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By D. W. Fuerstenau

The United States is rapidly approaching an acute shortage of minerals. According to some observers, this situation may even be more severe than the energy crisis because the US consumes nearly 30% of the world's mineral production. With the mounting competition for world mineral supplies, we can no longer take for granted that foreign mineral products will be available to the US as abundantly as in the past. In 1973, US consumption of mineral raw materials was $35 billion. Of this, $7 billion consisted of imports. Without a drastic reorientation towards the mining of domestic mineral resources, the 1973 $7 billion deficit might mushroom to $100 billion by the year 2000. Thus, a return to domestic sources of raw materials will become absolutely necessary-and will be bucking the trend of the past 25 years, characterized by capital investment for growth in multinational overseas mining operations.

Jan 11, 1975

By Fred A. Anderson, George B. Wallace, Edward J. Slebir

As the reservoir rises behind an arch dam, it presses the arch into the canyon walls and valley floor. To compute the stresses, deflections, and arch reactions, it is necessary to know how much the rock deforms under the combined loads of gravity, temperature, and water. Under sustained loading for a period of time, the rock continues to deform or "creep" slightly, causing changes in the stresses and shape of the arch. Under reduced loading, the elastic arch tends to resume its unstrained position and it is important to know that the rock will follow it and maintain tight contact along the abutment. It is, therefore, important to know the "set" characteristic of the rock. Past attempts to define load-deformation properties of rock, including creep and set, have largely l been limited to laboratory test of intact cores and in-situ jacking test utilizing bearing plates 2 to 3 ft in diam. In both cases, a uniaxial load has been applied to a volume of rock too small to fully reflect the influence of cracks, joint systems, foliation, and other inherent weaknesses within the rock mass. This chapter describes a large radial jacking test which more nearly duplicates the action of an arch dam by penetrating the load deep into the rock mass, bringing into play more of the factors which contribute to prototype deformation. Results obtained from the large radial test are compared with those from smaller uniaxial test. DESCRIPTION OF RADIAL TEST The radial jacking test described in this chapter was performed in an adit off of an exploratory tunnel driven into a proposed dam abutment

Jan 1, 1972

By Lloyd Jones

THE strip industry made rapid strides in regard to both width and gage until about 1922, when the maximum width was about 20 in. In the hot mills, strips of thin gages in wide widths could be pro-duced, but only at a sacrifice of length of coil, con-siderable increase in power, and decreased tonnage out-put on the min. Hot mills, like the Bray sheet mill and the modern strip and skelp mills, demonstrated that . wider strip could readily be rolled, if such a product was desired. The strip manufacturers were really limited in width, not by the hot strip, but by the cold rolling of wide strip. This factor was roll pressure. The diameter of rolls increased as the width of strip increased. For cold rolling strip 20 in. wide, cold rolls were approximately 20 in. diameter. As the rolls became larger, the arc of contact between the rolls and the strip increased approximately, according to the formula, b=\/Rd where b = the width of contact R = roll radius d = total reduction Now the crushing strength of cold steel of low carbon varies from 70,000 to 100,000 lb. per sq. in. With each succeeding pass the crushing strength increases and may approach 250,000 lb., according to the total reduction. The total reduction in cold rolling of low-carbon steel will range from 40 to 70 per cent of the original thickness, usually in about three or four passes. On this basis, assuming a reduction of 20 per cent in one pass, the crushing strength of steel assumed at 120,000 lb., we have a total pressure between rolls, which (assuming that the rolls do not deform) equals: Roll Strip Strip Diameter, Width, Thickness, In. In. In. Pounds 10 10 0.100 380,000 20 20 0.100 1,080,000 30 36 0.100 2,375,000 42 50 0.100 3,900,000 These figures are approximate, but you will note that in increasing the width five times, or 50 in., the total pressure between the rolls increased approximately ten times. In other words, the pressure per inch of width increased from 38,000 to 78,000 lb. for the same per-centage of reduction.

