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By Scott Turner

PRESIDENT SCOTT TURNER officiated as chairman of the opening session on mining methods, Monday morning, Feb. 15. The first paper was that of Max H. Barber on open-pit mining in the Lake Superior district, read by C. F. Jackson. W. L. Taylor stated that liquid-oxygen explosive had been tried but that the operators are satisfied with the explosives used; lb. of dynamite breaks one ton of rock. Although black powder is considered the most effective explosive, the factor of safety has decided in favor of high explosives. Lucien Eaton told of the Michigan Lime and Chemical Co.'s large L. 0. X. plant at Rogers City, Mich. Mr. Turner referred to this plant, also said that the Bureau of Mines is shortly to issue Bulletin 349 completely covering the use of L. 0. X. from its inception. W. D. B. Motter, Jr., gave some items concerning Chuquicamata, Chile. Twelve years ago the blasts moved 850,000 to 1,000,000 tons, but average shots now move 250,000 to 300,000 tons. One pound of L. 0. X. breaks 3.7 tons, against 3.1 tons by dynamite. Another speaker told of some large dynamite blasts in open pits, one of 360,000 Ib. in March, 1931, and another of 430,000 lb. to be shot in 1932. Tests have been made to determine the effect of large shots on the surrounding country. Mr. Taylor said that 4- and 5-yd. shovels were preferable to those of 8 yd.; their performance is better and cost is lower.

Jan 1, 1932

By AIME AIME

PLANS for the Semi-centennial Meeting have almost reached completion, although in any undertaking of such magnitude a few changes are always to be expected at the last moment. As worked out up to the time of going to press, the program of the meeting is shown below. The headquarters will be the Irem Temple, in Wilkes-Barre.

Jan 1, 1921

By KENNETH CITTISGHAM

During the year 1945, the total number of wells drilled in Ohio, including the. non¬productive wells, was 1034. For the 10-year period ending with 1945, the average completions per year were 1125, the least number being 910 in 1938, and the maximum 1561 in 1941. The proportion of dry holes in 1945 was 37 per cent?about the same as for the year 1944 but exceeding the 10-year average, which was 34 per cent. The number of oil wells completed exceeded the number drilled in 1944 by about 30 per cent, the average initial per well being the same in both years. Gas-well completions were approximately 10 per Cent less than in 1944, with the average initial per well about 5 per cent greater. The difference in number of oil and gas wells in the two years was in Berea and Clinton completions, as the number of oil wells increased about 30 per cent, owing to greater activity in the two sands named, while the number of gas wells decreased because of fewer completions therein.

Jan 1, 1946

By David R. Maneval, H. B. Charmbury

At the turn of the century, iron and coal were the keys to industrial prosperity. At that time, Pennsylvania was the leading mineral producer in the Country, producing 200,000,000 tons of coal in a typical post-World War I year. Close to 90% of the Nation's energy was generated by coal. Today, coal accounts for but 22% of the energy generated, leading to a drastic reduction in Pennsylvania coal output. In contrast with the post-World War I days, 1964 figures showed a drop in production to 93,846,480 net tons. This huge decrease has created a serious economic problem in a state so much dependent on the coal industry. Along with the reduction in coal production, employment has dropped considerably through these years (the number of miners alone has gone from 280,011 to 43,149), thus causing a heavy burden of supporting the unemployed on the taxpayers. As far back as 1939, coal operators and economists saw the problems on the horizon and the need for improvement. Realizing that research was a way to improve and stabilize the industry they demanded legislative action. The Legislature reciprocated with a small grant to The Pennsylvania State University with the provision that each State research dollar be backed with one from the industry. This initial grant laid the foundation for the further growth in coal research.

Jan 3, 1966

By Kennedy, Joseph E.

MANY things must be considered in coming to a decision as to what type of crushing and grinding equipment is to be used for preparing run-of-mine ore for concentration, amalgamation, flotation, or cyanidation. This paper is intended in no sense as a thorough discussion of these various factors, but rather contains a few observations based on a wide experience in the manufacture and application of milling machinery. One important governing factor is the tonnage to be dealt with. In large plants treating several thousands of tons in eight hours, a large jaw or gyratory crusher of substantial construction is standard equipment for reducing ore to 8 or 10 in. or finer. In open-pit work, using power shovels with 4- or 5-yd. dippers, a jaw crusher should have a receiving opening of 48 by 60 in., or 60 by 84 in., so that dynamite or hand sledging will not be required for breaking to a finer size before primary crushing. The coarsely broken ore is then conveyed by a belt to vibrating screens, the oversize from which is again conveyed by belts to secondary crushers having openings of 12 or 14 in., the product from which is conveyed to storage. From storage, the material goes to a finishing gyratory for reduction to 1/4 in. or finer.

