will find that it is possible to hoist 11,000 tons in two shifts at Snow Hill or over 3 million tons in 300 days per year and over 2 million tons at Potash. Other mines have records which can be pointed to with pride. When a record for one year is established, it doesn't stand too long as here in America the ambition of all the men connected with the operation forces them on to larger tonnages. T. C. WIEDENHOEFER MANAGER MINE HOIST DEPARTMENT NORDBERG MANUFACTNRING CO. MILWAUKEE, WIS.

Jan 1, 1952

By O. C. Blade

A comparison of certain properties of the commercial aviation gasolines produced in the United States during October 1949, 1950, and 1951 is presented in table 1, the 1951 data being average values from tables 3 through 6 of the present report. Trends in the average values of certain characteristics of these fuels may be seen in figures 1 through 4 which also indicate the minimum and maximum values. As may be observed, with the exception of changes in kinematic viscosities, the properties of commercial aviation gasolines have not varied greatly during recent years. A list of aviation-gasoline survey reports published by the Bureau of Mines is given on page 6. A summary of the average values of the properties on jet fuels is shown in table 2.

Jan 1, 1952

By G. B. Shea

In the search for possible future sources of petroleum products, the recovery and utilization of bitumens from the large outcropping deposits of bituminous sandstones in the United States have been under study by the Bureau of Mines for the past several years. The results of laboratory research in the adaptation of the hot-water separation process to recovery of bitumen from bituminous sandstones were reported in an earlier publication, on the Edna deposits near San Luis Obispo, Calif. This publication included analytical data on the cracked products and asphalts obtained from the extracted bitumen and on the cracking and distillation of the bitumen. Comparable hot-water separation tests have been run in the San Francisco, Calif., laboratory of the Bureau of Mines on quarry samples from the large deposits of bituminous sandstones near Vernal and Sunnyside, Utah. The bitumens were isolated from the extracts of the separation plant and analyzed at the Petroleum and Oil-Shale Experiment Station of the Bureau of Mines, Laramie, Wyo. Part I of this report summarizes the general characteristics of the Vernal and Sunnyside deposits and discusses the hot-water separation process as applied to those bituminous sandstones and the results obtained. The separation tests showed that these bituminous sandstones also respond to treatment by the hotwater method of separation. The bitumen in the Vernal sandstone was extracted more readily by the action of hot water than was the Edna bitumen, which greatly simplified plant operation. The Vernal sandstone contains no water-soluble iron compounds, such as were found in the Edna sandstone, and, therefore, required no preliminary water washing. Recovery of bitumen from the Vernal sandstone was about 96 percent.