Jan 3, 1928

By Ross C. Anderson, William I. Weisman

THE manufacture of anhydrous sodium sulphate or salt cake from natural deposits in the United States has been in general somewhat of a marginal undertaking. Competition from foreign sources and from large quantities of byproduct sodium sulphate produced domestically in the manufacture of hydrochloric acid and other chemicals has existed and continues. For example, most of the sodium sulphate produced is a byproduct or co-product in the manufacture of hydrochloric acid through the reaction of sodium chloride with sulphuric acid. In recent years, many manufacturers of rayon have installed equipment to recover sodium sulphate from waste spin bath liquors; today this is an important source. Before World War II large quantities of sodium sulphate were imported from Germany. In 1949 imported material from Europe again appeared on the domestic market. Natural sodium sulphate from Canada in substantial quantities also enters the United States markets. Despite this kind of competition, numerous attempts have been made to exploit various natural deposits of sodium sulphate in this country, but only a very few of these have survived economically over a period of years. One of these few operations is the plant of the Ozark-Mahoning Co. located 13 miles south of Monahans in West Texas. Several factors contributing to the successful life of this plant may be summarized as follows: 1—Geographical location. Monahans is reasonably close, freightwise, to the Kraft paper mills in Texas, Arkansas, and Louisiana; the Kraft paper industry is the greatest consumer of sodium sulphate in the United States. 2—Availability of natural gas as low cost fuel. Proximity of the natural gas fields of West Texas has been a tremendous asset, as the availability of low-cost natural gas is to all industry throughout the Southwest. 3—The nature of the deposit. The occurrence of sodium sulphate brines in southeastern New Mexico and West Texas has been very well described by Lang,' who writes that the brines are found in the Castile formation of the Delaware basin. Here weathering has altered the anhydrite so that a relatively porous gypsiferous zone overlies a dense impervious mass of anhydrite. This porous zone provides traps where percolating ground waters that have picked up soluble salts may lodge. These traps or pockets are the natural brine reservoirs exploited at Monahans. Although several hundred wells have been drilled, currently some 25 wells serve to supply brine to the plant. All are within 1 1/2 miles of the plant and are conveniently tied together by an electric power system serving electric motors driving the pumps. Having the raw material in the form of a brine which can be pumped from shallow wells makes possible much simpler and more efficient handling than if it were in form of solids. By contrast, other deposits of sodium sulphate, such as those in Arizona, Nevada, and North Dakota, are in the form of the solid minerals, thenardite and mirabilite, which present somewhat more of a mining and mineral dressing problem.' The largest producer of sodium sulphate from natural sources in the United States is at Searles Lake, Cal., and there a brine also is utilized. 4—Water. Substantial quantities are needed for cooling towers and for operation of gas engines. An area underlain with brine is not a promising source of fresh water, but fortunately, after a long search, an adequate supply was found nearly two miles from the plant. It may be appropriate to discuss briefly the grades of sodium sulphate offered on the market. Salt cake is the name usually applied to the grade of sodium sulphate used by the Kraft paper industry. It may be a low analysis byproduct, 95 to 97 pct sodium sulphate, with as much as l 1/2 to 2 pct residual acid, or it may be a natural product. Usually salt cake is considered a low grade product, but a great deal of a higher grade of material is marketed under this name. The specifications for glassmakers' salt cake are somewhat higher than those of the paper industry, usually requiring 98 pct sodium sulphate. Technical anhydrous sodium sulphate is a high grade material and usually exceeds 99 pct sodium sulphate. It finds the biggest market in the textile industry and is used as a builder in some synthetic detergents. Glauber's salt, Na2SO4. 10H20, is usually of high purity. Preferred for some uses, it normally has been recrystallized from an anhydrous salt. A unique manufacturing process has been developed at Monahans. This process results in the production of an exceptionally high grade of salt cake, and qualifies for nearly all uses, including many which specify the technical anhydrous grade. All of the finished product, which is very white, passes a 10-mesh U. S. Standard screen, and is retained on a 200-mesh U. S. Standard screen. It is over 99 pct Na2SO4 with main impurities being sodium chloride and magnesium sulphate. Iron content is less than 0.01 pct. As mentioned, the raw material at Monahans is a brine drawn from wells. Attention was first attracted to this location because a so-called alkali

Jan 1, 1954

By R. E. McNulty, R. D. Heidenreich, C. H. Gerould

TIIE electron microscope techniques and their application to magnesium alloys that are to be discussed in this paper are the result of research at The Dow Chemical Co. over the past three years. The view- point underlying the work is not wholly metallurgical, but has evolved through a variety of problems related to the physics and chemistry of surfaces. It can be said that the interest and stimulus were closely associated with a desire to employ effectively the electron microscope in research concerning the properties of solids as related to their physical structure. Only a portion of this work will be described here, and, consequently, many points of interest and debate are of necessity either only briefly recognized or are completely omitted. As has been pointed out by Desch,1 a great deal of the knowledge of the structure of metals has been obtained through microscopic studies of the surfaces of metallic specimens prepared in a particular manner. This information, when added to that obtained by means of thermal, electrical, X-ray, dilatometric and other methods, has been of great value both from a theoretical and a practical view- point. With the introduction of new instruments such as the electron diffraction camera and the electron microscope, it would be expected that still more information could be obtained. These two instruments have in many instances been very effective when directed toward the investigation of metallic surfaces, provided the proper methods of surface preparation were employed. Experience shows that the problems of surface preparation greatly limit the general application of these instruments to metallographic studies. It was found necessary to discard the standard metallographic procedures in applying the electron microscope to magnesium alloys and to work out new methods, which were required to satisfy certain conditions. These methods are unique to magnesium only in the specific chemical reactions occurring during etching. The presentation of the methods and results will be ordered as follows: I. Electron Microscopy of Surfaces. 2. Surface Preparation of Magnesium Alloys. 3. Discussion of Microstructures. a. Pure Magnesium and Single-crystal Studies. b. Types of Precipitation. c. Some Common Structures. d. Special Micrographs and Techniques. e. Structure of Mg-A1 Solid Solutions. 4. Correlation of "Fine Structure" with Corrosion Behavior of Magnesium Alloys.

