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By R. T. DeHoff, F. N. Rhines

The problem of determining the number of particles of a phase distributed randomly in unit volume c an opaque matrix from measurements made on random plane sections is closely investigated. A formalized derivation for the general case is presented. Applications of this formal result are made to specific types of aggregates dispersed in a matrix; circles, particles with flat circular ends, cylinders, and constant shape aggregates of ellipsoids of revolution. Quantitative measurements of the average geometric properties of these aggregates are developed. In 1953 R L. ullman' developed a technique for determining, from measurements made on random plane sections, the number per unit volume, Nv, of particles imbedded in an opaque matrix. Using meas urements previously developed for volume fraction, vV,' and surface area per unit volume, .SV,= he was able to determine average volume, average surface area, and average dimensions of particles, independent of size distribution, for spheres and circular disks,' as well as for uniformly sized cylinders of any axial ratio.4 The technique is mathematically rigorous for the cases studied. Recent studies of this problem have yielded rigorous solutions for the general case of cylinders, permitting evaluation of NV independent of size distribution and with all axial ratios intermixed. Generalization to shapes having flat, circular ends has been deduced. A solution has also been obtained for the ellipsoid of revolution, independent of size distribution, but requiring a constant axial ratio in the generating ellipses. In the following derivations, as in Fullman's work, it is assumed that the problem is purely a combination of geometry and probability theory,that is,eithe the particles to be measured are imbedded randomly in the matrix, or a sufficient number of random plane sections are taken to make the sample representative. A relationship exists between the number of par- ticles of a given size and shape that may be situated in unit volume (Ny) and the number of intersections a random plane can be expected to make with these particles (Na). It is the purpose of the following section to derive this relationship. The probability of Intersecting Convex Bodies. Consider a convex, closed surface, that is, one for which no two surface normals point in the same direction, situated at some arbitrary position and orientation in a cube of material that is one unit long at each edge. It is not necessary to assume that this surface is symmetrical. Let planes be constructed in the cube parallel to the top face and at random distances from it. Those planes which lie between the two planes which are just tangent to the top and bottom of the surface will intersect it. The probability that a plane will intersect the body may be defined as the limit of the fraction of planes that lie between the two tangent planes as the total number of planes constructed becomes infinite. This fraction, in the limit, is equal to the distance between the two tangent planes, DV, divided by the total length over which planes are constructed, which has been taken as unity. If the body is now rotated to a new orientation, applying the same argument, the probability of intersection is again numerically equal to the distance between tangent planes. Let DV ($, 0) be defined as the distance between tangent planes of a convex body as a function of orientation, Fig. 1. The probability, Pr, of intersecting the body with a randomly oriented and randomly positioned plane, is then equal to the average value of DV (0, $) over all possible orientations. It may be easily shown that for a system of spherical coordinates the probability of finding an orientation between Q and Q + de, $ and $ + d$ is (1/4n) sin $ dB d$, so that This relationship may be applied to each body in an aggregate of convex bodies. In particular, if the bodies in an aggregate are classified according to size and shape, and the ithsuch class is considered, Eq. 111 may be applied to this entire class, the probability of intersecting any body in this class being

Jan 1, 1962

By P. E. Phillips

INTRODUCTION Rocky Mountain Energy is the mining subsidiary of the Union Pacific Corporation, and has been active m uranium exploration since 1966. Since 1970, RME's uranium activities have been funded 50% by Mono Power, the fuels supply division of Southern California Edision. RME and Mono recognized the need for development of new methods for recovering low grade uranium deposits, and contacted Halliburton Services in regard to in situ leaching, because of Halliburton oil industry experience. As a result, a joint venture between the three parties was formed in 1974. BEAR CREEK The first ISL test was at the Bear Creek deposit in the Powder River Basin of Wyoming, about 60 miles north of Douglas. The mineralized sand is the tertiary Wasatch formation, which is an arkosic sandstone. The sand typically contains about 10% clay. The predominant clay being montmorillonite with some kaolin and illite. The roll front type trends occur along the flanks of old stream channels. This deposit was being considered for exploitation as an open pit mine. Routine amenability testing indicated it was quite amenable to a conventional acid leach, consuming about 50 lbs. of sulfuric acid per ton of ore in an agitation leach, while recovering more than 95% of the uranium. Hydrology testing indicated a permeability of about 1 darcy. The groundwater has a pH of about 8. 0 and TDS of about 1200. Test Facility The Bear Creek test was a summer-only operation, with no effort towards weather-proofing. Uranium was removed from the recovered leach solution by fixed bed ion-exchange columns. Loaded resin was eluted with strong sulfuric acid. The recovery plant was designed for simplicity of operation, with little concern for the grade or form of the yellowcake product. The plant was constructed of PVC piping and valves, fiberglass reinforced tanks, and stainless steel pumps. All waste streams were directed toward a lined evaporation pond. Wellfield The test pattern was a 30 foot radius 5 spot, with wells fully penetrating the 30 foot ore sand to a total depth of about 165 feet. The injection wells were three-inch PVC with slotted casing, while the producer was eight-inch I. D. , with a stainless steel screen. The eight-inch well was necessary in order to accommodate a six-inch stainless steel pump. Wells were cemented using cement baskets, at the top of the sand, to contain a latex acid resistant cement. Leach Conditions The lixiviant generally contained from five to seven gpl sulfuric acid, and about 0. 5 gpl hydrogen peroxide. The pH was generally 1.4 and EMF was usually -550 millivolts. Injection rates were rather variable from well to well, with a total injection rate between 12 and 25 GPM. Production rate was set at 25 GPM, since the production pump could not be throttled more than that. Approximately two pore volumes of solution were injected during the three-month test. Leach Results Injectivity was quite variable, but was generally rather poor, because of poor completion techniques. The problem stemmed from injecting before the injection wells had been completely cleaned of drilling mud. Pregnant solutions averaged about 50 ppm U308, with groundwater dilution varying from 10 to 50%, depending on the injection rate. Injectivity did decline with time, probably as a result of contamination of the well faces. A total of 1000 lbs U308 had been recovered when the test was shut down. This probably represented about 10% of the mineral under leach.

