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By James Barr

The process of manufacturing ferrophosphorus lies not alone in smelting a mixture of phosphates, silica, iron ore, with coke as fuel, but upon smelting this mixture with coke and air as chemical agents as well as heat producers and in conjunction with extraordinary procedures and manipulations making the process continuous and profitable. For best results there must be cool top, limited fusion zone, hot bottom, good mechanical condition of stock and perfect distribution of gases; also minimum air supply, high heats and excess of coke or carbon. FERROPHOSPHORUS, an alloy of phosphorus and iron or perhaps a physical mixture of definite compounds of iron and phosphorus, has become of increasing importance as the use of the basic open hearth has extended. Commercial ferrophosphorus is a porous, brittle metal of crystalline structure and often with a bluish metallic luster. A typical analysis is as follows: PER CENT. Phosphorus 18.0 to 22.0 Iron 80.0 76.0 Oxygen 0.2 Sulfur : 0.3 Silicon 0.1 Carbon : 0.1 Manganese 0.2 This alloy was first made in an electric furnace, as it was thought that the high heat of the electric arc together with the reductive effect of carbon was necessary. About 1898, J. J. Gray, Jr., was producing a high-phosphorus pig iron in his furnace at Rockdale, Tenn., by using high-phosphorus, local iron ores. As these high-phosphorus ores were difficult to obtain, he conceived the idea of increasing the phosphorus content of the pig iron by the addition of phosphate rock to the furnace burden. When the phosphorus content of the pig was kept low and only small amounts of phosphate rock were added to the burden, no particular troubles were encountered, but when he increased the amount of phosphate rock to produce an 18 per cent. alloy, his furnace troubles rapidly multiplied.

Jan 10, 1924

By R. C. Mahon

The Homer iron-ore mine is at Iron River, Mich. Because it covers a large area, 400 acres, and because there was a considerable depth of water in the glacial drift above most of the ore bodies, this mine has presented a number of unusual and often difficult operational problems. The purpose of this paper is to discuss a few of the most important problems encountered. Surface Water Mining on this property was started on one 40-acre tract. The ledge in this area was the highest on the property and, although it was a real problem, water could be handled by underground pumping. During the first period of mining, drifts were driven just under the ledge and holes were drilled through to the surface material. In later years these drifts have been extended wherever possible. Although these drifts assisted in the drainage problem, it was evident they would not be sufficient alone. Before mining on a large scale could be started on this property, the water problem had to be solved. A number of test holes were put down to the ledge, and showed that there was an average 150 ft. of water-saturated glacial drift above the ledge over the rest of the mine. The upper 70 ft. was gravel and sand and the lower 80 ft. clay and hard pan, with some narrow seams of sand and gravel. It was evident that the water above the clay seam could not be drained off through openings in the ledge and for that reason it was decided to install surface drainage wells. This method of drainage was entirely new at that time and was pioneered at the Homer mine. The sinking of the first well was started in June 1929. Immediately following the first well a number of wells were sunk, but only to the top of the clay seam. This was done with the idea that a greater capacity could be obtained in a shorter time from gravel seams where the water flowed more freely and that the cost of pumping the water above the clay seam would be considerably less than if the wells had all been sunk to the ledge at once. Before the drainage of surface water was started, 1200 gal. per min. against a head of 800 ft. was being pumped from the mine. By March of 1932, with five surface wells in operation, the mine pumping had been reduced to 150 gal. per min. At this time we were pumping 5000 gal. per min., under an average head of 110 ft. from the surface wells. The total pumping costs were 30 per cent less than in 1929 because the greater capacity was being pumped at a much lower head. From 1932 to 1936 the mine was idle and all surface pumping was stopped. Since 1936, 10 new wells have been added with a total pumping capacity of 8000 gal. per min. at the present time. The water level has been lowered 70 ft. since surface pumping was started in 1929, and it is now approximately at the top of the clay. To remove the water below the clay seam, 11 of the 15 wells were sunk to the ledge.

