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By C. G. Dunn

Nuclei that are formed during the recrystallization of a grain following plastic deformation generally have orientations that differ from the original. This aspect of recrystallization may be described in terms of orientation changes. It was shown previously, in an investigation on recrystallization in silicon ferrite,&apos; that the observed orientation changes were highly improbable on a chance basis, and that some mechanism must be involved. Analysis showed that the data could be satisfactorily classified into the smaller of two subclasses† of the 36 third-order twin transformations. No attempt was made at that time to classify the data within the special class of 12 third-order twin transformations, although results did indicate a greater probability for some of them to occur. A classification of this type would be more valuable, of course, with more data on orientation changes per old deformed grain. The present work fulfills this condition to some degree and also extends the scope of the investigation to include the effects of original orientation and the amount of plastic deformation. The effect of original orientation is of considerable interest because of the known effects of orientation on plastic deformation, particularly with regard to degree of fragmentation and the formation of deformation bands, and because the orientations produced by deformation should influence the orientations obtained through recrystallization. Barrett and Levenson2 reported that single crystals of iron deformed in the range 82 to 97 per cent could be put into three general classes; namely, crystals in class I maintained a reasonably sharp single orientation, those in class 2 transformed into two distinct orientations, and those in class 3 changed into a large number of major and minor orientations. Something similar would be expected for silicon ferrite crystals, and verification of this for a few cases of types class I and class 2 deformation has been reported elsewhere3 for single crystals of silicon ferrite cold-rolled in the range 10 to 20 per cent. As an example of class 2 deformation, it was shown that a crystal of silicon ferrite cold-rolled in a (110) plane in a <100> direction developed two layers, or deformation bands differing in orientation. When this happens it is possible to etch away one band and leave a single structure for use in studying the orientation changes that occur during recrystallization. when preferred crystal orientations are obtained in recrystallization, it is not particularly necessary to evaluate groups of common orientation, as was done previously (ref. I, Table 6); but an

Jan 1, 1947

By C. G. Dunn

Nuclei that are formed during the recrystallization of a grain following plastic deformation generally have orientations that differ from the original. This aspect of recrystallization may be described in terms of orientation changes. It was shown previously, in an investigation on recrystallization in silicon ferrite,&apos; that the observed orientation changes were highly improbable on a chance basis, and that some mechanism must be involved. Analysis showed that the data could be satisfactorily classified into the smaller of two subclasses† of the 36 third-order twin transformations. No attempt was made at that time to classify the data within the special class of 12 third-order twin transformations, although results did indicate a greater probability for some of them to occur. A classification of this type would be more valuable, of course, with more data on orientation changes per old deformed grain. The present work fulfills this condition to some degree and also extends the scope of the investigation to include the effects of original orientation and the amount of plastic deformation. The effect of original orientation is of considerable interest because of the known effects of orientation on plastic deformation, particularly with regard to degree of fragmentation and the formation of deformation bands, and because the orientations produced by deformation should influence the orientations obtained through recrystallization. Barrett and Levenson2 reported that single crystals of iron deformed in the range 82 to 97 per cent could be put into three general classes; namely, crystals in class I maintained a reasonably sharp single orientation, those in class 2 transformed into two distinct orientations, and those in class 3 changed into a large number of major and minor orientations. Something similar would be expected for silicon ferrite crystals, and verification of this for a few cases of types class I and class 2 deformation has been reported elsewhere3 for single crystals of silicon ferrite cold-rolled in the range 10 to 20 per cent. As an example of class 2 deformation, it was shown that a crystal of silicon ferrite cold-rolled in a (110) plane in a <100> direction developed two layers, or deformation bands differing in orientation. When this happens it is possible to etch away one band and leave a single structure for use in studying the orientation changes that occur during recrystallization. when preferred crystal orientations are obtained in recrystallization, it is not particularly necessary to evaluate groups of common orientation, as was done previously (ref. I, Table 6); but an

Jan 1, 1947

By O. Chanute

all of&apos; level atid grade curve excepting Nos. 897, 898, 899, and 900, because of the impossibility of grouping them in the same nlonner, no two Iraving the chemical composition and physical prnprtiw as tho only v:iriubles. Phosphorus ~tnits have been taken to show the relatia h:wdnes. Of these 16 groups, 10 decidedly indicate slower wear firr tile harder rail. Of the remaining 6 groups, three (groups 2, 9, and 15) come well within the limits of error (.25 pound per yard) inherent in the palculation of the data as pointed out by Dr. Dndley; they catrnot, therefore, be considered tls exceptions. Group 2 and 15 arc but little outside of the limita of error. Rail 923 of group 9 is obviously abnornlal. This is probahly due to its being overheated, which the chemical analysis shows might easily be the case, and which the physical tc~ts and its low specific gravity tend to c-nfirm. We inry say, then, thnt 13 of these 16 groups fairly indicate, if they do not definitely point out, th:kt, under the same condition% tbe harder rail gives the slower wear. Of course, these comparisons are too few to enable us to arrive at positive ~?o-nclusions; but indications thus obtained are, I believe, fir more vsluitblc and tresttvc~rthy than those fhzt averages rvottld give us, mnde up from any ~lumhr of rails under intloy difi~rcal, conditions. Finally, it, seems to me, that the conclusions arrived at earlier in my remarks, together with, and confirmed by the last, show strong evidence that, under the same conditions, the harder rail will givc the slower wear. 0. C~~anute, New York CITY: We are very much shlipl, 1 ant sure, to Mr. Sandberg for his paper upn " Rail Ryaifications and Rail Inspertion in Europe." We have in thp United States bithhrto Been inspecting mils sonlewl~at haphazard, and we ore glad to gtbt tfie results of Mr. Sandberg&apos;s long experience. We recognize that he was an)ong the first, if not the very firwt, to iipply mare rittional and scientific rules to the designing of irou rails., niora recently, to adapt tlitve rules to the clesigtting of steel I&apos;:I~]R, to conform to the oapaljilities and rqniretnento of this new n1:ttcrial. Although we have generally adopted, in this country, sectinns for hteel rails which marly of our railway men think even bottar than those of Mr. Snlldlmrg, he is neverthelees the Ieat3er in foot.iito11~ we have followed ever since it has been est~blished

