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By R. S. Dean

THE commercial production of electrolytic manganese on a small scale commenced in 1939. The writer made a short report on the progress of production and utilization in MINING AND METALLURGY for January 1941. Progress during the last two years, naturally, has been more rapid. In June 1940, Congressman (now Senator) Scrugham presented a proposal to Congress for an appropriation authorizing erection of a pilot plant for "the production of metallic manganese by electrolytic or other means." As a result of this, a pilot plant, having a capacity of a few hundred pounds per day was built at Boulder City, Nev. Subsequent appropriations permitted expansion of the plant to about a ton per day. Much has been learned in this plant about electrolytic manganese, and the product even in its small way has found war uses of considerable importance. Meanwhile, the Knoxville plant of the Electro Manganese Corporation, the sole commercial producer, was expanded to about 4 tons per day to supply needed material for the war program. Other proposals for commercial plants have not been favorably received by the War Production Board. The Bureau reported on one such proposal by the American Alloys and Chemical Corporation. PRACTICE AT BOULDER CITY The present practice, as carried out at Boulder City, which is being described in more detail in a current paper by the Bureau staff at Boulder City, starts with a manganese dioxide ore containing about 20 per cent manganese. Currently this ore comes from the Three Kids deposit near Las Vegas. The steps for the production of electrolytic manganese, in general, are as follows: I. The manganese in the ore is reduced to MnO. 2. The MnO is dissolved from the ore in spent electrolyte, which contains about 38 to 47 grams per liter of free sulphuric acid, 135 grams per liter of ammonium sulphate, and 10 to 12 grams per liter of Mn as sulphate. In the leaching step the manganese is built back up to 32 to 36 grams per liter of Mn as sulphate; the pH is adjusted to neutral by means of gaseous ammonia and passed through thickeners. 3. The electrolyte is purified by adding H2S, which precipitates the heavy metals. After filtering, ferrous sulphate is added to the solution and oxidized with air and the solution is filtered and clarified on a precoat filter. 4. The purified catholyte is electrolyzed in a diaphragm cell using stainless-steel cathodes and lead-silver anodes. The current density is about 45 amp. per sq. ft. Current efficiencies of 60 to 65 per cent are regularly obtained.

Jan 1, 1944

By J. E. Dutrizac, R. J. C. MacDonald, T. R. lngraham

When sintered disks of synthetic chalcopyrite (CuFeS2) were leached in acidified aqueous solutions of ferric sulfate, the following reaction stoichiometry was obtained: CuFeS2 + 2Fe2(SO4)3 = CuSO4 + 5FeSO4 + 2S Over the temperature range from 50º to 94ºC, the reaction displayed parabolic kinetics. The parabolic rate constant for the dissolution of copper is given by the equation: log.k(mg2/cm4-hr)= 11.850 - 3780/T The activation energy for the dissolution process is 17 ± 3 kcal per mole. The parabolic kinetics have been attributed to the progressive thickening of a sulfur film on the surface of the chalcopyrite. When the leaching solutions contain less than 0.01 molar Fe+3 , the Fe concentration influences the rate of leaching, probably through a mechanism involving the diffusion of ferric sulfate through the sulfur layer. At higher Fe+3 concentrations, the rate control in the leaching. reaction has been attributed to the diffusion of ferrous sulfate through the sulfur. The rate of reaction is insensitive to changes in acid concentration and in disk rotation speed. ThE reaction of acidic ferric sulfate solutions with various sulfide minerals is of practical interest for both bacterial and heap leaching. This leaching medium is generally used with low-grade ores that cannot be treated profitably by conventional means. In both bacterial leaching1-3 and heap leaching, the active agent for sulfide dissolution is ferric sulfate. Although the reactions of ferric sulfate with chalcocite, covellite, and bornite have been investigated,4*7 the kinetics of leaching chalcopyrite with ferric sulfate have not been thoroughly studied. This paper reports a study of that reaction. EXPERIMENTAL Reagent-grade sulfur was purified by the method of Bacon and FanelliB and then it was vacuum-distilled to remove any soluble magnesium salts that had been introduced during the purification procedure.9 From stoichiometric quantities of the purified sulfur and hydrogen-reduced electrolytic copper sheet (99.90 pct Cu), CuS was synthesized at 450°C in a vacuum-sealed, pyrex vessel. About 24 hr was required for the completion of the reaction. A similar procedure involving hydrogen-reduced iron wire (99.90 pct Fe) was used to synthesize FeS1.002. A 2-furnace arrangement was required. The iron was heated to 800°C while the sulfur was maintained at about 400°C. Although the reaction to consume the sulfur was rapid, the material required additional heating (1 week) in a sealed silica ampoule at 800°C before it was homogenized. X-ray powder diffraction analysis confirmed that the copper sulfide was covellite and that the iron sulfide was troilite. The composition of the iron mineral was confirmed by wet chemical analysis. The two sulfides were ground to minus 100 mesh, weighed in equimolar amounts, mixed thoroughly, and pressed into pellets at 80,000 psi. The pellets were vacuum-sealed in pyrex ampoules and then sintered for 3 days at 550°C after an initial heating at 450°C for a few hours. The pellets were then cooled, polished with 3/0 emery paper, rinsed in acetone, and stored. The material had the characteristic brassy color of chalcopyrite and was shown by X-ray diffraction to be CuFeS2. Microscopic examination of the polished surfaces revealed small inclusions of pyrite (approximately 0.5 vol pct) as the only impurity. The presence of small amounts of a second iron compound will not alter the amount of dissolved copper but might increase the amount of ferrous ion slightly. It was calculated that dissolution of all of the pyrite and 100 mg of Cu (a typical value) would change the expected ferrous concentration by only 4 pct. Microscopic examination of a pellet after leaching revealed that the pyrite was not preferentially solubilized; only those pyrite grains at the surface were attacked. Hence, the pyrite is unlikely to alter the rate of copper dissolution. The chalcopyrite disks were about 1.7 mm thick and 27 mm in diam. They were about 80 pct of theoretical density, and for this reason their true reaction area was somewhat larger than the 5.8 sq cm area presented by the polished face. The disks were cemented to lucite cylinders in such a way that only the polished face was exposed. The disks were then leached by methods previously described.6,7 RESULTS AND DISCUSSION Stoichiometry and Kinetics. The initial experiments were directed to the problem of resolving the stoichiometry of the leaching reaction. Disks of CuFeS2 were leached at 80°C for various periods of time in acidified ferric sulfate solutions that were protected from oxidation by a cover of flowing nitrogen. When the disks had been partly leached, they were removed, their soluble salts were washed out, and then they were treated with CS2 in a Soxhlet extraction apparatus. The ratio of elemental sulfur to dissolved copper thus obtained was approximately 2 to 1. After the extraction of elemental sulfur from the pellet, the residue consisted of unreacted chalcopyrite only. For runs in which an appreciable amount of copper was dissolved, the ratio of ferrous ion to cupric ion in the solution was