Jan 1, 1946

By Ralph Arnold

The lure of possible new oil fields shut off development of the older Montana fields during the year 1928, cutting down the total production to less than the 1927 mark. However, with the production of oil worth approximately $7,000,000 during the past 12 months, Montana oil producers pushed their total up to $44,000,000 taken from two principal fields during the past 8 years. The highest prices in the United States for oil of similar gravity gave northern Montana producers a decided advantage and brought up the value of the oil to about the same value as the 1927 production, although the oil runs were 1,000,000 bbl. less. Production by Fields The discovery of two new oil fields in the Sweetgrass Arch, in north central Montana, in 1927, diverted attention from the Kevin-Sunburst field. Many operators turned their attention to wildcats and practically all of the drilling contractors were called into the new Pondera and Bannatyne fields. Although discovered in 1927, the Pondera field developed its first commercial production in March of 1928. Mild interest was followed four months later by a rush of operators, and by August practically every string of tools had been moved from the Kevin-Sunburst field to Pondera. The fall months found 40 wells drilling in this field and as the year closed there were 100 producers. Pondera field is remarkable in that every well thus far drilled within the structure proper has been a producer. Dry holes have been drilled around the field, but none within an area of four sections. Experts estimate that Pondera will produce 5000 bbl, of oil to the acre. The producing area now outlined by wells drilled is expected to produce around 15,000,000 bbl. of oil. Pondera crude brings $1.75. This field has one pipe line and a second line is being laid. Lack of a pipe line delayed development of the new Bannatyne field, south of Pondera, and also in the Sweetgrass Arch. Bannatyne has 11 producing wells scattered over an area of approximately three sections. Bannatyne is a sand field with a 70-ft. sand in the lower Jurassic, which core-drill tests show to be 60 per cent. saturated. With the completion

Jan 1, 1929

By P. Barnes

SEVERAL interesting and important considerations may be based upon an analysis of the cost of producing Bessemer rails, and the facts thus set forth may be much more clearly emphasized by reducing each item of cost to its own fractional proportion of the sum total in each branch of the manufacture. The estimate given below is not absolutely complete in minuteness of detail, but it may be fairly taken as a basis for the suggestions to be noted : Perhaps the most striking comments that can be made upon this statement are those arising from a consideration of some of the smaller items, and even the least of these quite loses its insignifi¬cance when its proportion of the whole is expressed in actual money, as for the work of a year, or even for a single day. Thus, if the item A 5 be considered, the importance can be clearly

Jan 1, 1877

By W. C. Kerr

A BRIEF sketch only is here intended, with a few illustrations, in order to give a general notion of the character and structure of these veins. I have stated elsewhere, several years ago, that these veins were wrought on a large scale and for many ages by some ancient peoples, most probably the so-called Mound Builders; although they built no mounds here, and have left no signs of any permanent habitation. They opened and worked a great many veins down to or near water-level ; that is, as far as the action of atmospheric chemistry had softened the rock so that it was workable without metal tools, of the use of which no signs are apparent. Many of the largest and most profitable of the mines of the present day are simply the ancient Mound Builders' mines reopened and pushed into the hard undecomposed granite by powder and steel. Blocks of mica have often been found half imbedded in the face of the vein, with the tool-marks about it, showing the exact limit of the efficiency of those prehistoric mechanical appliances. As to the geological relations of these veins, they are found in the gneisses

Jan 1, 1880

By Walter Larsh

The Ruth orebody, so far developed, is roughly oval in plan, major and minor axes about 1600 ft. (487 m.) and 1200 ft. (365 m.) respectively, average thickness about 120 ft. (36 m.), and with a general dip of about 15° to the northwest. According to the generally accepted theory, the ore is a secondary enrichment of the porphyry due to leaching in situ by meteoric waters. Copper occurs as chalcocite, chalcocite coating pyrite, and a little chalcopyrite. The leached zone, or capping, covering the ore varies in thickness from 110 to 540 ft. (33 to 164 m.), average about 410 ft. (124 m.). The ore is blue gray in color, and the capping, light yellow to brown. Ore was first discovered in quantity in the workings from the Ruth shaft, now caved. Further development work was done by churn drills, and the outline of the orebody, as stated above, determined in this manner. Drilling is still being carried on. The porphyry area is entirely surrounded by sedimentaries, lime-stones and shales; and, in so far as the underground workings have shown, the porphyry' itself, the sedimentaries near the contact, and the country underlying the ore (igneous or sedimentary) is very soft, heavy, and in places swelling ground. All openings have to be carefully and stoutly timbered, and, after once driven, they have to be constantly eased off and repaired. The mine is at present served by two shafts, the Star Pointer, the main working shaft; a vertical, and the Ingersoll, an incline used for waste and material. The two shafts are about 2600 ft. (792 m.) apart, connected on the 500 level; a connection is now being driven on the 600 level. (Dec., 1917. This connection is now completed.) The 300 level, which is finished, and the 400 level which is nearly completed, are accessible only from the Ingersoll and connections with the 500 level.

