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A comprehensive, systematically structured mineral evaluation system is a prime requirement for objectively assessing mineral supply impacts on the economy. The Minerals Availability System developed by the US Bureau of Mines is a step in the right direction-a well-organized unraveling of the mineral supply network, structured in format to allow quick retrieval for rapid listing, supply analysis, and other uses.

Jan 9, 1977

By H. R. MacMichael

IN the Chihuahua district of Mexico the first smelting was that inaugurated by the early Spaniards for the production of silver bullion. The ores treated were high in silver and lead. Silver-lead bullion was first produced and the lead then cupelled to waste, with the slag produced running from 5 to 10 per cent in lead and about 25 to 40 per cent silica. The ore was mined by peon labor, often being carried out of the mine on the backs of the men, and then transported by burros. Furnaces were of primitive construction, adobe or rock being used in the earlier period and some iron later when it became available. Blast was originally supplied by bellows manually operated. Such smelting was carried on to a greater or lesser extent until the latter part of the last century, and so extensive was it that many hundred thousand tons of leady slag remained to be treated in modern furnaces. In 1908 a smelter was placed in operation at Chihuahua by the American Smelting & Refining Co. At that time the ore was practically all oxidized and by a little selection, self-fluxing mixtures could be made, but with an average lead content usually under 10 per cent and sometimes as low as 7 per cent. The bullion produced would carry from 150 to 300 oz. silver per ton. The slag made approximated 30 per cent SiO2, 42 per cent FeO, 21 per cent CaO, and perhaps 3 per cent Zn. The ores varied from hard, coarse material which was passed through a crusher, to a considerable quantity of fines which had to be smelted as it came, as crushing the coarse was the only preparation for smelting.

Jan 1, 1933

By R. J. Wasilewski

The resistivity of single crystal and of polycrystalline metal has been determined between 77° and 1500°K. At comparable purity, the resistivity of the poly crystalline metal differs from that calculated from the axial resistivity values. The observed anisotropy is briefly discussed. A zone refined titanium rod was obtained from F. J. Darnell.1 The hardness of this material varied between 52 and 64 Bhn, indicative of its high purity. Single-crystal specimens were cut out from a few large grains, elongated predominantly along <1120>. Resistance measurements were taken with specimens immersed in liquid nitrogen, solid CO,, and acetone, and at ice point, and then recorded continuously between room temperature and 1200" to 1250°C, throughout at least one full heating and cooling cycle at a rate of l° to 3°C per min. Four probe measurements were taken using spot-welded Pt wire 0.003 in. in diam as potential terminals, and the electromotive force drop was determined poten-tiometrically. A standard resistance, in series with the specimen, served to determine the dc current (generally about 80 ma) in the circuit. In order to eliminate spurious thermal electromotive forces, the circuit polarity was automatically reversed at 5-min intervals, and the mean of the two adjacent readings taken as the true electromotive force value. The discrepancy between the readings at reversed polarities was generally of the order of a few µ v, as all the spot-welded joints were contained within a volume of some 5 mm cube. The relative resistance (R T/R 273o) variation with temperature is tabulated below, and shown in Fig. 1, as measured during the first heating run. The discontinuous drop in resistance, corresponding to the a — ß transformation, started for both initial orientations at 1170°K; i.e., some 15°K above the accepted value of 882°C (1155°K). Judging by the temperature ranse of transformation—some 30°K for the single crystal and 20°K for the polycrystal-line specimen—this difference is believed caused by the solution of the oxygen present initially on the surface of the specimens. The general trend of the resistance vs temperature shown in Fi,;. 1 is in qualitative agreement with that previously reported for polycrystalline titanium metal. 2,3 Significant anisotropy is, however, indicated by the crossing over of the curves for <1120> and <0001> orientations at 273°, 650°, and 1000°K. This difference is most startlingly shown in Fig. 2, where instantaneous temperature coefficients (l/R273 x dR /dT) are plotted for both the first heat-

Jan 1, 1962

By F. B. Plummer

GEOLOGICAL engineering as applied to oil fields, or production geology as some prefer to designate the profession, is designed to fill in the border line between pure geology and pure petroleum engineering. Therefore the production geologist may, and frequently does, reach out to grasp much in both professions. Where once the oil geologist was content to pick samples out of a mud ditch and try to identify oil formations from the color or appearance of flakes of rock bailed out of a well, now he has to deal with graphic electrical logs obtained by intricate and highly technical electrical apparatus and to make exact determinations to the fraction of an inch of the depth and thickness of an oil sand 2 miles below the surface.