Jan 1, 1881

By J. P. Hirth, C. S. Hartley

A simple graphical technique is presented for rapidly determining the ratio of resolved shear stress on slip systems in single crystals to the applied stress (Schmid factor) for various simple states of stress. Master plots are presented which, when used in conjunction with an appropriate stereo-graphic projection, allow one to determine the Schmid factor for all slip systems in a crystal of given orientation. It is shown that one such plot suffices for simple tension or compression, flex-ural bending, and biaxial compression, while a combination of two different plots must be used for torsion and simple shear. STUDIES of the deformation behavior of single crystals are most valuable when the applied stresses are resolved onto the various slip systems of the crystal. A careful study of the available slip systems and proper selection of the state of stress used and the specimen orientation can then yield quantitative information about the probability of specific dislocation reactions, the strength of barriers to plastic flow, and the mechanisms of yielding and flow in general. It is the purpose of this work to analyze some common states of stress used in deformation studies and to present a simplified way of determining for these cases the ratio of the resolved shear stress on the slip systems to the applied stress (Schmid factor). I) CALCULATIONS We first define two orthogonal coordinate systems: one based on the specimen geometry (a right circular cylinder of radius, R, is considered here) and one based on the particular slip system of interest. The unit vectors and angles specifying the two coordinate systems and the relationship between them are listed in Table I. The relationships among the various angles and unit vectors are shown on a stereographic projection in Fig. 1. The transformation matrix between the (XI X2 X3) and (x&apos;1 x&apos;2 x&apos;3) system is x1 x2 x3 X&apos;1 a11 al2 a13 x&apos;2 a21 a22 a23 x&apos;9 a31 a32 a33 where in this case the direction cosines aij are given by a11 a12 a13 aij = a21 a22 a23 a31 a32 a33 sin 0 cos F cos 0 sin F cos 9 = -cos 0 cos F sin 0 sin F sin 6 [1] 0 -sin F cos (p In terms of Eq. [I.], the stresses sij, referred to the slip system, are determined from oij, referred to the specimen axes, by1,2

Jan 1, 1965

By James A. Barr

THE writer&apos;s first experience with flotation was during World War I, in the beneficiation of Alabama graphite schist ores. One plant used a cone with a peripheral overflow; dried ore was distributed by a rotating disk onto the surface of the water. No reagents were added. The flake-graphite film floated along with some fine silica. Dewatering screens separated the flake. Naturally unit capacities were low but results were satisfactory. Minerals Separation mechanical flotation machines were used in another graphite mill. Pine oil was added directly to the head cells, with very good results. All the graphite concentrates from both types of flotation were carefully refined after drying by passing through burr mills, followed by very close sizing on sifters. In March 1927, the writer ran a very successful flotation test on minus 28-mesh, deslimed, Tennessee brown phosphate "sand,"[t] using a mechanical laboratory machine built in the shops of the Anaconda Copper Mining Co., through the courtesy of Ernest Klepetko. The machine was soon converted to a subaeration type, with improved results. This phosphate research program was greatly enlarged and led to the formation of the Phosphate Recovery Corporation. The development of nonmetallic flotation followed rapidly. The first flotation machine for beneficiation of phosphate was installed in January 1928, in a pilot mill, by International Minerals and Chemical Corporation, at Mulberry, Florida. These machines were Minerals Separation, 24-in., subaeration types, with weir control of pulp between cells. The sand bleeders were very small-about 1 ¼ -in. diameter. As the feed contained 15 to 25 per cent plus 28-mesh material, the cells choked up frequently. This was corrected by cutting a 2 by 6-in. bleeder opening above the impellers, at grid level, which short-circuited the coarse fractions directly into the feed of the next cell. The matter of size classification and grinding was a considerable problem, which does not come within the scope of this article. Because of the troubles noted above, cascade-type roughers were installed in the second flotation plant, built about 1930. Experimental cells had been tried in 1928. General plant layout is shown in Fig. I and details of the cascade construction in Fig. 2. The cascade installation in plant No. 2 gave good results and eliminated troubles due to choke-ups and coarse feed. The capacities and metallurgical results equaled those of the modified 24-in. Minerals Separation flotation machines in plant No. I The objections to the cascade installation are the floor space and mill height required. In 1939 an improved Minerals Separation machine was tested in parallel with the cascades. Extensive tests indicated