Jan 1, 1952

J. Chipman (Massachusetts Institute of Technology, Cambridge, Mass.)—The fact that the experimental work has been applied to copper rather than iron and that the paper is presented to the Iron and Steel Division, I regard as rather significant. It shows the unity of metallurgy and the fallacy of trying to cut it up by metals. This result for the solubility of sulphur in molten copper correlates with Professor Schuhmann's finding that the published data on the other side of the copper-sulphur miscibility gap are also in error. I should like to ask the author to say a little bit more about the sulphur capacity of the slag. T. Rosenqvist (author's reply)-—I hope that Dr. Chipman will find the derivation of the expression for sulphur capacity more clearly explained in the printed version of the paper than in the oral discussion at the meeting. I feel that this quantity, which actually is the ratio of two activities, can be measured more easily than the individual activities. Even if the ratio CaO/CaS is chosen as the standard state, the expression can be used for any slag, even for slags completely free of lime, and it represents a way to put the desulphurizing power of all slag constituents into one bag. Some doubt has been expressed as to whether oxygen ions really exist in calcium oxide and in molten slags. From a thermodynamic view point that question is of minor importance. The term oxygen ion activity, or any activity for that matter, is defined rigorously by the equation: activity = exp u/RT, where u is the change in free energy connected with the transfer of one mol of ions from the standard state into the slag. Whatever happens to the ion in the slag is of no concern to the thermodynamicist. Regardless of whether the ion is "free" in the slag or not, or whether it is present in a very small amount, its activity can always be expressed, and for a thermo- dynamic calculation that is all we need. However, ionic activities will only be of some real value if they are simple functions of the slag composition, or can be measured easily. Concerning the real nature of the oxygen in the slag, my feeling is that the oxygen atom has a rather multiplex nature depending on how strongly it is tied by covalent forces or polarized by the other atoms or ions present. The oxides of iron, cobalt, and nickel differ from calcium oxide and blast furnace slags as to the amount of free electrons that can give rise to electronic conductivity. In slags we know that the conductivity is mostly ionic. The fact that reversible emf's can be obtained with oxygen electrodes in certain salt melts, indicates a significant amount of oxygen ions in these melts. But extended work, e.g. polaragraphic studies and measurements of transference number; are needed to obtain quantitative information about the real structure of the slags. D. E. Babcock (Republic Steel Corp., Youngstown, Ohio)—-With reference to the ion, it might be well to remember Dr. Moses Gomberg. All of his life he had no use for the ionization theory and he contributed greatly to the field of chemistry on the assumption there was no such thing as ions. I do not think we have to worry about whether the oxygen is ionic or not. I think one thing specifically should be brought to your attention and this I think is one of the important contributions of Dr. Rosenqvist. He pointed out what we know as oxygen potentials or what is described as oxygen potentials. I have used this concept for a long period of time and I want to state that if this concept is properly applied, it vitiates much of what we have in the literature, or makes our usual ideas regarding oxidation seem primitive. That one thing is more valuable than almost all the rest of the discussion as a fundamental basis on which to build a reasonable in-

Jan 1, 1952

By L. W. Kelly

The purpose of this circular is to show that, under the conditions prevailing in the Stony Point mine, wooden bolts were just as satisfactory as steel bolts or conventional timbering, with a reduction of approximately 80 percent in cost of material. This one fact alone is enough to justify an experiment with wood. This circular will show that production records have remained constant with both wooden and steel roof bolts. INTRODUCTION The successful use of steel roof bolts is now an accomplished fact in the coal industry, but the use of wooden bolts in place of steel is a rarity. The first experiment with wooden roof bolts was mane in 1949 in the Rio Verde mine, Norton Coal Corp., Nortonville, Ky.2/ This experiment was successful; the company obtained excellent results, but operations in the mine were suspended before enough time had elapsed to prove the value of wooden roof bolts conclusively. It is obvious that, if wooden roof bolts could be used in place of steel bolts or conventional timbering-methods, every mine operator would use word, as the cost is only a fraction of the other two methods. The Stony Point experience has proved conclusively that, under certain conditions and with a certain type of roof, wooden roof bolts can be used successfully.

Jan 1, 1952

By F. S. White, E. L. Kinsella

THE solids fluidization process brought out by the Standard Oil Development Co. in the early forties for catalytic cracking of petroleum enabled rapid transfer of large quantities of heat from gases to solids under closely controlled temperature conditions. As the technique developed, it was natural that applications for the process outside the petroleum industry would be sought. The calcination of limestone to ' quicklime requires the transfer of a large quantity of heat at high and closely controlled temperatures. For these reasons, this possible application was one of the first to be investigated. Laboratory work conducted by the Dorr Co. indicated that lime produced by solids fluidization would be of exceptional quality. Calculation of theoretical heat balances showed that savings in fuel could be expected. Accordingly, a pilot plant of 10 tons per day capacity was erected by the New England Lime Co. at Adams, Mass., to substantiate the process. Operation of the pilot plant, previously described,1 showed that excellent lime could be produced with a saving in fuel. The work also indicated the necessity of preparing feed to remove the finer fractions if dust losses were to be held to a tolerable level. A single-stage fluidized solids drier-stripper was developed to accomplish this sizing.2 Subsequently a commercial lime reactor was erected at Adams, beginning operation in June 1949, followed by air sizer facilities completed in November 1951. This paper will describe these two applications of solids fluidization. The raw material for preparation of lime by the fluidizing process is a high calcium crystalline lime stone having the approximate chemical composition shown in Table I. The stone is first prepared by crushing and grinding in conventional equipment to pass a 6-mesh sieve. After crushing, the material is sent to the FluoDry unit for drying and dedusting. Fluidized Drying and Sizing As shown in Fig. 1, the FluoDry unit consists of a 1/4-in. cylindrical steel shell, 9-ft ID by 22-ft 6-in. overall height. Attached horizontally at the bottom is a 6-ft diam combustion chamber, approximately 7 ft 6 in. long. A refractory constriction dome, consisting of first-quality fire brick shapes, partitions the inside into two sections. The lower section is called the windbox and the upper the bed compartment. The vertical wall of the shell is lined with 6-in. first-quality rotary kiln blocks. In the bottom of the windbox a hoppered section is formed with castable refractory to permit the windbox to be readily cleaned via a 6-in. screw conveyor. The combustion chamber is similarly lined with refractory shapes. On the 6-in. diam section, 4 1/2 in. of insulating brick, HW-26, are used. The inner combustion chamber is lined with 4 1/2-in. Korundal brick to withstand the high flame temperatures encountered. An 8-in. steel feed pipe, entering vertically through the top and fitted with an air-operated cone valve,