Jan 1, 1944

By Reynold Q. Shotts

Data are presented which show that fractions of varying densities-from the same coals are oxidized at different rates by nitric acid. From oxidation data, the approximate quantity of "bright" and "dull" components may be calculated. Definite relationships between oxidation rate and rank are shown for density fractions from at least two bituminous coals.

Jan 1, 1950

By J. B. DILWORTA

Ix estimating the investment-value of a mining-property or plant, the value of which decreases with operation, it is often necessary to know the sum which must be set aside periodically from earnings either to renew the plant at the end off its life, or to return the capital involved when the earning-power of the investment has ceased. In such cases it is manifestly incorrect simply to divide the total amount to be retired by the number of periods (usually years) during which the investment shall be active, as the sums periodically withdrawn have a certain interest-earning power varying with the length of time they are held and with the opportunities afforded for their investment. But though the interest may be considered to accumulate regularly year by year, yet the rules and tables for compound interest do not apply, owing to the addition of a fixed amount of new money to the total at the beginning of each period. Briefly stated, the problem of sinking-funds is, to find the amount of money which must be periodically set aside at a certain rate of interest, regularly compounded, to yield a certain sum in a given, length of time. In most practical cases the interest is compounded and the fresh amount added at the same time-usually once a year-and on this basis the problem may be solved as follows

Nov 1, 1909

By D. M. Dunbar, H. C. SCHLILTZ

HE Chile Exploration Co.&apos;s mine and reduction plant are at Chuquicamata, Chile, on the eastern edge of the Atacama Desert, 163 miles northeast of Antofagasta, 80 miles from the Pacific Ocean, and 70 miles from the Bolivian border and at an altitude of 10,000 ft. Guggenheim Brothers started prospecting in -4pri1, 1912, and purchased the main property, in October of that year; the property is now controlled by the Anaconda Copper Mining Co., which acquired it in January, 1923. The orebody strikes north 10" east, and has a dip of about 64" west. The length, as far as developed, is approximately 1% miles; the width averages sightless than mile.

Jan 1, 1925

Secretary&apos;s Note.—M. Beutter&apos;s introductory remarks, being simply a resume of Prof. Hone&apos;s paper, with a disclaimer of the intention to criticise it, have been omitted, and the remainder of this contribution lias been somewhat condensed, without omitting any essential statements. M. Beutter, St. Etienne, France* (communication to the Secretary†):—Professor Home observes in $ 47 of his paper that the fluid-compression of steel diminishes piping by forcing the liquid and viscous metal, while the pipe is forming, or after it has formed, into the space so produced, and proceeds to study the different systems of compression, alluding to the four principal ones: Whitworth, Illingmorth, Williams, and Harmet. Having now for many years had in use the last-named process, that of M. Henri Harmet, I shall attempt to explain the difference between it and the three other processes named. The Harmet process does not diminish or close the pipe after it has formed, but prevents it from forming; and I may add that it is the only one of the four which produces this result. It is true that the Whitworth process diminishes the interior hollow by localizing it around the central axis of the ingot, which is afterwards removed by trepanning the hollow forg-ings for which this process has been, I believe, always exclusively used. The Illingworth and Williams processes, which, in my opinion, are applicable only to small ingots, of necessity allow the pipe to form, or (to use the expression of Professor Howe) they allow the surface of the lake of liquid steel to sink into the interior of the ingot, since they involve waiting a certain time before compression, so that (in the case of the Williams process) the ingot-mold can be opened without danger, or (in the case of the Illingworth process) so as to avoid the risk that the steel may enter the grooves produced by the removal of the distance-bars. During the time required for

