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By Arthur F. Taggart

Tins paper proposes two units (believed to be new) for designating, respectively, capacity and efficiency for primary and intermediate crushers. CAPACITY Operators know that the tonnage of rock that can be put through a particular crusher in a give time depends primarily upon the crusher setting (open setting for jaw and primary gyratory types; closed setting for cones and rolls), and secondarily on the lithological character of the feed, its maximum size, the amount of fines that it carries, its moisture content, and the method of feeding. The proposed capacity formula states a relationship between all of these factors in one empirical equation, as follows: TR = TR80KK'K" [I] in which Ts, called "reduction tons per hour" is the new capacity unit; T = tons of feed per hour requiring crushing; R80 = "80-per cent reduction ratio"; and K, K', and K" are constants that approximate quantitatively lithological character, moisture content and method of feeding, respectively. T is the hourly weight of material in the feed that is coarser than the coarsest discharge of the crusher. (For marked departures from the normal specific gravity range - 2.5 to 2.7 - T should. be adjusted to corresponding volumes.) In the crushers under consideration, discharge from the crushing zone is limited by the minimum dimension of a particle. Passage of the particle through a square-mesh screen is limited by its intermediate dimension. Sheppard1 has shown that the ratio of maximum: intermediate: minimum average dimensions of broken rock fragments may be taken as about 1.5 : 1 : 0.6 for nonslabby rocks such as are normally crushed, while for slabby rocks, such as thin-bedded sedimentaries, foliated schists and gneisses, and the blocky metamorphics, the ratio is of the average order I.3:I:0.3, with the possibility of an even smaller relative value for the thickness. Gaudin2 finds the corresponding ratio for 150 + 200-mesh pyrite to be 1.4: 1 : 0.6. The hourly tonnage of crusher-plant feed, therefore, should be diminished by the amount thereof that would pass a square-mesh screen of which the aperture is S times the crusher setting, where S is a shape factor with the value 1.7 for rocks that break exceptionally cubic; 2 for ores and rocks that exhibit no appreciable tendency to slabbiness; and 2.5 to 3, or, in exceptional cases, even larger, when the tendency to slabbiness is marked. R80 is the quotient of the theoretical square-mesh aperture that would pass 80 per cent of the feed, divided by the aperture that would pass 80 per cent of the product. It is best determined from screen tests.

Jan 1, 1941

By J. B. Gayle

PRESENT flotation processes depend almost entirely on the buoyant properties of air bubbles to effect separations of mineral and gangue, but there is no convenient method for measuring aeration in flotation cells. Consequently the relation between aeration and cell performance has not been definitely established. In connection with a comparative study of kerosene and froth flotation processes, a modification of the displacement method' was developed in which air escaping from the cell surface is measured with a wet-test meter. The arrangement shown in Fig. 1 is satisfactory for use with a number of cells and is not critical, except that care must be taken to use connections and tubing large enough to avoid excessive back pressures. Attempts to use smaller tubing resulted in back pressures that caused displacement of the liquid level in the receiver, so that some air was lost around the sides, an occurrence readily visible when measurements were taken with only water passing through the cell. With the arrangement shown, the manometer indicated an average reading of about 1/2-in. water pressure. No loss of air was evident. Fig. 2 presents typical data for a standard Denver sub-A 24x24-in. unit cell operated at impeller speeds from 0 to 710 rpm. Each plotted point represents the average for five readings distributed over the entire cell surface. Readings were taken with only water passing through the cell. Results of this type have been used in a study of the influence of operating variables on cell performance and in detecting inherent differences in the operating characteristics of cells of various kinds. Although the information presented in Fig. 2 resulted only with water passing through the cell, the apparatus is suitable for use during flotation of coal when various reagents are present. Possible modi- fications that might prove convenient under special circumstances include varying the size and shape of the receiver and substituting a calibrated, inverted, water-filled carboy or other gas-measuring device for the wet-test meter. If such modifications result in excessive back pressures, some form of draft should be applied at the discharge end of the measuring system.