Jan 1, 1979

By Arthur Clark, Terrill

THE Tri-State field is now believed to be the largest zinc district in the world. It has a potential production sufficient to supply the entire zinc demands of the country. It is estimated that a train 70 miles long would be required to haul the district's 1919 output of 60 per cent. zinc concentrates. For many years,, including 1917, Missouri surpassed Oklahoma in zinc production. In 1918, however, Oklahoma led by a wide margin and, in 1919, Missouri produced less than one-sixth of the total for the three states, when Kansas replaced Missouri in second place in the production. The district is now producing one-third of the zinc of the world and twice as much as ever before at the present price.

Jan 1, 1920

By P. S. Roller

P. S. ROLLER,*College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should he defined as the distance between an F line and a y line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and y lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. x between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling. It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz." This would require many data to prove, but no proof is given; any consideration would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.," Cambridge, Mass. Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids., Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystalline solids. Andreasen and Berg' have ground

Jan 1, 1941

By P. S. Roller

P. S. ROLLER,* College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should be defined as the distance between an F line and a [?] line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and [?] lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. I between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling. It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz," This would require many data to prove, but no proof is given; any considera¬tion would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.,* Cambridge, Mass.¬Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids.5 Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystalline solids. Andreasen and Berg8 have ground

Jan 1, 1941

By Shiou-Chuan Sun

SHIOU-CHUAN SUN (author's reply).-Thanks are extended to Professor Hassialis for his recognition of the usefulness of a new tool, the cataphoretic cell, as an adjunct to the bubble machine and the abstraction test. His questions concerning adsorption, dielectric constant, floatability, activation, and depression are interesting and worthy of discussion. I According to Mueller,28 it is questionable that any of the present formulas gives absolute values of zeta potentials which are completely accurate. The problem is further complicated by the question" of whether the "dielectric constant" has any meaning at all for such a microscopically thin layer. With these in mind, the results in this paper should be considered in the relative sense as a first approximation and not as absolute quantities. Adsorption Referring to item a of Fig. 8, the dielectric constant and the viscosity of the xanthate solutions are the same for galena, sphalerite and quartz. If dielectric constant and viscosity (but not the adsorption) were the first-order variable of the cataphoretic system, as claimed by Professor Hassialis, then the magnitudes of zeta coefficient of galena, spha)erite and quartz would have been similar to each other. These are not. The big difference of zeta coefficient of these three minerals in the same xanthate solutions is not altogether unjustified to be attributed to their difference in adsorption of xanthate ions. This is supported by the results of chemical abstraction tests reported by Ince.14 Again, the big difference of zeta coefficient between wollastonite, calcite, and fluorite in the same laurylamine hydrochloride solutions (p. 118, Figs. 58, 59, and 60 of ref. r5) also indicates that the cataphoretic system is not governed chiefly by the dielectric constant. Beside the data shown in Fig. 8, other experimental evidences15 which support the idea of correlation between adsorption and zeta coefficient are: (I) there is a qualitative parallelism between the zeta coefficient curve and the chemical abstraction curve of sphalerite in solutions of various concentrations of sodium cyanide, (2) there is a qualitative parallelism between the zeta coefficient curve and the chemical abstraction curve of sphalerite in solutions of constant concentration of copper sulphate at different pH values, and (3) there is a qualitative parallelism between zeta coefficient curve and chemical abstraction curve of quartz in solutions with constant concentration of barium chloride at different pH values. The above mentioned experimental results are generally in agreement with Gouy's double layer theory,1-5 and support this idea that the concentration and the kind of ions adsorbed at the mineral surface and in the adjacent surrounding liquid chiefly govern the electrokinetic phenomena. Dielectric constant and viscosity of the liquid medium may be of somewhat lesser importance. It follows that the correlation between the coefficient and adsorption as claimed in this paper is justified. Dielectric Constant The term "dielectric constant " was first introduced into the electrokinetic equations by Perrin.30 In a footnote of his publication he states that Holmheltz has forgotten to include the dielectric constant. The conversion of the moment29 to the zeta potential involves an additional difficulty and confusion not involved for the moment, namely, assigning a value to the dielectric coefficient. This is conventionally assumed to be constant and equal to the dielectric constant of the bulk liquid. 21 The published literature 10, 12 indicates that the values of dielectric constant for dilute chemical solutions have been assumed as those of pure water.