Jan 1, 1970

By O. C. Wheeler

The reports for the years 1942 and 1943 were written in 1943 and 1944, respectively, and were summarized at the meetings of the Institute in February of those two years, but the manuscripts were held until released for publication in 1945. 1942 Lack of transportation facilities during 1942 resulted in a sharp curtailment of producing operations in Colombia at the middle of the year, and production of crude for export was thus greatly affected. Total production in 1942 amounted to only 10,617,008 bbl. Of this amount the Tropical Oil Co. produced 9,453,995 bbl., which includes 126,607 bbl. of petroleum condensate, and the Colombian Petroleum Co. produced 1,163,013 bbl, from its Barco property. The following summary, showing a comparison of the number of barrels of crude produced during the years 1941 . and 1942, reflects the extent to which producing operations were reduced during 1942: These above figures do not include any production recovered from Condor's (Shell's) Casabe wells during testing operations. An active exploratory drilling Program was carried out by several of the major companies, and the year 1942 established a record for wildcatting activity in Colombia. Excluding development operations on the Shell's Casabe wells on the Yondo concession, 13 wildcats were completed. Eight of these were in the Magdalena Valley, and all were abandoned; five were on the Barco Concession, and of these four were completed as producers. Particularly notable among the petroleum developments during 1942 was the continued success of drilling operations on the Shell's Yondo Concession of Compania Colombiana de Petroleos El Condor opposite Barranca-Bermeja. In 1941 the discovery well, Casabe No. I, was drilled and completed with an estimated initial production of approximately 480 bbl. of 19 1/2 A.P.I. gr. oil per day. In 1942, seven stepouts from the discovery well were successfully completed with initial production ratings ranging from 150 to 1230 bbl. per day. An extensive wildcat program on the Barco Concession of the Colombian Petroleum Co. yielded favorable results with the completion of four producers in the Socuavo and Tres Bocas areas; a fifth test, Sardinata Norte No. I, was abandoned at 500;" feet. Wildcatting operations in the lower Magdalena Valley area were initiated in 1942 With the drilling of a deep test, El Dote No. I, by Compania de petroleo Shell de Colombia. The well was a failure and was abandoned at a depth of 10,046 (?) feet. Wildcats on which drilling operations were being conducted at the end of the

Jan 1, 1945

By Bruce W. Gonser, Robert I. Jaffee

Recent work by the authors1 established the constitutional diagram of the gold-germanium system. Of particular interest in the simple euctectiferous system was the eutectic alloy at 12 per cent Ge, which melted at the very low temperature of 356°C and had an unusually fine eutectic structure. It was of interest to investigate the effect of third components on the melting point, microstructure, and strength of this alloy, and in this work the effect of copper is discussed rather completely, with less complete data on the effects of silver, platinum, nickel, and zinc. No reference in the literature has been found on ternary systems with gold-germanium. Materials and Alloying -4s in the work on the binary system, gold-germanium, the germanium used in this investigation was the cyanide-reduced, 99.9 per cent product, produced by the Eagle-Picher Co. Fine gold sheet from Handy and Harmon, pure platinum sheet from the American Platinum Works, copper magnet wire from Anaconda, nickel wire assaying 99.6 per cent Ni. and electrolytic zinc were used for alloying. To make the alloys, charges of the pure metals were put in a graphite crucible and melted in an atmosphere of pure hydrogen. The temperature was brought up to about I050 C, and the metal allowed to freeze in the crucible. Except with alloys containing zinc, melting losses were consistently less than a few milligrams. When zinc was a component, weight losses up to half the weight of the zinc present were incurred, and the alloy compositions were calculated by assuming the entire weight loss to be that of zinc. In all other cases the composition of the alloys could safely be taken as that of the charge. The ingots, as solidified, were segregated, the heavier primary phases being at the bottom and sides of the ingots and the lighter primary phases at the top, but by sectioning the entire ingot vertically for microscopic examination, this factor could be evaluated. Ingots for microscopic examination weighed one gram, and those for thermal analysis weighed three grams. Thermal Analysis In this work, the eutectic line composition and thermal arrests were desired, and, after microscopic exploration had established the position of the eutectic line at a given percentage of the third component, a thermal analysis was run at that composition. The method of thermal analysis used in the work on the binary Au-Ge system,' a method particularly well suited for a small quantity of material, was used in this work. Clearly defined arrests were obtained, although with some alloys containing platinum