Jan 1, 1918

By K. K. Kershner, W. A. Calhoun, C. W. Funk

Future resources of aluminum may require the utilization of low grade ores to provide a more permanent protection for the nation. Aluminum minerals such as cloy, shale, and high iron laterites may become important sources. An evaluation of reserves is given. New laboratory results are reported on the Vereinigte desilication process. THIS report on future resources of aluminum metal emphasizes the necessity for investigating utilization of low grade potential ores to provide more permanent protection for the nation. Partial or complete interruption of the importation of high grade aluminum ores must be given consideration. Ultimately, aluminum minerals such as clay, shale, and high iron laterites may become the principal sources of aluminum. The development of extractive processes for treating these low grade materials is a step in long-range national preparedness that should receive more attention. In formulating such processes, the application of beneficiation methods for the purpose of classifying the several types of minerals before subjecting each type to the more expensive extractive procedures should be considered, and a careful survey of the basic factors of power sources, transportation facilities, and sources of raw material should be made. Domestic Bauxite Resources Of the domestic supply of current, treatable-grade bauxite, 95 pct comes from Arkansas. The present reserve of Arkansas bauxite suitable for metal production, including abrasive-grade, Bayer-grade, and modified Bayer-grade ores, is 34,000,000 tons. The total reserve of bauxite of all grades is approximately 75,000,000 tons. Reserves are estimated on a mined and dried basis. Data available to the Bureau of Mines in 1948 to 1949 were used in preparing Table I, which shows a breakdown by grades of bauxite for about a half of the reserves of the state. The modified Bayer ore, 10,555,500 long dry tons, is based on a cutoff of 8 to 15 pct silica. The sub- marginal 10,074,700 long dry tons is assumed to contain more than 40 pct gibbsite, more than 15 pct silica, and less than 32 pct available alumina. A considerable amount of this low grade material was consumed in 1949 and 1950. Additional tonnages of bauxitic and kaolinitic clays are indicated by the Bureau of Mines drilling project.' Applying a 50 pct factor to arrive at long tons mined and dried, the results of this project are approximately 31,500,000 tons kaolinitic clay at +35 pct Al²O3 and 11,800,000 tons kaolinitic clay at 20 to 35 pct A1,0,. Total reserves of kaolinitic clays are probably quite extensive—several times the reserves shown in the Bureau's work. Total reserves of bauxitic and high iron clays from the drilling program and previously established figures are 63,980,000 tons. The mineral dressing methods,' the lime-sinter process, he lime-sinter modification of the Pederson process," the lime-soda sinter process,"-' and the Alcoa combination processa. have been described in the literature and will not be elaborated upon here. The less familiar Vereinigte Aluminium-Werke process used by the Germans during World War 11 may require some amplification. It was used to bene-ficiate boehmite ores containing 12 to 18 pct silica for a feed to the European Bayer process. Ore was