Jan 1, 1936

By Carl M. Fellman

LABOR disputes caused considerable turbulence in the coal mining industry during 1946. As an outcome of these disputes, a definitely fundamental change in safety procedure was instituted: establishment of safety regultions for coal mines through the agency of agreements between the miners' union and the agency or agencies operating the coal mines. Although a temporary plan at present, the future will determine the. extent to which it may become permanent. Notwithstanding the turmoil in the industry extending through much of the year, the safety record for the first eleven months of 1946 was the best in the history of the industry for that period. The fatality rate for 1946 is lower than in any other year in the history of coal mining in the United States. This follows a pattern inasmuch as new low coal-mine fatality records were also made in 1945, 1944, and 1943. The five-year period 1942 to 1946 inclusive has a coal-mining fatality rate far lower than any other five-year period in the history of the industry. It is significant that this period exactly coincides with the operation of Federal coal mine inspection which got under way early in 1942.

Jan 1, 1947

By V. T. BERNER

THIS apparatus was designed primarily to meet the demand for a quick, efficient stopping to seal off the burning area temporarily during a mine fire, but it can be used in any circumstance where an immediate brattice is required for ventilation control. In using the old style brattice, mine-rescue workers have found it difficult, in smoke-laden drifts and stopes, to erect an effective stopping with wooden frame and canvas. Stulls must first be cut to the proper length, the canvas tacked in place, and the numerous irregularities of the walls and back must be caulked. This operation requires the employment of axes, saws, and hammers, which cannot easily and safely be handled in dense smoke. The fire fighters have often been forced to abandon their brattice, unfinished, because of the rapid approach of the fire, or because of the exhausted sup- ply of oxygen in their breathing apparatus. It is neces-

Jan 1, 1930

By C. C. Hawes

SILICA is the main impurity in iron ore. It is intimately associated with the iron oxide, sometimes free but more often in the combined state, as a mineral silicate. Its separation and purification sometimes present difficult problems to the analyst. In the absence of any definite knowledge to the contrary, orthosilicic acid. H4SiO4, is supposed to be the only silicic acid made, all other forms, such as metasilicic acid, H2SiO3, representing different degrees of its dehydration. These acids are exceedingly weak but form the most stable compounds found in nature. They are known as the silicate minerals and are formed by the combination of the different metallic oxides

Jan 1, 1936

By C. A. LINDBERG

Several new methods have been developed on the Iron Range to remove the material adhering to truck bodies in freezing weather. A machine known as a Gradall, incorporating the features of digging both above and below the horizontal, digging angle of blade control, full revolving, travel almost anywhere, and operated by one man, has proven successful in trials at body cleaning. Several of these units are in use with great saving in box-cleaning costs and improvements in efficiency of the truck operation with full capacity boxes. In addition to their winter use, the same machines have been used for ditching and excavating in the summer with great saving. The latest method of combating this freezing problem is to carry the exhaust from the engine through the body, making use of the tremendous heat that is now being wasted. To do this, the present wood sandwich which is used between the bottom of the box and the wear plate is removed and replaced with angles or channels run lengthwise of the box and spaced to permit gases to flow between or through them and also provide a good sup- port for the wear plate. The channels or angles are cut a little short on each end to form a cross passage that permits the gases to flow through the area. In the front end, openings are made in the wear plate and connect to a manifold created by welding a plate across the whole front of the box, spaced from the original end by the original vertical reinforcing angles at that point. The body being described is for a Euclid truck which is the only truck that has been adapted to this body-heating idea to date. At the rear of the truck, the holes from the cross passage are made in the bottom plate of the body and lead into a triangular-shaped body cross member under the scoop end of the truck.