Jan 1, 1945

By Robert J. White

Strong&apos;s equation for calculating bottom-hole pressure reduction due to gas cutting of drilling mud is corrected, resulting in a simpler equation which is easier to use. Use of the equation is illustrated by examples. It is shown that reduction of weight of a 10 lb/gal mud by 50 per cent at the surface due to gas cutting produces only n 3 per cent decrease in hydrostatic head at the bottom of a 5,000-ft well. It is also shown that at the surface 5 volumes of air per volume of mud are required to reduce the apparent mud weight from 8.5 to 6.5 lh/gal in a 3,000-ft well. Illustrative graphs are included to show the effect of various degrees of gas cutting on apparent mud weight at depths of 2,000, .5,000 and 10,000 ft and to show variation of apparent mud weight, with depth for a 100 lb/cu ft (13.37 lb/gal) mud gas cut to 90 lb/cu ft (12.03 lb/gal) at the surface. INTRODUCTION In 1938, Strong&apos; published the following equation for calculating reduction in bottom-hole pressure due to gas cutting of drilling mud h = 1/D [p + pn/100 In p + p(1-n/100)/p(1- n-100)] where: h = depth in feet P = pressure in atmospheres at depth h due to mud column only D = hydrostatic pressure in atmospheres of a col-umn of uncut mud 1 ft high p = back pressure at well-head (in atmospheres) n/100 = fraction by volume of gas in mud at well-head at back pres-sure p This equation was recently found to contain an error due to an incor-rect expression for variation of vol-ume fraction of gas with pressure. Values calculated by the original equation are close to the correct values for small percentages of gas for which it was derived but become increasingly erroneous as the gas content increases and are completely unusable at the very high gas con-tents employed in aerated mud. DERIVATION The nomenclature used is the same as in Strong&apos;s paper. The pres-sure dP exerted by a lamina of thickness dh at a depth h ft is ex-pressed by Eq. 1 where x = volume fraction of gas at depth h dP = D (l - x)dh . . . (1) When we express x in terms of the wellhead per cent gas n, the well-head pressure p and the pressure P at depth h, the result is Eq. 2. x = (np/p+p)/(100-n) + (np/P + p) When this is substituted in Eq. 1 and rearranged to separate the var-iables the result is dh = 1/D [dp + np/100-n dp/(P + p]. Integration between appropriate lim-its results in or if the wellhead is open so that p = 1 atm and we rearrange Eq. 4, the result is loss in head = hD — P = n/100-n In (P + 1) ...(5) Eq. 5 can be solved easily by successive approximation using hD as a first approximation for P on the right-hand side and calculating a first value for loss in head. This can then be used to get a new value of P using the left-hand side of the equation. Values converge very rap-idly even for quite large values of per cent by volume gas n. EXAMPLES The effect on bottom-hole pressure of 50 per cent gas cutting of the mud in a 5,000-ft well using 10 Ib/ gal mud will be calculated, we have: h = 5,000 ft 10 X 7.48 144 X 14.7 = 0.0353 atm/ft hD = 176.5 atm n = 50 Substituting these values in Eq. 5 and using hD as a first approxima-tion of P on the right, we have, approximate loss in head = 50/50 In (176.5 + 1) = 2.30 log 177.5 = 5.17 atm. We calculate a new value of P as follows:

Jan 1, 1958

By M. E. Wadswort, A. S. Peck

Infrared data disclose that phenacite reacts with oleic acid to form a chemisorbed oleate monolayer on the mineral surface. The absorption band characteristic of the antisymmetrical C = 0 oleate structure appears at a spectral wavelength of 1623 cm-&apos;. Chemisorption also results in the displacement of OH- and silicate from the mineral surface. Infrared absorbance at 1623 cm-1 is correlated with float-ability at different pH for -60-mesh phenacite particles. Phenacite, Be2SiO4, is an ore mineral often associated with quartz, microcline, and chrysoberyl in pegmatites and mica schist. Experimental phenacite beneficiating processes have been confined principally to the flotation of low grade ores occur-ing in several mountain ranges of Utah and Nevada. The U.S. Bureau of Mines &apos; has recently reported on the separation of phenacite from ore deposits of Mount Wheeler, Nev. Infrared techniques were previously applied by the authors to characterize the role of sodium oleate and oleic acid in the flotation of several non-sulfide minerals.23 A parallel study was projected to the flotation of the orthosilicate mineral phenacite with oleic acid. Infrared spectra provided a &apos;molecular fingerprint&apos; of the mineral surface and permitted identification of the reaction mechanisms. Infrared data were complemented with flotation and chemical tests. EXPERIMENTAL PROCEDURES The experimental procedures used in this investigation, for which data are presented in graphical or spectral form, are described in chronological test order. Phenacite crystals from Mt. Antero, Colo., were obtained from Ward&apos;s Co., Monterey, Calif. The crystals were fragmented in a porcelain mortar to -60 mesh and then pretreated with 0.5 molar HC1 for 15 min to remove iron. Then, the mineral particles were washed free of acid with distilled water and dried. Part of the dried material was ground to a fine powder with a Fisher Grinder using an agate mortar and pestle. Next, the powder was placed in a graduated cylinder and mixed with 500 ml of distilled water. The mineral in suspension was decanted from the settled fraction after 5 min, and the colloidal suspension was used for adsorption tests in which a high specific mineral surface area was required for adsorption and infrared analysis. Conditioning was performed in 50 ml glass centrifuge tubes for both infrared and flotation test determinations. Twenty ml of colloidal suspension was used for the infrared adsorption tests. One and a half g of deslimed -60-mesh HCl pretreated phenacite was added to 20 ml of distilled water for flotation conditioning tests. Six mg of oleic acid was used for each infrared adsorption test, and 0.2 mg of oleic acid was used for the flotation conditioning tests. HCl and NaOH were used for pH modification. The pulps were stirred at a speed of about 1500 rpm for 3 min. Then, the solids were separated out and either dried at 50°C or immediately transferred to a float cell. The pH of the residual liquid was determined with a Beckman Zeromatic pH meter. The flotation procedure is described as follows: 1) A 325-1111 glass micro-float cell was filled with distilled water. 2) Conditioned mineral was placed in the cell and the pulp was aerated through the glass fritted bottom at 7 psi pressure. 3) Floated mineral was skimmed from the cell with a glass slide until flotation ceased. (The end point was reached in all tests in less than 2 min.) 4) The flotation products were dried, weighed, and the recovery was calculated. Samples for infrared analysis were made into mulls using either oleic acid or series 13-21 Halocarbon oil obtained from Halocarbon Products Corp., Hacken-sack, N.J. The mulls were placed between two sodium chloride windows. A characteristic phenacite infrared band at 1868 cm-&apos; was used as an internal standard for quantitative determinations in which absorbance at this position was quantified by the baseline-tangent method4 and related to absorbance at 1623 cm-&apos;. The ratio obtained by dividing absorbance at 1623 cm-&apos; by absorbance at 1868 cm-&apos; was defined as an absorbance quotient for use in graphical presentation.