Jan 1, 1952

By Elmer A. Jones

Development work in mines of St. Joseph Lead Co., Southeast Missouri, using trackless loading equipment shows definite advantages: Speed of cleaning, ability to work on steep grades and sharp crosscuts, good later tracking conditions, and possibility of wasting rock in old abandoned stopes. THE Desloge mine is one of a group of mines owned by the St. Joseph Lead Co., in the so-called Lead Belt of Southeast Missouri. It has been operating approximately 40 years producing a daily tonnage ranging from 1600 to 3800, depending on various factors. On an average, the present mine is 300 ft deep with an extreme extent of 2.7 miles by 1.7 miles. Most of the present mining is in the lower horizons of the Bonne Terre dolomite, a fairly soft rock requiring no support except occasional channel iron and roof pins. Due to the contour of the underlying sandstone and porphyry, some development work has been through short stretches of these different formations. In some areas the lower part of the dolomite is quite sandy and generally is considered abrasive. Sizes and Methods of Drifting Have Changed: Development work has, like other phases of mining, changed considerably in the last 20 years, from cleaning by hand to the use of a mechanical method. The conception of a standard drift has also changed. Before the advent of mechanical loading in stopes there were many pieces of main line drift averaging 7 ft high x 9 ft wide. With the coming of the St. Joe shovel the drift was increased to 8 ft high x 10 ft ELMER A. JONES, Member AIME, is Assistant General Mine Superintendent, St. Joseph Lead Co., Bonne Terre, Mo. AIME Columbus Meeting, September 1949. TP 2859 A. Discussion (2 copies) may be sent to Transactions AIME before July 31, 1950. Manuscript received Sept. 16, 1949. wide. Now with the Joy loader and shuttle car, a 9x11 ft drift is coming to be standard size. Methods of cutting drifts have changed considerably over the last few years. Some changes in drilling equipment have been made such as reducing the weight of column, clamp, air feed, and fittings to go with the HC10 Cleveland jackhamer. Other changes have been made in drill steel such as the use of detachable bits; throw-away variety, screw type, and the tungsten-carbide insert bits. Later developments have introduced spiral-welded pipe with victaulic couplings, millisecond exploders, bulky powder, and the use of electric cap lamps. Experimental work is going on continually in various phases of development work, some to be tried and discarded, others gradually to become standard practice. To sum up, the standard method of development now in use at Desloge is as follows: The work is done on a two-shift cycle 7:00 p.m. to 3:00 a.m. and 7:00 a.m. to 3:00 p.m. Two men known as development drillers comprise a crew on each shift. These two men arrive at their work about 7:30 to find a round blasted by the previous shift. After scaling down back and walls they proceed to the cleaning operation, using either a Joy loader and shuttle car, a drag hoist mounted on a slide ramp, or an Eimco Finlay to load into a 21/2 ton mine car. Cleaning may take from 2 to 4 hr depending on conditions such as length of round broken, length of haul, and other variables. After fitting up pipe and setting up columns and drills they are ready to drill a round which should measure 9 ft high x 11 ft wide and break from 5 to 7 ft advance. This may also take from 2 to 3 hr. Then they are ready to charge and

Jan 1, 1951

By Elmer A. Jones

Development work in mines of St. Joseph Lead Co., Southeast Missouri, using trackless loading equipment shows definite advantages: Speed of cleaning, ability to work on steep grades and sharp crosscuts, good later tracking conditions, and possibility of wasting rock in old abandoned stopes. THE Desloge mine is one of a group of mines owned by the St. Joseph Lead Co., in the so-called Lead Belt of Southeast Missouri. It has been operating approximately 40 years producing a daily tonnage ranging from 1600 to 3800, depending on various factors. On an average, the present mine is 300 ft deep with an extreme extent of 2.7 miles by 1.7 miles. Most of the present mining is in the lower horizons of the Bonne Terre dolomite, a fairly soft rock requiring no support except occasional channel iron and roof pins. Due to the contour of the underlying sandstone and porphyry, some development work has been through short stretches of these different formations. In some areas the lower part of the dolomite is quite sandy and generally is considered abrasive. Sizes and Methods of Drifting Have Changed: Development work has, like other phases of mining, changed considerably in the last 20 years, from cleaning by hand to the use of a mechanical method. The conception of a standard drift has also changed. Before the advent of mechanical loading in stopes there were many pieces of main line drift averaging 7 ft high x 9 ft wide. With the coming of the St. Joe shovel the drift was increased to 8 ft high x 10 ft ELMER A. JONES, Member AIME, is Assistant General Mine Superintendent, St. Joseph Lead Co., Bonne Terre, Mo. AIME Columbus Meeting, September 1949. TP 2859 A. Discussion (2 copies) may be sent to Transactions AIME before July 31, 1950. Manuscript received Sept. 16, 1949. wide. Now with the Joy loader and shuttle car, a 9x11 ft drift is coming to be standard size. Methods of cutting drifts have changed considerably over the last few years. Some changes in drilling equipment have been made such as reducing the weight of column, clamp, air feed, and fittings to go with the HC10 Cleveland jackhamer. Other changes have been made in drill steel such as the use of detachable bits; throw-away variety, screw type, and the tungsten-carbide insert bits. Later developments have introduced spiral-welded pipe with victaulic couplings, millisecond exploders, bulky powder, and the use of electric cap lamps. Experimental work is going on continually in various phases of development work, some to be tried and discarded, others gradually to become standard practice. To sum up, the standard method of development now in use at Desloge is as follows: The work is done on a two-shift cycle 7:00 p.m. to 3:00 a.m. and 7:00 a.m. to 3:00 p.m. Two men known as development drillers comprise a crew on each shift. These two men arrive at their work about 7:30 to find a round blasted by the previous shift. After scaling down back and walls they proceed to the cleaning operation, using either a Joy loader and shuttle car, a drag hoist mounted on a slide ramp, or an Eimco Finlay to load into a 21/2 ton mine car. Cleaning may take from 2 to 4 hr depending on conditions such as length of round broken, length of haul, and other variables. After fitting up pipe and setting up columns and drills they are ready to drill a round which should measure 9 ft high x 11 ft wide and break from 5 to 7 ft advance. This may also take from 2 to 3 hr. Then they are ready to charge and