Jan 1, 1946

By Robert I. Jaffee, Bruce W. Gonser

Recent work by the authors1 established the constitutional diagram of the gold-germanium system. Of particular interest in the simple euctectiferous system was the eutectic alloy at 12 per cent Ge, which melted at the very low temperature of 356°C and had an unusually fine eutectic structure. It was of interest to investigate the effect of third components on the melting point, microstructure, and strength of this alloy, and in this work the effect of copper is discussed rather completely, with less complete data on the effects of silver, platinum, nickel, and zinc. No reference in the literature has been found on ternary systems with gold-germanium. Materials and Alloying -4s in the work on the binary system, gold-germanium, the germanium used in this investigation was the cyanide-reduced, 99.9 per cent product, produced by the Eagle-Picher Co. Fine gold sheet from Handy and Harmon, pure platinum sheet from the American Platinum Works, copper magnet wire from Anaconda, nickel wire assaying 99.6 per cent Ni. and electrolytic zinc were used for alloying. To make the alloys, charges of the pure metals were put in a graphite crucible and melted in an atmosphere of pure hydrogen. The temperature was brought up to about I050 C, and the metal allowed to freeze in the crucible. Except with alloys containing zinc, melting losses were consistently less than a few milligrams. When zinc was a component, weight losses up to half the weight of the zinc present were incurred, and the alloy compositions were calculated by assuming the entire weight loss to be that of zinc. In all other cases the composition of the alloys could safely be taken as that of the charge. The ingots, as solidified, were segregated, the heavier primary phases being at the bottom and sides of the ingots and the lighter primary phases at the top, but by sectioning the entire ingot vertically for microscopic examination, this factor could be evaluated. Ingots for microscopic examination weighed one gram, and those for thermal analysis weighed three grams. Thermal Analysis In this work, the eutectic line composition and thermal arrests were desired, and, after microscopic exploration had established the position of the eutectic line at a given percentage of the third component, a thermal analysis was run at that composition. The method of thermal analysis used in the work on the binary Au-Ge system,' a method particularly well suited for a small quantity of material, was used in this work. Clearly defined arrests were obtained, although with some alloys containing platinum

Jan 1, 1946

By D. B Kibler

The Florida hard-rock field extends from Suwanee and Columbia Counties in northwest Florida to south of Croom, Florida, in Hernando County. This area is approximately 100 miles long and varies from 2 to 30 miles in width. Florida hard-rock phosphate was discovered at Hale Springs, about one mile northwest of the town of Dunnellon, Marion County, in the year 1888, by Alber-tus Voight, who lived within a few hundred yards of the spring. Soon after this discovery there was a great influx of land speculators and wildcat operators into the vicinity of Dunnellon. A few years later phosphate was found in Alachua, Levy, Citrus and Hernando Counties, and a great number of small crude plants were built from Newberry, Alachua County, all the way to Croom, Hernando County. Phosphate was found in many places as "outcrops." Looking for outcrops was the first method used by the early prospectors in searching for deposits. Several years passed before any systematic method of prospecting was used. Early Mining The early method of mining hard-rock phosphate was with picks and shovels, wheelbarrows, mules and wagons. The pick-and-shovel method was used altogether until about 1900, when some of the larger operators began to use railroad-type steam shovels of from ½ to I-yd. capacity Picks and shovels, with negro labor loading the phosphate ore (called "matrix") into wooden side-dump skip-cars, was continued until about 1907. From the discovery of hard-rock phosphate until approximately 1905, overburden was removed with mules and scrapers. At some places, the overburden was loaded into side-dump skip-cars by negro laborers and the cars were then pulled up an incline with a steam hoisting engine. The maximum depth of overburden that could be economically removed by hand labor, mules and scrapers was approximately 15 feet. About 1905 one of the companies erected a producer-gas power- I house near Dunnellon and generated its own electricity and about the same time the larger companies began using hydraulic methods for the removal of overburden. From the year 1888 through the year 1939 approximately 75 companies were operating at one time or another in the hard-rock territory, on approximately 500 locations. The hard-rock deposits are more irregular and have higher mining costs than the pebble-rock deposits, therefore 1 they have not been competitive in the domestic market, but because of the higher grade of the hard rock it is in demand for I the export trade. Discovery of phosphate outside of the United States has made it increasingly difficult in recent years for the hard-rock companies to survive, so today i only three companies are operating. These

Jan 1, 1942

By D. B. Kibler

The Florida hard-rock field extends from Suwanee and Columbia Counties in northwest Florida to south of Croom, Florida, in Hernando County. This area is approximately 100 miles long and varies from 2 to 30 miles in width. Florida hard-rock phosphate was discovered at Hale Springs, about one mile northwest of the town of Dunnellon, Marion County, in the year 1888, by Alber-tus Voight, who lived within a few hundred yards of the spring. Soon after this discovery there was a great influx of land speculators and wildcat operators into the vicinity of Dunnellon. A few years later phosphate was found in Alachua, Levy, Citrus and Hernando Counties, and a great number of small crude plants were built from Newberry, Alachua County, all the way to Croom, Hernando County. Phosphate was found in many places as "outcrops." Looking for outcrops was the first method used by the early prospectors in searching for deposits. Several years passed before any systematic method of prospecting was used. Early Mining The early method of mining hard-rock phosphate was with picks and shovels, wheelbarrows, mules and wagons. The pick-and-shovel method was used altogether until about 1900, when some of the larger operators began to use railroad-type steam shovels of from ½ to I-yd. capacity Picks and shovels, with negro labor loading the phosphate ore (called "matrix") into wooden side-dump skip-cars, was continued until about 1907. From the discovery of hard-rock phosphate until approximately 1905, overburden was removed with mules and scrapers. At some places, the overburden was loaded into side-dump skip-cars by negro laborers and the cars were then pulled up an incline with a steam hoisting engine. The maximum depth of overburden that could be economically removed by hand labor, mules and scrapers was approximately 15 feet. About 1905 one of the companies erected a producer-gas power- I house near Dunnellon and generated its own electricity and about the same time the larger companies began using hydraulic methods for the removal of overburden. From the year 1888 through the year 1939 approximately 75 companies were operating at one time or another in the hard-rock territory, on approximately 500 locations. The hard-rock deposits are more irregular and have higher mining costs than the pebble-rock deposits, therefore 1 they have not been competitive in the domestic market, but because of the higher grade of the hard rock it is in demand for I the export trade. Discovery of phosphate outside of the United States has made it increasingly difficult in recent years for the hard-rock companies to survive, so today i only three companies are operating. These