Jan 1, 1968

By Robert L. Frantz

Progress and improvement continue to be the bywords of a dynamic coal industry. The industry continues to gain strength and expand its horizons in the face of competition from atomic energy and the problems of restrictive legislation. The direct contribution of coal mining and transportation to the national economy is about $3 billion a year ac- cording to a study made by the National Coal Policy conference. Coal remains the backbone of the power industry producing almost two-thirds of all the electrical energy generated by fossil fuels and one-quarter of all the energy consumed in the nation. Power supplied by electric utilities increased 6.9% last year. The chairman of the Federal Power Commission predicted a 7.5% growth rate until 1980 for the nation&apos;s electric output. Some 80,500 MW of fossil fired power plant capacity are on order or under construction before 1974, compared to a present capacity, all sources, of about 290,000 MW. The comparable new additions of nuclear power plants is expected to total about 51,000 MW.

Jan 2, 1969

By Bartlett W. Paulding

Since its introduction to the petroleum industry in 1949 by Clark,1 hydraulic fracturing or "hydrafracing" has developed into a widely respected and useful technique for increasing production from petroleum reservoirs. Even then, Clark described production increases of from 6 to 100 barrels of oil per day (bopd). Since that time, others have described treatments in which production increased from 12 to 456 bopd, from only a show to 40 bopd;2 from 48 to 219 bopd; 3 from 500 to 1400 bopd and from a show to 192 bopd.4 Although representative of only a few of many successful hydraulic fracturing operational6 these increases in production attest to the past and potential value of the hydrafracing technique to the petroleum industry. The technique has received a considerable amount of theoretical and experimental attention.1, 5-20 Unfortunately, however, the degree of general understanding of the mechanics of hydraulic fracturing is still inadequate to solve many recurring problems. The practicing engineer is plagued with difficulties in predicting such factors as: (1) the required pressure and viscosity of the fracturing fluid to not only induce the fracture but to control the extent of its growth, (2) the strength characteristics and amount of proppant to include in the fluid so that the fracture is held open once the fluid pressure is decreased, and (3) the volume of the reservoir influenced by the induced fractures and the resulting increase in production. A major cause of these difficulties is the uncertainty associated with determining the orientation of the hydraulically induced fractures. One&apos;s natural conclusion in considering the problem is that the orientation of the fracture would depend on several variables, as, for example, the relationship between the viscosity and injection rate of the fluid and the

Jan 1, 1968

By C.A. Zapfee, C. E. Sims

An investigation of the behavior of inclusions in steel several years ago1 led to the conclusion that some of the commonly occurring inclusions in steel have appreciable solubilities, particularly in liquid steel. In the belief that further study and clarification of this matter was warranted, a laboratory study of oxides in liquid steel was started. These oxides are generally the result of deoxidation treatment, and the basis of an understanding of their behavior must be, therefore, the behavior of the deoxidizer in liquid iron. The Fe-Si-O system was selected for the first study, and the apparatus used and experiments conducted are described herein. Apparatus The apparatus is shown in Fig. r and the system is shown by diagram in Fig. 2. Furnace, A lmosphere Commercial hydrogen (1) passes through an orifice (2) into a manometric system that measures the tankside pressure (3) and the back pressure (4) in the system. A water trap (5) guards against excess tankside pressure. A by-pass with stopcocks (6 and 7) is provided, so that the hydrogen can be routed directly from the tank to the furnace if desired. When stopcock 7 is closed and stopcock 6 is open, the gas proceeds through the conditioning train, of which the first unit is a furnace (8) that contains platinized asbestos at 350°C. The temperature is measured by a thermocouple and potentiometer (9) and is controlled by a slidewire resistance (10). In this furnace fres oxygen is converted to water vapor. The exiting gas passes over a water bath (11) kept slightly below boiling temperature by an auxiliary heating unit (12). This unit also heats a surrounding copper coil (13) containing water, and the heated water rises through a collar (14)