Jan 1, 1908

By D. H. Gieskieng

THE advent of curved jaw crusher wearing plates made an approach other than segmental layout analysis desirable for prediction of capacities. For some time it had been known that the drawing board capacities of crushers using these plates had to be considerably modified by complicated experience factors to achieve agreement with results. Because these apparent capacities could be readily increased severalfold by minor crushing chamber shape changes, it was necessary that the utmost precaution be taken in predicting capacities of jaw plates modified for nonchoking, special wear characteristics, or any other reason. To this end the laboratory and field tests outlined by the author in a previous paper1 were made on Blake-type jaw crushers. The results of these tests were summarized in a simple first degree equation applicable to crushers using either straight or curved jaw plates. This equation first outlines the maximum capacity potential of a given crusher, then reduces this figure in accordance with installation circumstances by means of a realization factor. It was found subsequently that this equation, with the addition of an eccentric throw factor, is applicable to standard types of single-toggle or overhead eccentric jaw crushers as far as maximum capacity potential is concerned. However, these crushers have realization factor curves somewhat different from those outlined for the Blake type. While this paper is concerned principally with standard type single-toggle crusher capacities, the evaluation of data obtained with these machines is simplified by comparative reduction to the 10 x 7 in. Blake-type equivalents upon which the summary of the preceding paper was made. Convertibility of data from one type of crusher to the other also tends towards confirmation of both. The agreement of these data is sufficient to be considered complimentary. Consequently the feed factors, f, previously reported for Blake crushers are slightly adjusted to an average with the single-toggle crusher results. Blake-type equation: C = f.d.w.y.t.n.a.r [I] Single-toggle type equation: C = f.d.w.y.t.n.a.e.r [21 where C is the capacity in short tons per hour through the crusher, f is a feed factor, dependent upon the presence of fines in the feed, and the surface character of the jaw plates used. Values of f : Smooth Plates Corrugated Plates With normal fines 0.0000414 0.0000319 Fines scalped out 0.0000368 0.0000252 Large pieces only . 0.0000312 0.0000215 d is the apparent density of the broken product in pounds per cubic foot. (If the true specific gravity of the feed is known, 40 pct voids may be assumed and d becomes 37.4 times sp gr). w is the width of crushing chamber in inches. y is the openside setting of the crusher, in inches. In the case of corrugated jaw plates it is measured from the tip of one corrugation to the bottom of the valley opposite. t is the length of jaw stroke in inches at the bottom of the crushing chamber. It is the difference between open and close-side settings. n is rpm, or crushing cycles per minute, a is the nip-angle factor. It is unity for 26" and 3 pct greater for each less nip-angle degree. A nip-angle of 20" has an a value of 1.18, and an angle of 30" has an a value of 0.88, see Fig. 1. r is the realization factor. It is unity for perfectly uniform choke feeding and usually less for actual operating conditions according to the method of feeding used and the probabilities of hang-ups involving the size of feed and crusher opening. Approximate values are given by the curves in Fig. 2. These values are further reduced by intermittent feeding, e is the throw or diameter of gyration of the single-toggle crusher eccentric in inches. As evident in Fig. 1A, variation of feed size will generally have little effect on nip-angle if both jaw plates have flat areas. Jaw plates having continuous curvature, as in Fig. 1B will have different nip-angles, depending upon the size of feed. For test work as described in this paper this effect was accounted. For general compilation of capacities for average feeds it is suggested that the nip-angle be taken at the various settings computed, at an arbitrary level, such as is indicated in Fig. 1C. Data Evaluation To bring the Blake and single-toggle type crusher capacity test results to common terms for evaluation, all data are converted to terms of 10 x 7 in. Blake-type performance at conditions of 100 lb per cu ft, 10 in. chamber width, 250 rpm, 0.65 in. stroke, 3-in. openside setting, and 18° nip-angle. (The nip-angle of the 10x7 in. Blake is 18° at 3-in. setting.) The single-toggle crusher performances are also divided by the eccentric throw to bring this effect to unity. As outlined,&apos; laboratory and field tests made on Blake-type crushers ranging from 10 x 7 in. to 60 x 48 in. were summarized along the foregoing conditions of speed, stroke, etc. This resulted in groups of data which correspond to feeds with fines, feeds