Jan 1, 1942

By J. D. Milburn, H. J. Hill

In quantitative interpretation of electrical logs the presence of clay minerals introduces an additional variable which further complicates an already complex problern. Although recognizing the difficulties introduced as a result of the heterogeneity of natural sediments and despite the present incomplete state of knowledge regarding electrochemical behavior of shales, dissemirzated clay minerals and concentrated electrolytes, it was felt that useful empirical correlations might be obtained from experimental investigation. Six typical sandstone formations, having a wide variety of petrophysical properties, were selected for the study. Approximately 45 samples from each formation were selected to satisfactorily represent the range of pore size distribution within the particular formation, As a matter of general interest, four limestone formations were also included in the investigation. Previously proposed equations relating to resistivity, SP and interrelationship of the two phenomena have, where possible, been tested with data obtained in this investigation. These equations do not satisfactorily describe experimental behavior of samples through all degrees of shaliness or throughout the range of brine solution resistivities normally encountered in logging practice. An empirical equation has been developed which quantitatively relates formation resistivity factor to satirrnting solution resistivity, porosity, and "effective clay content." This relation is indicated to be uniformly applicable to clean or shaly reservoir rocks. It is shown that both the SP and resistivity phenomena of shaly samples are related to the sample cation exchange capacity per unit pore volume. The independent chemical determination of this parameter is thus a means of determining the "effective clay content" of samples. Some implications regarding theory and electric log interpretation of shaly sands are discussed. INTRODUCTION The use of electrical resistivity logs as a means for estimating formation porosity is based upon the original work of Archie1. It was shown by empirical methods that the resistivity of a rock, when 100 per cent saturated with brine solution, could be related to rock porosity through the equation, F = Ro/ RW = f-m.......(1) where F, the formation resistivity factor, is defined as the ratio of the resistivity of the 100 per cent water saturated rock, Ro, to the saturating solution resistivity, Rw; f is the ratio of pore volume to bulk volume; and m is the slope of the average line when the data are plotted on log-log paper. This relation was established for relatively shale-free formations containing water having high salt concentrations, and under these conditions F may be considered constant for a given rock sample. Applications of Eq. 1 to interpretation of logs of shale-free formations has been quite successful when satisfactory formation water resistivity data could be obtained. The most readily available means of estimating formation water resistivity is the self-potential log. Conventional methods of SP log interpretation involve correct.

Jan 1, 1957

By E. P. Abrahamson II

The effect of transitiorz element bitzary solid solution additions upon the recrystallization temperature of columbium has been investigated. The elements Mn, Fe, Co, Ni, W, Re, and 0s OWY the yecystallization tempel-aturye, while Ti, V, CY, zr, Mo, Ru, Rh, Pd, Hf, Ta, W, Ir, and Pt mise it. A correlation is kloted between the rate of chatzge of recyystallization tewaperature with atomic hercent solute and the free atom electron configuration of the solute element. The work of Abrahamson and Grant1 has indicated a correlation between the free atom ground-state electron configuration and the change in brittle-ductile transition temperature per atomic percent solute in chromium base alloys. A similar correlation was observed for the change in the recrystallization temperature of iron base alloys by Abrahamson and Blakeney and Abrahamson. It was also observed that the limit of rapid change in recrystallization temperature also Correlated with the electronic configuration of the free solute atom. A further study by Abrahamson4 established the generality of the phenomenon by yielding similar results for vanadium base dilute alloys. It has been inferred that a peri- odicity might be expected with changes in the free atom electron configuration of the solvent element. This study is the third set of transition base metal systems to be studied. It is the object of this and further studies on transitional metal systems to detect any systematic variations in the observed correlation which might contribute to the understanding of the electron interactions which are evidently exhibiting themselves. PROCEDURE All alloys were made using 99.9 pct Cb with 0.030 Ta, 0.016 0, 0.0025 H, 0.003 Mo, 0.021 Fe, 0.005 C, and 0.002 N. The solute elements were 99.9 pct pure. According to the published binary phase diagrams5 and metallographic examinations at X750, all additions were in solid solution. The alloys were arc melted and remelted 6 times in the form of cubic 400-g buttons under an argon atmosphere. They were then hot forged at 1200oC to 0.6-in. diameter, annealed for 3 hr at 1200o under argon, and furnace cooled. The specimens were then machined to 0.400 -in. diameter. Grain size was checked and found to remain essentially constant at 150 i 30 grains per sq mm. The specimens were then cold swaged to 0.187-in. diameter, yielding 46 i 1 pct cold work. All alloys were chemically analyzed for the major solute. Analyses of random samples indicated that the impurity content remained at the values of the starting material. Tungsten contamination was held to 0.007 pet.