Jan 1, 1959

By Karl V. Lindell

THE Jeffrey mine of the Canadian Johns-Manville Co., Asbestos, Que. has operated for a number of years, supplying the parent company, Johns-Manville, raw material for asbestos products. The mine is situated in the 70-mile-long serpentine belt running from Asbestos, Que. on the west to East Broughton on the east. This district is the most important producing area in the world of quality spinning-fiber asbestos, with 11 operating mines supplying 70 pct of the world's needs. The Jeffrey mine, located 90 miles from Montreal and approximately 35 miles southwest of Thetford, accounts for 50 pct of the Canadian asbestos output. Production from the Canadian serpentine belt is the chrysotile asbestos, which is considered the best available for spinning. Consumers in the United States prefer the clean, fine fibers of the Canadian chrysotile of the type produced by the Jeffrey mine. Mining In 1881, when the mine was first opened, all work, including drilling, was done by hand. The deposit, located on a hillside, was first worked as a sidehill cut by hand drilling 15-ft holes and blasting with black powder. The fiber was chiselled from the broken rock. Hand derricks, later powered by horses, were used to raise the rock as the pit deepened. Eventually, the pit became so deep that derricks were, by necessity, replaced by aerial tramways. In the late 1930's, further deepening of the open pit, with the resulting constriction of the pit bottom brought about the need for altering the mining system then in use. The high ratio of overburden to be handled to mill ore mined and the value of surface structures, not company property, that would be destroyed in expanding the area of the pit, definitely indicated vertical rather than horizontal expansion would have to be used. One exception to this was that a small area on company property to north and east of the pit could be opened up. The mining problem then resolved itself into two phases; make the small lateral expansion and as rapidly as possible choose a mining method that would expand the mine vertically. Several methods for vertical expansion were studied and rejected. Methods considered included shrinkage stoping, block caving, glory holing, and various modifications of open workings. Block caving was finally accepted as the most suitable method of underground development. Cost factors and production were important factors in the decision, also, the asbestos mines in the Thetford area were already having success with this method. Original plans provided for the lateral expansion of the open mine where possible and beginning the underground operation. This program was based on the open pit being worked out in 1954, with the underground development and production gradually increasing until the open mine was depleted. In 1949, the plan was changed to allow underground development to progress at maximum so that underground output would be 550 tons per hr by 1951. Any deficit in ore would be mined from the open pit. This change permitted extending the life of the open pit through 1964 and in this time would make available a large quantity of readily attained ore. Block Caving at The Jeffrey Mine Plans for the underground operation were formulated on a 12,000-ton-per-day basis, and the adaptability of block caving methods to high production presented a desirable situation. As shown in the main-haulage plan, a continuous loop was laid out for smooth operation of the underground railroad from the loading points to the underground crusher. A feature encountered at the Jeffrey mine, as well as other asbestos mines using caving methods, is the large size of caved material, requiring that ore be crushed before hoisting.

Jan 1, 1952

By Thomas Leonard Watson

SULPHIDE ore-bodies of more or less lenticular shape occurring in metamorphic crystalline schists, gneisses, and. slates, and conforming closely in strike and usually in dip to the inclosing rock, have been recognized and described along the, Atlantic seaboard from Quebec to Alabama. The ore-bodies do not entirely conform, in all cases, to the structure of the inclosing schists, but they sometimes cut across the foliation in dip, though conforming more closely with the direction of strike. In composition the principal sulphide mineral is usually pyrite or pyrrhotite; or both, accompanied by copper, frequently gold, and almost always some sphalerite and galenite. Other. metallic ores are, so far as these sulphide bodies have been investigated, less common; silver is not uncommon, and in some of the pyrrhotite masses nickel is present in quantity sufficient to be commercially valuable. Cobalt frequently accompanies .the nickel, usually in smaller amount, and the two vary much in relative proportion. Quartz and calcite are probably the most common of the, non-metallic minerals. A number of silicate minerals are frequent representatives in one place or another; garnet, amphibole, biotite, pyroxene, zoisite and others. The sulphide bodies of Virginia, comprising both pyrite and pyrrhotite, are among the most extensive along the Atlantic seaboard. Among the pyrite bodies may be mentioned the well-known and extensively worked mines of Louisa and Prince William, counties, located on' the main pyrite belt (A, Fig. 1) of the middle and northern, eastern Piedmont region in Virginia, and among the pyrrhotite bodies is the famous " Great Gossan Lead "(B, Fig. 1) of the Floyd-Carroll-Graysou coun-