Jan 1, 1946

By A. P. Svenningsen

IN the early stages of design it was not considered necessary that separate crushing plants be built for the new sulphide concentrator and smelter until sometime in the future. The plan was to use the existing crushing facilities for both oxide and sulphide ore. A few additions were contemplated for the existing plants, such as increased bin capacity, and possibly two new secondary crushing units. The more the problem was studied and discussed with the plant operators, the more it became evident that it was complex. It involved the classification of different kinds of ore from the open pit mine -sulphide, oxide and mixed-and how best to separate them so that each kind of ore was given the proper processing and treatment. It also involved the problem of keeping the different ores from being contaminated in bins, hoppers and chutes. Added to these, transportation became complicated and would involve additional handling and loading of ore from crushing plants to conveyors, to bins, and finally to railroad cars which were to be hauled to the concentrator and dumped into the fine ore bin. General In the early part of 1951 it was decided that the concentrator be constructed with ten grinding units instead of seven as originally authorized. The smelter was to be increased proportionally and naturally also the overall tonnages of ore to be handled by the new sulphide plant. Due to this increase in plant capacity and the larger tonnages involved, the difficulties which would arise by using the existing crushing plants were increased to a point where it became evident that the building of new crushing plants for sulphide ore exclusively was technically, as well as economically, advantageous. Authorization was, therefore, given by the company to construct new crushing plants to handle 30,000 tons of ore per day, and capable of reducing the run-of-open-pit ore to the proper size feed for the 10x14-ft rod mills in the concentrator. The ore, mined in the open pit, sometimes comes in pieces as large as 6 to 7 ft diam. The rod mills may call for ore crushed to 3/4 in. The large .size of ore delivered from the open pit determined that a 60-in. gyratory crusher be used as primary breaker. Such a crusher will have a capacity considerably in excess of 30,000 tons per day. The crusher will be a single discharge unit driven by a 500-hp electric motor through a tear coupling and a floating shaft. This type of drive has proven successful at a number of other crusher installations which our company has operating in the United States, Mexico and South America. The tear coupling will protect both the crusher and motor against damage in case of overload. No new features are incorporated in the design of the crusher itself, except that the, discharge chute is made the full width of the crusher with parallel sides instead of the usual converging sides. This change in detail should eliminate, a feature which has been a bottleneck in some of the operating plants and has caused loss of production due to ore hanging up and blocking the chute. The secondary crushing plants will have three 7-ft standard Symons cone crushers and six 7-ft short head Symons crushers. Between the primary and secondary crushing plants a coarse ore bin will be constructed with a nominal draw-off capacity of 30,000 tons of ore. The standard Symons and the short head Symons will be in separate buildings. All the crushing plants and the coarse ore bin are interconnected with conveyor belts for transporting the ore to the crushers at the tonnage rate desired. The final product of the new crushing plants is produced by the short head crushers. It will be delivered onto a conveyor belt leading to the top of the fines ore bin in the concentrator. A separate conveyor belt running the full length of the fines ore bin and provided with a movable tripper of rugged design will discharge the sulphide ore into the bin. The concentrator bin is planned and designed so that the installation of this additional conveyor will not interfere with the operation of the two railroad tracks on which crushed ore is brought from the existing oxide plant. Thus when completed the bin can be filled simultaneously by ore from the new crushing plant and by ore from the existing leaching plant.

Jan 1, 1952

By H. H. Reamer, B. H. Sage

The formation volume and the relative volume of the liquid Phase of mixtures of Condensate and gas from five different parts of a field in the Louisiana area have been established experimentally. These studies included gas-condensate ratios as high as 25.000 cu ft/bbl at pressures up to 5,000 psi. The gravities of the tank condensates were between 53.9° and 63.9° API. The compositions of the condensate and gas samples inves-ligated are presented in tabular form. The formation volume. the relative volume of the liquid phase. and the specific volume of the mixtures are available. A rough correlation of the formation volume as a function of pressure. temperature, and gas-condensate ratio has been presented. This correlation i-considered applicable only to the range of conditions covered in this investigation for systems of similar nature and com-position. INTRODUCTION Often a knowledge of the influence of temperature. Pres sure, and relative quantities of condensate and gas upon the volumetric behavior of fluid*. from a given field has been employed in estimating the Conditions existing in the reservoir. The early work of Beecher and Parkhurst' and the more rrcent Studies of the Dominguez' and San Joaqnin Valley Field;: are examples of experimental studies of this character. Correlation of the volumetric behavior of naturally occurring hydrocarbon mixtures has Proved To be feasible' . at low gas-conden-sate ratios-; for fluids*. from a single geographic area. Progres 5 is being made in the Prediction of the volumetric behavior of distillate systems at high gas-condensate ratios and pressure.. However. the estimation of the retrograde". ' dew-point pressure still remains uncertain and probably can best be established by experiment. For this reason it appears of interest to make available experimental information about the volumetric behavior of mixtures of condensate and gas from five different part.; of a single field. This material shows a rough similarity to data for mixtures of condensate and gas from San Joaqnin ' Valley Fields. In the present paper. 110 attempt has been made to coordinate the available information on tbe volumetric behavior of fluids from condensate fields. It appears hopeful that progress can be realized by the utilization of recently developed equa-tions of state.!' However. the primary problem in any such correlation is the proper identification of the components and constituents of naturally occuring systems. Additional experimental information may be required to permit the determination of the values of the constants of the equation of state for References given at end of paper. Manuscriptt received in the office of the Petroleum Branch June 11. 1951. *For present purposes low gas- condensate ratios will he considered as those below 3,000 cu ft/bbl, and the region of high gas-condensate ratio will be limited to compositions containing more than 5,000 cu ft of pas per bbl of condensate. Mixtures Containing between 3,000 and 5.000 CII ft OF gas per bbl will be considered of intermediate gas-condensate ratio.

