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By L. A. Rapoport, W. J. Leas, C. W. Carpenter

A program of scaled flow model experiments has been undertaken to study the performance of five-spot water floods. The modeling procedures are discussed and the construction and operation of the flow models are described. The tests were specifically designed to investigate the effects of injection rate and of oil-to-water viscosity ratio on oil recovery. The areal sweep of five-spot water floods was also studied, and direct comparisons between five-spot and linear flooding behavior have been established to permir better utilizatfon of core analysis results for purposes of field waterflood evaluations. INTRODUCTION For a long time reservoir engineering procedures have been confined to the consideration of idealized, one-dimensional or linear systems.11* However, direct application of the linear waterflood mechanism to the evaluation of actual field performance is justifiable only for the purposes of rough approximation. To obtain a realistic analysis of field behavior, it is necessary to account for reservoir geometry and well distribution, and to take into consideration the configuration of the flow of fluids between injection and producing wells. In the past it has been attempted to account for the complex nature of the fluid movement in the reservoir by associating each well pattern with a certain "sweep-efficiency constant", e.g., a sweep efficiency of 72 per cent for the conventional five-spot pattern.' More recently, methods of waterflood evaluation have been improved by the introduction of variable sweep-efficiency factors, defining the areal coverage as a function of cumulative water injection and viscosity or "mobility" ' ratio.""' These variable sweep-efficiency factors, derived from miscible fluid displacement experiments, require the assumption of piston-like displacement of oil by water. This implies that only one fluid, either oil or water, is moving at any one point of the reservoir ht any moment. Such an assumption, however, is incompatible with the bulk of the experimental evidence and also theoretical considerations which indicate that usually in the area invaded by water, both oil and water are flowing in varying proportions, in accordance with the relative permeability characteristics of the porous medium. Another recent approach toward the evaluation of five-spot flooding behavior involves correlations of the areal sweep efficiency (observed in immiscible fluid experiments) with a "mobility ratio," based on the consideration of the "average" relative permeability to the displacing fluid. Combination of these correlations with a modification of the Buckley-Leverett frontal-drive equation leads to a computational procedure for predicting the recovery performance of areal water floods. While of considerable interest, this approach has no complete theoretical justification and, therefore, does not entirely clarify the features of five-spot water floods. Rigorous, theoretical treatment of waterflood behavior in two- or three-dimensional systems as encountered in the reservoir, including proper consideration of relative permeability and capillary-pressure functions, is as yet beyond the scope of analytical or computational methods. The, behavior of such water floods can, however, be studied by means of scaled flow model ekperiments, which is the approach followed in these investigations. The flow model experiments described in this paper simulate a uniform, horizontal formation of constant thickness with fully penetrating wells. The system under consideration is entirely liquid saturated and not influenced by compressibility effects, which is representative of the situation achieved after gas fill-up under practical conditions. Gravity segregation effects are assumed to be negligible, which implies a relatively thin reservoir formation and/or low vertical permeability. THEORETICAL CONSIDERATIONS For a flow system of uniform thickness with no appreciable effects of gravity segregation, the analysis of water-oil displacement can be confined to a two-dimensional treatment based on the consideration of a "slice" of unit thickness. Referring to previously established derivations," the displacement of oil by water in such a system can be described by the following, dimen-sionless equations:

By Albert C. Harding

For several years barite (BaSO4) production has topped such better known minerals as feldspar and fluorspar in annual tonnage and is now well established among our foremost nonmetallic industries. Description Of widespread occurrence in the United States, barite is relatively soft (hardness, 2.5 to 3.5), inert, heavy (specific gravity, 4.5), and is also called "barytes" (sometimes pronounced "bear-i-tease"), "heavy spar" and "tiff." It occurs with granular and earthy structure but usually has coarse, orthorhombic crystallization. It is commonly white, and although occurring in various shades of red, yellow, and gray, barite always gives a white streak and can be scratched by the ordinary prospecting pick. Its softness and weight are usually sufficient to establish field identification. Another aid is the "rotten egg" odor of H2S which sometimes emanates from freshly fractured surfaces. Barite can be readily distinguished from two other minerals which have somewhat similar occurrences and characteristics. The relatively rare celestite, strontium sulphate, has a specific gravity slightly less than 4.00 and gives a distinctively crimson color in a flame test. The very rare wither-ite, barium carbonate, has a specific gravity of 4.3 but readily effervesces with dilute hydrochloric acid. Occurrence All known commercially important barite deposits in the United States occur as re- placements or cavity fillings in schists, shales, limestones, dolomites, and quart-zites, or as residual deposits from such formations. In Georgia, Tennessee, North Carolina, and Missouri, high-grade barite is recovered as fragments from residual clays covering quartzites and dolomites. In Arkansas, Nevada, and California, barite, usually of lower grade, is mined from replacement deposits. These seven states contributed the entire production reported by the U. S. Bureau of Mines for 1945 as given in Table I. Table i—Crude Berile Sold or used by Producers in 1945 State Short Tons Arkansas................... 260.660 Missouri.................... 225.467 Georgia..................... 110,393 Tennessee................... 32,812 Nevada..................... 28,919 Other States............... 37,811 Total..................... 696,062 ('other states,, included only california and North Carolina. Such a large part of this undistributed production came from California that that state probably outranked both Nevada and Tennessee. For the second consecutive year Arkansas was the leading producing state, particularly worthy of note because there was no commercial production of barite there prior to 1941 and because the entire production of Arkansas has come from one large deposit which is being operated by two companies. Consumption The 1945 consumption as reported by the Bureau of Mines was: 482,442 tons as ground barite 139,288 tons in making lithopone 99,173 tons in producing barium ------- chemicals Total 720,903 tons