Jan 1, 1962

By W. Joseph Schlitt

The characteristics of successful heap and dump leach operations are first reviewed in a general sense. Then attention is focussed on solution management as the key operating parameter in leaching. The first part of this study includes a detailed conceptual analysis of fluid flow. This identifies the relationships between the flow parameters and the chemical and physical properties of the rock. Then these relationships are extended to practical systems in order to establish optimum application rates and solution distribution methods for heaps and dumps. The study shows that there is a maximum effective application rate, qmax, that can be used without causing solution short circuiting. This rate is a function of the minimum permeability in the heap or dump. The expression for qmax is qmax (gal/ft2/hr)[~]0.9 k (darcies). Here k is the appropriate permeability coefficient. Using this equation, qmax is found to be on the order of 1 gal/ft2/hr (4 cm/hr) for a typical heap or dump leach operation.

Jan 1, 1984

By R. T. Hukki

The purpose of this paper is to present an analysis of the unit operation of grinding and the circulating load, of the unit operation of classification and the circulating load, and of the two superimposed into one operation to clarify the overall trend of events relating the circulating load to the capacity of the closed circuit grinding system, to the specific surface area of the final fine product, and to the energy consumption in the closed circuit grinding process. The capacity of the closed circuit is frequently, if not always, determined by the classifier, not by the mill." With these words a startling idea was expressed in 1934 by Dorr and Anable.1 Today, some 30 years later, grinding in closed circuit is universally accepted. The basic curve presenting the capacity factor versus circulating load by Dorr and Anable has obviously been so convincing that not a single authoritative investigation regarding its validity seems to have been conducted. The only major subject of discussion appears to have been the relative merits of the 'high' and 'low' circulating loads. It seems to be universally agreed that much still remains to be hoped for regarding sharpness of classification. BASIC CURVES FOR CONSTANT MILL LOADING Fig. 1 presents a series of basic curves for constant mill loadings in function of new feed rate and circulating load. The mill loading indicated by each curve is the sum of hourly tonnages of new feed and circulating load. For example, at point 0 where the new feed rate is 100 t/h and the circulating load 300%, or 300 t/h, the mill loading is 400 t/h. In closed circuit grinding, the new feed rate corresponds to the rate of finished classified material produced. Although at this stage the effect of classification is excluded in construction of the basic curves, the trend indicated by these curves is strongly in favor of closed circuit operation under relatively small circulating loads. A study of the curves reveals that, e.g., a mill operating under a load of 200 t/h at point P where the new feed rate is 50 t/h and the circulating load 300% will also operate under the same total load at point Q where the new feed rate is 100 t/h and the circulating load 100%. In other words, a reduction of circulating load from 300% to 100% would lead to doubling of the new feed rate and doubling of the rate of production of finished material, without a change in the total mill loading. The obvious question is, is it also possible by a reduction of the circulating load in a closed industrial grinding circuit to increase the production of finished classified material and, if so, how and to what extent? GRINDING TESTS BY DORR AND ANABLE In a series of open circuit grinding tests reported by Dorr and Anable1 limestone was ground at various feed rates in a 3-ft ball mill. The original data are reproduced in Table I. It can be seen that with the increased feed rate production of -65-mesh material increased and the energy consumption per unit weight of the -65-mesh product decreased in the reverse ratio. The data shown in Table I have been reproduced for our purposes in Table II for a scale corresponding to grinding in a large industrial mill. This reproduction

Jan 1, 1968

By Erminio Ferraris

The development of the mining industry in Sardinia dates from the application of the mining law of 1859, which, following the example of the French mining law of 1810, declared prospecting to be free, and suppressed the liens which had previously attached the mine to surface ownership. The first ore-dressing plants were installed by German engineers upon the classic models of the Harz and of Freiberg; but in 1880 Sardinia commenced to open a new path in that industry, utilizing the refuse from the mining of rich, selected minerals, of poor deposits, and of mixed ores. The first experiments were attempte to adapt the existing apparatus to local conditions, but progress was rapid, as was demonstrated already in 1887 by the calamine-dressing mill of Monte-poni, which exhibited a large plant designed in accordance with the experience gained, and constructed entirely in the machine-shops of the country, after new models and designs.' This example was followed two years later by the Sociékté Malfidano, with a large plant of the same sort, and with the reconstruction of the old-fashioned dressing-mill, built in 1880 by the Austro-Belgian Company. In 1900, almost all the mines of Sardinia had adopted the new designs, and the island contained 32 concen-trating-works, employing 2,000 h.p. and 1,924 workmen. Since that time, there have been added a large dressing-mill for blende at Gennamari, and two for calamine, at San Giovanni and at Masua, while the establishments at other mines have been enlarged and perfected. For the treatment of intimately mixed ores, a large establishment has been built upon the new