Jan 1, 1952

By D. J. Rose, W. M. Armstrong, A. C. D. Chaklader

The wetiability of single crystals of MgO by specimens of vacuum-cast iron was studied using the sessile drop technique in vacuo at 1550ºC. Formation of FeO at the liquid-vapor interface caused the contact angle (6) to decrease from 117 to 65 deg during the first minute. After cooling, all specimens possessed a peculiar annular interfacial deposit of Fe,O,. Within the annulus the interface showed no sign of chemical attack. Chemical reaction occurred where the iron was alloyed with Ti, V, Cr, Nb(Cb), Ta, cmd Zr. Vnriation of 6 with alloy concentration was studied. Although vanadium md chromium improved the wettability of MgO by iron, the effect of Zr, Ti, Nb, and Ta was indeterminable because the 8 derived from sessile drop considerations was that of the metal against a restrictive peripheral volume of liquid oxide wetting the substrate. INTEREST in the nature of metal-ceramic interactions has been stimulated by progressive development of metal-ceramic combinations. One of the more valuable methods used for bond investigation is the sessile drop technique. Recent attempts to improve wettability through solute additions to the drop have revealed that the solute may a) react with the oxide creating new compounds at the solid-liquid interface, b) adsorb at the interface in a monolayer formation, or c) distribute throughout the drop uniformly causing no wettability variation. This work, part m in a series1,2 of investigations of interface reactions between metals and ceramics, embraces a study of the Fe-MgO system. MgO possesses a large negative free energy of formation (-173 kcal per g mole 0, at 1550ºC) compared to FeO at this temperature (-73 kcal per g mole 02). Hence the electronegativity difference between the drop and substrate allows selection from an extensive group of metal solutes with intermediate electronegativity differences that would, in the absence of chemical reactions, be expected to adsorb pre- ferentially at the solid-liquid interface. However, the high oxygen pressure of MgO in vacuo creates an oxidation environment which influences the solute behavior. Recent studies of the Fe-MgO systemJ74 have been restricted by chemical reactions that obscured observations. In view of the current interest in liquid-phase sintering of metal-ceramic combinations under partial oxidizing conditions,5 a more comprehensive study of the Fe-MgO system seemed beneficial. EXPERIMENTAL PROCEDURE 1) Specimen Preparation. Optical-grade MgO single crystals and Ferrovac vacuum-cast iron were used in all experiments. A spectrographic analysis of the iron indicated 0.008 C, 0.05 V, 0.005 Mo, 0.01 Ti, 0.002 S, 0.01 Cr, 0.001 Mg, 0.01 Si, 0.003 A1 (wt pct). The MgO crystals were cleaved along (100) planes into plates approximately 20 by 20 by 1 mm. Following annealing in vactto at 1100°C for 3 hr, surface irregularities were removed by a chemical etch in phosphoric acid at 100°C. The iron rod was machined into small cylinders approximately 0.250 in. in diam and length and these were carefully cleaned in organic and acidic solutions to remove surface impurities. All specimens were weighed to the nearest 0.1 mg. 2) Apparatus. The apparatus described in detail by previous authors1,2 was modified for this work. A Polaroid Land Camera was incorporated into the optical system so that drop measurements at ten-second intervals after melting were possible with ultra-high-speed self-develop ing film. 3) Procedure. The iron cylinders were placed upright on the MgO plates and positioned in the molybdenum susceptor. After furnace assembly the system was pumped down to a vacuum of 1 µ and flushed with H2 at 800°C. The system was again evacuated to 5 x 10-5 mm of mercury and the temperature raised to 1550°C within 2 min. The molten drops were photographed at appropriate time intervals. TWO min after melting the iron vapor pressure caused gas discharge, or corona, within the tube. The specimens were then slowly cooled to room temperature, sectioned and examined metallographi-cally. X-ray identification of interfacial reaction products was attempted by the powder diffraction technique. Alloying elements (Ti, V, Cr, Zr, Ta, Nb) were added to the iron in various concentrations up