Jan 1, 1952

HIGH-GRADE manganese ore, from which manganese is obtained commercially, is not found in large quantities in any major steel-producing nation in the free world. The U. S. is a "have not" nation with respect to deposits of directly mineable high-grade manganese ore. Known resources of 48 pct Mn or better grade ore amount to less than 200,000 tons. In 1950 the U. S. steel industry consumed 1.8 million short tons of metallurgical grade manganese ore that contained about 800,000 tons of manganese. About 16 pct of the manganese content was lost in processing, so that about 650,000 tons, or 13 pounds per ton of steel actually entered into steel production. Under present practices use expands directly with steel output, and by 1975 the demand in both the U. S. and the rest of the free world is expected to be roughly 60 pet greater than in 1950. In peacetime about 80 pet of manganese consumption goes into steel production; high-manganese steel, dry cells, and chemicals account for the remainder. The manganese supply problem centers around high-grade ore for ferromanganese production. Use of ores containing less than 35 pet Mn sharply increase the costs of making ferromanganese. Use of ferro-manganese of grade below 70 pet in turn requires changes in steelmaking that increase steel cost. Under normal conditions the present small domestic production cannot be expected to increase. Major resources in the U. S. consist of 12 low-grade deposits. The cost of mining and treating these ores to extract a product as good as that yielded by imported ores is at least twice and in some cases more than four times the 1951 price of foreign ores delivered to the U. S. However, as long as trade relations and overseas shipping are not interrupted, deposits in India, Africa, and Brazil can meet steadily increasing demand at approximately present costs. Cost considerations indicate that the U. S. should continue to rely upon overseas sources for its peace-time supply, and that this situation is satisfactory. But, this does not take into account the question of how the U. S. will be able to meet its needs in war. Position of the Rest of the Free World In 1950, free world steel producers outside the United States, with a steel output of 70 million ingot tons, consumed about 1.3 million tons of metallurgical-grade ore. Their manganese ore demand, expected to increase directly with steel production, will by 1975 be about 2.3 million tons. Russia possesses over half the known manganese ore reserves of the world and is producing twice the tonnage of any other country. It supplied more than a third of the U. S. manganese requirements up to 1938 and again in 1948, but by 1950 Soviet manganese exports to the free world had virtually ceased. The free world's supply of manganese now comes mainly from India and Africa. Somewhat over 10 pet of U. S. imports came from Brazil and Cuba. Security Considerations In the event of war the U. S. might be substantially cut off from 90 pet of present sources. Reduction in manganese specifications might cut consumption by over 10 pet without seriously affecting steel quality. By elimination of losses in the production of ferromanganese savings as high as 10 pet might be possible. But, wartime manganese requirements cannot be met through conservation alone. To meet possible future emergencies the U. S. should continue its comprehensive security program for manganese, including stockpiling and research on the economic use of low-grade ore, domestic ores, the recovery of manganese from slag and the reduction of manganese requirements in steel production. If this work, including additional pilot plant operation is pursued vigorously, it should be possible in an emergency to get an adequate supply of manganese from domestic sources. The national stockpile then can be looked upon as a source of supply during the period of at least 2 years required to reach full-scale production from low-grade resources. Ferromanganese Smelting In comparison with smelting of pig iron, ferro-manganese smelting is a very wasteful process. Under present ferromanganese blast-furnace smelting practice, about 8 pet of the manganese in the furnace charge is lost to the slag, and roughly the same amount is lost to the stack gases; the total loss approaches 15 pct. Present practice is a compromise between excessive slag loss and excessive stack loss. In fact, it may be seriously questioned whether conventional blast furnace design is suitable for manganese smelting. U. S. Resources The known manganese deposits of the U. S. contain a total of 3500 million long tons of raw material and 75 million long tons of metallic manganese. More than 98 pct of this contained metal is in 12 large low-grade deposits of which the most important are those at Chamberlain, S. Dak; Cuyuna, Minn.; Aroostook County, Maine; and Artillery Peak, Ariz. Reserves of high-grade ore (48 pct Mn) amount to less than 200,000 tons. About 20 million tons of ore average over 15 pct Mn, and when grade is decreased to 10 pct Mn reserves amount to about 100 million long tons. If cut-off grade is decreased to 5 pet Mn, resources amount to 800 million long tons. Many of these low-grade ores may be beneficiated by flotation or other concentration methods. Pyrometallurgical Methods For smelting ferromanganese, it is essential to have an ore containing at least 50 pct manganese, with an Mn:Fe ratio of about 8:1. Direct smelting of 20 pct Mn concentrates is not promising. The only method that offers any promise involves two-step smelting.