Jan 1, 1951

By B. F. Oliver

A levitation zone melter is described which has the necessary characteristics for processing metals of controlled purity. No crucible is required; control of the solidification variables is stable; the environment of the charge is chenzically controlled. Rotation of the solidifying interface and mixing in the liquid by the magnetohydrodynamic stirring produce an eight-fold increase in freezing velocity over former melting units with better resulting solute redistributions. The smooth appearance of processed specimens, resistivity ratios as a hnction of position, as well as a segregation curve, are presented as evidence of smooth continuous solute fransport over distances in excess of ten zone lengths. THE difficulties of the floating zone-purification of higher density metals in large quantities under conditions of lowered surface tension are well known and have hindered fundamental research programs for some time. The inordinate amount of time required to obtain a small reproducible yield of high-quality metal has excluded the use of such material for many important material research programs. However, with the use of controllable induction levitation forces' to accomplish stable liquid zones with nearly flat solidifying interfaces, and with good physical and chemical control of the processes involved, these difficulties can be overcome. A zone melter that is to be used for both high-purity and controlled composition work in metal systems must have the following three characteristics: 1) The process must avoid the use of containers. 2) Physical control of the solid-liquid interface and of the mixing within the liquid phase must be stable. 3) The chemical environment of the melt chamber must be controllable. A description of an apparatus that accomplishes these objectives follows.* APPARATUS One of the basic features that is necessary for good process control of large sections is that the charge, the coil, and the motion mechanisms be within the same chamber. This arrangement permits a close coupling between the work coil and the load, which then provides a good flexibility for adjustments of coil dimensions, for controlling the amount of levitation needed for stable operation at a given power level. An overall view of the melter is shown, with the melt-chamber door open, in Fig. 1. The radiation shields surrounding the single-turn induction coil are of water-cooled copper. The chamber is stainless steel, as are most of the components within the chamber. The control panel for the scanner is located at the immediate lower right of the melt chamber. This panel controls the operation of the scanner. The operations controlled are the interface rotation rate, the stretch-squeeze or diameter control, the up and down scanning speeds, and the induction power. All of the drive motors are constant-torque variable-speed motors. The scanning speeds are set with precision helipots and continuously observed with the voltmeter on the control panel and checked with a mechanical indicator within the melt chamber. The rotation rate is also set with a precision helipot. The up-scanning speeds are 2 to 300 in. per hr.