Jan 1, 1947

By P. S. Roller

P. S. ROLLER,* College Park, Md.-The author introduces the idea of a grind limit, or lower size limit of the particles formed directly by grinding, and this is stipulated to be 0.70 micron. I wonder where all the particles come from that lie between the resolving power of the microscope and 0.7 micron, which are seen under a microscope in examining a sample of grind? There is also introduced the idea of excess fines; this is all right in itself but the author's quantitative definition seems to be virtually impossible to justify, and, moreover, is obscure. Just why E, the excess fines, should be defined as the distance between an F line and a y line is hard to follow, and just where the distance should be measured does not appear to be stated, as it should be since the F and y lines intersect and the distance between will depend upon where that distance is measured. The basis of the presentation is the relation of log weight against log size that was "discovered" by Schuhmann. While this relation is implied to fit, the only evidence that I can see is a slightly irregular line in Fig. I between 2 per cent and 18 per cent by weight and then a precipitous drop at 18 per cent away from any sort of straight line. Some of the conclusions are rather startling, It is stated that E, the excess fines, remains constant for all the comminution products of any particular material, except that it may decrease for some coarse crusher products. No reason is given for this assertion. One would, on the contrary, expect that excess fines, if such a concept is permissible, would generally vary with different conditions instead of being constant. Again, "as much net energy is required to reduce a ton of quartz to particles of the grind-limit diameter (namely, of course, 0.7 micron) by grinding as is required to melt the quartz." This would require many data to prove, but no proof is given; any consideration would reveal the impossibility of such a result. Still further, ore samples "ground to widely different sizes, and with different circulating loads and different size distributions or product, show the same production of new surface area per revolution of the mill, when this grind limit is used." This statement is confuted by ordinary experience, which teaches that grinding efficiency varies generally with the condition of material. For example, the cushioning effect of fines is widely accepted, while the inefficiency of a ball mill when the feed is too coarse is also well known. R. SCHUHMANN, JR.,* Cambridge, Mass.-Contrary to Mr. Bond's stand, there is a great deal of miscellaneous experimental evidence that comminution does not stop at around 0.7 micron, nor does it cease consuming energy at that point. Comminution is in fact recognized as an important method of producing colloids.5 Mechanical comminution methods were used by von Weimarn6 to produce colloidal dispersions of a variety of solids, including the minerals chalcocite, galena, hematite, stibnite. Utzino7 gives average particle sizes ranging from 0.02 to 0.09 micron for colloids produced by mechanical comminution of several crystal-line solids. Andreasen and Berg8 have ground

Jan 1, 1941

By I. W. Alcorn

The cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 121/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8 -in. surface casing and digging a 7 3/4 - in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 It., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. AS opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By M. R. Annis

Drilling mud rheological and gel property changes due to elevated temperatures frequently cause problems in drilling deep wells. A laboratory investigation of the effects of time, temperatures up to 3001; and mud composition on these properties was made with concentric-cylinder, rotational viscometers of the Farm type. The viscometers were sealed from the atmosphere to prevent dehydration of the sample. Gel strengths were found to be more sensitive than viscosities to changes in temperature and mud cornposition. The high-temperature gel strengths of bentonitic clay drilling fluids were not predictable from low-temper.atl4t-e measurements. At high temperatures, [lie clay flocculates. The floccu-lation temperature depends on the concentration and type of solids in the mud and the degree of chemical defloccula-tion. As the bentonite content of a mud increases, floccrrla-tion temperature decreases. Lignosulfonnte treatment raises the flocculation temperature. To minimize adverse high-temperatrrre effects on mud properties, the mud should be maintainted at minimum solids content, properly treated with. a defloccullant which is stable at the maximum ten?perature encountered and maintained at a pH high enough to insure activation of the deflocculant for the duration required. The Methylene Blue Test for cation exchange cnpacity of a clay was found to be a useful guide in deter.niinirig when a mud contains a dangerously high clay conWnt. INTRODUCTION In drilling a well it is desirable, for several reasons, to know the effects of time and temperature on the rheological and gel properties of the drilling fluid in the hole. Problems associated with drilling fluids normally occur down-hole: so it is reasonable to be interested in the properties of drilling fluids at the conditions which exist down-hole. It has long been observed that high pump pressures are often needed to start circulation after a trip and that the "bottom-up" mud that is circulated back to the surface is usually more viscous than it was before "cooking" at the bottom of the hole for several hours. Burk-hardt' showed that a knowledge of rheological proper-ties of muds under down-hole conditions is necessary to predict pressure surges. He showed that the viscosity of a mud greatly influences the magnitude of pressure surges and the likelihood of lost returns. McLean et aL~ showed that the rheological properties of a mud under down-hole conditions are very important to cementing operations. This paper presents data showing the effects of time and temperature on rheological and gel properties of water-base drilling fluids. Measurements of gel strengths and rheological properties were made with two different concentric-cylinder rotational viscometers of the Fann type. One was the conventional six-speed viscometer modified to operate at temperatures up to 350F and at the vapor pressure of the liquid at test temperature. The other viscometer was newly designed to operate at temperatures up to 500F and pressures up to 1,000 psi. Tests with both instruments were made under essentially the same operating conditions. The new viscometer is shown in Fig. I. In this investigation, great emphasis has been placed on the rheological properties of clay suspensions, since clays are largely responsible for a mud's thixotropic and gel-

By I. W. Alcorn

THE cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 12 1/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8-in. surface casing and digging a 7 3/4-in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 ft., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. As opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By I. W. Alcorn