Jan 1, 1909

By Ramesh Malhotra, Hubert E. (Deceased) Risser

THE WORLD Coal, as a source of energy and as a source of coke for the smelting of iron ore, has contributed significantly to the development of every major industrial nation of the world A number of other materials-oil, natural gas, wood, waste materials, water, and, in some applications, nuclear materials -can be used satisfactorily for the generation of heat and energy, but no generally acceptable and economic substitute has been found to supplant coke in iron ore smelting in conventional blast furnaces. In the development of many nations, coal provided the initial source of energy for industrialization Although coal provided more than 60% of the world's energy each year between 1902 and 1929, oil and natural gas have surpassed it as a source of energy in more recent years (Ref. 38, pp. 439, 441). Despite the rapid rise in importance of other fuels, a considerable amount of the total energy consumed is still supplied by coal. In 1972, coal supplied 32.1 % of the total world energy (Ref. 49, pp. 2-3). To produce this energy, 2.4 billion metric tons of coal were consumed-a 16.0% increase over the amount consumed in 1962. A further worldwide increase in the amount of coal used to supply energy is projected. According to Dr. Karlheinz Bund (Ref. 37, p 145), by the year 2000 world coal consumption could amount to 5 billion mt. Coal Resources and Production The world's coal resources are estimated to be about 10,872 billion mt. Of this amount, 1,297 billion mt are classified as measured reserves (Ref. 50, p. 57). In Table 13.A.1 the estimates of measured reserves and total resources, recent production, and the ratio of measured reserves to production of major coal-producing countries are shown. About 90% of the world's total coal resources are estimated to be in the USSR, the United States, and China. The United States ranks second in the world in coal resources, but in measured reserves, it ranks first. Because of the difference in rate at which available coal resources are being exploited, the ratio of measured reserves to actual production varies widely among areas and countries. France and East Germany have the lowest reserve to production ratios, 49.5 and 60.1, respectively. In Mexico, where currently less than 5 million mtpy of coal are produced, the ratio is 1257 For all countries producing more than 200 million mtpy of coal (except East Germany), the ratio ranges from 350 to 670. For the United States, the ratio is 669.7. In Table 13.A.2, distribution of the world's coal resources by area is shown. More than 97% of the total resources lie in the northern hemisphere, with Asia, including European USSR, having 63.4%. Only 1.9% of the total coal resources are estimat-

Jan 1, 1976

By P. E. Fitzgerald, Dana G. Hefley

An apparatus has been developed which permits selective acidizing of producing formations and determination of the relative permeability of a formation by fluid injection. This apparatus, known as the "Electric Pilot," is an electrical fluid-interface locator, the electrical circuit of which is completed when the elec-tcodes in the well are in contact with a conductive fluid, but is not completed when immersed in a nonconductive fluid. Results are given for a number of typical acidizing applications in which the interface locator was used to control the acid. Description of the method used in making a permeability survey and the data obtained are also included. Introduction Early in the development of the acidizing industry it was recognized that one of the most important problems to be solved in the chemical treatment of any well was the determination of what part of the exposed section in the well should be treated and the selection of the technique that would insure treatment of the desired section. One of the first techniques adopted for assisting in proper placement of the acid, which is still widely used, is the method of confining the acid to the producing section by means of a column of oil, which is known as the Carr1 method of treatment, named for its inventor. In the Carr method it is not possible to selectively acidize a specific horizon in a multiple pay zone, so many improvements and modifications of the basic Carr technique were devised to accomplish this result. In order to overcome the possibility of the acid entering an undesirable section at the bottom of the well, organic gel plugs2 were introduced into the bottom of the well, which served to halt the entrance of the treating solution into these zones.3 These chemical gels later liquefied because of bacterial action and were removed from the well with the produced fluid. In fields where a gas cap existed, as in the Goldsmith field of Ector County, Texas) it was imperative that a treating technique be evolved that would confine the acid to the oil-producing zone below the gas cap. By so doing the oil-productive capacity of the wells could be improved without greatly increasing the gas production and this would lower the gas-oil ratio of the well. ~h, technique employed4 to achieve this effect utilized two pumps, one for pumping the chemicals down the tubing, the other to simultaneously force oil down the annular space to confine the acid to the oil pay, thus preventing the chemicals from entering the upper gas-bearing zones. Experience has indicated that gas wells and certain types of oil wells respond better to ''stage treating.'' This technique utilizes the introduction of a certain volume of acid and then allows the well to ''clean up99 before introduction of the next and usually successively larger stages. Where such methods are employed steps must be taken to keep a disproportionate share of the acid from entering the most permeable and previously treated Zones. A special tool

Jan 1, 1944

By P. E. Fitzgerald, Dana G. Hefley

An apparatus has been developed which permits selective acidizing of producing formations and determination of the relative permeability of a formation by fluid injection. This apparatus, known as the "Electric Pilot," is an electrical fluid-interface locator, the electrical circuit of which is completed when the elec-tcodes in the well are in contact with a conductive fluid, but is not completed when immersed in a nonconductive fluid. Results are given for a number of typical acidizing applications in which the interface locator was used to control the acid. Description of the method used in making a permeability survey and the data obtained are also included. Introduction Early in the development of the acidizing industry it was recognized that one of the most important problems to be solved in the chemical treatment of any well was the determination of what part of the exposed section in the well should be treated and the selection of the technique that would insure treatment of the desired section. One of the first techniques adopted for assisting in proper placement of the acid, which is still widely used, is the method of confining the acid to the producing section by means of a column of oil, which is known as the Carr1 method of treatment, named for its inventor. In the Carr method it is not possible to selectively acidize a specific horizon in a multiple pay zone, so many improvements and modifications of the basic Carr technique were devised to accomplish this result. In order to overcome the possibility of the acid entering an undesirable section at the bottom of the well, organic gel plugs2 were introduced into the bottom of the well, which served to halt the entrance of the treating solution into these zones.3 These chemical gels later liquefied because of bacterial action and were removed from the well with the produced fluid. In fields where a gas cap existed, as in the Goldsmith field of Ector County, Texas) it was imperative that a treating technique be evolved that would confine the acid to the oil-producing zone below the gas cap. By so doing the oil-productive capacity of the wells could be improved without greatly increasing the gas production and this would lower the gas-oil ratio of the well. ~h, technique employed4 to achieve this effect utilized two pumps, one for pumping the chemicals down the tubing, the other to simultaneously force oil down the annular space to confine the acid to the oil pay, thus preventing the chemicals from entering the upper gas-bearing zones. Experience has indicated that gas wells and certain types of oil wells respond better to ''stage treating.'' This technique utilizes the introduction of a certain volume of acid and then allows the well to ''clean up99 before introduction of the next and usually successively larger stages. Where such methods are employed steps must be taken to keep a disproportionate share of the acid from entering the most permeable and previously treated Zones. A special tool