Jan 1, 1949

By A. U. Seybolt

As is well known, many questions affecting the properties and uses of a metal cannot be answered without careful consideration of the state of purity realized in the various operations of preparation, refining, or special treatment that have brought the product to its final state. The occurrence of an impurity in minute quantity offers no guarantee that its effect will be inconsiderable because dissolved substances often modify the properties of a metal quite out of proportion to their concentration. In such cases the determination of solid solubility must be accepted as a problem of considerable difficulty, requiring the use of especially purified materials and methods not ordinarily used in the metallurgical laboratory. This aspect of the equilibrium between metals and their oxides is especially important and a search for fundamental data reveals little except .conflicting opinion even concerning the systems of common occurrence and interest. Rhines and Mathewson1 in 1934 studied the solubility of oxygen in solid copper by annealing to equilibrium in air and gave reasons for the failure of other methods, notably the ordinary procedure of heat-treatment followed by microscopic examination, to yield reliable results. In this paper, a similar study has been made of the solubility of oxygen in cobalt. The simple theory of this method is outlined below.

Jan 1, 1935

By W. H. Brandt

THERE has been much progress in the last two decades in understanding the hardenability of steel. Roughly, the progress has been along two lines, which may be designated as empirical and fundamental. This empirical work has progressed to the point where steels may be manufactured to the desired characteristics quite closely and heat-treatment and application problems can be approached with considerable confidence. Particularly noteworthy among the empirical work is that of Grossmann,12 who has given formulas for calculating hardenability as a function of composition, and the establishment of the isothermal transformation curve by Davenport and Bain.13 The latter does not fall entirely in the class of an empirical study, for while it provides a means of presenting a tremendous amount of empirical data in the form of a single curve, it also outlines clearly the problems that face the fundamental investigator. The isothermal transformation curve shows clearly that the phenomena occurring in the neighborhood of the knee of the curve are those that limit the hardenability of steel. Above the knee of the curve, the product of decomposition of austenite is lamellar pearlite and fundamental progress has consisted chiefly of a detailed study of structure and formation of pearlite, largely at the hands of Mehl and his collaborators.1-5 They have shown that formation of pearlite proceeds by a process of nucleation and growth and they measured both the nucleation and growth rate. However, there is still a wide gap between our empirical knowledge and fundamental knowledge, owing largely to the fact that there is no satisfactory theory of nucleation rate or of growth rate. This paper is being written in an attempt to fill in a small part of the gap. A method was presented recently14 for calculating the velocity of edgewise growth of pearlite using an expression for the carbon concentration that satisfies the diffusion equation. The edgewise velocity of pearlite growth in a plain carbon eutectoid steel was calculated, using linear extrapolations of the ferrite and cementite solubilities. While the results are in fairly good agreement with the experimental facts, the calculated velocity of edgewise growth is smaller at 600°C. than G, the radial growth velocity of pearlite nodules, whereas one would expect it to be larger. In this paper the calculation will be repeated using a new extrapolation for the ferrite solubility, which gives better agreement between G and V, the calculated velocity of edgewise growth. A calculation will also be made of the effect of carbon content on velocity of edgewise growth, and the problems of calculating the effects of other alloying elements on growth rates will be discussed. It will be assumed that the mechanism of pearlite growth, established by Mehl and