The cost of drilling in the past few years of proration and ever decreasing allowables has received increasing thought and study. It seems to parallel the strides made with respect to pumping problems in 1932 and 1933. Low profit margins stimulate the search for economies. Any discussion of slim-hole drilling involves a multitude of abstract values, and the problem can be attacked only by a study of records and rational thinking. In a measure, traditional methods must be studied with a questioning attitude. The words "slim hole" may be somewhat misleading, as the term is entirely relative. For that reason I wish to define it as used in this discussion. On the Gulf Coast, I would define a normal hole as one in which 10 3/4-in. surface casing has been set and below which a 9 7/8-in. hole is dug. With this as a datum, holes of larger diameter below surface would be called "large holes," with 121/4 in. as the "extra large" class and holes of smaller diameter classed in the slim-hole range, 6 1/4 in. representing about the minimum size. The optimum size has not been definitely determined, but the preference is for a program that would call for setting 8 5/8 -in. surface casing and digging a 7 3/4 - in. hole. Different sections of the country might very properly deviate from this practice because of differences in depth, formations, regulations, or other factors. This paper will be confined to a discussion of the drilling of exploratory wells, as the drilling of this class of wells is quite a different problem from that of drilling a proven field, although many of the same considerations apply in both cases, as will be apparent. Although for many years shallow slim-hole and core-test holes of small diameter had been dug, it was not, I believe, until the fall of 1937 that serious consideration was given to its application in drilling wells 7000 ft., 8000 ft. and 9000 ft. deep on the Gulf Coast. This does not mean that nearly all such wells drilled on the Gulf Coast now are of the slim-hole variety, but only that efforts have been aimed in that direction with very encouraging results and there has been a very definite trend to reduce the size of holes, a trend that probably will continue. As to exploration policy, there appear to be two distinct and opposite views. The one usually followed in the past was more or less designed to "shoot the works" on one well. Generally on the Gulf Coast, the depth at which production might be obtained is limited by the drilling equipment and economics. With this thought in mind, wildcat wells are often planned with the idea of carrying the depth to 10,000 It., or even 15,000 ft. All of the possible hazards are anticipated and the general drilling and casing programs are so predicated. Such wells contemplate the drilling of large-diameter holes; the setting of at least one protective string of casing and an extensive coring policy, all of which finally adds up to an expensive well. AS opposed to this general policy is the slim hole. This, on the Gulf Coast, recognizes that new discoveries represent gam-

Jan 1, 1941

By T. R. Ingraham

The thermal decomposition of cupric sulfate and of cupric oxysulfate has been examined by deter-mining the equilibrium gas pressure generated over each pure compound. The equilibrium data have been used to calculate the thermodynamic properties of both compounds, and the results have been combined with previously published data to establish a predominance-volume diagram for the Cu-S-O system over the normal range of roasting temperature and gas composition used in the treatment of copper minerals. Copper oxysulfate recrystal-lizes at 1100°K with an apparently irreversible en-dothermic heat requirement of 2.9 * 0.3 kcal per mole. WHEN mixed sulfide minerals containing copper are roasted, either of two procedures may be followed, depending on the nature of the other metals. In the first procedure, the roasting is controlled to produce a maximum amount of water-soluble copper sulfate which may be removed selectively from the water-insoluble compounds of the other metals. In the second, control is directed toward the production of water-soluble sulfates of the other metals and an insoluble salt of copper. Effective control of either of these processes would be enhanced by knowledge of the thermal stability of cupric sulfate and of cupric oxysulfate. In this paper, established data for the thermal stability of cuprous and cupric sulfide and of cuprous and cupric oxide will be combined with new data for the therma1 stability of cupric sulfate and cupric oxysulfate, to designate the volumes of stability of the significant compounds in the Cu-S-O system in the temperature range 750° to 950°K. EXPERIMENTAL Materials, Apparatus, and Procedure. The copper source material used in all experiments was Fisher Certified Reagent Grade "anhydrous" cupric sulfate, for which the following analysis was supplied by the manufacturer: chloride, 0.002 pct; alkalies and earths, 0.2 pct; insoluble, 0.002 pct; iron, 0.020 pct. Cupric oxysulfate was prepared from the normal sulfate by two procedures. In the first, equimolar quantities of cupric oxide (prepared by decomposition of the sulfate) and cupric sulfate were ground together and heated under vacuum in a sealed silica vessel at 800°C. This procedure was subject to failure because of explosion of the silica vessels due to the pressure of the final traces of water released from the sulfate on the initiation of its decomposition. The second procedure involved roasting the cupric sulfate, with frequent rabbling, in a muffle furnace at 725°C for a period of 48 hr. This produced an excellent product, which was confirmed as copper oxysulfate by X-ray and chemical analyses. The experimental technique used was that of measuring the total pressure exerted by the sulfur trioxide, sulfur dioxide, and oxygen produced by the following reactions: 2 CuS04 = cuo • CuSO4 + SO3 [1] cuo . CuSO4 = 2 cuo + SO3 [2] SO3= SO2 + 1/2 O2 [3] The apparatus used in the experiments has been described previously.1-3 Its main feature of a U-tube mercury manometer in which the mercury is protected from corrosion by enclosure in a flexible Pyrex bellows. Before beginning a run, the interior of the apparatus, including the bellows-manometer and all connecting tubing, was thoroughly washed with hydrochloric acid and distilled water to remove any