Jan 1, 1963

By C. R. Holmgren

A large consolidated sand core was used in this investigation. Four groups of experiments including gas drives, water drives, and water drives combined with gas injection, were made. The results of the gas-drive experiments indicate that greater ultimate recovery may be obtained by higher pressure gradients. However, other considerations challenge the absolute veracity of this conclusion. Results indicate that with connate water present and a low pressure gradient during a gas drive, less gas is required to produce a given amount of oil. The saturation gradient determined during a gas drive is presented. Results of the water-drive experiments indicate that the recovery of oil is not a function of either gradient Or input rate. The presence of initial gas saturation resulted in a slightly lower final oil saturation. Injection of sufficient gas to maintain the initial gas saturation during a water drive resulted in lower final oil saturation. Introduction With the demand for petroleum products reaching an all-time record high and the location and exploitation of new reserves becoming more difficult and expensive, the oil industry must look to more thorough and efficient methods of producing the petroleum already located. Much of the previous work along these lines is exploratory and suggests a more complete investigation. For example, a great amount of laboratory work has been done on flooding sands initially saturated roo pct with oil. Since sands saturated roo pct with oil rarely, if ever, exist in nature, and because the water content is a vital factor in the ultimate recovery from the sand, laboratory investigations on cores without water must be supplemented by comparable work on cores with water present.1'2 Other investigators have done a considerable amount of laboratory work using gas and water drives.3 Krutter and Day4 have reported results of investigations using cores saturated IOO pct with oil and have concluded that " the ultimate recovery obtained, using air drive, from a completely oil-saturated core depends on the pressure gradients used. The higher the pressure gradient, the greater will be the recovery until a certain Pressure range is reached above which the reward of increased recovery is much smaller for similar increments in pressure." Yuster and Day6 conclude that less gas and less time are required to produce a given quantity of oil at high pressure gradients than at low pressure gradients; they also conclude that an increase in the viscosity of the oil requires a proportionate increase in the gas needed for recovery. Their results indicate that the gas required to produce a given quantity of oil is greatly reduced by the presence of connate water. R. C. Earlougher2 concluded, based on results of water-flooding a large number of

Jan 1, 1949

By C. Richard Tinsley

INTRODUCTION Project financing has been defined by various authors in this book. It is founded on reliance upon the project's own future cash flows and secondarily on the collateral value of the asset base developed. It is chiefly provided by the banking community who, after an assessment of the risks involved (See Chapter 9) analyse the company's presentation (See Chapter 10) in order to structure a financing package often at a leverage of up to 80% debt. This paper describes the various support and structuring techniques used to delimit and abate the risk components perceived from the banks' experience on project development especially for the period before cash flow is generated (pre completion). ADVANTAGES The advantages of project financing extend beyond the offloading of risks to banks. Among the advantages are the following: 1. Longer loan maturity than conventional, unsecured financing; 2. Loan disbursement and repayment tailored to specific mine; 3. Ability of a joint venture without a "track record" to raise financing; 4. Obtain financing where further direct borrowing (corporate credit) limited by existing indentures; 5. Possibility of arranging finance off-balance sheets; 6. Leverage return on equity; 7. Obtain tax savings with certain project financing structures; and 8. Obtain flexible loan repayment conditions. It can also be said that from the point of view of the company, project financing has certain disadvantages which can be summarised as follows: 1. Perceived higher cost (risk premium) compared with straight corporate credit; 2. More documentation; and 3. Need to negotiate risk-sharing aspects. The main attraction of project financing is the ability to pass certain risks on to the lenders within a flexibly funded transaction. (For a more detailed description of the specific risks see the author's paper at the beginning of Chapter 9. The same risk system is followed here.) The typical project financing has corporate support prior to completion and will usually carefully control the resultant cashflow after completion to ensure the bank does not unnecessarily assume further risks of an equity nature. WITHIN THE COMPANY'S CONTROL Operating Risk This is broken into three interrelated segments for ease of description. Technical Component: This must be specifically addressed if new technology or unproven adaptations are to be used. The banks will have scrutinized the sampling and bulk testing work, perhaps grilling the geological, mineral processing, metallurgical, and mining engineering staff or consultants directly. A number of supports are available such as: TECHNOLOGY GUARANTEE: Either the company itself or the process developer will guarantee that the technology will work for a period of, say, six months to two years after startup. If not, then financial penalties will apply although these are rarely sufficient to repay the loan. Many process developers are weak financially and may have little substance to back such a guarantee.