Jan 1, 1943

By C. A. Zapfee, C. E. Sims

An investigation of the behavior of inclusions in steel several years ago1 led to the conclusion that some of the commonly occurring inclusions in steel have appreciable solubilities, particularly in liquid steel. In the belief that further study and clarification of this matter was warranted, a laboratory study of oxides in liquid steel was started. These oxides are generally the result of deoxidation treatment, and the basis of an understanding of their behavior must be, therefore, the behavior of the deoxidizer in liquid iron. The Fe-Si-O system was selected for the first study, and the apparatus used and experiments conducted are described herein. Apparatus The apparatus is shown in Fig. r and the system is shown by diagram in Fig. 2. Furnace, A lmosphere Commercial hydrogen (1) passes through an orifice (2) into a manometric system that measures the tankside pressure (3) and the back pressure (4) in the system. A water trap (5) guards against excess tankside pressure. A by-pass with stopcocks (6 and 7) is provided, so that the hydrogen can be routed directly from the tank to the furnace if desired. When stopcock 7 is closed and stopcock 6 is open, the gas proceeds through the conditioning train, of which the first unit is a furnace (8) that contains platinized asbestos at 350°C. The temperature is measured by a thermocouple and potentiometer (9) and is controlled by a slidewire resistance (10). In this furnace fres oxygen is converted to water vapor. The exiting gas passes over a water bath (11) kept slightly below boiling temperature by an auxiliary heating unit (12). This unit also heats a surrounding copper coil (13) containing water, and the heated water rises through a collar (14)

Jan 1, 1943

By Harrison Souder

THE object of this paper is to bring to. the notice of engineers a safety detonating fuse by the use of which misfires in blasting may be eliminated and safety in blasting operations promoted. This new detonator is a French invention and is known as Cordeau detonant or detonating fuse and is sold under the name of Cordeau-Bickford. It consists of a lead tube 5 to 6 mm. in diameter, filled with trinitro-toluene. While applicable to all classes of mining it will appeal especially to those who have to do with deep-hole blasting in open-cut clines or quarries, or any operations where a large number of holes are to be shot at one time. Since the introduction of the deep-hole blasting in the Cornwall iron mines, much trouble has been experienced from misfires and many experiments were made to determine the best and safest method of blasting. Before the introduction of the detonating fuse it was found that to insure the best results, two or more high-power electric exploders, not less than No. 6, with specially insulated conductors, should be placed in each hole; the holes wired in parallel, and for a large number of holes the current to be supplied by a live wire to the bus or lead wires at about the middle of the bench. The voltage in general use is 110 to 220, but as our material is a magnetic iron ore and the holes are generally damp, we found it advisable to reduce this to about 30 volts to prevent short circuits. A No. 7 or No. 8 special insulated detonator was adopted. But with the best of care and attention to the minutest detail, we were not able to eliminate occasional misfires and the consequent ever present danger of accidents.

Jan 10, 1914

By Y. Kondo, A. M. Gaudin

A combined method for sizing in the sub-micron range was developed using centrifugation and gamma emission. Its essential features are: 1) Centrifugation of suspended mineral in water carried out at 4°C where water has its maximal density and zero temperature gradient of density 2) Gamma activition of the sample suspension induced by irradiating it in a nuclear reactor prior to centrifugation 3) Measurement of the activity of a beam of suspension as it is sedimenting in the centrifuge, using a scintillation counter provided with a discriminator. Albite, the sodium feldspar, was used as feed to the comminution step, and the induced gamma activity of Na24 was measured. Dissolved sodium ion caused a a non-settling activity in the measurement. Other "noise" is due to scattered gamma radiation from albite particles in the centrifuge outside of the line of sight. These two activities were obtained sep arately and subtracted from the total activity counted. The relative solid concentration in the suspension under the collimating slit is proportional to this net activity. The cumulative weight per cent of undersize particles was calculated and plotted against size on logarithmic scales. The relationship was found to be linear down to about 30 mµ. The specific surfaces calculated from particle size distribution agree with those measured by krypton gas adsorption. The measurement of particle size distribution of comminution products has been of interest in clarifying the mechanism of the process. The Gaudin-Schuhmann equation1,2 relates the cumulative percentage undersize, y, to relative particle size x/k. The particle size, x, and the size modulus, k, together with the distribution modulus, m, characterize the particle size distribution, m is the slope of a plot of log (y) against log (x/k). This equation has been verified for ground particles down to 1 micron in size.3&apos;4 For particle sizes smaller than a few microns, the precision and facility of sedimentation and microscope techniques become questionable. Use of a centrifuge which shortens settling time from weeks to hours is attractive. Many centrifugal methods have been proposed; however, determination of the solid concentration of the suspension at a reference level as a function of time is difficult, and sampling the suspension is likely to introduce errors. Measurement of particle size distribution by X-ray absorption in gravitational5 and in centrifugal6 sedimentation have also been tried. In these methods, the absorption coefficient for the solid is independent of size, since the wave lengths of the X-rays and gamma-rays are orders of magnitude smaller than the particles. Radiometric techniques7 in size distribution measurement have been developed within the last several years. They offer advantages over usual gravimetric and optical methods. PRINCIPLE OF METHOD The combined method of centrifugation and gamma emission for sizing in submicron range described in the present paper has the following essential features. 1) Centrifugation is carried out at 4 ± 0.1°C. Convection currents in the water phase due to density changes caused by temperature variation may disturb completely the settling system. To avoid this difficulty, the unique property of water having zero temperature gradient of density at 4°C is used. 2) Prior to size measurement the sample suspension is irradiated in a nuclear reactor and gamma activity induced. The gamma activity of the particles sedi-mented radially is measured by a scintillation counter provided with a discriminator. By this method, interruption of centrifugation by withdrawing samples is avoided and the counting rate at any time represents the solid concentration in the suspension at the level of the collimating slit of the centrifuge at the particular time at which the measurement is made. 3) By proper selection of the feed material and of the comminution device, gamma activities of foreign