Jan 1, 1963

By E. A. Kilinski

THE mineral production of Dutch Guiana has been of little consequence. This is due to several rea-sons. Inaccessibility is probably the greatest factor, for the Guianas are off the main commercial route. More or less closely related causes are the tropi-cal climate and the temperament of the inhabitants. Small quantities of gold have come from Dutch Guiana for many years, and recently bauxite has entered the list of exports. It is likely that other minerals may be found in commercial quantities. The Colony is on the northeast coast of South Amer-ica. It is known by two names, that of Dutch Guiana, and that of Surinam. - British Guiana (Demerara) adjoins it on the west, French Guiana (Cayenne) on the east, and Brazil on the south. It covers an area of approximately 58,000 sq. miles. The population, in the. neighborhood of a hundred thousand,, is small, but the variety of races and the sub-sequent mixtures of these is almost incredible. This unusual and colorful population readily falls into three groups, (1) those of the bush, (2) those of the planta-tion, and (3) those of the towns. The bush people con-sist of two groups, the Indians and the bush-negro. The Indians, of whom there are but a little over two thousand, keep entirely to themselves, lead a. very primi-tive existence, and rarely come into the towns. The bush-negro gives the touch of human life to the thick jungles. A descendant of escaped slaves of by-gone days, he is at present a proud, stalwart, happy, clean, moral, carefree and ease-loving individual-a distinct personality. Work is not his pal, although he does cheerfully that which is necessary for his existence. He will transport passengers and goods in his dugout and guide the stranger through the jungles, but that is all. The plantations employ some negroes, but more Javanese, imported as contract labor from the East Indies. Paramaribo, the only large town in Dutch Guiana, is the center of all government, commercial, and social activities. It has a motley population of 40,000, in-cluding negroes native to Dutch Guiana and many from all parts of the West Indies. In addition, there are Javanese, Chinese, and some Hindus. The two latter groups control most of the retail trade. The Dutch and a few scattered Germans are the only Europeans. All of the higher government offices are held by the Dutch, the others are filled from the ranks of the natives. Marriages and intermarriages among these various groups have produced types that defy almost any effort to decipher an individual's forbears. Dutch is the official language, but among the people a jargon known as Tahki-Tahki holds sway. English is taught in the schools, consequently many speak it, and many more can make themselves understood in that language.

Jan 2, 1928

By R. A. Graham, R. W. Rohde, J. R. Holland

In a shock wave compression study of a martensitic Fe-30 pct Ni alloy, Graham, Anderson, and Holland' found a region of unusual compressibility extending from a few kilobars up to about 20 kbars. This region was characterized by a continuously decreasing shock wave velocity with increasing stress. In this note, we report the results of a study initiated to determine the source of this unusual compressibility. In earlier shock-loading experiments on martensitic Fe-30 pct Ni, Loree et a1.' determined that a phase change occurred at about 80 kbars. Subsequently, Bowden and elly found that samples of a similar material shock-loaded at about 100 kbars experienced a martensite-to-austenite transition. Their crystal-lographic analysis showed the transition was a direct reversal of the established martensite system to the austenite phase. The shock loading in the present experiments was accomplished by precisely controlled plane impact between two discs of the sample material. This configuration created a state of uniaxial strain in the two the front of a projectile which was accelerated to various velocities. Planarity between the impacting discs was controlled to 5 x 10-4 rad, and impact velocities were measured to 0.5 pct. From a knowledge of the Hugoniot relationships of the material and the impact velocity, the shock-wave stress can be calculated. Samples were shocked at 18, 50, 70, and 100 kbars. The specimen discs shocked at 18 and 50 kbars were instrumented with quartz gages to observe the stress time profile. A more detailed description of the experiment has been given previously.4 The shock-loaded specimens were recovered intact by decelerating them in a catcher filled with soft wood. Discs of Fe-30 pct Ni nominally 12.7 mm thick and 64 mm in diam were cut from bar stock. Chemical analysis of the material provided the following composition in weight percents: 29.5 Ni, 0.50 Mn, 0.21 Si, and 0.10 C, with traces of Cr, Cu, Mo, Al, and Co. The discs were austenitized at 650° C for 2 hr at a pressure of less than l0-' torr, and furnace-cooled. Martensite was then produced by quenching the discs at -196" and holding 168 hr. The initial martensitic structure formed is shown in Fig. l(a). Typical microstructures of samples shock-loaded at 18 and 100 kbars are shown in Figs. l(b) and l(c). The banded structure with the austenite regions running parallel to the shock propagation direction was characteristic of samples shock-loaded at all stresses. No such banding was found in samples that were not shock-loaded. Comparison of Figs. l(b) and l(c) shows that, although under the optical microscope the two microstructures are quite similar, the martensite plates of the specimen shocked at 100 kbars etched in a distinctly different manner. This unique etching behavior of martensite in samples shock-loaded above the 80 kbar transition point has also been observed by Leslie el al.' These investigators determined from X-ray and electron microscope studies that fcc martensite, similar in appearance to normal bcc martensite, had been formed under shock loading. A quantitative determination of the amount of austenite present after shock loading was made by X-ray techniques. Both the methods of Lindren' and Miller' were utilized. The results are shown in Table I. The error limits are those experimentally observed from two independent analyses of each sample and represent the precision of the experiment. The absolute error, which should be similar for each sample, is believed to be about 10 pct. It is seen that as the shock stress increases the amount of austenite present also increases. On the basis of the X-ray and microscopic observations, we conclude that shock loading below the 80 kbar ersal will initiate a martensitic reversal: the amount of reversal that occurs increases with in-