Jan 1, 1945

By William C. Hagel

Oxygen-18 exchange between gaseous oxygen, held at a pressure of 125 mm Hg in a PL-IORh chamber, and splzeres of a Fe2O3 containing three or less grains was determined from 9000 to 1250°C. Isotope equilibrum on crystal surfaces appears to be quickly established, and the rate-determining factor is self-difiusion confbvrning to the relation D = 2.04 esp(- 77,900/RT) sq cin sec-'. Changing sphere diameters caused no detectable variation in diffusion coefficients. Oxygen ions diffuse faster than cations at low temperatures, but a crossover may occur at about 1150°C. INVESTIGATIONS by Davies et al.' and Himmel et al2 have suggested that the mechanism of iron oxidation above 600°C involves a predominance of cation migration in the wüstite (FeO) and magnetite (Fe301) layers and oxygen-ion migration in the hematite (a Fe2O3) layer. Evidence for little cation migration during hematite growth consists of: 1) the disappearance of radioactive-silver markers, presumed to vaporize from the oxide-gas interface, and 2) a large discrepancy between parabolic rate constants for magnetite oxidation and those calculated from Lindner's3 data for cation self-diffusion, D:, using Wagner's theory.4 Reversing the calculations gave anticipated self-diffusion coefficients, D:, for oxygen ions of 3.6 x 10-9 sq cm sec-1 at 1100°C and 5.9 x 10-10 sq cm sec-1 at 1000°C. Since then Kingery et a1.5 reported three values for oxygen-ion diffusion in polycrystalline, sintered hematite. These were obtained near 1200°C, at an oxygen pressure of approximately 150 mm Hg, and could be represented by the Arrhenius expression, Dg = 1 x 10 11 exp(-146,000/RT) sq cm sec-1. Extrapolation to temperatures below 1100°C infers that oxygen ions are less mobile than cations. This contradicts the aforementioned oxidation evidence and also coble7s study of the initial sintering of hematite. From the latter, apparent diffusion coefficients for the slowest rate-controlling species could be calculated, and fair agreement was found with an extension of Lindner's data for cation diffusion. Fig. 1 depicts such comparisons graphically. Three points have been included for the determinations of Himmel et al. on cation diffusion in large, natural hematite crystals surrounded by pure oxygen at atmospheric pressure. No dependence on crystal log raphic direction was noted. Since the value of 10-l4 sq cm sec-' at 1000°C actually represents an upper limit, one might interpret that there is a small decrease in D: when single crystals are employed. On the whole, however, cation diffusion appears relatively well-defined,* while the few results tension to lower temperatures. After having gained experience measuring oxygen-ion diffusion in certain glasses7 and in the corundum oxide, a Cr2O3, the purpose of this work was to determine D; values for hematite and thus possibly assist a mechanistic understanding of the kinetics of the oxidation and reduction of solid iron oxides. I) EXPERIMENTAL PROCEDURE A) Material. Hematite powder of moderately high purity was received from the Columbian Carbon Co. An accompanying lot analysis listed 0.03 H2O, 0.008 weight percents, with the remainder Fe2O3. Subsequent treatments were ball milling, drying, prepressing, and hot-pressing for 60 min at 1350°C under 5000 psi. Since graphite dies coated with a slurry of alumina were used for hot-pressing, the surrounding atmosphere at temperature was sufficiently reducing to form some magnetite on the outer surface of each cylinder. After removing any attached alumina, a stabilizing anneal of 50 hr at 1300°C in oxygen was added. The final product was nonmagnetic and exceeded 97 pct of the density (5.24 g cm-3) usually given for FezO3. On polishing cylindrical ends, counts were made of the number of grain boundary intercepts per

Jan 1, 1967

By Fred C. Bond

Many grindability tests have been made in the Allis-Chalmers laboratory since the tabulated results were last published.' The lists of standard ball-mill and standard rod-mill grindability tests are brought up to date in this paper, and considerable new material is added. This includes a tabulation of open-circuit grinding tests, with impact and compression-crushing tests. The standard ball-mill grindability tests (Table I) were made in a standard laboratory ball mill, 12 by 12 in. inside, running at 70 rpm, with a charge of 285 iron balls ranging from 1½ to 3/4 in. in diameter and weighing 20,125 grams. Seven hundred cubic centimeters of minus 6-mesh stage-crushed dry feed was used, and the circulating load was maintained constant at 250 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 28 to 200, and the grindability is the number of net grams of screen undersize produced per revolution. The table includes 479 tests. It has been calculated that the ball mill does 52.3 joules of useful work on the ore per revolution2 with 93 joules input to the mill. The standard rod-mill grindability tests (Table 2) were made in a laboratory tilting rod mill, 12 in. in diameter by 24 in. long inside, with a wave-type lining, running at 46 rpm. The grinding charge consisted of six I1/4-in. diameter and two 13/4-in. diameter steel rods, all 21 in. long and weighing 33,380 grams. The mill was rotated level for eight revolutions, then tilted up 5' for one revolution, down 5° for one revolution, and returned to level for eight revolutions. It was discharged by tilting downward at 45° for 30 revolutions. Feed consisted of 1250 cc of minus in. stone and the circulating load was maintained constant at 100 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 3 to 65, and the rod-mill grindability is the number of net grams of screen undersize produced per revolution. The table includes 105 tests. The impact and compression-crushing tests (Table 3) are given in a table that is an extension of a previous publication8 in which the method of testing was described in detail. Two equal hammers weighing about 30 lb each were arranged to strike simultaneous blows on opposite sides of the test specimen. Broken pieces of rock passing a 3-in. square and retained on a 2-in. square were set up with the smallest dimension between the hammers, and this distance was measured. The height of the hammers was increased slightly before each blow until the specimen broke, and the foot-pounds per inch of thickness was calculated. Ten pieces