Jan 1, 1965

By P. E. Baker

The commercial neutron-gamma log measures the intensity of gamma-rays produced by the interaction of neutrons with nuclei in the formation and in the bore hole. This measurement gives information concerning the distribution of neutrons around the neutron source, and in particular gives information on porosity or fluid content. Another type of information is contained in the frequency or energy spectrum of the neutron capture gamma-rays. The capture gamma-rays of each element, in fact of each isotope, have only those energies characteristic of that isotope.'" The neutron capture gamma-ray spectrum, if obtainable, would permit the identification of the elements producing the spectrum. Such information would have considerable value in oilfield exploration work if it could be applied practically in well logging. Experiments have been carried out at the La Habra Laboratory of California Research Cop. to determine the feasibility of a method of well logging in which neutron capture gamma-ray spectra are measured down hole. In these experiments, spectra were obtained for artificial formations under simulated borehole conditions, using actual borehole geometry. This paper will describe the experimental apparatus and methods and will present the spectrum curves for several formations. THEORY Basic Formulas The neutron capture cross section of an element is a measure of its tendency to capture neutrons. The macroscopic cross section, or cross section per cubic centimeter, of a material composed of several elements is the sum of the cross sections of the nuclei making up 1 cc of the material. This is expressed in the equation SB = Sn1s2, ..........(1) where SB is the macroscopic cross section, s, is the cross section of a nucleus of type i, and ni is the number of nuclei of type i/cu cm. If N is the number of neutrons absorbed per second by the composite material, then the number captured by element j is Ni = N nisi/SB = N nisi/Si ni si ......(2) When the nucleus captures a neutron, a new nucleus is formed, which is usually in an excited state and falls to its ground or unexcited state by emitting one or more gamma-ray quanta.* The emission spectrum has been measured for many elements (see Table 1). The intensity of gamma-rays produced by an element is proportional to the number of neutrons captured per unit time by the element, which in turn depends on the capture cross section of that element and the capture cross sections of the other elements present which are competing with the first element for thermal neutrons. In a small sample of material the number of neutrons, N, captured per second is N = ?SB, ..........(3) where $ is the thermal neutron flux and SB is the macro-scopic cross section defined above. The number of neutrons captured by element j is N1 = ?ni si = N ni si/SB . ......(4) The number of gamma-ray quanta of energy, E, pro-duced per unit time by element, j, is I(E)=Nni si/SB P(E) .......(5) where P(E) is the emission probability of element, j, for this gamma-ray energy. Borehole Effects The capture gamma-ray spectra of many elements have been measured accurately in academic and government laboratories. These measurements are made on small samples, and from these measurements the emission spectra of the elements are obtained.

Jan 1, 1958

By Gerald C. Mathis

FERGUS County, Mont., shown in Fig. 1, is known for its once famous gold mines near the old towns of Gilt Edge, Maiden, and Kendall. But at Heath, a small farming community near the foothills of the Snowy Mountains, thousands of tons of gypsum have been mined since 1916. The value of the gold taken from the mines of Fergus County runs into millions of dollars; the value of the products made from the gypsum of the Heath mine is also in the millions. The gypsum mine will continue to produce for many years, while the gold mines are closed today. The Heath gypsum mine was developed by the Northwest Gypsum Co. in 1916. The products of the small mill built by this company were plaster, agricultural gypsum, and raw gypsum for cement manufacturing. The United States Gypsum Co. prospected the adjacent property in 1916 and 1917 by means of two adits and some exploration drill holes. During World War I, the company abandoned operations at Heath, but again became active in this area in 1928 hen it purchased all the property of the Northwest Gypsum Co. A long range program of expansion was begun with the building of a block plant and the purchase of additional land and mineral rights. A modern wall board plant was built in 1936. When the old wooden mill purchased from the Northwest Gypsum Co. burned in 1937, an all-metal calcining plant located at Blue Rapids, Kansas, was dismantled, moved to Heath, and erected on the site of the mill. In August 1951 a new mine was opened 3000 ft north of the mill. Rock is conveyed from the mine to the mill by an 18-in. belt conveyor. The Heath gypsum deposit lies at the base of the Ellis formation of Jurassic Age. The Ellis at Heath lies unconformably on the black fissle shale of the Heath formation of the Big Snowy Group, which is Mississippian in age. The strike of the gypsum bed is N45°E and the dip approximately 3" NW; the regional strike and dip of the area. The gypsum lies in a series of six layers separated from each other by thin seams of shale. The lower layer is composed of' gypsum and shale intermingled and is not of commercial grade. The upper five layers are high in purity, but only the middle four layers are mined on the advance, as the top layer is left for roof support. This top layer is mined on retreating operations at the same time the pillars are robbed. Total thickness of the gypsum layers varies from 12 to 14 ft. Fig. 2 shows the average thickness and arrangement of the layers. Mining Methods The mine was originally developed by driving an adit into the gypsum horizon. The Northwest Gypsum Co. mined in the top four layers, the top layer being split, leaving about 12 in. for roof support. A modified room and pillar method was used: entries were 15 ft wide and rooms varied between 20 and 30 ft and were turned on 90" from the entries. Crosscuts between the rooms left pillars from 10 to 20 ft square. The rock was loaded by hand into 1-ton mine cars and pulled to a central gathering point by mules; from the gathering point, mules or horses were used to pull the string of cars to the mill tipple. This method was continued by the United States Gypsum Co. from 1928 to 1937.