Jan 1, 1949

By Davis Baker

The direct-metal cars or ladles of the Maryland Steel Company have a capacity of 18 tons when filled within 12 inches of the top. On account of this large capacity, the formation of skulls in these ladles is by no means as rapid as in smaller ladles; but a skull once formed is a very unwieldy piece of metal to handle, and especially to remove from the ladle. We have found it cheaper to clean ladles often rather than make very heavy skulls. We aim to keep the ring of chilled metal that forms on the surface of the iron chipped off, so that the skull shall form principally at the bottom until it becomes so thick as to render the ladle unfit for service. Sometimes it happens that a ladle may be held for several hours filled with molten metal. This invariably leaves a large amount of chilled metal in the ladle. Frequently the skull will weigh several tons, and will be of the full size of the ladle inside, with a top crust, having perhaps a hole 12 inches square, through which the molten metal was poured. In order to remove skulls that cannot be easily chipped out with a bar, we use what may be called a skull-extractor, consisting of a hydraulic cylinder mounted in the center of a heavy cast-iron plate, standing vertically on a suitable foundation. Figs. 1 and 2 show the construction and operation of this apparatus, Fig. 1 being a view of the extractor with pistons extended to full stroke, and Fig. 2 a view with ladle in position and pistons about to commence the stroke. in both figures, A is the hydraulic cylinder, B the vertical cast-iron plate, and C the foundation. The water-pressure, admitted through the pipe, P, is 750 pounds per square inch, and the inside diameter of A is 14 inches. There are two pistons: D, fitting the bore of the cylinder, and E, 84 inches in diameter, moving inside of D. Both pistons move together when

Jan 1, 1893

By C. WIFLARD HAYES

(New Haven Meeting, February, 1909.) I DESIRE to make it perfectly clear at the outset that I fully realize the hazardous nature of any attempt to estimate the quantity of iron-ore or any other mineral in the ground. I should not have had the temerity to venture upon an inventory of the iron-ore supplies of the United States except for the requirements of the Conservation Commission. During the preparation of this inventory I have been impressed by the wide diversity in estimates of those best qualified to. hold an opinion, and I am therefore prepared in advance for criticism of my figures. Inasmuch as this report was prepared in large part for unprofessional readers, I considered it necessary to classify and describe the commercial iron-ores in a somewhat elementary manner, in order to explain the reasons for the large element of uncertainty in the estimates. I have therefore given a chemical and a geological classification of iron-ores, which place the commercial deposits in six classes: (1) magmatic segregations in basic igneous rocks; (2) contact-deposits formed in connection with igneous intrusions; (3) concentration-deposits; (4) replacement-deposits; (5) bedded-deposits, and (6) gossan-deposits. The iron-ores are further distinguished under two classes available and not available. This classification is in large measure arbitrary, and it, is evident that it will vary front time to time. In the available class are placed those ores which can be worked at a profit under conditions essentially as they exist at present. Since it depends wholly on the question of costs, the various factors which affect cost will determine the classification of any particular deposit. Actual production, * Abstract of Report prepared for the Conservation Commission. t Chief Geologist, United States Geological Survey. .