Jan 1, 1949

By Fred C. Bond

MANY grindability tests have been made in the Allis-Chalmers laboratory since the tabulated results were last published.1 The lists of standard ball-mill and standard rod-mill grindability tests are brought up to date in this paper, and considerable new material is added. This includes a tabulation of open-circuit grinding tests, with impact and compression-crushing tests. The standard ball-mill grindability tests (Table I) were made in a standard laboratory ball mill, 12 by 12 in. inside, running at 70 rpm, with a charge of 285 iron balls ranging from 1 1/2 to 3/4 in. in diameter and weighing 20,125 grams. Seven hundred cubic centimeters of minus 6-mesh stage-crushed dry feed was used, and the circulating load was maintained constant at 250 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 28 to 200, and the grindability is the number of net grams of screen undersize produced per revolution. The table includes 479 tests. It has been calculated that the ball mill does 52.3 joules of useful work on the ore per revolution2 with 93 joules input to the mill. The standard rod-mill grindability tests (Table 2) were made in a laboratory tilting rod mill, 12 in. in diameter by 24 in. long inside, with a wave-type lining, running at 46 rpm. The grinding charge consisted of six 1 1/4-in. diameter and two 1 3/4-in. diameter steel rods, all 21 in. long and weighing 33,380 grams. The mill was rotated level for eight revolutions, then tilted up 5° for one revolution, down 5° for one revolution, and returned to level for eight revolutions. It was discharged by tilting downward at 45° for 30 revolutions. Feed consisted of 1250 cc of minus 1/2-in. stone and the circulating load was maintained constant at 100 pct by adjusting the number of mill revolutions for each grinding period. Tests were made at all mesh sizes from 3 to 65, and the rod-mill grindability is the number of net grams of screen undersize produced per revolution. The table includes 105 tests. The impact and compression-crushing tests (Table 3) are given in a table that is an extension of a previous publication3 in which the method of testing was described in detail. Two equal hammers weighing about 30 lb each were arranged to strike simultaneous blows on opposite sides of the test specimen. Broken pieces of rock passing a 3-in. square and retained on a 2-in. square were set up with the smallest dimension between the hammers, and this distance was measured. The height of the hammers was increased slightly before each blow until the specimen broke, and the foot-pounds per inch of thickness was calculated. Ten pieces

Jan 1, 1947

By E. S. Machlin

The wzodel of Blandin and Diplunt;, generalized to include a phase factor, is applied to the carbon-carbon interaction in iron. Darken&apos;s "energetic" model is generalized to include not only first neighbor interactions but further neighbor interactions as well. On the bases of these generalized models relations are derived for the activity of carbon in both austenite and ferrite in terms of the carbon-carbon Pair interaction energies. A single function then yields the pair interaction energies consistent with the experimental activities of carbon in both ferrite and austenite. Thus, a simple explanation is given for the observation that the nearest-neighbor interaction between carbon is repulsive in austenite and attractive in ferrite. Certain consequences of this approach are explored. OnE object of the present paper is to attempt to take into account the consequences of electrostatic contributions to the carbon-carbon pair interaction energy for carbon as a solute in iron. Friedel&apos; has shown that oscillations in electrostatic potential are to be expected about a solute atom in a metallic solution. Blandin and 6lant6&apos; have shown that such oscillations yield an interaction energy between pairs of solute atoms that obeys the relation: W{ = A cos(2ftFri + 4>)/(kFri)3 [l] where kF = Fermi wave vector, ri = distance between solute atoms comprising the pair7 <p = phase factor dependent only on electronic nature of solute and solvent, A = coefficient dependent only on electronic nature of solute and solvent. Machlin3 found that Eq. [I] accurately described the pair interaction energy derived from short-range order measurements based on field ion microscope observations of dilute alloys of platinum. He also found that the value of the phase factor $ derived from residual resistivity measurements agreed well with that obtained from the analysis of the short-range order data. Harrison and paskin4 were able to predict the long-range ordering energy of 0 brass using Relation [I] and residual resistivity values to predict the value of the phase factor $. Machlin5 has repeated their analysis and applied it to the prediction of the long-range ordering energy in AgZn and AgCd with excellent agreement between prediction and experiment. Both A and $ are independent of the crystal structure. The Fermi wave vector depends uniquely upon the conduction electron concentration per unit volume in the spherical approximation of the Fermi surface. Thus, Eq. [I] is expected to apply to both fer- rite and austenite with only one set of values of A and $. Mossbauer studies6 yield the result that iron has one 4s electron. We shall make an assumption found to hold previously for platinum3&apos;7 and nikel, which is that only the s electrons are involved in shielding the perturbing potential of carbon. With this assumption, kF = 1.35 A-l. Although A and $ may be obtained from certain mdels&apos;&apos;&apos; we shall take A and $ to be empirical constants in the spirit of Kohn and osko.&apos; Thus, Eq. [I] involves two adjustible parameters. Consequently, two independent relations in A and $I are required in order to evaluate them for carbon as a solute in iron. We may use a recent analysis of Aaronson, Domain, and poundg who showed that Darken&apos;s energetic model,1° as well as others, can be used to describe the activity-temperature data for carbon in iron in both the aus-tenitic and ferritic phases. Darken&apos;s model takes into account only first neighbor pair interactions. For our needs, all neighbor pairs need to be taken into account. It is convenient to generalize Darken&apos;s model. The result for the partition function for austenite is: over the temperature range 800" to 1200°C and where the uncertainty corresponds to one standard deviation. Eq. [4] effectively yields only one relation. Another relation is required to obtain unique values for A and $. One property of Eq. [4] is that it is independent of crystal structure. Hence, data for a iron can be used to obtain another relation. To arrive at this relation we must generalize Eqs. [2] and [3] so that they may be applied to the bcc a iron. The result is that:

Jan 1, 1969

By Roy L. Zuber

Identification and analysis of mining hazards involves the correlation of accident, injury and illness information from all segments of the mining industry. It would not be feasible to process the volume of data involved without substantial reliance upon a computerized information system similar to that maintained by the Health and Safety Analysis Center. Numerous statistical programs are applied to the computerized accident-injury-illness data base to identify those mining tasks and associated activities which are responsible for most of the injuries and occupational illnesses. These programs are used to flag those mining tasks or conditions at specific mines or segments of the mining industry where either a significant increase in the injury/illness trend has been reported or where the number of injuries/illnesses related to specific mining tasks or conditions are high. Upon identification of those mining hazards which warrant further analysis, a computerized search of the coded accident data is performed to correlate the predominant categories by cause of injury or illness, source of injury, nature of injury, type of accident, etc. From this correlation, those factors most common to the conditions responsible for the hazard may be identified. Based upon analyses of this form, recommendations are made regarding the appropriate engineering, training, and/or operational improvements necessary to alleviate the hazard. In addition to mining hazard analysis, the computerized data base also has application in the area of optimum resource allocation. Computer programs have been developed to assist management in allocating resources based upon organizational requirements and historical accident - injury trends.

Jan 1, 1983

By E. H. Phillips, A. F. Keenan

6.2-1. Historical Development. Hammering on hand-held drill steel was the earliest type of percussion drilling. It was not until 1838 that Singer developed a steam-operated drilling machine that lifted a drilling bar and let it drop by gravity. J. J. Couch of Philadelphia patented the first percussion drill in 1849. His machine was a steam drill which literally hurled the drill bar at the rock and retracted it by means of a ratcheted wheel. Later J. W. Fowle added automatic rotation of the steel. In 1871, Simon Ingersoll patented a rock drill mounted on a tripod which permitted drilling holes at any angle, vertical to horizontal. Up to the turn of the century, percussion drills in use were piston-type drills. The drill steel was solid and attached to an extension of the drill piston. Hole cleaning was accomplished by the pumping action of the steel in the hole. Down-holes were drilled wet and cleaned with a blow pipe while up-holes were drilled dry, allowing the cuttings to drop out by gravity. The forerunner of our present hammer drill was the free-piston drill which was first patented in 1884, by Sargent. Because solid drill steel was used, it did not successfully clean down holes. J. G. Leyner solved the problem of applying the hammer drill principle to down-hole and horizontal-hole work in 1897. The key was to pass com- pressed air and water through hollow drill steel to clean the hole. Leyner&apos;s work on valves and rotation mechanisms increased the blows per minute of hammer drills from 300 or 400 up to 1800 and more blows per minute and made possible the first lightweight drilling machines. The development of automatic feeds followed. 6.2-2. Percussion Drills. TYPES OF DRILLS. There are two types of percussion drills in use today; the piston drill and the hammer drill. In the

Jan 1, 1968

By Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (I), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profit- ability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus mini- mum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12 ) Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-low- grade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1985

By W. G. Jewitt

YELLOWKNIFE, the most northerly Canadian gold mining district, is once more in the throes of a boom. Touched off by spectacular and well-publicized diamond-drilling results on the property of Giant Yellowknife Gold Mines, Ltd., mining companies and the general public are competing for mining claims and for the shares of companies owning mining claims in the area, with little regard for values. Such a stock-market boom has various unfortunate consequences but undoubtedly one beneficial result is that considerable amounts of money are made available for serious development. The Yellowknife mining area lies on the north shore of Great Slave Lake, in the Northwest Territories, at a latitude of approximately 62°30&apos; north. The topography is a low-relief phase of that usual in the Canadian Laurentian shield. Roughly, one third of the surface is lake-covered, one third muskeg or overburden, and one third is bare

Jan 1, 1944

By R. J. Wygal

A method is presented by which homogeneous, isotropic, and reproducible aggregates of particles such as glass beads, sand, ground glass, etc., may be prepared simply and quickly by using a "particle distributor". Several applications of the packing method to the construction of model reservoirs are described. The flow characteristics of these packings me given. Some factors which limit the usefulness of unconsolidated packs for simulating reservoir rock are overcome through artificial cementation of the aggregates. INTRODUCTION Many laboratory studies of porous media are based on experiments using test packs of sand, glass beads and other granular materials. More confidence can be had in the results of such studies if the test packs are known to have uniform properties throughout, to be mechanically stable and to be accurately reproducible. This paper presents a method by which such packs of a variety of shapes and sizes may be constructed using a particle distributor. Some factors which limit the usefulness of unconsolidated packs as a means for simulating reservoir rock are described. A method of overcoming these limitations, which involves artificial cementation of the packs, is presented. DEVELOPMENT Particles packed by the tamping or vibrating techniques are usually either non-uniform or mechanically unstable. An invention by Currie and Gregoryl overcomes these difficulties to some extent. The invention is based on the use of a stream of falling particles impinging on and passing through a distributor matrix of several layers of closely spaced spheres. The particles are well dispersed by numerous collisions with the spheres. They become homogeneously distributed at some distance below the distributor matrix and at that position are collected in the sample container. This apparatus, however, was found to be ineffective when small particles were used since small particles accumulate on the top surfaces of the spheres. An improved distributor matrix was made, first in the form of a stack of five sieves, and then in the more adaptable form of spaced layers of wire screen. Fig. 1 shows a longitudinal view through a five-layered distributor of eight-mesh wire screen. This device gives excellent results.* Fine particles pass through the distributor with very little hold up. The fine structure of the wire matrix with its many deflecting surfaces ensures good particle distribution. There are no dead areas immediately below the screen as were sometimes observed to occur below a matrix of spheres. Using such a device, 20-30 mesh sands were packed to a porosity of 32.1 per cent at rates varying from 0.17 to 6 lb/rnin. The lowest porosity achieved by tamping small increments of similar sands into a container was 33 per cent. Motion pictures at 3,000 frarnes/sec of the surface of a growing pack of 2-3/4 mm glass beads have helped to show the mechanism by which packs of low porosity are obtained. The beads, although packed at the rate of 5 lb/min, were observed to hit the pack surface singly, for the most part, and at well spaced points of impact. Part of their energy is transferred to surface beads which are thereby