Jan 1, 1954

By E. C. Skinner

Cardox, which consists essentially of a steel tube containing carbon dioxide compressed to the liquid state, is a trade name designating a device used principally in coal mines to break down coal. The use of Cardox to break down coal was developed in the Middle Western coal fields around 1927 or 1928. At first, a large, heavy tube was used, requiring many parts. Through experimentation and technologic improvement, the weight of the tube has been brought down from 88 lb. to 20 lb. and the new, light tube makes use of larger volumes of CO2 at lower pressures than formerly was thought possible. This blasting method is extensively used today in coal mines and probably one tenth of the bituminous coal produced in the United States is prepared with Cardox. Construction and Application As used in the coal mine, Cardox looks like a piece of steel pipe 2 to 3 in. in diameter and 3 to 5 ft. long. When it is necessary to fire a shot, the tube is wired with two leg wires that are long enough so that when the tube is pushed the desired distance into the hole the wires will protrude 5 to 6 in. It has been found that No. 18 insulated copper annunciator wire is the best material for leg wires. A two-conductor rubber-covered shot-firing cable of sufficient length is then attached to the leg wires. A 2 to 6-volt current is applied, either by means of shot-firing batteries or the standard type of shot-firing generator. Upon the application of the current, the charge in the cartridge is set off. Because the breaking action upon the coal is supplied by the relatively slow physical expansion of the carbon dioxide gas rather than by the rapid chemical formation of gases as in more familiar explosives, the result is a slow heaving action, somewhat similar to that produced by black powder, rather than a shap detonating effect. Permissible powder does its work at roughly ten times the pressures in use with Cardox. The escaped gas from a Cardox shot apparently works along the planes of weakness in the coal face, such as the coal cleats, the edges of the cut, and along the top of the seam, breaking off the coal toward the free faces and heaving the mass of coal outward. Cardox tubes are manufactured in various sizes to take care of variable conditions of seam size, undercut depth, cleatage, coal hardness, and so forth. The tubes range in diameter from 2 to 3 in. (Table I) and the normal CO2 charge is from 24 to 96 oz. For any one tube size, several sizes of shear disks are available, thus excellent control of the type of shooting desired is possible. In general, local experience and experimentation dictate the size of tube chosen for any particular job. For high coal, and deeper or wider cuts, a larger size is chosen to start with, and the opposite is true for low coal, softer coal, or small cuts. After the tube size has been experimentally selected, tests are run in the mine to determine how many shot holes are needed and where the holes should be placed. The size of the disk,

Jan 1, 1944

By E. C. Skinner

Cardox, which consists essentially of a steel tube containing carbon dioxide compressed to the liquid state, is a trade name designating a device used principally in coal mines to break down coal. The use of Cardox to break down coal was developed in the Middle Western coal fields around 1927 or 1928. At first, a large, heavy tube was used, requiring many parts. Through experimentation and technologic improvement, the weight of the tube has been brought down from 88 lb. to 20 lb. and the new, light tube makes use of larger volumes of CO2 at lower pressures than formerly was thought possible. This blasting method is extensively used today in coal mines and probably one tenth of the bituminous coal produced in the United States is prepared with Cardox. Construction and Application As used in the coal mine, Cardox looks like a piece of steel pipe 2 to 3 in. in diameter and 3 to 5 ft. long. When it is necessary to fire a shot, the tube is wired with two leg wires that are long enough so that when the tube is pushed the desired distance into the hole the wires will protrude 5 to 6 in. It has been found that No. 18 insulated copper annunciator wire is the best material for leg wires. A two-conductor rubber-covered shot-firing cable of sufficient length is then attached to the leg wires. A 2 to 6-volt current is applied, either by means of shot-firing batteries or the standard type of shot-firing generator. Upon the application of the current, the charge in the cartridge is set off. Because the breaking action upon the coal is supplied by the relatively slow physical expansion of the carbon dioxide gas rather than by the rapid chemical formation of gases as in more familiar explosives, the result is a slow heaving action, somewhat similar to that produced by black powder, rather than a shap detonating effect. Permissible powder does its work at roughly ten times the pressures in use with Cardox. The escaped gas from a Cardox shot apparently works along the planes of weakness in the coal face, such as the coal cleats, the edges of the cut, and along the top of the seam, breaking off the coal toward the free faces and heaving the mass of coal outward. Cardox tubes are manufactured in various sizes to take care of variable conditions of seam size, undercut depth, cleatage, coal hardness, and so forth. The tubes range in diameter from 2 to 3 in. (Table I) and the normal CO2 charge is from 24 to 96 oz. For any one tube size, several sizes of shear disks are available, thus excellent control of the type of shooting desired is possible. In general, local experience and experimentation dictate the size of tube chosen for any particular job. For high coal, and deeper or wider cuts, a larger size is chosen to start with, and the opposite is true for low coal, softer coal, or small cuts. After the tube size has been experimentally selected, tests are run in the mine to determine how many shot holes are needed and where the holes should be placed. The size of the disk,