Jan 1, 1985

By O. M. TUPPER

REQUIRING a finer ground clay than that obtainable with a dry pan or hammer mill, the Clay Corporation of California installed a five-roller, low- side Raymond impact mill at its Lincoln plant in 1925 and since then has been economically grinding fire clay with a moisture content as high as 15 per cent. This was one of the first, if not the first, of this type of mill to be used in the grinding of clay on the Pacific coast. The pulverizer consists of a die-ring, lying in a horizontal plane, enclosed in a sheet-iron housing. A vertical shaft supports a spider from which five rolls are suspended in such a way as to always have them bearing on the die-ring. The vertical mill shaft is driven by means of beveled gear and pinion from a horizontal shaft which is connected by belt with a 60-hp. motor with suitable pulleys to give a speed of 175 r.p.m. A large fan which has a speed of 1100 r.p.m. and is belt connected to a 40-hp. motor is connected to the mill housing. The discharge end of the fan is in turn connected to a cyclone dust collector.

Jan 1, 1929

By J. G. Balkestein, J. W. R. Baerts

During an attempt to develop a method for accurate, rapid, continuous analysis of ash content of wal, the Dutch State Mines Laboratory found that the absorption coefficient for X-rays was related to ash content. This relationship was utilized in the development of an apparatus called CendreX, used to measure the absorption and in turn the ash content. The principle, procedure, accuracy, and applications of the CendreX automatic determination of ash in coal are discussed in detail. Until recent years coal preparation research was concentrated on the possibilities and efficiency of preparation equipment. Ever increasing demands for quality in coal, however, have caused researchers to become concerned with quality control. One of the characteristics strongly influencing quality (calorific value) and, consequently, the price of coal is ash content. Both customer and producer require that the ash content be controlled. To make this possible it is necessary for the producer to collect samples from the coal stream in the washery and to determine their ash content. On the strength of the results obtained the production process can then be adjusted, if required. Until recently incineration was the only method for determining ash content. When coal is incinerated with oxygen, it takes at least 15 min to determine the ash content. Even then some time may be lost in passing on the data. Therefore, this discontinuous and time-consuming method does not make a suitable starting point for checking, re-adjusting, or modifying the production process. Since it was not possible to accelerate incineration by mechanization and automation, another type of rapid continuous analysis had to be found. This would have to depend on some coal characteristic closely linked with ash content. At the Central Laboratory of Dutch State Mines, Dijkstra and Sieswerda found that the absorption coefficient for X-rays was suitable for this purpose. An apparatus based on the measurement of this characteristic was developed. Named CendreX,* this patented equip- *A Stamicarbon N. V. registered trademark. ment, used in combination with a sampler and conditioning apparatus, permits continuous determination of ash content with satisfactory exactness to within 3 min.t At the American Mining Congress meeting of May 1961, Log A. Updegraff, surveying aspects of continuous analyaia, reported on the CendreX appartus (Mining Congress Journal, August 1961). Since the first Condition for correct ash determination, whatever the method used, is always a sample taken in the right way, some practical aspects of sampling will be discussed first. SAMPLING In coal handling operations, each grain has its characteristic properties. In both the cross-sectional and longitudinal direction of the product stream the mixture of these grains is imperfectly blended. The elimination of the factor of segregation between the cross-sections from the direction of the stream is the principal aim of continuous analysis. Therefore, the samples to be analyzed are withdrawn in various places from the segregated product stream in such a way that they form a semi-continuous flow. In order to be truly representative of the product stream, the sample flow must fulfil two conditions: 1) The sample flow must be sufficiently large to insure that adventitious differences between the composition of the product stream and sample flow will be as small as possible. 2) All grains through the entire cross section of the product stream must have an equal chance to get into the sample flow. Since there is always a segregation through the cross section the sample flow will have a bias if the sampler collects only part of this section. In order to prevent the volume of the sample flow from becoming too large, the cross sections are taken discontinuously. However, this discontinuity gives rise to the adventitious differences that arise between the composition of product stream and the sample flow, and these differences become greater and are more deranging as the fluctuations follow each other more quickly. Moreover, the intervals between the morrnents of sampling mean a delay in the analysis. The aim will, therefore, be to collect small samples, or in other words thin cross-sectional increments, taken at sufficiently high frequencies. The following points should also be taken into account: 1) The smallest dimension of the openings in the sample collector must be equivalent to at least 3 d in