Jan 1, 1952

By J. P. Newell, R. W. Storey

AUGER mining is a form of continuous mining in that it completely replaces with a one-cycle operation the older conventional cut, drill, shoot, and load method of mining. Relatively new, having been utilized only within the last few years, it is even more continuous in its production of coal than some other forms of continuous mining being practiced in this country today. There are at present coal recovery .augers in operation working along strip-mined highwalls from the outside. This paper will deal with underground mining, specifically with an experiment which was carried out with a machine developed for this purpose. The location of this experiment was in Pike County in eastern Kentucky in the upper split of the No. 3 Elkhorn seam. Coal height was 32 in. and roof conditions were very bad, with 10 in. of soft draw slate coming down with every cut in conventional mining. The coal is high-grade bituminous, excellent for metallurgical coking. Underground auger mining fills a need for a method of recovering coal in thin seams, under bad roof conditions or in split seams. In many cases it might save money by eliminating the handling of excessive quantities of rock. In fact, if the coal can be recovered, as in the case of this experiment, without mixing with any impurities outside the seam, it may be possible to load or ship a satisfactory product with little or no preparation or cleaning expense. The need for some economical method for recovering coal from thin seams is becoming more pressing as the more profitably mined seams are exhausted. Basic Principles of Auger Mining The principles of the auger, or screw conveyor, have long been used in drilling small diameter holes for blasting and other purposes. These same fundamentals apply to large diameter drilling. However, considerably more attention was given to designing a cutting head that would not pulverize the coal and consequently would require a minimum of power to drive it. The head, as designed for this job, consists of a cylinder 25 in. in diam with steel bits set in 3 positions spaced around the forward rim. A conventional coal drill bit, at the head of a bursting cone in the center of the cylinder, trails the forward rim slightly, see Fig. 1. The bursting cone and auger connections are anchored to the cutter barrel by 3 arms spaced at 120 This head design proved to be very satis-