Jan 1, 1944

By Nicasio del Castillo

INTRODUCTION The tax structure of foreign mining investment into the United States has a significant impact on the financing and the profitability of the operation. For these reasons, it is critical for the foreign investor to develop and implement a well planned tax strategy that will enable its venture to: (1) reduce its U.S. income tax liability (2) lower U.S. withholding taxes when funds are transferred from the United States, and (3) decrease or defer home country taxation on its U.S. earnings ALTERNATIVE FORMS OF INVESTMENT A threshold business and tax consideration is the form in which the investment is to be made. The most common alternatives include: forming or acquiring a U.S. mining corporation, directly establishing a branch mining operation, or entering into a joint venture or partnership with a U.S. mining company. U.S. Corporations The most common structure through which foreign mining companies engage in business in the United States is to form a U.S. mining corporation. Forming a U.S. Corporation. If the investor decides to establish a new U.S. corporation the organization is relatively simple. No distinctions are made in the United States between "public" and "private" forms of incorporation. Minimum share capital is not usually required and share capital may consist of shares with no par value. Since corporate law is the responsibility of the individual states, the foreign investor must choose a state in which to incorporate. If appreciated mining property is transferred to a corporation solely in exchange for stock or securities in that corporation, and the transferors have control of the corporation (defined as at least 80 percent of all voting shares and at least 80 percent of all other shares) immediately after the exchange, no gain or loss will be recognized. However, if the transferors receive cash or other mining properties in addition to stock, any resulting gain may, in certain situations, be taxed to the extent of the cash and the fair market value of other property received. Acquiring an Existing U.S. Corporation. Acquiring an existing U.S. mining corporation is not as straight-forward as forming a corporation. This alternative may be accomplished through either a stock or an asset acquisition. Generally, with a stock acquisition the purchaser inherits all of the business and tax attributes of the former mining corporation. If the investor is able to acquire mining assets, however, the purchaser will usually be able to avoid any unknown or hidden liabilities that could be attached to an existing corporation. Another principal advantage in acquiring assets is that the purchase price may be directly allocated to the acquired mines or other mineral-related assets with the result that the assets' depreciable base may be increased. Assuming the domestic seller and the foreign purchaser can agree, it is generally advisable to allocate the purchase price in the sales agreement to specific individual assets. From the foreign investor's standpoint, it is preferable to allocate the purchase price to depreciable assets such as buildings and equipment versus an allocation to mineral land holdings or intangibles such as goodwill, which are generally not depreciable. It is not uncommon, however, for a seller mining company being acquired to insist for tax reasons that the investor purchase shares rather than assets. If the foreign

Jan 1, 1985

By A. P. Pipilen, M. Weintraub, W. F. Hosford, A. A. 230-000-000-006 Orning

In a study of the principal factors affecting the transport of coal-water mixtures through a centrifugal pump and a pipeline, the interrelation between solids concentration, velocity, and pressure drop was established for a minus 2-in. bituminous lump coal in concentrations up to 48% by weight. Optimum concentrations were found to exist for maximum capacity of a given pipeline and for minimum energy requirements per ton of coal. Coal size degradation by particle fracture took place in the pump; abrasion and attrition to form fines took place in both pump and line. A novel system of studying flow of solid-liquid mixtures in pipelines by the determination of relative sound intensities was developed. The hydraulic transport of minerals is in part an established art and in part a relatively new field of industrial technology. The established art developed especially where the mineral was under water, where water was used in hydraulic mining, or where large quantities of water had to be pumped from the mining site to expose the mineral. Power requirements and limited life of pumps and pipelines, under such conditions, were relatively unimportant. Examples may be found in dredging, and in hydraulic mining and transportation of phosphate rock.' In the design of a coal pipeline the most needed information relates to flow characteristics, power requirements, pressure drop, size degradation, and size segregation as a function of solids-water ratio and size distribution of the mineral. A considerable amount of information has become available concerning pressure drop and energy requirements.2'3 These topics were therefore not studied intensively. Coal degradation, however, has not been clearly delineated. Lammers and coworkers4 showed that coal may be severely degraded in size but did not distinguish between degradation produced in the pump and in the line. It is the primary purpose of this paper to outline techniques used to acquire some of the much needed design data. A full report of the investigation is to be published by the Bureau of Mines at a later date. APPARATUS AND PROCEDURE The apparatus for circulation tests (hereafter referred to as Type III tests) consisted essentially of a loop and a centrifugal solids handling pump, with attendant instrumentation, and coal weighing and feeding equipment. It also included an adequate water supply, as shown in Fig. 1. The selection of 6-in., schedule 40, carbon steel pipe was based on the top size of coal to be used. Assuming that the top size of coal should not exceed 1/3 the inside diameter of the pipe, a 6-in. pipe would accommodate 2-in. coal, a realistic size for transport from the mine face. An 872-ft loop was constructed first. Later a few tests were made with a lengthened circuit of 1361 ft by superimposing an additional loop on the original one. All bends used were of 3-ft radius. A noisemeter microphone, taped to the pipe at various points, but generally about 50 ft downstream of the pump, was used to sense segregation of the coal and the length of slugs of coal in water. To measure the stream velocity, radioactive sources were encapsulated in plastic cubes. The specific gravity of the blocks was adjusted to 1.3 and 1.8 to simulate low ash and high ash coal particles. Their passage through the pipeline was detected by a Geiger counter. Because the repeated circuits of the Type III tests did not permit coal degradation in the line to be distinguished from that produced by the pump, the two