Apr 1, 1909

Jan 1, 1899

By S. A. Mersol, C. T. Lynch, F. W. Vahldiek

The microhardness of single crystals of tungsten disilicide has been investigated by the Knoop method. The average random room-temperature hardness of the WSi, matrix was 1350 kg per sq mm. Hardness crnisotropy was noted with respect to plane and indenter orientation as determined by single-crq.stal X-rny studies. Annealing at 1600" and 1800°C decreased the average hardness to 1310 and 1230 kg per sq tnm, respectively, and produced a second phase identified by X-ray diffraction and electron-microprobe analysis to be wSio.7. Ball-impact experiwzents produced rosettes at 850°C. Optical and electron microscopy showed evidence of slip and cross slip and twinning produced by microhardness indentations. Prismatic (100), [001] slip was found and cor~elated with hardness data. THE present study was undertaken to investigate the hardness anisotropy of as-grown and annealed single crystals of tungsten disilicide. The existence of the silicide WSiz in the W-Si system has been well-established and its structure thoroughly investigated zachariasen2 found WSi, to have a tetragonal C type of structure, similar to that of MoSi, with lattice parameters a = 3.212A, Kieffer et al. studied the W-Si system and measured the density and microhardness (at a 100-g load) of both polycrystalline WSi, and WSi,.,. The values found were 9.25 g per cu cm and 1090 kg per sq mm for WSi,, and 12.21 per cu cm and 770 kg per sq mm for WSi0.7, respectively. According to Samsonov et a1.5 the microhardness of polycrystalline WSi2 is 1430 kg per sq mm (at a 120-g load). EXPERIMENTAL The WSi, single-crystal boules investigated in this paper were grown by a Verneuil-type process using an electric arc by the Linde Division of the Union Carbide Corp.6 The largest specimens were 8 mm in diameter by 16 mm long. The crystals had an average density of 9.01 g per cu cm with a tungsten • silicon content of 99.9 wt pct. The major impurities were: 87 ppm O, 41 ppm N. 54 pprn C, 500 ppm Zr, 50 ppm Na, and 50 ppm Mn. The crystals were silicon-poor, the average silicon content being 22.20 pct (stoichiometric value is 23.40 pct), and tungsten-rich, the average tungsten content being 77.70 pct (stoichiometric value is 76.60 pct). As-received single crystals were ground and analyzed by powder X-ray diffraction technique using Cu Ka radiation. Laue and layer line rotation patterns were obtained on cleaved sections of WSi, single crystals. Electron-microprobe traverses of representative crystals were done using a Phillips-AMR electron microanalyzer. Carbon replicas were used to prepare electron micrographs. This work was done with a JEM-6A electron microscope. Prior to the metallographic examination, the specimens were mounted in Lucite and then polished for short times on polishing wheels using 9-, 3-, and 1-p diamond-grade pastes. Finally they were fine-polished with Linde A powder for 24 hr on a Syntron vibratory polisher. The samples were etched with 4H 2 O:1HF:2HNO3, which is a medium fast-acting etchant. The combination 1HF:2HNO3:5 lactic acid is also a satisfactory etchant. Annealing runs for selected specimens were made at 1600" and 1800°C for 3 hr at 1.0 to 3.0 x 10-5 mm Hg. A Brew tantalum resistance furnace with WSi2 powder for setters was used. The WSi2 powder was the same as that used for the crystal growth. Temperatures were measured with a calibrated W, W-26 pct Re thermocouple and a microoptical pyrometer. Powder X-ray diffraction, emission spectrographic, and electron-microprobe analyses were done after the annealing runs. For microhardness measurements a Tukon Microhardness Tester Type FB with a Knoop indenter was used. Although measurements were taken at loads ranging from 25 to 1000 g, the 100-g load was chosen as the standard load. All measurements were taken at room temperature. Only indentations of cracking classes 1 and 2 were considered.' DISCUSSION OF RESULTS Powder X-ray diffraction analysis showed the as-received crystals to be single-phase WSi2. Laue and layer line rotation patterns obtained on cleaved sections of WSi2 single crystals proved them to be tetragonal WS 2 2 The results also indicated that the c axis of the crystal was oriented parallel to the boule or growth axis. Electron-microprobe traverses across the matrix of the as-grown crystals showed them to be homogeneous WSi,. Optical and electron microscopy of etched crystals, however, revealed that they contained minute amounts of the "golden" and the "blue" second phases as opposed to the "white" or WSi2 phase. These two second phases were concentrated in inclusion and etch-pit

Jan 1, 1965

By Meissner, C. A.

THE tendency all over Europe, just as it is with us, is to go to the use of turbines for new construction or replacement of old steam or even gas engines. 'The lower construction cost and the lower operating cost, when they can have high steam temperatures and pressures, together with their efficiency, is the cause of this change to turbines, both high and low pressure. The present trend is to equip plants with modern, efficient steam boilers, turbines, etc., and gradually work into high-pressure superheated steam equipment to accomplish a saving. One company has just completed a 375 to 400 lb., 700 deg. superheated steam installation and turbo-blowers with this end in view. Some of our new steam plants are using 350 lb. working pressure.

Jan 1, 1928

By A. C. Veatch

Jan 1, 1921

By D. C. DERINGER

METALLURGISTS and mill operators appreciate, in a general way, .the economic or commercial relationship between recovery and grade of product but few have correlated in detail fluctuating metal prices with the metallurgy. They know that many factors such as metal price, dedusiions for returning charge, transportation, percentage of metal in head ore, etc.; all affect the commercial relationship between recovery and grade of product. In practically every mill some one at some ' time has calculated the amount of recovery that can be sacrificed for each percentage increase in grade of concentrate and from this calculated ratio the commercial or ., economic tail has been established. Such a ratio finally becomes a well-entrenched rule by which the progress of the metallurgy is judged.. When once established, these-, ratios are but seldom corrected to changes of metal prices for each possible grade of concentrate that may be produced. The calculation is tedious when often repeated. Other methods have been suggested and used to correlate the commercial with the metallurgical phase. A factor .derived ,by multiplying the grade of concentrate by recovery is often used, but is of little importance because it does not possess a single commercial quantity. The percentage monetary recovery calculated by dividing the gross value of the ore into the net smelter return gives a good economic index and, is to be commended, but it is difficult to correlate exactly and quickly the work of each day on a commercial basis.