Sep 1, 1907

By H. E. Strauss

The time-to-fracture test has been applied to pure platinum and to two alloys of platinum under the special conditions of small cross-sectional area of the specimens and of a test temperature above the annealing temperature. While the results are similar in character to those secured with large specimens of steel at temperatures below the annealing temperatures, two differences are noted: (I) an upward inflection in the curves for log stress vs. log time for the alloys, and (2) a sharp break in the curve for log stress vs. log time for pure platinum when single crystals filled the wire. The experiments showed that: (I) platinum and 10 per cent rhodium-platinum maintained their strengths at IIOO°C. better than did to per cent iridium-platinum, and (2) the rhodium-platinum after a time exceeded the iridium-platinum in strength at IIOO°C. Platinum alloys in their service at high temperatures encounter conditions somewhat similar to those that exist in the usual form of the "time-to-fracture"l test with steel. This paper is intended primarily to present results obtained by a "time-to-fracture" test under conditions approximating actual service for platinum (at the temperature IIOO°C. under oxidizing conditions) with three materials, pure platinum (99.99 per cent), a platinum alloy with 10 per cent iridium, and a platinum alloy with 10 per cent rhodium. The experimental procedure differed from that previously used in this type of measurement in that the samples were of small diameter (0.010 in.), were fully annealed, and were tested at a temperature definitely above their annealing points (600°C., IOOO°C. and 1050°C., respectively, for 60 per cent cold reduction and 15 min. heating time2,3). The experimental arrangement was similar to that previously described,' using as heating unit a small, horizontal, platinum-wound furnace, ½ by 12 in., maintained at constant temperature by an automatic controller. The samples were 0.010-in. wires, which had been annealed at 1200°C. by drawing them slowly through the furnace. The time-to-fracture was measured electrically. Elongation was not determined because of the difficulties inherent in the experimental arrangement. Life—that is, the time-to-fracture—was considered the important property in these first tests. No attempt was made to measure the reduction in area, but the types of fracture were noted qualitatively; in the text they are referred to as (I) the normal, plastic, or necked fracture, (2) the knife-edge fracture, and (3) the brittle or intercrystalline fratture with small reduction of area. Results for different constant applied stresses are plotted in Fig. I. In all cases at least two trials were made for each stress, although they cannot always be indicated separately on the curves. The breaking Stresses drop very rapidly to a small frat-286