By F. D. Weeks

The Salida smelting plant, owned by the Ohio & Colorado Smelting & Refining Co., is situated at an altitude of 7,000 ft., about 2 miles west of Salida, Colo., and 215 miles southwest of Denver. Salida is a division point on the Denver & Rio Grande R. R. where the narrow-gage line serving the San Juan region joins the through standard gage. This narrow-gage road also serves the coal and coke companies at Crested Butte and the lime quarries and mines at Monarch and Garfield, which are only 20 miles distant. These facts, together with the possibility of an appropriate site, doubtless influenced the builders to locate at this point. The smelting plant and yard occupy 80 acres of land on the northeast side of the Arkansas River. The slope of the ground is such as to permit of the ore being delivered by cars at a point 25 ft. above the feed floor; and in front of the furnaces there is a fall of 50 ft. for the slag dump. A recent purchase of river bottom has provided slag room for 20 years. As originally constructed, the plant consisted of four blast furnaces, 40 by 144 in. at tuyéres, two copper-matting furnaces, dust chamber, stack, power plant, blowers, sampling works, and one railroad trestle about 16 ft. from bed floor to rails. Later 12 hand roasters of excellent design were added and the practice of smelting leady matte in a copper-matting furnace was abandoned. Since the plant passed under the present management the equipment has been increased so that the following is a brief description of the same at this date: Two lead blast furnaces, 48 by 180 in. at tuyéres, 17 ft. 6 in. tuyéres to feed floor; four lead blast furnaces, 40 by 144 in. at tuyéres, 17 ft. 6 in. tuyéree to feed floor; two copper-matting furnaces, not used; blast-furnace flue, 900 ft.; area 245 sq. ft.; blast-furnace stack, 150 ft.; above furnace floor, 190 ft. When smelting a 500-ton charge per day, the gases approximate 63,700 cu. ft. per minute and travel at a speed of 260 ft. per minute. The slag is hauled by a 10-ton electric locomotive; the slag shells and settler refuse are elevated to a bin on the feed floor by an electric hoist. The matte is tapped in 1,600-lb. cakes and loaded into cars

Jan 1, 1916

By Frank H. Kingdon

and eventually we would have trouble. That is, these little localized overstresses may not be important if the structure is loaded a few times, but may be important if the structure is loaded many thousands of times. The existence of localized stress has been demonstrated by direct experimentation with delicate extensometers by the elaborate investigations of Coker. Localized stresses four, five, and six times those that we would get by the ordinary laws of mechanics exist in most of our structures. These stresses, ordinarily, under steady load do no damage but if they are repeated many times may start progressive failure. So far, we have seen no failure that could not be explained by localized high stress at a nick, a crack, a point where one heat treatment stopped and another began, or where there was local faulting, or any one of a dozen causes. In the case of drill rods, the actual distribution of stress must be different from that which any mathematical analysis could predict. It would seem that the spread of damage from the localized overstresses, so far as we can see now, explains the phenomenon known as fatigue. The Ideal Drill Steel By Frank H. Kingdon The essential qualities of a drill steel are: (I), It must be easily forged; (2), the forged bit end must permit easy heat treatment to obtain hardness to resist chipping; (3)) the bar or body must be stiff to resist bending or twisting and yet tough to resist shock and vibration, with resulting breakage, (4), the forged shank end must be easily heat treated to obtain hardness with great toughness. Minor requirements are that the bit and the shank are in alignment with the body; that the shank is of the proper shape and length and the shank collar or lugs of the proper diameter and length; that the hole is free from obstruction; that the striking end of the shank is flat and square; that the bit is of the proper shape, with the cutting and reaming edges formed full and to the required size; that the reaming edges are concentric with the axis of the steel, and that there are no sharp corners at the shoulder where the bit blends into the body. The essential qualities of a bit, are: That its shape be such that maximum cutting speed can be maintained for as great a distance as possible before wear of the gage and cutting edge reduces the speed of penetration to a point where a change of steel is made necessary; that the size or diameter of the drill hole corresponding to the gage of the bit can be maintained with the least possible reduction as the depth of the hole increases, and that the bit can be correctly and readily formed and heat treated. The following features of bit design require attention: Shape,