Jan 1, 1969

By William Kelly

An exploring-drift, on the bottom level of the Curry mine, at Norway, Mich., on the Menominee range, cut a stream of water, which increased considerably the expense of pumping. The hope that the supply would drain off was not realized. It was decided to build a dam in the exploring-drift, as it was estimated that the cost would be recovered by the saving in fuel in less than two months.

Jan 1, 1898

By A. M. Ross

DEURING the past year the uranium milling industry has seen the installation and initial operation of solvent extraction circuits in the Climax Uranium Co. mill at Grand Junction, Colo.; the Kerr-McGee Oil Industries mill at Shiprock, N. M.; and the AEC pilot plant at Grand Junction, operated by National Lead Co. Inc. The Vitro Uranium Co. mill in Salt Lake City is converting to solvent extraction, and the new Union Carbide Nuclear Co. mill to be constructed at Rifle, Colo., will also employ this method. Solvent extraction circuits will be incorporated in two mills now under construction -the Texas Zinc Minerals Co. mill at Mexican Hat, Utah, and the Gunnison Mining Co. mill at Gunnison, Colo. In Canada the expansion of the Eldorado mill at Port Radium includes installation of a solvent extraction circuit. Solvent Extraction Process: In simplest terms, the organic solvent is mixed with leach liquor and then the organic and aqueous phases are allowed to separate. By this contact the solvent selectively extracts uranium from the liquor, leaving a raffinate, or barren leach liquor, which may be discarded. The pregnant solvent is stripped of its uranium by con- tact with a stripping solution, such as aqueous sodium carbonate, which has a greater affinity for uranium than the solvent does. The barren solvent from the stripping step is available for further leach liquor extraction, and the pregnant stripping solution, which usually contains 30 to 60 g U3O8 per liter, is ready for precipitation of a final product, or uranium concentrate.

Jan 9, 1957

By A. G. Moncrieff

In the paper the authors show [(Fig. 12)] in diagramatic form a "typical vacuum filtration system." This indicates a drum filter with two filtrate receivers-one for mother liquor and one for wash liquor. Have the authors ever found a drum filter that will adequately separate mother liquor from wash liquor? I recall in about 1957 being superintendent of a major copper leaching operation treating oxide flotation concentrates. The plan was being expanded, and filters installed to replace CCD following acid leaching and primary thickening of the pulp. The design of the expansion allowed the mother liquor to join the thickener overflow as feed for the electrolytic tankhouse. The water balance required that only pan of the wash liquor could be electrolysed direct, and the remainder had to be treated by lime precipitation. The design assumed the mother liquor would have the same copper content (about 50 gpl) as in the leach pulp solution, and the wash liquor about 50% of this (25 gpl from memory). You can imagine my horror when, on starting up, the solution from the mother liquor receiver contained somewhat under 40 gpl and, from the wash liquor receiver, somewhat over 40 gpl. The filters were provided by two of the most reputable manufacturers and, in the end I believe they agreed that never before had they supplied rotary drum filters that would segregate the filtrates. By changes to the valve design we made minor improvements, but eventually decided that insufficient segregation of the filtrates was possible to warrant trying to do it. On further consideration I concluded that, owing to the time delay between solution passing through the cake and entering the valve, segregation was impossible, though radical alteration of the internal piping of the filter might have improved it. I would be interested to know if anyone has found a solution to this problem using a drum filter. If not, why go on publishing information that infers that it is possible?