Jan 1, 1964

By A. F. Henderson, C. F. Cornell, A. F. Dunyon, D. A. Dahlstrom

Ossi E. Palasvirta (Development Engineer, Oliver Iron Mining Diu., U. S. Steel Gorp.)—The authors are to be congratulated for their interesting article, which thoroughly illustrates the variables inherent in filtration of taconite concentrate. The work and the conclusions based thereon largely parallel the test work done by the writer at the Pilotac plant" and the experience gained with a commercial size agitating disk filter in the same plant. At Pilotac, however, a thorough study was also made of the effect of depolarizing (demagnetizing) the filter feed, and it is the purpose of this discussion to comment on the merits of depolarization of the magnetite concentrate prior to filtering. The work at Pilotac was done in three phases: 1) preliminary laboratory testing with a circular filter leaf of 0.047 sq ft, followed by 2) plant testing using a 4-ft diam, single-disk agitating filter that was purchased on the basis of the pilot tests on the 4-ft model. In the laboratory tests depolarization was achieved by slowly withdrawing&apos; batches of thickened concentrate from a coil producing an alternating field of about 300 oersteds. In plant tests the standard Pilotac procedure&apos; was employed, wherein the pulp falls freely through the depolarizing coil. The preliminary tests in the laboratory at first seemed to indicate that although depolarization of the filter feed decreases the cake moisture, it also tends to decrease the thickness of the cake, thus decreasing filtering rate. The tests with the 4-ft disk filter soon showed, however, that the compactness of the cake, attained during the form period because of depolarization, permitted a considerable decrease in drying time without any sacrifice in final moisture content. Thus, the filter could be operated at a much higher speed, and the overall capacity was higher than with magnetized feed. Because of the great compactness of the cake there was little shrinkage during the drying period, which prevented cracking and subsequent loss in vacuum. This in turn permitted operation with as thick a feed pulp as the diaphragm pumps could handle, eliminating the necessity of pulp density control. On the basis of these findings, the 6-ft agitating disk filter has been operated at 2 rpm, using feed pulps varying from 65 to 73 pct solids. Initially Saran 601 was used as medium, but it was later replaced with a relatively open, tight-twist nylon cloth. Filtering rates up to 750 lb per ft- er hr can be attained with feeds averaging about 70 pct- 270 mesh, and there is no trouble because of cracking. The cake moistures vary between 8.5 and 9.5 pct. To recapitulate, the merits of depolarizing the filter feed may be summed up as follows: 1) The well dispersed pulp shows less tendency to settle in the filter tank. 2) The homogeneous filter pool results in more even cake formation. 3) Because of absence of flocs, great compactness of cake is attained during the form period. 4) The cake does not tend to crack during the drying period. 5) A drier cake is produced. 6) A shorter drying period is necessary, permitting higher operating speed, which in turn results in increased capacity. 7) Density of the feed pulp can be kept as high as the equipment can handle. This increases capacity, since it is directly proportional to the percentage of solids in the pool. A few tests were also made to study the effect of chemical flocculants on filtration efficiency. Results showed that the effects of chemical and magnetic floc-culation were quite similar. Thus the use of a floccu-lant would impair rather than improve the filtering of magnetite concentrate. A. F. Henderson, C. F. Cornell, A. F. Dunyon and D. A. Dahlstrom (authors&apos; reply)—We want to thank O. E. Palasvirta for his comments, particularly in view of the encouraging results obtained with demagnetized taconite concentrate. In our studies an attempt was made to evaluate the effects of depolarizing the feed to the plant filters by passing the slurry through a coil, similar to the method described by Palasvirta. Unfortunately, in our experiments there were no startling improvements in performance level, neither cake rate increase nor cake moisture reduction. However, when slow filter cycle speeds were employed, the filter cake tended to crack and the vacuum level dropped, resulting in an increase in cake moisture content. When demagnetized feed was used during slow speeds, no cake cracking was evidenced and the vacuum level remained constant. Thus the depolarizing coil was found necessary only in cases of cracking. It should be noted that most of our test work concerned a feed of 85 to 90 pct —335 mesh and about 60 pct by weight solids concentration. This contrasts with 70 pct —270 mesh and 65 to 73 pct by weight solids as noted by Palasvirta. Reviewing both sets of results, it might be concluded that depolarizing may be successfully employed to alleviate cake cracking tendencies and may markedly improve cake rates and moistures on the coarser taconite concentrates. Further investigations may disclose the exact relationship of grind and pulp density to the depolarizing action.