Jan 1, 1922

The U. S. Navy Department has perfected plans for the enrollment and training of considerable numbers of engineering officers. A school for this purpose, the U. S. Navy Steam Engineering School, has been established by the Department at Hoboken, N. J. The school is open to men between the ages of 21 and 40 who meet the physical requirements of the Navy, who are of thorough ability and officer-like character and who have completed the mechanical or electrical engineering course at certain recognized technical schools, or who possess an education and experience adjudged to be an equivalent thereof. Enrollment in the U. Naval Reserve Force of men properly qualified may be made at any Naval enrolling office, notation being made of the applicant&apos;s qualifications and desire to be detailed to this school. Applicants will be enrolled as Chief Machinist Mates, and during the course of instruction will draw the pay of this rating, $83 per month, plus $60 per month paid as subsistence. Upon graduation, men will be commissioned as Ensigns in the U. S. Naval Reserve, with a salary of $1700 per year. The duty to which a graduate of this school will be assigned will be that of an engineer officer in the auxiliary service of the Navy. Special provision has been made for the continuance of the school with proper material by a Navy regulation which permits undergraduates of the Freshman, Sophomore, and Junior Classes in recognized engineering schools to enroll in the Naval Reserve Force with the rating of Seaman 2d Class, and continue their courses at the institutions where they have matriculated. Such men will be called into active service after their graduation, and can at that time, if they are qualified to pass an officers&apos; physical examination, apply for admission to the U. S. Navy Steam Engineering School. Men who are registered in the draft, either graduates or undergraduates, may enroll with the proper enrolling officer by securing from their draft board a letter of release, which in all probability can be obtained for this purpose. The aim of this school is not simply to produce engineers, but to produce naval engineering officers in the true sense of the word. The course of instruction covers a period of five months, arranged in four periods.

Jan 10, 1918

By Jose A. Ruiz

INTRODUCTION For heterogeneous non-catalytic solid-gas reactions carried out in fluidized bed reactors, nominal bed temperature is an average value between gas and solid. For highly exothermic reactions, actual particle temperature can be up to hundred degrees above termocouple temperature, since heat dissipate at slower rate than heat generation during some steps of transformation. These micro instabilities at particle level can displace thermodynamic equilibrium from the desired phase to others. This effect is important when relatively narrow range of phase stability exists, like in sulphation roasting of sulphides, where bed discharge contains, in addition to the desired sulphate others metastable phases such as basic sulphate and ferrite. For example, for copper-iron sulphide concentration roasted below 680°C, no cupric ferrite is formed, as shown by Jakob et al., therefore the separate Kellogg&apos;s stability phase diagrams Cu-S-O represent the roasting (Fig.l). At temperature above this one, like 760°C for example, phase stability changes from CuSO4 to the basic sulphate Cu0•CuSO4. Solid state reaction between Cu0•CuSO4 and Fe203 also occurs to form cupric ferrite, Cu0•Fe2O3, therefore the Cu-Fe-S-O Kellogg&apos;s phase diagram of Fig. 1 (continuous line) represents this condition of roasting. Although both Cu0•CuSO4 and Cu0•Fe2O3 are metastable phases under normal operating conditions of sulphation roasting(1-10% SO2, 1-10% 02 and 680°C, points A and B,Fig. 1) sulphation of these species and particularly ferrite are slow and limited due to diffusional control. This explain why reactors discharge contains several phases: CuSO4, Cu0•CuSO4, CuO•Fe203 plus small amounts of CuS and FeS due to the baclonixed conditions of fluidized bed reactors. The presence of cuprous ferrite due to reducing conditions inside individual particles also occurs, but in very limited extent at the reaction interface, as it is demonstrated by the low soluble iron levels of calcines. On the other hand, the presence of cupric ferrites can be explained only due to the higher temperature achieved by particles, since below 680°C cupric ferrite formation does not take place. [ ] This phenomena also occurs for others non-ferrous metal sulphides concentrates roasted under sulphation conditions, such as 7n, Co and Ni. Therefore the prediction of discharge composition of fluidized beds is not possible based only on kinetics and mechanistics models alone or purely thermodynamical considerations, since all factors should be integrated to describe the overall phenomena. SINGLE PARTICLE TRANSFORMATION Assuming a first order reaction with respect to the gas and zero for solid, such as