Jan 1, 1943

By H. E. Strauss

The time-to-fracture test has been applied to pure platinum and to two alloys of platinum under the special conditions of small cross-sectional area of the specimens and of a test temperature above the annealing temperature. While the results are similar in character to those secured with large specimens of steel at temperatures below the annealing temperatures, two differences are noted: (I) an upward inflection in the curves for log stress vs. log time for the alloys, and (2) a sharp break in the curve for log stress vs. log time for pure platinum when single crystals filled the wire. The experiments showed that: (I) platinum and 10 per cent rhodium-platinum maintained their strengths at IIOO°C. better than did to per cent iridium-platinum, and (2) the rhodium-platinum after a time exceeded the iridium-platinum in strength at IIOO°C. Platinum alloys in their service at high temperatures encounter conditions somewhat similar to those that exist in the usual form of the "time-to-fracture"l test with steel. This paper is intended primarily to present results obtained by a "time-to-fracture" test under conditions approximating actual service for platinum (at the temperature IIOO°C. under oxidizing conditions) with three materials, pure platinum (99.99 per cent), a platinum alloy with 10 per cent iridium, and a platinum alloy with 10 per cent rhodium. The experimental procedure differed from that previously used in this type of measurement in that the samples were of small diameter (0.010 in.), were fully annealed, and were tested at a temperature definitely above their annealing points (600°C., IOOO°C. and 1050°C., respectively, for 60 per cent cold reduction and 15 min. heating time2,3). The experimental arrangement was similar to that previously described,' using as heating unit a small, horizontal, platinum-wound furnace, ½ by 12 in., maintained at constant temperature by an automatic controller. The samples were 0.010-in. wires, which had been annealed at 1200°C. by drawing them slowly through the furnace. The time-to-fracture was measured electrically. Elongation was not determined because of the difficulties inherent in the experimental arrangement. Life—that is, the time-to-fracture—was considered the important property in these first tests. No attempt was made to measure the reduction in area, but the types of fracture were noted qualitatively; in the text they are referred to as (I) the normal, plastic, or necked fracture, (2) the knife-edge fracture, and (3) the brittle or intercrystalline fratture with small reduction of area. Results for different constant applied stresses are plotted in Fig. I. In all cases at least two trials were made for each stress, although they cannot always be indicated separately on the curves. The breaking Stresses drop very rapidly to a small frat-286

Jan 1, 1943

By R. E. Somers

I. Introduction...........................604 1. Location, Topography, and Climate...............604 2. Scope of Work and Acknowledgments...............606 3. History and Mining......................606 11. General Geology.........................609 1. Introduction...........................609 2. Granite. :..........................610 3. Quartz Porphyry......................... 612 4. Monzonite........................... 614 5. Volcanic Breccia........,................ 618 6. Gravels............................. 619 111. Economic Geology......................... 619 1. Introduction........................... 619 2. Mineralogy........................... 619 3. Occurrence of the Ores...................... 622 4. Oxidation and Downward Enrichment............... 630 5. Rock Alteration........................639 6. Genesis of the Ores........................ 643 I. Introduction 1. Location, Topography, and Climate The Burro Mountains are located in the southwestern part of New Mexico, in Grant county. The group is made up of two distinct mountain masses, known as the Little Burros and the Big Burros, the former lying about 6 miles southwest of Silver City, and the latter about 9 miles farther in the same direction. They are separated by the valley of the Mangas River, a tributary of the Gila, and a stream only in name, except after a heavy downpour of rain. The Little Burros attain an elevation of 6,625 ft. above sea level, and the Big Burros 8,054 ft., with the surrounding country averaging about 6,000 ft. To the south of these mountains stretches the desert plain, while to the north are the more frequent elevations of the mountainous country.

Jan 1, 1916

By C. S. Anderson

Georgia, since 1829, has produced nearly $18,000,000 from her gold mines, but in late years the output has dwindled to insignificance. In view of present universal efforts to increase gold production, it seems pertinent to record some recent observations in that field, briefly review its past and weigh probabilities of reviving gold mining there. Little in this paper can be claimed as new or original, and it does not present results of a detailed or profound investigation of the whole region. The Allatoona gold mine was reopened in January, 1932, and explored at greater depth during the period from January to June of that year. A number of other abandoned mines were visited, but since the workings, in common with most such mines in Georgia, were flooded or caved, underground conditions could not be examined. For a wider knowledge of the subject, a search was made through available pertinent literature and statistics, a bibliography of which is appended hereto. To all the sources obligations are gratefully acknowledged. Production The discovery of gold in Georgia, in 1829, led to an influx of 6000 to 7000 miners and a production of over $200,000 in the following year. A branch of the United States Mint, established at Dahlonega in 1838, accounted for $6,115,569 up to the time of the Civil War. The output fluctuated greatly from year to year, with an all-time peak of $582,782 in 1843. In only one other year, 1845, was the half million mark exceeded. After the war mining began again, but 1883 is the last year for which an output of $200,000 is reported. Placers, always the main source of production, were yielding less and less, and in spite of increased quartz-mining activities the downward trend continued. The last year in which there was a production of $100,000 was 1902. In 1914, $16,270 came from 26 placers and 11 deep mines, with the latter reporting 1750 tons of $3 average value. In 1925, one placer and four lode mines that mined 8100 tons assaying $1.19 per ton produced $9500. One placer and one quartz mine, operating in 1930, reported a total output of $4194. Table 1 pictures the waning output.