Jan 1, 1931

By G. M. Schwartz, D. M. Davidson

THE Duluth gabbro outcrops containing sulphides of copper, nickel, and iron are located on both sides of State Highway No. 1 an airline distance of 8.5 miles southeast of Ely in northeastern Minnesota. The region of known sulphide occurrences includes parts of sections 5, T. 61 N., R. 11 W., and parts of sections 25, 26, 32, 33, and 34, T. 62 N., R. 11 W. These sections, given in Fig. 1, are all in Lake County, Minnesota. Part of the area, which lies entirely within the Superior National Forest, is shown on the topographic map of the Ely quadrangle. The original discovery was made in 1948 when a small pit was opened in weathered gabbro rubble for use on a forest access road. A shear zone had caused unusual decomposition in this glaciated area, and the resulting copper stain was noted by Fred S. Childers, Sr., an Ely prospector, who began searching the outcrops along the base of the intrusive. He was joined in further exploration by Roger V. Whiteside of Duluth. In the summer of 1951 a small diamond drill was moved into the area and a hole 188 ft deep was drilled, passing through 11 ft of glacial drift into sulphide-bearing gabbro. This paper is a preliminary report on the geology of the newly discovered ore. The Duluth gabbro is one of the largest known basic intrusives and may be defined as a lopolith.1 It extends northeastward from the city of Duluth as a great crescent-shaped mass that intersects the shore of Lake Superior again near Hovland, 130 miles to the northeast, see Fig. 2. The distance around the outside of the crescent is nearly 170 miles. The form of the intrusive is simple at Duluth where it ends abruptly north of the St. Louis River; at the east end, however, the gabbro splits into two elongated, sill-like masses separated mainly by lava flows and characterized by minor irregularities. The outcrop reaches a maximum width in the central part where it is about 30 miles across, and a maximum thickness of about 50,000 ft. It may be significant that the sulphides occur at the base of the thickest part. The lopolith has segregated into rock types ranging from peridotite to granite. The most abundant types are olivine gabbro, gabbro, troctolite, anorthosite, and granite. Of lesser importance quantitatively are peridotite, norite, pyroxenite, magnetite gabbro, and titaniferous magnetite. Grout estimates that two-thirds of the gabbro at Duluth is olivine gabbro. Variations in the percentages of plagioclase, augite, olivine, and magnetite-ilmenite constitute the only essential differences found among the basic rock types. The predominant mineral is plagioclase, mainly labradorite. Plagioclase and olivine seem to have crystallized early, and the olivine rich rocks, usually troctolite, are found in the lower part. Segregations of titaniferous magnetite are abundant near the base of the gabbro along the eastern part and also occur far above the base. These have recently been described in detail by Grout' Near the top, segregation has produced a gradation to granite, or "red rock," as it is known locally. This consists of quartz, red feldspar, and hornblende. The red rock forms a. zone with a maximum width of nearly 5 miles but is quantitatively unimportant from Duluth northward for 35 miles. In Cook county, where the gabbro splits, each of the two sill-like masses has a red rock top somewhat thicker in proportion to the gabbro below than in the main central mass. The intrusive ranges from coarse to medium in grain size and from granitoid to diabasic in texture. Throughout much of the Duluth gabbro in Minnesota banding and foliation are well developed, as Grout has emphasized! The bands are mainly a result of variation in the percentage of minerals, as in troctolite with alternating bands high in olivine and in plagioclase. A few bands may consist largely of one mineral, as is true of some segregations of magnetite. Many of the banded rocks show a clearly developed parallelism of platy plagioclase crystals, and both banding and foliation are believed to conform to the floor of the lopolith. Throughout its extent in Minnesota the Duluth gabbro dips east and south toward Lake Superior. It is generally believed to extend beneath Lake Superior and is found as a smaller mass exposed along the north side of the Gogebic district in Wisconsin and Michigan. The dip at and near the base ranges along most of its length from 20 to 40°, but at places the internal banding dips even more steeply. The dip of the upper part is much less, and if it is assumed that the flows along the north shore of Lake Superior are a dependable indication, it does not exceed 15º. The formations shown in Table I which are intruded by the gabbro range from Keewatin to Middle Keweenawan in age. They present a significant picture. At the top, the gabbro and its accompanying