Jan 1, 1984

By G. L. F. Powell, L. M. Hogan

Samples of silver and Ag-O alloy, 0.12 wl pet, have been undercooled to a maximum of 250°C by melting in a slag of commercial soda-lime glass. Grain refinement occurred in undercooled silver samples whet the degree of undercooling exceeded a critical value which varied from 153° to 175°C, depending on the melting conditions. In under cooled Ag-0 samples, the grain structure was fine and equiaxed at degrees of undercooling larger than 50oC. The grain refinement in both silver and Ag-0 alloy samples was the result of recrystallization during or immediately following freezing. When the oxygen content was increased, the degree of undercooling at which recrys-tallization occurred decreased. INVESTIGATIONS of the grain structure of strongly undercooled melts of nickel and cobalt1-4 have revealed a change in room-temperature grain size as a function of undercooling. At large degrees of undercooling the grain size was fine, but samples undercooled by smaller amounts exhibited a much coarser grain structure. The reported undercooling at which this grain size transition occurred in nickel varied between 140° and 175&apos;C undercooling. Walker1 attributed this grain size change to an enhanced nuclea-tion rate. Cavitation at a melt-solid interface was con sidered to generate a high-pressure wave. This, in turn, produced nucleation in the liquid as a result of an elevated freezing point increasing the undercooling in the melt. A similar grain size transition was observed in undercooled silver between 133° and 153°C undercooling.5 The grain size transition in undercooled silver was found to be dependent on oxygen con tent. More recently, the authors5 showed that a simila transition occurred in an undercooled Cu-O alloy, =0.08 pet, but not in oxygen-free copper undercooled more than 200°C. The transition was attributed to recrystallization of the solidified grains after freezing; the greater the degree of undercooling, the more complete the recrystallization. In view of this effect of oxygen content on grain size transition, additional undercooling experiments were carried out with silver. This paper reports the results of these studies on silver and Ag-O alloys undercooled more than 50°C EXPERIMENTAL 1) Materials. The silver was supplied by Matthey Garrett Pty., Ltd., as fine granulate, silver mini&apos; mum 99.9 pet, and contained the trace impurities shown in Table I. The silver originated from two different refineries resulting in the difference in impurity content shown in Table I. However, the variation in composition had no discernible effect on the undercooling behavior. Samples of the silver were melted in an open-ended, vertical, cylindrical resistance furnace, wound with Kantha] wire, the temperature of which was adjusted by means of a variable transformer, 2) Melting and Undercooling. Bulk samples of silver were undercooled by preoxidation and removal of oxidized impurities in a glass slag, as previously described,5 but special precautions were necessary to allow comparison between oxygen-free silver and a Ag-O alloy. Samples free of oxygen were prepared by first melting weights of granulate varying from 200 to 400 g in clean vitreous silica crucibles in air for approximately 15 min. A sample was frozen and a complete glass cover formed over the top surface by adding granulated glass rod. The sample was then melted and frozen several times until no oxygen bubbles appeared in the glass slag. Subsequent thermal analysis indicated that the oxygen content was negligible. The glass acted as an efficient barrier to absorption of oxygen from the air. Ag-O alloys were prepared by melting silver in silica crucibles with only a partial cover of glass, so that the top surface of the melt was in contact with air. The silver was held molten in air for several hours at 1000° to 1050°C, in the expectation that the oxygen in solution would reach equilibrium with atmospheric oxygen. At T - 1000°C the equilibrium oxygen content was calculated as 0.12 wt pet O, using Sieverts law and Eq, [l], derived by Mizikar et a!.7 from the data of Sieverts and Hagenacker8 for 1 atm of oxygen:

Jan 1, 1970

By William G. Christy

EFFORTS at smoke abatement date back to the year 1273 in England when a law was passed prohibiting the use of "sea cole." The law was not enforced, so King Edward I, 33 years later, appointed a commission "to enquire of all such who burnt sea cole in the city or parts adjoining, and to punish them for first offense with great fines and ransomes, and for the second offense to demolish their furnaces." Many other commissions have since been named to investigate the smoke problem in England, but it still persists in many British cities. In the United States the movement to abate the smoke nuisance started in certain Middle Western cities, and has made considerable progress. Until recently, the coal-producing industry co-operated little in smoke¬abatement programs, taking the attitude that smoke elimination is the dealer&apos;s or user&apos;s responsibility; and the dealers, in turn placed the responsibility entirely on the user.

Jan 1, 1942

By R. E. Cech

Particles -3 to 30 µ diam of a 29.5 pet Ni, 70.5 pet Fe alloy after being melted and solidified while falling through a hydrogen atmosphere were found to contain a distortion-free body-centered-cubic phase. In most particles, the amount of body-centered-cubic phase approaches or equals 100 pet. It was shown by a series of experiments that this phase forms directly from the liquid. It is pointed out that a mechanism previously suggested in recent papers on the theory of the liquid-solid transformation can account for the solidification of body-centered-cubic phase in a temperature range where face-centered-cubic is the stable phase. IN the course of an operation to produce powders for a study of the martensitic transformation in Fe-Ni alloys, an unexpected phenomenon was observed. Powders of a 29.5 pet Ni alloy, which has a subzero M, temperature, were found to contain considerable amounts of a body-centered-cubic phase at room temperature. The experiments to be described were designed to elucidate the mode of formation of this unexpected phase. Experiments and Discussion An alloy containing 29.5 wt pet Ni, balance iron, was vacuum melted from high purity electrolytic iron and carbonyl nickel shot and solidified in the crucible. The metal was converted to oxide, ground to 100 pet <37µ, pressed to a cake, and homogenized 20 hr at 1350°C. To insure homogeneity the grinding, pressing, and homogenizing heat treatment was repeated a second time. The oxide was then ground to 100 pct <37µ and hydrogen reduced at 500°C. The reduced powder was given a spheroidizing treatment to produce spherical cast particles. The treatment consisted of dispersing the particles in hydrogen and allowing them to drop through a vertical hydrogen furnace set to a temperature higher than the liquidus of the alloy. This operation, commonly known as shotting, has been described more fully elsewhere.&apos; The powder produced by this method was found to contain approximately equal parts of body-centered-cubic and face-centered-cubic phases. A magnetic separation technique, previously described,&apos; was applied, and it was found that particles were either completely face-centered-cubic or nearly completely body-centered-cubic. The microstruc-

Jan 1, 1957