Jan 1, 1962

By T. Cantwell, N. C. Rasmussen, H. E. Hawkes

Beryl is a mineral that may be difficult to distinguish from quartz by casual field inspection. The easily recognized green color and hexagonal crystal form of coarse-grained beryl are by no means universal, even in beryl from pegmatitic deposits. If it occurred as a fine-grained accessory mineral in an igneous rock, it would almost certainly escape detection unless samples were submitted for petrographic or chemical analysis. There may be substantial deposits of some beryllium mineral, other than beryl, that has been overlooked because that mineral also closely resembles the common rock-forming minerals. A reliable and simple method of identifying beryllium minerals and determining the beryllium content of a rock would be helpful in exploration. This article describes preliminary experiments in applying nuclear reaction to the qualitative identification of beryl and to the semiquantitative determination of the beryllium content of rock samples. Gaudin,1,2 the first to apply a nuclear reaction in detecting beryllium minerals, developed a method that irradiates the sample with gamma rays, which react with beryllium nuclei to produce neutrons. The neutrons are then measured with standard equipment. The cross section for this reaction is about 1 millibarn. The cross section is a measure of the probability that a reaction will take place, for example, between a beryllium nucleus and an incident gamma ray or alpha particles.3-5 At 1-millibarn cross section for the reaction, satisfactory performance required a source strength of the order of 1 curie (3.7 x 10"' disintegrations per sec, where each disintegration releases one or more gamma rays). The reactions will not take place if the gamma radiation is below a minimum energy, in this case 1.63 mev. The size of the source and the energy of the radiation made heavy shielding necessary for these experiments, both to reduce the background count of the neutron counter and to safeguard personnel. The original discovery of the neutron by Chad-wick in 1932 resulted from experiments with another nuclear reaction, induced by bombarding beryllium with alpha particles in which the products are carbon-12 and neutrons. The equation for this reaction is as follows:' " ,Be" + ,He'? 6C12 + 8,n' [1] re-particle neutron In the above nuclear equation (Eq. 1), the sub- script number indicates the number of protons in the nucleus (the atomic number) and the superscript the total number of neutrons and protons (approximately the atomic mass). For the alpha-neutron reaction the cross section is about 250 milli-barns, or 250 times that of the gamma-neutron reaction used by Gaudin. The positively charged alpha particle is repelled by the positive charge of the beryllium nucleus; it must, therefore, have a certain minimum energy in order to approach close enough to the beryllium nucleus to react. For reaction with the beryllium nucleus, the lower limit of the alpha-particle energy is 3.7 mev. The alpha-neutron reaction, with polonium-210 as an alpha source, was selected for the present experiments. Alpha particles are emitted by polonium-210 at 5.30 mev, which is adequate for the reaction with beryllium. Furthermore, this isotope of polonium emits alpha particles with negligible associated gamma radiation, thus eliminating the necessity of shielding. The half-life of polonium-210 is 138 days. Inasmuch as alpha particles carry a possible charge and are large compared with most nuclear particles, their energy is rapidly dissipated in passing through matter. Their range in standard air is 3.66 cm,3 and they penetrate only a few tens of microns into a mineral sample. The short range in air can be minimized by preparation of a flat sample surface that can be brought very close to the alpha source during analysis. On the other hand, short range of alpha particles in air lessens the radiological health hazard and makes it possible to use this method without shielding. It must be emphasized, however, that the alpha emitters are potentially very dangerous if they enter the human body. Polonium must be handled with extreme caution. The literature has reported experiments on the yield of neutrons from reaction of alpha particles with beryllium nuclei. Feld" reports that in intimate mixtures of polonium and beryllium, 3 x 106 eutrons per sec are produced per curie of polonium. Elsewhere in the same reference it is stated that a sandwich-type source yields about one third as many neutrons as an intimate mixture. A table of neutron yields for full energy polonium alpha-particles on thick targets as reported by Anderson7 is the basis of Table I. From Table I it can be deduced that the elements most likely to interfere, i.e., those that also produce neutrons when bombarded by alpha particles, are boron and fluorine. These data also show that it will probably not be possible to determine very small quantities of beryllium in rocks because of the masking effects of the major elements, sodium, magnesium, and aluminum. The neutrons emitted in the alpha reaction are detected by another nuclear reaction. Either of the

Jan 1, 1960

As knowledgeable men in the industry know, a mineral occurrence is found through prospecting but an ore deposit is "made" -made through, first, imaginative and effective exploration; thence through efficient engineering, sales and management. There are two main phases of exploration-reconnaissance and target investigation-each of which can be further divided into two additional stages (see accompanying table). The search for a new ore deposit within a known and geologically mapped mining district involves only the last two stages, whereas the search for a new orebody in a virgin geologic environment involves at least three stages and often the full sequence. Of course, some of the exploration stages may be bypassed depending on available results of previous mineral surveys and/or exploration by individuals, business concerns, and governments.