Jan 1, 1952

By K. P. Gupta, A. K. Pal, L. Chandrasekaran, U. P. Singh

The Mn-Sn binary system, investigated at the high-manganese end and between 500° and 1000° C, shows four phases at temperatures below 727"C, namely the u Mn, the p Mn, the Mn3 Sn, and the Mn, Sn phases, while at higher temperatures only the last three phases remain stable. The solubility of tin in a Mn is very small and the maximum solubility of tin in P Mn phase appears to be abollt 10 at. pct Sn. The solubility range of the Mn,Sn and the Mn, Sn phases is 23.0 to 26.0 and 37.0 to 40.0 at. pct Sn, respectively. The lattice parameter of the 13 Mn phase increases with increasing tin content. The Mn, Sn thase is hexagonal with and appears to be basically of the Ni3 Sn type structure except that for quite a few X-ray diffraction lines the calculated and observed relative intensities do not agree well. The Mn-Sn binary system has been studied by several investigators.''~ Their results indicate that three intermediate phases, namely, the Mn3Sn, Mn,Sn, and MnSn, phases, exist at high concentrations of tin. However, so far the proper phase equilibrium has not been established at the manganese-rich end and very little data is available for the composition range between pure manganese and Mn3Sn. Moreover, earlier investigators differ in their opinion about the exact composition range at which the Mn3Sn and Mn,Sn phases appear and some doubt has been cast by some investigators3 regarding the structure of Mn3Sn phase which has been reported to be isotypic with Ni3Sn (Mgscd type) structure. In this investigation an attempt has been made to establish the proper phase equilibrium between pure manganese and Mn + 50 at. pct Sn composition in the temperature range of 500" to 1000°C. PROCEDURE The raw materials used were from three different sources. For exploratory work five alloys containing 5, 10, 15, 20, and 25 at. pct Sn were prepared using 99.9 pct pure manganese and tin supplied by E. Merck & Co., Germany. The rest of the alloys and one more 25 at. pct Sn alloy were prepared using 99.9 pct Mn supplied by Gallard Schlesinger Chemical Mfg. Corp., U.S.A., and 99.999 pct Sn supplied by Semi Elements Inc., U.S.A. Weighed amounts of manganese and tin were melted in recrystallized alumina crucibles in an inert gas (argon) high-frequency induction melting furnace. By careful control of temperature and time of melting the losses were reduced to below 0.2 pct in all cases. Since the losses were very small, no attempt was made here to analyze the samples chemically. Alloys were wrapped in molybdenum foil and sealed in small evacuated fused silica capsules. The alloys were annealed at different temperatures, controlled within + l°C, for sufficiently long periods to attain proper phase equilibrium, and subsequently quenched in cold tap water. The annealing periods used at different temperatures were 15 days at 500°C, 7 days at 600°C, 5 days at 700" and 750°C, 3 days at 800°, 850°, and 900°C, 2 days at 968"C, and two alloys annealed at 968°C were reannealed for 10 hr at 1000° C. From each annealed specimen, a part was utilized for metallographic study while another piece was used for X-ray diffraction study. 1.0 pct HNO, solution and oxalic acid solutions of concentrations 0.05 to 1.0 pct were used for etching Mn-Sn alloys above and below the MnsSn composition, respectively. Since all alloys were brittle, X-ray specimens were prepared using the as-crushed -325 mesh alloy powders. Only one 25 at. pct Sn alloy powder was reannealed in an evacuated silica capsule at 800°C for 5 min and water-quenched. X-ray diffraction patterns . for the Mn3Sn phase with the as-crushed and the reannealed powders did not show appreciable change. Un-filtered iron radiation at 25 kv, 15 ma was used with either Norelco 114.6-mm-diam Debye Scherrer Camera (for phase identification) or Norelco 12-cm-diam symmetrical focusing camera (for lattice parameter determination of the 6 Mn phase). The estimated accuracy of lattice parameter determination for the focusing camera was * 0.001A. RESULTS AND DISCUSSIONS The results of metallographic and X-ray diffraction study made with different alloys are shown in Fig. 1. The variations in lattice parameter with composition for the p Mn and the Mn3Sn phases are given in Tables I and 11, respectively, and the lattice parameter as a function of composition for the p Mn phase is shown in Fig. 2. The lattice parameter of the p Mn phase increases with increasing tin content while for the Mn,Sn phase the data obtained from two two-phase alloys and one single-phase alloy indicate increase in a,, and decrease in c, parameters with increasing tin content. The results, Fig. 1, indicate that the solubility of