Jan 1, 1959

By G. R. M. Del Giudice

A survey of the flotation literature of the past 10 years indicates an increasing use of the microscope as a tool for investigation. Thus, the metallurgical microscope has been used by Tucker and Head;&apos; Taggart, Taylor and Ince;2 Ince,3 and del Giudice4 to examine mineral surfaces that had been subjected to the action of various flotation reagents and to the action of slime pulps. It has also been used by Gaudin, Groh and Henderson;5 Martin;6 Morrow and Griswold;7 and others to examine bakelite briquets of mill products. The microscope with transmitted light, both with bright-field and dark-field illuminations, has been used by Taggart, Taylor and Knoll8 and by Ince9 to observe Brownian motrement and flocculation in flotation pulps. The use of the microscope at low-power magnifications with incident illumination for the examination of mill products has been known for years.10 Finally, del Giudice11 has used a relatively new type of illumination, the Ultropak, for the examinntion of slime-coatings.

Jan 1, 1935

By Roger C. Ilsley

The $2.1 billion Milwaukee Water Pollution Abatement Program (MWAP) requires the construction of several miles of underground openings at depths up to 350 feet for the conveyance and storage of sewage and storm water overflows. The system will include approximately 27km (17 miles) of deep tunnels ranging from 3m (10 f t.) to 9m (30 f t.) in diameter, 26 shafts and a pumping station cavern. The facilities will be located primarily in Silurian Age dolomites of the Niagara Group. The rock is characterized as competent with vertical to sub-vertical orthogonal jointing and predominantly thin shale bedding planes. The site investigation was carried out incrementally, commencing on an area wide basis, becoming site specific as facility plans were developed. To achieve cost savings through uniformity and continuity, a computerized data bank was established. Software was developed during the initial stages of the program for the production of borehole logs, structure contour maps, geologic profiles and summaries of data for statistical evaluations. This paper discusses in general terms the collection, compilation and presentation of data necessary for the characterization of the rock mass.

Jan 1, 1984

By Darwin Soler

This chapter will outline the operating procedures at Freeport Kaolin Co.5 kaolin mining operations in Georgia, with emphasis on explaining new techniques recently introduced to reclaim mined-out areas for reforestation. Kaolin is found in the central Georgia area (See Fig. 1) in lenticular, sedimentary deposits of high purity. The beds are Cretaceous in age and vary from 4 to 60 ft in thickness, under a cover of sedimentary overburden ranging from 5 to 150 ft. The kaolin beds contain the mineral kaolinite (a hydrous aluminum silicate) and have a particle-size distribution ranging from about 20µ to the submicron level and averaging about 60% below 2µ. The only contaminants in the deposits are coarse mica, quartz particles, and minor amounts of iron and titanium oxides. These kaolin deposits vary widely in quality from deposit to deposit and also vary within individual deposits. Some of the controlling quality specifications which are determined during exploration are viscosity in water suspension, optical brightness index as measured by reflectance, amenability to bleaching, and particle-size distribution. For this reason, only a small portion of the total kaolin reserves in the Georgia belt are suitable for wet processing for the present markets being served. Kaolin clays are selectively mined by different mechanical methods and are beneficiated by both wet and dry processing to produce a number of products used as fillers and coatings in the paper industry; fillers or additives in the paint, plastic, and rubber industries; and as basic raw materials for the ceramic industry. The State of Georgia is the largest kaolin producer in the United States, with 3,206,454 tons produced in 1966, according to U.S. Bureau of Mines statistics. Freeport Kaolin Co’s processing plant is located in Gordon, Ga., about 20 miles east of Macon, Ga.