Jan 1, 1934

By H. M. Chance

Jan 1, 1887

By N. A. Gokeen, M. Ohtani

EFFECT of elements on the solubility of carbon in liquid iron is useful in calculating and correlating a number of thermodynamic properties as shown elsewhere in detail.' Kitchener, Bockris, and spratt2, and later Turkdogan and Hancock3 have investigated the effect of sulfur on the solubility of carbon at various temperatures and concentrations. For any chosen sulfur concentration, the solubility of carbon reported by Kitchener et al. is considerably lower than that by Turkdogan and Hancock. It was therefore felt that a reinvestigation of this problem was necessary. EXPERIMENTAL METHOD An alloy of Fe-C-S .was prepared with electrolytic iron and ferrous sulfide melted by induction in a graphite crucible, 4 in. high 1 1/2 in. ID under a CO-atmosphere. The temperature, held at 1500" ± 5°C in all experiments, was measured with a Pt-Pt + 10 pct Rh thermocouple immersed into the melt. In order to assure saturation with carbon, the melt was stirred with a graphite rod. After approximately 1 hr at 1500 °C, a sample was obtained by sucking the melt into a silica tube, 4 mm ID. The sample was then analyzed for carbon and sulfur by usual methods. RESULTS The results are presented in Table I and plotted in Fig. 1 with the individual values of Turkdogan and Hancock.3 Scattering in the data is equivalent to 3 pct or less in carbon analysis. The broken line represents the empirical equation obtained by Kitchener et al. from their data. It is evident that the present data are in good agreement with those of the former investigators but in disagreement with the latter. The cause for this discrepancy cannot be readily offered since the experimental methods in the three investigations were nearly identical. Additional evidence corroborating the author's results may be presented in the following manner. The difference between the solubilities of carbon in the ternary Fe-C-X melt with a small concentration of any element X, and in the binary Fe-C melt is %?C = %CIII-%CII [1] where %?C is the change in solubility, and %CIII and %CII are the solubilities in the ternary and the binary systems respectively. The value of %Cii

Jan 1, 1961

By Arthur Knapp

Oil-well depletion curves, to be of value, should show when a well or lease may no longer be operated at a profit. The difference, at any time, between the total expenditures and the total income of a lease or well may be called the lease status. Plotting this lease status against time will give a curve subject to more accurate and different interpretations than the barrels-time curves. According to the hypothetical barrels-time curves shown in Fig. 1, well A, at the end of 16 mo., is producing twice as much oil as well B and will continue to produce for another year whereas well B will cease producing in about 6 months. Data fob Lease Status-time or Dollars-timb Curve Time Lease Months Status Lease purchased for $3000.............................. 0 — $ 3,000 No expenses chargeable to lease for 3 mo................. 2 — 3,000 Well A is started; cost of drilling for month is $16,000..... 3 — 19,000 Well is completed at additional cost of $10,000............ 4 - 29,000 Well A is brought in and flows 10,000 bbl. first month (we Fig. 1). Necessary to invest in tanks, boilers, pipe lines, etc.; difference between expenditures and receipts gives profit for month of $5,000............................ 5 — 24,000 Investment is small, cost of operating well A is small, so profits are $15,000................................... 6 — 9,000 Well ceases to flow so there is additional investment for pumping equipment and additional operating expense; net earnings from 6,000 bbl. of oil produced is $4,000. 7 ... Lease operation nor becomes normal and curve becomes smooth........................................... 8 — 500 9 2,000 10 4,500 11 6,000 12 7,700 13 8,700 14 9,000 15 9,500 16 10,000 A hypothetical lease status-time, or dollars-time, curve of this lease is shown in Fig. 2. At zero time, the lease is purchased for $3000, which is plotted below the zero dollar line, All subsequent entries are plotted