Jan 1, 1944

By H. I. Smith

CONTENTS [PAGE Problems of Mining and Processing Mineral Aggregates (TP 1649, by NATHAN C. ROCKWOOD, Min. Tech., Nov. 1943. Discussion by H. I. SMITH, IRVING WARNER and the author) t Mica in War (TP 1749, by RUSSELL G. WAYLAND, Min. Tech., July 1944. Discussion by H. S. SPENCE) 3 Three Roofing-granule Plants in Pennsylvania (TP 1787, by RICHARD M. FOOSE, Min. Tech., Jan. 1945. Discussion by H. I. SMITH)5 Methods of Prospecting and Mining Optical Calcite in Montana (TP 1896, by E. W. NEWMAN, Min. Tech., Nov. 1945. Discussion by W. C. STOLL and FRANK C. ARMSTRONG, and the Glass Sand and a Glass Industry in Puerto Rico (TP 1939, by HOWARD A. MEYERHOFF and J. EARL FRAZIER, Min. Tech., Nov. 1945. Discussion by A. M. SHORT, the author and PHILIP The "Bonanza" Mica Operation of Purdy Mica Mines, Limited, Mattawsn Township, Ontario (TP 2154, by HUGH S. SPENCE, Min. Tech., March 1947. Author's supplement and discussion by JOHN A. BAKER, E. J. LINTNER, and the author) 7 Process for Manufacture of Dead-burned Magnesite and Precipitated Calcium Carbonate from Dolomite (TP 2155, by ROBERT D. PIKE, Min. Tech., March 1947. Discussion by 0. C. RALSTON and the author) 9] Problems of Mining and Processing Mineral Aggregates H. I. SMITH*-Mr. Rockwood has presented a very interesting paper on problems of the mineral aggregate industry in meeting specification requirements, which apparently are ever decreasing in tolerance. May I call attention to a method of mineral separation that may be of value in separating the flat aggregate particles from the more angular ones and possibly the soft particles from the hard. The method is the old Pardee spiral coal separator, extensively used many years ago in the anthracite region for separating rock and bone from anthracite coal.1 The success of the process depends on the relative coefficient of friction of the aggregate particles to be separated rather than on their relative specific gravity. The flat and the soft particles normally will slide more slowly and stay on the inner spiral until they reach the delivery spout at the bottom whereas the round and the angular pieces will work to the outer edge and fall off into a spiral of larger diameter. The capacity of the separator is increased materially by having three inner spirals, with a separate feed for each, on which the particles travel more slowly, and one outer spiral from which the more rapidly traveling pieces slide to the bottom. IRVING WARNER*-I have long been convinced that much of the equipment and methods developed for mineral dressing might be advantageously employed in the mineral aggregates industry, and more particularly in the production of sand and gravel. The two industries are not well acquainted with each other or their problems. They have separate trade publications. The difficulty with the situation is that there are few large sand and gravel plants and few provided with technical services in their own organization or by retainer. This does not refer to engineering organizations but rather to improvement of plant technic in processing.

Jan 1, 1947

TWO American students entered the École des Mines in 1856, Joseph Lesley of Philadelphia and Thomas Egleston of New York. Lesley remained there only one year, but Egleston completed the whole curriculum and received, in 1860, the certificate given to foreign students in lieu of the official appointment and degree awarded French students. Returning home he received, in 1861, an appointment by Secretary Henry of the Smithsonian Institution to sort, label, and arrange the collections of mineralogical and geological specimens which had been brought to Washington by the various government exploring expeditions in the West. Thomas Egleston, born in New York City on Dec. 9, 1832, was the son of Thomas Jefferson Egleston, who came to New York from Massachusetts some time before 1829. The first American ancestor of the family had settled near Dorchester in 1630. T. J. Egleston and Joseph Battell established an iron-merchandizing business* at 166 South St., New York, which prospered exceedingly; so much so that when Thomas J. Egleston died, July. 12, 1861, his estate had a book value of over $500,000. Since an older brother was expected eventually to take over this business, Thomas, who had a liking for science, went to Yale, where he completed the regular four-year course in 1854, and then remained another year for advanced work in chemistry under Silliman. One biographer says he went to Europe, early in 1856, on account of his health. It was the custom at that time for wealthy young men to make a tour of Europe after completing their academic education. Whatever the precise reason, he began scientific work at the Jardin des Plantes in Paris but soon trans-