Jan 1, 1952

Jan 1, 1922

Jan 1, 1922

By T. A. Rickard

"The chief aim in style ought to be absolute precision", said Flaubert. "There is only one noun that can express your idea, only one verb that can set that idea in motion, and only one adjective that is the proper epithet for that noun." Let this be your motto. The engineer tries to be exact in all his measurements; he should measure his language with similar exactness. The technical term is a word of precision. It is not only precise, it is a word-saver. To the unscientific the earth is "a ball slightly flattened at the poles, something like an orange". To the scientific it is "an oblate spheroid". A similar contrast between the precision of the technical term and the vagueness of common words is afforded by a conversation between Ivanhoe and Rebecca: "&apos; What device does he bear on his shield?&apos; asked Ivanhoe. &apos;Something resembling a bar of iron, and a padlock painted blue on a black shield.&apos; &apos;A fetterlock and shacklebolt azure&apos;, said Ivanhoe "-using the technical terms of heraldry. It is good style in writing, however, to adapt your manner of expression to the intelligence and experience of the person or persons to whom the writing is addressed. The description of a mine should be worded differently according as it is intended to inform a mining engineer, a director, or a bishop. REMEMBER THE READER. It is as incorrect to write in a technical way for a non-technical reader as to write in a popular way for a technical or scientific reader. That style is best which enables the writer to place himself in the closest mental touch with his reader, to establish the maximum of sympathetic understanding. Let your precision be\ proportioned to your accuracy. You should not try to be precise in matters that do not admit of

Jan 1, 1931

By George F. Kunz

THE recently discovered occurrence of diamonds near Murfreesboro, Pike county, Ark., was brought to. our attention by Mr. Samuel W. Reyburn (Trustee for Messrs. C. S. Stifft, A. D. Cohn, August Zinsser, Jr., and himself, owners of the property), through whose courtesy we investigated the locality of the deposit. Prospecting with the diamond-drill is now in active progress ; but mining operations on a large scale have not yet been ordered. Geology.-The geology as well as the petrography of this interesting locality has been well described by Messrs. J. C. Branner and R. N. Brackett.1 Briefly summarized, the igneous rock in which the diamonds are found is a vitreous peridotite, forming a stock or volcanic neck, which has broken up through the Carboniferous and Cretaceous quartzites and sandstones. After an extensive period of erosion, during which an unknown portion of the neck and presumably a previously existent volcanic cone have been removed, the surface was covered with thin beds of Post-Tertiary conglomerate. The volcanic intrusion was accompanied by the formation of several small dikes of a rock much like that of the main body. One of these dikes cuts across the stock, while another cuts the Cretaceous sandstone, but is overlain by the conglomerate, thus giving a datum for the period of intrusion. So far as known, there was little, if any, metamorphism of the country-rock by the igneous magna, which probably followed an approximately vertical course, so that a more or less vertical extension downward of the igneous body to indefinite depths may be expected.