Jan 1, 1952

By Edward I. Salkovitz

THE determination of the orientation of metal single crystals and the studies of plastic deformation are greatly facilitated by the use of the stereographic projection. To draw a standard projection of the crystal it is necessary to have a tabulation of the angles between the various crystal-lographic planes of the particular crystal studied. Bozorth' has published such a table for the cubic system which can be found in many reference books; such convenient tabulation for the hexagonal metals, however, does not appear to be in the literature, although scores of workers have calculated these angles. Therefore, Table I of angles for magnesium, zinc, and cadmium has been compiled. Since the c/a ratio is different for each of these metals, three separate calculations are necessary. The angles 4 between (HKIL) and (hkil) were calculated by means of the formula: Hh+Kk+½(Hk + hK)1+~3aLl cos The c/a ratios were calculated from data given in the 1948 edition of Metals Handbook. Fig. 1 is a typical (0001) standard projection of zinc. For all practical purposes, it serves for cadmium as well, as can be seen from the table. For magnesium, however, it is necessary to construct another net. R. M. Bozorth: Physical Review. (1925) 26, 390.

Jan 1, 1952

By Wilbur Jurden

THE writer's first experience with a nonferrous reduction plant of great magnitude was at the Washoe reduction works of Anaconda some 35 years ago. Here was a plant which had been planned with remarkable skill and foresight considering the time and the state of development of copper-plant practice in the year 1902. The designer utilized topography to fullest extent to provide proper sequence of operations and, what is most remarkable, to leave adequate space for future developments, most of which at that time were unknown. However, the practice then was to locate the various units of the reduction works at the most advantageous points of the existing terrain with little regard for tramming or other auxiliaries and then connect these various units by the essential trackage, conveyor systems, piping, etc., as the need developed. This occasionally led to undesirable track arrangements, sharp curves, and steep grades, especially when it became necessary to extend various portions of the plant. Conveyor systems also became rather complicated, running as they did at various angles, and such items as piping and electrical distribution were often found to be in the wrong place, entirely inadequate in size, or awkwardly arranged for any kind of extension. This condition was not peculiar to Anaconda, for all copper plants at that time were built in the same manner and it was the constant association with these difficulties which, in the year 1925, influenced the layout of the Andes Copper Mining Co. plant. In that plant all trackage was laid out straight and level, all conveyors at right angles to each other with minimum length and number of transfers. All buildings were placed parallel and the main structures were complete for all purposes so that auxiliary buildings and dog houses would not be added later. Piping and electrical work was provided for in the original layout and carefully designed to avoid additions and alterations, and careful study given to every movement of material throughout the entire plant so that it would be accomplished with the least possible effort. Naturally it was hardly expected to attain all these objectives perfectly but our efforts did succeed in creating a plant which was unique and outstanding for its time-1927. It was also most gratifying to find that these design principles contributed to considerable savings starting right in the drafting room, carrying through the construction and ultimately yielding savings in operations and manpower. Not only that, but such a plant gives the observer an impression of symmetry and order, is more attractive to the workmen, and unquestionably eliminates many accident hazards. However, the Andes plant buildings were fitted to the existing terrain instead of having terrain created to fit the buildings-an item which we found advantageous to correct on the next large plant. At Morenci in 1939, all of the desirable features of the Andes plant such as parallel buildings, etc., were incorporated; but we went one step further-power shovels were brought in to make the terrain fit the reduction works. The result at Morenci is well-known and needs no elaboration here, but the success achieved by the design methods used for this and previous plants naturally influenced and guided the layout of the Chuquicamata sulphide plant which is the largest yet conceived. Chuquicamata Plant Design At Chuquicamata several factors not encountered previously complicated the problem to a great extent. The most desirable location for the smelter would allow smelter gases to blow directly into the open-pit mine already producing 60,000 tons of oxide ore per day and employing 1550 men. This, of course, would be a serious condition and, therefore, we were forced to move the smelter to a less desirable location but followed our previous experience at Morenci and made the terrain fit the job. The most difficult problem, however, was the provision for receiving various types of ore both by rail and conveyor. These consisted of: 1-Sulphide-bearing residue from the stockpile from which oxide copper had previously been leached. 2-Sulphide-bearing residue coming direct from the leaching vats. 3-Sulphide ore crushed at the existing crushing plants and hauled to the concentrator in cars. 4-Sulphide ore from the new crushing plant adjacent to the concentrator. 5-Sulphide ore obtained

Jan 1, 1952