Jan 1, 1982

By H. Rumfelt

Surface excavators for mining coal and such non- metallic minerals as quarry stone, gypsum, phosphate and the like have advanced tremendously in the last 50 years, and their advancement has paralleled that of excavators used in the construction industry during the same period. Up to a certain stage, the construction industry led the way with this development and even pioneered in how the equipment was used. Afterwards, however, new mining machinery and related operating techniques progressed along a different path than their construction counterparts because of the different demands of the two industries. The word "stage" as used above quite generally connotes a transition period, possibly of several years' duration, during which mining development followed the construction industry. Later, mining branched out and responded to its own needs. Most readers are familiar with the events which led to the development of the power shovel. The first dredge was recorded as having been used in England in the latter 1700's and was no doubt developed along the lines of some notes written in the late 1400's by Leonardo da Vinci.

Jan 1, 1971

By C. R. Fischrupp, M. D. Helfrick, W. A. Straw

In the manufacture of telephone apparatus a number of nonferrous sheet metals arc blanked and formed to produce a wide variety of parts, which are generally small in size because of space and weight requirements. Many are given sharp bends in forming, so that they may be assembled into compact units. This study covered sheet metal of the alloys of brass, phosphor bronze and nickel silver that are most extensively used by our company. In general, the sheets range from 0.005 to 0.100 in. in thickness and from 1/4 hard to extra spring temper in hardness. As many parts function as flat springs, and others require the maximum stiffness for structural reasons, a large percentage of the metal is used in the hard and spring tempers. So, because of the necessity for the minimum radius of bend consistent with the maximum in hardness, there has been a need for information as to the forming characteristics of these sheet metals in the various alloys, tempers and thicknesses. The literature yields relatively little regarding such forming characteristics, so that designers of apparatus and tools have been largely dependent upon past experience or tool tryouts for guidance. In presenting tabular data intended as a basis for selecting the temper of the metal and radius of the bend, we recognize the influence of many other factors which are not covered in this study. For example, the angle of the bend and conditions of the metal, such as grain size and presence of impurities, as well as conditions of the fabricating process, such as rate of forming and adjustment of the tools, may all affect the result. In addition, a bumping operation subsequent to forming, which is frequently used to set the bend, increases the strain on the metal. We wish to emphasize the fact that this is not considered a scientific investigation of the formability of materials or an attempt to correlate forming characteristics with various other properties of the metals, except in so far as such properties are controlled by commercial limits on composition and temper. It is expected, however, that the data presented may serve some useful purpose in estimating the relative formability of different thicknesses

Jan 1, 1931

By Bernard Lallement

INTRODUCTION Geostatistics have been used for five years at the B. R. G. M. to compute the ore reserve estimation. At present this technic is also used at the beginning of the exploration project to determine the best way to prove the orebody. The best way is a compromise between the cost of the exploration and the precision of the estimation. It is clear that geostatistics cannot be used without any data, but very often an exploration project is decided on an orebody which has already been drilled. Old data can therefore be used. This was the case this year for two orebodies: an iron deposit and a lead deposit. The iron deposit was explored by drifts and has now to be proved. We used geostatistics to determine the best drilling pattern. The variograms computed with these old data showed that an anisotropy existed between the two directions of the orebody, which meant that the drill pattern had to be rectangular. The lead deposit was explored by drill holes and a short drift in part of the orebcdy. The problem was to choose between a regular pattern of drill holes or a long drift and some added drill holes. The case of this study will be exposed afterwards. On the other hand, if the orebody is completely unknown, the exploration project generally forecasts two exploration steps (the first step

Jan 1, 1977

By Ali Altway, M. E. Tunison, T. W. Chapman

Simulations of various contactor configurations for simultaneous extraction or stripping of two metals reveal the dependence of conversion and selectivity on operating and design parameters A two-film mass transfer model is used to provide the rates in countercurrent column and stirred-tank cascade calculations Even in these mass-transfer-limited cases the stoichiometry of the extraction reactions introduces nonlinearities and interactions between the two metal transfer rates, but these can be represented In a suitable model One conclusion is that a countercurrent column should be considerably more efficient than a cascade of mixer-settlers of comparable size.

Jan 1, 1981

By Adam L. Wesner, A. C. Richardson

This paper describes the development of a continuous float-and-sink process to produce coal low enough in ash content to be suitable for production of electrodes. The cleaned coal had a combined iron and silicon content of 0.239 pct. The efficiency of recovery was 84 pct. DURING World War I1 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 pct iron plus silicon in the coke. For this work, the specifications were that the coal should contain less than 1 pct 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 — --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 pct 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, 1953