Jan 1, 1921

By R. W. Baldwin

On page 309, column 1, the second equation should read: Ma-M1 = dlog_Pa M2-M1 d log P2

Jan 1, 1961

Jan 1, 1968

By E. J. Kenaedy, E. H. Robie

THOUGH the fall meeting of the Coal Division was held in the heart of the anthracite section, at Hazleton, Pa., the bituminous industry was well represented also, and the two groups found much common ground for discussion. The' technical part of the program was confined to two sessions in the morning and afternoon of the first day, Friday, Oct. 14, the evening and the following day being reserved for field trips and social functions. The attendance more than met the expectations of those in charge of the arrangements; there seemed no hint of depression here.

Jan 1, 1932

By J. E. Harris

By elimination of other possible nucleation processes, it has been demonstrated, for commercially pure magnesium and a Mg-Al alloy, that at stresses less than that necessary for triple-point cracking creep cavities are most probably nucleated heterogeneously at second-phase particles. Evidence for the existence of such particles is presented and the role of grain boundary sliding in the nucleation process is examined quantitatively. DURING creep deformation of polycrystalline magnesium and its dilute alloys in the temperature range 0.45 to 0.7 Tm (Tm = absolute melting temperature) cavities form mainly along grain boundaries oriented normal to the stress axis and failure eventually occurs by their growth and impingement.' The possibility of low-ductility creep failure of magnesium-alloy reactor cans during service has prompted a number of studies of cavity formation in these materials.2-5 In this paper divergent theories of cavity nucleation6-10 are examined critically and conclusions drawn regarding the most probable mechanism of formation of cavities in magnesium. The paper is divided into two parts. The first section consists of the review and quantitative assessment of the possible role of grain boundary sliding in the nucleation process. In the second part an experimental program is described where an electron-microscope technique has been used to determine if second-phase particles exist in as-received commercially pure magnesium and in a magnesium alloy 1) THEORETICAL REVIEW Possible nucleation sites for creep cavities are: a) high-angle grain boundary triple points,11-14 b) subboundary/grain boundary intersections,15-18 c) ledges in the grain boundary, 19-25 d) second-phase particles. 9,26-30 Each of these possibilities is now considered with special reference to deformation of magnesium. 1.1) Triple-Point Cracking. The stroh31 condition to nucleate fracture is: " > vr [1] where ss = applied shear stress, ? = surface energy, C = shear modulus, L = length of sliding interface. This relationship is independent of the nature of the interface and it is possible to calculate the stress concentration at the end of a sliding high-angle grain boundary and thus deduce, approximately,* the minimum applied stress necessary to nucleate triple-point cracking. Such theoretical predictions for Nimonic alloys have been made by McLean13 and found to agree quite closely with the minimum stress determined by experiment. Stresses capable of nucleating triple-point cracking should also be sufficient to nucleate fracture at a grain boundary due to intragranular slip, for the lengths of sliding interface in the two cases are approximately equal. Thus the minimum stress for triple-point cracking should represent a demarcation between failure due to general "cracking" and that due to cavitation. Hauser, Landon, and Dorn32 have demonstrated that for polycrystalline magnesium in the temperature range 78° to 298°K the fracture stress is inversely proportional to the square root of the mean grain diameter. As pointed out by stroh31 this result indicates that Eq. [I] describes the fracture characteristics of magnesium at low temperatures. Heal3 has reported that for rapid deformation (112 pct hr-1) of polycrystalline magnesium the transition from cracking to cavitation occurs in the temperature range 175" to 275°C. He also presented the fracture-stress values for the same material tested under similar conditions. From Eq. [I] taking G = 1.3 x 10" dynes , ? = 600 ergs . cm-2,33 and L = 0.017 cm, a Stroh-McLean13,31 tensile stress of -266 kg. cm-2 is calculated, corresponding to a transition temperature of -190°C; this is in