Jan 1, 1969

By Thomas P. Forbath

THE sudden realization that known sulphur reserves amenable to mining by the Frasch hot water process are nearing exhaustion focused attention on widely scattered surface deposits throughout the world. These deposits are not necessarily of lower sulphur content than ores located underneath Louisiana or Texas salt domes which usually average about 30 pct sulphur disseminated in limestone matrix. Their near surface occurrence, however, renders exploitation by the Frasch process impossible. As is well known, the Frasch process depends on the presence of 500 to 1000 ft of overburden and cap rock above the sulphur deposits to permit melting underground sulphur in place by diffusing hot water under pressures of 200 to 600 psig in the formation and raising the molten sulphur to surface by air lift. This process renders possible the production of pure sulphur which is 99.5 pct pure without any subsequent treatment. Surface deposits contain sulphur in the same range of concentrations as the salt dome deposits, i.e., from 10 to 50 pct sulphur, associated with various gangue materials such as silica, limestone, and gypsum. The pirincipal distinction, then, does not lie in the percentage of sulphur contained in the ore, but in the geological nature of the deposit. A recent study' of the world sulphur supply situation estimated 1950 sulphur production in the free world countries at 5.6 million long tons, of which the United States produced 5.2 million tons, or 93 pct of the total. While America's domestic needs alone would have been covered by national production, about 1.4 million tons were exported during the same year. Despite all the steps which are being taken to restrict use of elemental sulphur and to force the fullest possible development of alternate sulphur sources here and abroad, the deficit in elemental sulphur production will probably increase with time. As a result of intensive prospecting for oil throughout the Gulf Coast area discovery of significant new salt domes is held unlikely. With the growing scarcity of sulphur and what appears to be an inevitable rise in price, recovery from deposits not amenable to Frasch-process mining assumes greater economic importance. Untapped Reserves The most important deposits in this category are located in Sicily, where elemental sulphur occurs in Miocene limestone and gypsum formation. Sulphur content of these ores ranges from 12 to 50 pct with an estimated average of 26 pct. Although quantitative estimate of these reserves is not available it is held that they exceed 50 million tons of sulphur. Similar deposits occur also on the mainland which contribute about one-third of Italy's total current annual production of 230,000 tons, but these are known to be nearing exhaustion. Significant surface deposits of volcanic origin are located in South America, Japan and western United States, silica being characteristic gangue con-stituent. The largest of these deposits are in South America. More than 100 extend over a zone 3000 miles long, paralleling the west coast of South America. 'Total sulphur content of these deposits has been estimated to be as high as 100 million tons. The main islands of Japan also possess at least 40 known volcanic sulphur deposits with probable reserves of 25 to 50 million tons.' Prospected reserves in western United States might amount to 2 million long tons, principal deposits being located in the northwestern part of Wyoming, southern Utah, and eastern California. Volcanic deposits of lesser importance are found around the Mediterranean, in Turkey and Greece, and in Africa, Egypt, Abyssinia, and Somaliland. Beneficiation Methods Different methods of beneficiation have been used in these various locations. In Italy the Calcarone kiln and Gill regenerative furnaces are used exclusively. Both utilize heat liberated by burning part of the sulphur in the ore to liquify or vaporize the remaining sulphur, which is recovered by solidification or condensation. The Calcarone kiln is of conical shape, generally 35 ft in diam at base and 18 ft high. A kiln of 25,000 cu ft capacity burns for about two months and yields about 200 tons of sulphur. The Gill furnace consists of a series of chambers with domed roofs. Sulphur is burned and melted in one chamber at a time and the hot combustion gases are used to preheat the ore charge in the subsequent cell. These furnaces operate on a cycle of 4 to 8 days. The recovery yield of both systems is about 65 pct. Sulphur losses amount to 25 pct through the combustion to sulphur dioxide; about 10 pct is retained in discarded calcines. Ores containing less than 20 pct are not considered suitable as furnace feed. These methods are not only wasteful because of the low recovery obtained, but represent a serious atmospheric pollution problem. Sulphur produced ranges from 96 to 99 pct purity and thus does not match Texas or Louisiana sulphur. Owing to the present shortage, sulphur in the Middle East sells

Jan 1, 1954

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951