Jan 1, 1941

By J. Szekely, R. G. Carr

A method is presented which allows the extension of the Ergun equation for the description of gas flow through packed beds in a nonisothermal regime. The pressure profile in the packed bed is calculated by the establishment of a differential energy balance and the iterative integration of the integral equation thus obtained. The predicted values of the pressure distribution were found to be in excellent agreement with those measured experimentally, under carefully controlled conditions. MANY metallurgical processes are carried out in systems in which a gas stream and particular solids are contacted in a packed bed arrangement under strongly nonisothermal conditions. Ore sintering, the firing of pellets, the operation of regenerators, blast furnace technology may be quoted as examples. Although many aspects of these systems have been extensively studied, an important variable, the pressure distribution and the relationship between pressure head and flow rate, received only little attention to date. The purpose of the paper is to present a brief survey of relevant previous work, followed by a more rigorously formulated model, together with experimental results that afford ready comparison of theoretical prediction with measurement. PREVIOUS WORK A great deal of information is available in the chemical engineering literature on flow phenomena in porous media.' The experimental results pertaining to gas flow were generally interpreted in terms of models based on the analogy with the flow through a pipe of an incompressible, isothermal fluid. Of the numerous, semiempirical correlations that have been proposed to relate pressure drop to the volumetric flow rate, the Ergun equation2 is generally regarded as the most satisfactory: The strong nonisothermality of the metallurgical systems previously mentioned would render the direct applicability of Eq. [I] rather doubtful as both density and viscosity are temperature-dependent. Ergun has used Eq. [I] for estimating the pressure drop in the stack region in the blast furnace by neglecting the viscous term and evaluating the gas density at the integral mean temperature.3 His calculations showed a rather variable agreement with measurements and it was uncertain whether the discrepancy found was attributable to the model or to the inaccuracies often inherent in measurements on an industrial scale. Other work on nonisothermal systems includes studies of the pressure distribution in sinter beds, notably by Voice et a1.4 and Mitchell.5 Mitchell developed a method involving the subdivision of the bed into several zones and assigning average property values (temperature, gas density, and so forth) to each zone. The pressure drop was then calculated for each zone by using an expression analogous to Eq. [I]. This method was an improvement over that described by Voice et al., but even here prediction and measurement differed by as much as 30 pct in certain cases. It would appear that no really satisfactory methods are available for the prediction of pressure drop in packed beds with a strongly nonisothermal regime. The method involving subdivision of the bed into small enough sections, that may be regarded as isothermal, would show promise; however its validity has to be tested by both a rigorous formulation and comparison with experimental data obtained under controlled conditions. In the following a formulation is presented together with experimental data allowing ready practical evaluation of the theory. FORMULATION The mechanical energy balance over an infinitesimal section of a packed bed may be written as follows:

Jan 1, 1969

By V. C. Williams

Some of the physical, chemical, and electrical processes for conversion of saline water to potable or industrial water are economically surveyed from an engineering viewpoint. Since all these processes require energy for drive and equipment for containment, the correlative economic factors are developed which indicate directive influences in the choice of particular regional processes. The supply of natural waters and its distance also affect decision. Any one process will probably not prove dominant in the field because auxiliary considerations such as the saline water source; types and continuing availability of fuel; electric power use or recovery; area economic status and advancement; and the political pressures of population, group demands, and land use tend equivocally to obscure capital and operation cost decisions. Basic engineering considerations, data, and economic factors are presented to assist in the direction of these decisions. An exploding world population, increasing industrialization, advancing standards of living, and the desire of less-privileged nations for betterment focus attention sharply on a major problem: water. *19 Up to now, in retrospect, people have had it relatively easy in the handling of this problem. All the better dams in the most advantageous sites, the better aquifers, the shortest aqueducts have been built. In another phase of the problem, concern is evident that wastes cannot indefinitely be disposed of merely by keeping them dilute and discharging them promiscuously. 7-9 And, perhaps, as past civilizations have done,l5 water, watersheds, streams, and irrigation may have been mismanaged or, at the least, not adequately studied.3,5,36,37 In this last is perhaps the core of the problem. As Gross states, "Ignorance and too often, indifference are contributing factors. Archaeology and theology both furnish ample testimony to the existence of rich lands where deserts now stand; it was man who ravaged his land. Unless education is a companion to water development, development might as well be forgotten. But without water, there is no beginning."13 The U.S. is showing increasing concern about its water for predictions are that by 1980 the daily withdrawals will be 494 billion gal, a figure nearly equal to the dependable supply.Is This is based on a conservative projected population of 230 million. The major categories of withdrawals are: To make available this per capita average of 2150 gal per day will require an expenditure of $219 billion over the next 20 years. The U.S. is not alone in this concern. The United Nations shows as arid zones of the world: all of Africa north of the equator and south of the 20's parallel; all of the Arabian peninsula; all of the middle east and Iran, Iraq, Pakistan, Afghanistan, northern and central India; a great band about 1000 miles wide along the 40'~ parallel from the Caspian Sea east across Russia through China to the Pacific Ocean; all of Australia except the coastal plain; the Caribbean Islands; the western nations of South America; and the western third of the United States and of Mexico. With one quarter of the earth's 57,500,000 sq miles of land thus suffering from lack of good water, increasing attention goes to the treatment of brackish and sea waters. The U.S. has been a leader in this field4,12, 16123,24 through its Office of Saline Water in the Dept. of Interior because even now some of its cities and regions are short of potable water. 11j'7,M Industrial water is also of vital concern as a result of ever higher industrialization1,14122 Other nations, among them JaPan, Israel,13188 Germany, Union of South Africa, Australia, Netherlands, France, Yugoslavia, Russia, and groups such as the Organization for European Economic Cooperation (OEEC)' are also diligent. The objective is low cost water, which means that both technology and economics have prominent roles in saline water conversion processes. TECHNOLOGY: SALINE WATER CONVERSION A number of reviews of methods have been made, principally by staff members of the Office of Saline Water (U.S. Dept. of Interior). Jenkins,31'32 Gillam,34p

Jan 1, 1962

By L. O. Millard

Stockpiles in the minerals industries serve a wide variety of purposes. Usually they are for surge between stages of processing, for a dependable plant feed in anticipation of delays, or to provide for single or double shift mining in conjunction with round-the-clock milling. Often surge piles are necessary to insure a uniform feed to continuous process equipment where the material is received as an intermittent or variable flow. Extensive use is made of stockpiling for the primary purpose of blending. The significant fact about all stockpiling of bulk materials is that storing and reclaiming are two separate and distinct operations and they may present widely different problems in the selection of methods and tools.

Jan 10, 1959