Jan 1, 1965

By F. Claisse, R. Angers

Self-diffusion in iron has been measured during rapid a-r transformations using a variant of the Kryukou and Zhukhovitskii diffusion method. The study was performed by thermally cycling iron foils (1 to 6 cpm) through the transformation (=910°C). Some foils have been subjected to over 1000 cycles and some have spent more than 15 pct of their total diffusion time in the process of transformation. The experimental results show that the a-r transformation has no measurable effect on self-diffusion in iron. The study is completed by a quantitative analysis of mechanisms which can affect the diffusion rate during the transformation. The analysis confirms the experimental results. SINCE diffusion is an important factor in many solid-state transformations, it is of interest to study how it is affected by the stresses generated during these transformations. Clinard and Sherby1'2 were the first to make a study along these lines. They measured diffusion coefficients in Fe-FeCoV couples subjected to slow thermal cycling (1.5 cph) through the a-r transformation range. They found an enhancement of diffusion by a factor of about two. The purpose of the present paper is to report measurements of the self-diffusion coefficient of iron during much more rapid thermal cyclings (1 to 6 cpm) through the a-r transformation (-910°C). These more rapid cyclings produce higher strain rates during the transformation and should emphasize any possible influence of transformation upon diffusion. EXPERIMENTAL Iron foils, 25 to 35 µ thick, were cold-rolled from 99.92 pct pure iron and annealed in pure helium for 2 hr at 870°C; the resulting grain diameter was about 150 µ. Specimens 0.5 by 8 cm were cut from the foils and I7e55 was vapor deposited on one of their surfaces. A 38 gage alumel-chrome1 thermocouple was spot welded in the middle of one of the specimen long edges, Fig. 1. Two 38 gage chrome1 wires were also spot welded along the same edge on each side of the thermocouple; they were placed 2.5 cm apart and used for electrical resistance measurements. In order to prevent twisting and crumpling, the specimens were pinched between two quartz plates 0.1 by 1 by 7 cm and the assembly was close fitted into a 1 cm ID quartz tube. Four holes were drilled through the tube to let the 38 gage wires out: these were connected to the recording equipment by means of extension wires. 20-gage nickel wires fixed at both ends of the specimens were used to thermally cycle the foils by Joule heating. The above described device was placed in a 2.7 cm ID quartz tube which in turn was placed in a tubular furnace. Either a pure helium atmosphere or circulating hydrogen was used during the experiments. Specimens were subjected to thermal cycles between a minimum temperature To and a maximum temperature Tm at rates ranging from 1 to 6 cpm. This was obtained by maintaining the furnace at a constant temperature near the minimum temperature To and periodically passing an electric current through the specimen. Cooling was achieved by heat losses to the surroundings. The forms and periods of cycles were varied from one specimen to another; however, each specimen was subjected to one type of cycle only. The temperature and electrical resistance variations of the specimens were recorded as a function of time. The temperature curves were used for diffusion calculations while the electrical resistance curves were used to monitor the transformation and to determine its starting point and its approximate duration. Diffusion was measured by the method developed by Kryukov and zhukhovitskii3 and modified by Angers and Claisse.4,5 In this method a metallic foil is coated on one side with a radioactive isotope and the activity is measured periodically on both sides during the diffusion anneal. The following equation then holds: where: I1 Activity on the surface on which the deposit is made. I, Activity on the opposite surface. t Diffusion time. B Constant. D Diffusion coefficient. d Foil thickness including the deposit. G(t) A function of time; it is a second order correction term which is given graphically in Refs. 4 and 5. The diffusion coefficient D is found by plotting ln[(Il - I2)/(I1 + I,)] -G(t) against t; the resulting slope m leads to an accurate calculation of D through Eq. [2]. The effect of the a-r transformation on diffusion is expressed by the ratio DT/DU where:

Jan 1, 1970

By J. L. Huiatt, D. G. Foot

The Bureau of Mines devised a direct flotation method for treating low-grade carnallite (KCl*MgC12*6H20) and sylvinite (a mixture of KC1 and NaCl) ores characterized by a high content of water-insoluble impurities. The procedure consists of flocculation-depression of the insoluble slimes; decomposition leach (carnallite only); potash rougher, cleaner, and recleaner flotation; and final product leaching. The technique eliminates the necessity of removing the water-insoluble slimes by mechanical or flotation desliming. Laboratory results demonstrated that the direct flotation procedure, when compared with conventional techniques, yielded equivalent or improved flotation results with equal or less reagent consumptions. The direct flotation method recovered 81.6% of the solid phase KC1 from camallite ore and 79.7% of the KCl from sylvinite ore in final products containing 60.5 and 60.9% K2O, respectively. A flowsheet incorporating the described flotation procedure is presented.

Jan 1, 1983

US 4,190,434 - In the thermal production of magnesium metal, a mixture of calcined dolomite, an iron- silicon-aluminum alloy as a reductant, and residual slag from the production of ferrochromium is smelted to evolve magnesium vapor, which is then condensed. The use of bauxite is lessened or eliminated because of the relatively high alumina content of the residual slag. R. Bonfils, A. Mena, C. Payn, and L. Septier, assigned to Ste. Francaise d’Electrometallurgie (SOFREM). Feb. 26, 1980. 6 pp. (75-10R). Filed: 6-14-78 (915,387). Priority: France, 6-24-77 (77/20,227).

Jan 1, 1982

By David B. Burrows, Francis R. Conley

INTRODUCTION A prerequisite to the evaluation and management of oil properties is accurate information derived from the analysis of core samples; but reliable data, such as porosity and permeability, can result only if all oil and water have been removed from the core samples analyzed. The most common method of core cleaning used to date has been Soxhlet extraction—a somewhat static technique, which depends upon the slow diffusion of hot solvent through the core sample. Soxhlet extraction performs satisfactorily only with highly permeable samples and at best requires long cleaning periods. The centrifuge method was developed in 1951 by Continental Oil Co. to overcome the inefficiencies inherent in the Soxhlet method. The new method is basically an operation in which hot solvents are caused to flow through the samples under the influence of centrifugal force. The centrifuge unit has proved so successful that samples which formerly required days for partial extraction can now be completely cleaned in approximately one hour. The original idea was conceived in 1950 by Elmer S. Messer, whose early experiments consisted of flowing solvent over sandstone samples suspended in a revolving basket and driven by a vertically mounted vacuum cleaner motor. Results were so outstanding that more permanent units were constructed. A modified model using this basic design was recently built by Core Laboratories, Inc., and is now used regularly in their operations. DESIGN Fig. 1 shows a photograph of the two extractors routinely used to clean small core plugs in the Production Research Div. of Continental Oil Co.'s Ponca City laboratories. For illustrative purposes, the basket from one of the units has been removed, and because both extrac- tors are of identical design only one is described in the following discussion. Each extractor unit consists of a centrifuge and a solvent still. The core basket of the centrifuge is made of mild cold rolled steel 7 1/2 in. in diameter and 1 1/8 in. deep. The basket holds 18 1-in. samples (cubical or cylindrical) and is attached directly to a cone fitted to the shaft of the motor. Both the outer rim and the hollow hub of the basket contain perforations for the passage of solvent and extract. The solvent is delivered to the perforations by means of a copper tube inserted into the hollow hub. The motor is a vertical tool post grinder motor; (115 v, 5.6 amp, 0-60 cycles, AC-DC, rated at 1/3 hp). It has a rated speed of 10,000 rpm under full load and 15,000 rpm with no load. Variation of speed is obtained by the use of a variable transformer. The solvent still was constructed from one end of a stainless steel oxygen bottle by inserting a heating element, sight glass, and piping as shown on the schematic diagram, Fig. 2. The heater used was a three-position, 1,000 w, 115-v, immersion type. Intermittent flow of solvent to the centrifuge is obtained by means of a siphon bottle inserted between the condenser and the centrifuge.

Jan 1, 1957