Jan 1, 1969

By R. K. Iyengar, G. C. Duderstadt

An investigation was carried out to determine the effect of Ca-Si additions on the dissolved oxygen content of liquid steel. An apparent equilibrium was reached after holding the melt for some time when the total oxygen content of the melt was identical with the dissolved oxygen. Results of the investigation show that for deoxidation with silicon and manganese the apparent equilibrium is reached after 8 to 12 min and the oxygen content of the melt is in good agreement with values reported in the literature for similar steel compositions. Ca-Si additions decrease the dissolved oxygen content appreciably below that obtained with Si-Mn deoxidation. It is postulated that some calcium dissolved in liquid steel at the time of its vaporization combines with dissolved oxygen to form CaO which then fluxes the manganese silicate, thereby lowering the activity of the deoxidation products. Slow flotation of calcium silicate inclusions is attributed to their slower growth rate. An addition of aluminum (to yield <0.005 pct Alsol) prior to introduction of Ca-Si improves the kinetics of removal of resulting deoxidation products. WITH the introduction of continuous casting of billets and blooms, application of calcium-containing alloys for deoxidation purposes has gained new interest. More so than in conventional ingot casting, the degree of control of the dissolved and total oxygen contents in steel for continuous casting can determine the success or failure of the operation because these affect both surface (pinholes) and internal quality of the billets. Of the available deoxidizers, only silicon has so far found wide application. However, in the typical range of application (0.20/0.30 pct Si), it is too weak a deoxidizer to suppress pinhole frequency to the level desirable for all but the least demanding appli-cations (approximately 10/50 pinholes per sq ft, de-pending on degree of subsequent reduction). On the other hand, Vincent&apos; has shown that it requires ap-proximately 0.007/0.008 pct A1 soluble in the steel to suppress pinhole formation to <10 pinholes per sq ft. In view of the difficulty of consistently maintaining this aluminum level and the associated problem of tundish nozzle constriction2 when aluminum is added to the ladle, steelmakers have searched for other de-oxidation alloys to circumvent the problem. Calcium-containing alloys offer such a possibility. When employed as a partial substitute for silicon, CaSi (30 pct Ca/60 pct Si) additions are reported to yield adequate pinhole contro1.3 Its use simultaneously preserves the "fluidity" of the steel and ensures good castability.4 Oxygen contents as low as 0.007 to 0.014 pct have been obtained with CaSi additions varying between 3 to 8 lb per ton.5-7 Due to the high degree of vaporization of calcium at the temperature of molten steel and the different methods used to introduce it into the liquid, the reported results show considerable scatter and lack of consistency. According to several investiga- deoxidation tors, the flotation characteristics of CaSi products are not as favorable as those found for alumi-num but are similar to those established for silicates with low alumina content.678 Thus, cleanliness ratings were improved when CaSi was replaced by a com-bination of Si + A1 followed by CaMnSi additions.5,9 In view of the lack of a comprehensive description of the deoxidation potential of CaSi alloys, a study was undertaken of the effect of CaSi additions on oxygen content in comparison to Si + Mn and Si + Mn + Al. The alloy chosen for this study contained 63 pct Si, 32 pct Ca, <3 pct Al, and 2.5 pct Fe. EXPERIMENTAL PROCEDURE The experiments were divided in three groups as shown in Table I. In each group, the CaSi addition was increased successively while maintaining the melt composition constant. The experiments in the third group were designed to determine the effect of prede-oxidation with small amounts of aluminum. Ingot iron, containing 0.02 to 0.03 pct C and 0.03 to 0.05 pct Mn, was melted in a 100-lb magnesia crucible. Pig iron was added to the melt under air in an induction furnace to attain the desired carbon content of 0.05 to 0.15 pct and thereby control the initial oxygen content of the melt between 0.04 and 0.015 pct. After removal of the slag, the furnace was sealed and argon was introduced through an opening in the graphite cover, Fig. 1. When the desired temperature was reached, electrolytic manganese was added to the melt and pin samples were taken (7 mm silica tubing) to establish initial oxygen content. To obtain maximum efficiency, CaSi and metallic silicon were introduced in a sealed steel pipe into the bath. This procedure assured minimum loss of calcium through vaporization as the deoxidants were always released at the bottom of the melt. Samples were taken at regular intervals and quenched in water. Temperature of the bath was measured with a Pt/Pt-10 pct Rh thermocouple. Portions of the pin sample were used for oxygen analysis; five determinations per sample were made to obtain average oxygen content. Oxygen analyses were made by the inert carrier gas fusion method with frequent cross checks with the vacuum fusion method.

Jan 1, 1970

By S. R. Maloof

In 1945, Hollomon&apos; showed that after plastic yielding and prior to necking under simple tension, both ferrous and nonferrous materials are approximated by an equation of the following form: where, a = true stress K = strength coefficient 5 = true strain n = strain hardening exponent If a material is stress free, a plot of log o vs log 6 would be expected to be a straight line, the slope of which is the strain-hardening exponent (n) and the intercept on the ordinate axis for 6 = 1. is the logarithm of the strength coefficient K. It can be easily shown that the strain-hardening exponent in the above equation is equal to the total uniform elongation prior to necking. Thus, one can determine the total uniform elongation prior to necking from such a plot and, as pointed out by C. W. MacGregor, 2 it is the amount of strain prior to necking which determines the extent to which a material may be deep drawn. In the view of the current interest in the fabrication of beryllium shapes, it was thought that a knowledge of the strain-hardening exponent of cross-rolled beryllium sheet would be of general interest. Load-strain data after yielding and prior to necking was taken from the work of Greenspan,3 converted to true stress and true-plastic strain, and plotted on log-log paper as shown in Fig. 1. The value of Young&apos;s modulus E for beryllium was taken as 43 . l06 psi. The slope of this plot gave a value of 0.212 for the strain-hardening exponent (n) and the intercept on the ordinate axis a value of 105,000 for K. The total uniform elongation is thus 21.2 pct, in excellent agreement with the value estimated by Greenspan (about 20 pct). Furthermore,

Jan 1, 1960

By Robert R. Walden, LeMont West

K ENTUCKY&apos;S first acid-grade fluorspar flotation Kmill, shown in Fig. 1, was placed in operation Aug. 1, 1952, by the Pennsylvania Salt Manufacturing Co. at Mexico, Ky. During 1951 a critical shortage of acid-grade fluorspar developed and it became necessary for the company to produce part of its fluorspar requirements. In 1951 the company purchased the Crider and Lennen metallurgical fluorspar mill and an adjacent stockpile of tailing at Mexico, Ky. The tailing, assaying about 26 pct calcium fluoride, originated from log washer and jig mill concentrators in the area. As the mill did not have a fluorspar dryer or convenient rail facilities, the company leased the Burleson concentrator at Marion, Ky., where these facilities were available. At Marion, Ky., on the Illinois Central Railroad, dry acid-grade fluorspar is transported direct to the company&apos;s Calvert City, Ky., hydrofluoric acid plant. The Mexico flotation plant was reconditioned and designed to produce acid-grade fluorspar from the jig tailing stockpile. The new plant flowsheet is shown in Fig. 2. The ore feed, old tailing, which contains 10 pct of waste rock and coarse gravel assaying approximately 8 pct CaF,, is bound together by sticky clay

Jan 1, 1955