Mar 1, 1908

By D. Walton

The effect of small solute additions of Cu on the activation energies for creep A1 single crystals were determined over the range from 78° to 850° K. Below 240°K and above 800°K activation energies were unchanged. Between 240°K and 400°K strain aging causes the creep rate to become vanishingly small and the flow stress was independent of the temperature. Between 500° and 50°K the activation energy was increased to 41,000 cal per mole. NUMEROUS investigationsL have shown that solid solution alloying invariably increases the flow stress. A typical example is documented in Fig. 1. Such increases in flow stress can arise from two general effects: Alloying can increase the long-range stress fields through which dislocations must move by causing appropriate clustering; modification of dislocation densities, and so forth. Alloying can also modify the short-range stress fields that determine the activation energies for thermally stimulated migration of dislocations, such as occurs in some solute atom pinning processes, and so forth. Actually both long- and short-range stress-field effects can occur simultaneously. It was the purpose of this investigation to evaluate the effect of alloying on the short-range stress fields by determining the effect of alloying on the activation energies for creep. The resulting observations also shed some light on the effect of long-range stress fields on dislocation processes. EXPERIMENTAL TECHNIQUE Single-crystal bars (6 in. long 3/8 in. wide and l/4 in. thick) of dilute solutions of Cu in A1 were seeded, so as to provide extensive single slip, with the pole of the tensile axis located in the standard stereographic projection as shown in Fig. 2. Each crystal was grown under an argon atmosphere in a graphite mold containing a large reservoir of molten alloy for the purpose of obtaining rather uniform composition of Cu along the length of the bar. The nominal compositions of the alloys which were produced from 99.99 pct Cu and high-purity A1 containing 0.001 pct Fe and 0.001 pct Si, are given in Table I. Strains, measured by linear variable transformers attached by quartz rods to a 3 in. gage section were sensitive to and the stresses were ap- plied by direct loading of the specimen. Temperatures were determined by thermocouples attached to the gage section of the specimens. Below 400°K the apparent activation energies for creep, Q, were determined by the effect of small, rapid changes in temperature of about + 10°K from T1 to T2 on the corresponding creep rates ?1 and ?2 during primary creep according to where R is the gas constant.&apos; Above 400°K, Q was evaluated from Eq. [1] for the secondary creep rate using temperature changes of about ±20K. A typical set of results is shown in Fig. 3. RESULTS AND DISCUSSION The previously obtained effect of temperature on the apparent activation energy for creep of single crystals of high-purity A12 is shown by the broken curve of Fig. 4. The datum points on the same graph illustrate the effect of dilute a solid solution alloying with Cu on the activation energy between 600" and 750°K. Each plotted point is the average of not less than five independent test values between 600° and 750°K and not less than three at other tempera-

Jan 1, 1962

By H. J. Rose

Sixty-five million net tons of coal were carbonized in the by-product and beehive coke ovens1 of the United States during 1924. This tonnage represented 13.4 per cent. of the bituminous coal which was mined in that year. Table 1 shows production and disposition.2 Table 1.—Production and Disposition of Ocen Coke in 1924 By-product Beehive Total Coal charged into ovens, tons................ 49,061,339 15,914,310 64,975,649 Coke produced (exclusive of screenings and breeze), tons............................. 33,983,568 10,286,037 44,269,605 Average yield of coke, per cent ............. 69.3 6.46 68.9 Disposition of coke, per cent.: Blast furnace.......................... . . 82.6 Foundry............................... . 13.3 6.6 Domestic.............................. . 1.4 6.7 Water gas manufacture and all other uses.. 4.6 . 4.1 Cokes having somewhat different combinations of chemical and physical properties are desired for the various uses listed above, and these properties depend primarily upon the types of coal used. American coals have a wide range of purity and coking quality, and the chief subject of this paper will be the selection of coals for the manufacture of blast furnace, foundry, domestic, and water-gas coke, in by-product coke ovens. It will obviously be necessary to include a discussion of those physical and chemical characteristics that are generally considered essential or desirable for each type of coke. Suitable Coals The examination and development of new sources of coking coal are important because of the irregular and limited distribution of high-grade

Jan 1, 1927

By AIME AIME

PRESIDENT Edwin Ludlow and Secretary Bradley Stoughton made their first-trip of the year 1921 to visit Local Sections of the Institute, and talk over Institute interests and policies, during the early part of April. The main purpose of the trip, as usual, was to bring back to headquarters suggestions or comments on the Institute&apos;s work on the part of, members. The trip was made a brief, one by the necessity of remaining in New Fork until Apr. 1, and of the attendance of President Ludlow at a meeting of American Engineering Council and. the organization of a Philadelphia Section of the Institute in Philadelphia on Apr. 16.