Jan 1, 1964

By Joseph L. Gillson

BRAZIL has had an industry based on ocean beach deposits of heavy minerals containing monazite, zircon, rutile, and ilmenite for well over 40 years, but except at the very earliest period, prior to 1906, and again during World War 11, has this industry been at all comparable in size with similar operations in other countries. Limiting factors have been neither a lack of reserves nor lack of markets (except for zircon). Although poor local transportation and shipping facilities have handicapped the development in Brazil, these conditions were perhaps equally bad (when mining operations started) in the Indian State of Travancore, and in New South Wales and Queensland, Australia, in which states the principal production of beach deposits of heavy minerals has occurred. The outstanding restraints on the Brazilian industry have been the lower quality of the Brazilian minerals, the relative small size of individual deposits that were known until recently, the lack of interest of Brazilian investors in mining ventures, and the unfavorable climate for investment of foreign capital under Brazilian mining law. The lower quality of the minerals is indicated by the fact that the Brazilian monazite carries only 5 to 6 pct Tho2*, and the ilmenite 561/2 pct TiO2, whereas the corresponding Travancore minerals carry 9 pct Tho, and 60 pct TiO, respectively. The Brazilian zircon is a dirty brown color (which can be removed readily) as compared with the gleaming whiteness of the Australian zircon. An inadequate market for the abundant zircon that will be available in any operation in Brazil is certainly a heavy restraint, since otherwise it could be an important byproduct. The oxide of zirconia, baddeleyite, that has come from Brazil is from other sources.' Since the Brazilian beach sands carry only a little rutile, that mineral has not been available as a valuable byproduct to help finance the operations. Substantial tonnages of rutile have been exported from Brazil. The sources of this mineral are stream placers and residual deposits over gneisses in the States of Goyaz, Minas Geraes, Ceara, etc. (Leonardos,' and Chambers". With these handicaps, Brazilian production from beach sands has been marginal, and there has been only one consistent producer. This is a company, originally French, which was called Companie de Franco-BrazilliCne, but, with the downfall of France in 1940, became Brazilian under the name Monazita e Ilmenita do Brasil ("Mibra"). This company has operated several deposits in the vicinity of Guara-pary, in the State of Espirito Santo. Its office and plant has been at Guarapary. Recently a company called, "Fomil" (Fomento Monazita-Ilmenita) started operations. This company was organized by the Foote Mineral Co. of Philadelphia, and such share as a foreign company can hold under Brazilian law has now been acquired by the Lindsay Light and Chemical Co. of Chicago. When the writer went to Brazil in the fall of 1940, he found that no general survey of the beaches had been made except for a thorough study of part of the area by Abreu.4 The paper by Miranda6 had not then been written. The federal Geological Survey and Department of Mines knew only of the existence of the deposits in Espirito Santo south of Vitoria, and had a hazy idea that others might be found north of Vitoria. One deposit in Bala was known in Rio, but the larger one at Guaratiba was not known, although a Frenchman had prospected

Jan 1, 1951

By Joseph L. Gillson

BRAZIL has had an industry based on ocean beach deposits of heavy minerals containing monazite, zircon, rutile, and ilmenite for well over 40 years, but except at the very earliest period, prior to 1906, and again during World War 11, has this industry been at all comparable in size with similar operations in other countries. Limiting factors have been neither a lack of reserves nor lack of markets (except for zircon). Although poor local transportation and shipping facilities have handicapped the development in Brazil, these conditions were perhaps equally bad (when mining operations started) in the Indian State of Travancore, and in New South Wales and Queensland, Australia, in which states the principal production of beach deposits of heavy minerals has occurred. The outstanding restraints on the Brazilian industry have been the lower quality of the Brazilian minerals, the relative small size of individual deposits that were known until recently, the lack of interest of Brazilian investors in mining ventures, and the unfavorable climate for investment of foreign capital under Brazilian mining law. The lower quality of the minerals is indicated by the fact that the Brazilian monazite carries only 5 to 6 pct Tho2*, and the ilmenite 561/2 pct TiO2, whereas the corresponding Travancore minerals carry 9 pct Tho, and 60 pct TiO, respectively. The Brazilian zircon is a dirty brown color (which can be removed readily) as compared with the gleaming whiteness of the Australian zircon. An inadequate market for the abundant zircon that will be available in any operation in Brazil is certainly a heavy restraint, since otherwise it could be an important byproduct. The oxide of zirconia, baddeleyite, that has come from Brazil is from other sources.' Since the Brazilian beach sands carry only a little rutile, that mineral has not been available as a valuable byproduct to help finance the operations. Substantial tonnages of rutile have been exported from Brazil. The sources of this mineral are stream placers and residual deposits over gneisses in the States of Goyaz, Minas Geraes, Ceara, etc. (Leonardos,' and Chambers". With these handicaps, Brazilian production from beach sands has been marginal, and there has been only one consistent producer. This is a company, originally French, which was called Companie de Franco-BrazilliCne, but, with the downfall of France in 1940, became Brazilian under the name Monazita e Ilmenita do Brasil ("Mibra"). This company has operated several deposits in the vicinity of Guara-pary, in the State of Espirito Santo. Its office and plant has been at Guarapary. Recently a company called, "Fomil" (Fomento Monazita-Ilmenita) started operations. This company was organized by the Foote Mineral Co. of Philadelphia, and such share as a foreign company can hold under Brazilian law has now been acquired by the Lindsay Light and Chemical Co. of Chicago. When the writer went to Brazil in the fall of 1940, he found that no general survey of the beaches had been made except for a thorough study of part of the area by Abreu.4 The paper by Miranda6 had not then been written. The federal Geological Survey and Department of Mines knew only of the existence of the deposits in Espirito Santo south of Vitoria, and had a hazy idea that others might be found north of Vitoria. One deposit in Bala was known in Rio, but the larger one at Guaratiba was not known, although a Frenchman had prospected

Jan 1, 1951