Jan 1, 1921

By John J. Collins

Page Purpose and scope..................... 521 Core sampling......................... 521 Sludge sampling....................... 527 Combining core and sludge assays ...... 533 Deviation and surveying of drill holes. . . . 537 Conclusions........................... 544 Acknowledgments...................... 545 Bibliography.......................... 545 Purpose and Scope The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described,25,26,31,32,50,65 but more informa- tion is needed on the results shown by actual mining operations. Core Sampling The accuracy of diamond-drill-hole Sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says; in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change; often indicated by a color change in the sludge water, should be a a for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By John J. Collins

[CONTENTS PAGE Purpose and scopeI Core samplingI Sludge sampling.....7 Combining core and sludge assays13 Deviation and surveying of drill holes17 Conclusions24 Acknowledgments25 Bibliography25] PURPOSE AND SCOPE The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described 25,26,31,32.50,65 but more information is needed on the results shown by actual mining operations. CORE SAMPLING The accuracy of diamond-drill-hole sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says, in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change, often indicated by a color change in the sludge water, should be a signal for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By Leslie Burris, J. C. Hesson

The diffusivity of uranium in liquid cadmium has been measured as a function of temperature by the capillary -bath technique. The diffusivity measurements were made at 4509 500°, 575º, and 650°C, over which temperature range the diffusivity of uranium varied from 1.3 to 2.3 x 10-5 sq cm per sec. The diffusivity of uranium in cadmium was less than that predicted by the Einstein-Stokes equation. VERY little information is available on the diffusivity of uranium in liquid metals. Bonillal has reported a single measurement of the diffusivity of uranium in bismuth. Measurement of the diffusivity of uranium in cadmium was undertaken, not only to supply data in an area in which a dearth of information exists, but also because of a need for this physical property measurement in connection with mass transfer studies now underway in the U-Cd system at Argonne National Laboratory. The capillary-bath method used in this work has been successfully used by other workers2"4 to obtain diffusivities in liquid metal systems. Although considerable experimental technique is required to obtain reliable diffusivity measurements, the basic information required is only that of temperature, capillary tube length, the initial and final compositions of the alloy in the capillary tube, and the composition of the bath into which the solute element is diffusing. From these measurements, diffusivities may be calculated using Fick&apos;s law. The diffusivity equations are presented in the Appendix. Equations which are often used to predict diffusivities in dilute solutions5 are the Einstein-Stokes equation: D= (kT)/(6pµr) and the Eyring equation: D = (?,kT/?y?zp), where D= diffusivity (sq cm per sec), k = Boltzman&apos;s constant, T= temperature ("K), p = viscosity (poises), r = radius of diffusing molecule or species (cm), and A,, Ay, and A, are the diameters of the diffusing molecule in the three axial directions. For round molecules, the Eyring equation gives values which are 37r times larger than the Einstein-Stokes equation. In this work, the experimental values were compared to the Einstein-Stokes equation. Experimental values of diffusivities in liquid metal systems usually lie between those predicted by the Einstein-Stokes and Eyring equations. The viscosity values for cadmium were obtained from the Liquid Metals Handbook.6 EXPERIMENTAL APPARATUS AND PROCEDURE The general procedure employed consisted of rapidly immersing in a molten cadmium bath containing a negligible amount of uranium a capillary

Jan 1, 1963

F. E. BACHMAN, Buffalo, N. Y. (Communication to the Secretary) : The investigation so fully described by Mr. Laudig was undertaken with the idea of determining if it is possible to learn by expesiment, without an actual trial in the furnace, what economical results can be expected from the use of a given ore-mixture. The research was necessarily preliminary, and the results obtained are only to be depended upon so far as they relate to the samples treated. With the exception of the ores obtained from the stock-piles of the Buffalo Furnace Company, and several samples taken at a neighboring furnace, there is no certainty that the ores represented the average shipments of any mine. For this reason the publication of the names of the ores might do producers a great injustice. The names of the ores arc therefore withheld, and their identity is further .veiled by leaving out of the analyses the phosphorus-contents. The investigation was begun with the preconceived idea that the rate of reduction would be found to be dependent upon the accessibility of the particles of the ore to the action of the reducing-gas. In other words, a spongy, brown hematite or a very fine Mesabi would be the easiest reduced. I had formed a scale with No. 21 first, No. 30 second, No. 1 third, Nos. 5, 24, 25 and the remaining Mesabis fourth, etc , which is sufficient to show how much actual results differ from a judgment based on physical characteristics. The ore expected to be most easily reduced proved to be 25th in the list, while the one placed

Jan 1, 1897