Jan 10, 1967

By Cleland N. Conwell

APPLICATION of seismic methods to the location of buried channels in placer mining is fairly common knowledge, yet some of the facets of seismic refraction surveys have remained obscure in their application to mining. This description by a member of the U. S. Bureau of Reclamation team investigating bedrock at the Oso Dam site, may offer a lead to wider application of the method for bedrock determinations in engineering of mine and mill plants, tailing disposal, and water supply. It may also offer assistance in planning geological exploration in areas covered by overburden. Pertinent, and perhaps more than passingly significant, is that the seismic survey of the Oso dam site started at noon and was completed the afternoon of the next day. Location Oso dam site is on the Navajo River, Archuleta, County, Colo., 30 miles southeast of Pagosa Springs, Colo., 10 miles northwest of Chama, N. M. Bedrock at the proposed dam site is Mesa Verde Formation, Cretaceous age sandstone. The nearest good sandstone exposure is about 5/8 of a mile upstream from the location of the seismic determinations. The material above bedrock is sand and gravel with sandy loam soil covering benches above the river. The site appeared favorable, after a geologic and topographic reconnaissance, but overburden thickness was unknown. Exploratory drilling was not contemplated in the near future so a short seismic investigation was selected as the most suitable method of determining depth to bedrock. Field Procedure Geophysically, a seismic depth investigation is simply the measurement of depth from ground surface to the top of a layer which transmits seismic waves at a higher velocity. In this case the high-velocity seismic layer tentatively correlates with the Mesa Verde sandstone. Field procedure required a seismic explorationline or "spread" with 12 geophones set at measured intervals. Explosive charges were buried at few feet in the ground and detonated at shotpoints near the ends of the spread, and at intermediate points. The time of travel of the seismic waves from the shotpoints through the ground to each of the geophones along the spread was measured and recorded photographically with equipment housed in a field truck. The seismograph on each spread provided a time-travel graph. From it were computed the speed of seismic waves in various layers below the surface and depth to each different seismic layer below the shotpoint. In the case of the Oso dam site it was possible, in addition, to determine the depths below certain geophones, because there were only two seismic layers involved. Interpretation In seismic exploration waves are refracted from the top of layers which often correspond to divisions between subsurface formations. When drill holes are available the geologic logs offer a means of correlating stratigraphic breaks with seismic layers, improving reliability of seismic interpretation in terms of depth to bedrock or other geologic terms. There were no drill holes available in the case of the work at the Oso dam site. Geologic conditions appeared favorable for seismic technique because of excellent contrast in seismic wave velocity between the overburden and the Mesa Verde formation. The upper layer had a velocity of approximately 2500 fps and was interpreted as overburden, the lower layer had a velocity of approximately 9000 fps and was assumed to be the sandstone. These velocities are of the order expected in these materials. The seismographs obtained at the Oso dam site were excellent. The energy from the shot came in strongly at each geophone, giving an easily read record. Time-travel graphs plotted from wave arrival times were uniform and made interpretation comparatively easy. In Spreads 1 and 2 near the river the lower velocity layer was 12 to 15 ft deep. For the spreads on higher ground, or the bench south of the river, it was from 15 to 29 ft deep.

Jan 1, 1952

By E. P. Mathewson

THE experiments described in this paper were made at the Washoe Reduction Works, Anaconda, Mont., for the purpose of determining the relative speed of elimination of the iron, sulphur and arsenic during the process of bessemerizing copper-mattes. The samples were taken from a horizontal "barrel" converter, 8 ft. in diameter and 12 ft. 6 in. long, having an average charge of 9 tons of matte, and blown with a 16-1b. air-pressure. All samples were taken with the blast on, turning the converter from its normal blowing-position to nearly verti¬cal, so that a conical mold, with a long handle attached to a horizontal rod, could be introduced from a platform above the converter into the mouth of the converter and a sample taken. The difficulties of taking a sample with the blast on are appar¬ent, and the irregularities of the lines in Figs. 1, 2 and 3 can be readily accounted for by the amount of slag in the samples. On the whole, the plotting of the results gives comparatively uniform lines, considering the conditions. The samples were taken by the regular sampling-force of the works at 10-min. intervals, as shown in Tables I., II and III, which also give the chemical analysis of the samples so taken. The test shown in Table I. was made with a new converter; that of Table II with a converter that had made three charges and had been cleaned out; that of Table III with a converter that bad blown one charge to white metal, and had then been " washed " out and used for the test. In plotting these results, it was necessary to take, as a basis, the percentage of copper in the samples. This basis does not exactly represent the true condition, since the volume is constantly decreasing and a certain quantity of copper passes into the slag, but it is as near as can be obtained in practice. Take, for instance, in Table I., the sample No. 5, which shows : Cu, 59.9;

Jan 1, 1907

By Charles F. Manness

In the great increase of product which has been accomplished in the American Bessemer plant, through improvements in machinery and refractories, and by more perfect skills in their manipulation

Jan 1, 1881

By James H. Boettcher

INTRODUCTION There are many and varied sources of finance available to the mining industry for exploration, development and/or the operation of mining projects and companies. These sources include equity from major as well as "penny" stock exchanges, precious metal mutual funds, private investor groups (from "grubstakes" to tax-oriented limited partnerships) and banks. The variety is representative of the fact that the world's capital markets are segmented and stratified according to risk-reward characteristics. This interrelationship is summarized in Exhibit I and 2 which relate (very approximately) various sources of finance to a typical mining project's life cycle and risk-return characteristics. The objective of this paper is to introduce, as it applies to the mining industry, a somewhat recent phenomenon represented by one of these capital markets segments referred to as "venture capital." Venture capital investing per se, which today is a multibillion dollars per annum investment business, goes back as far as Queen Isabela's backing of Columbus' "wild ideas." The real institutionalization of the venture capital process really began in earnest, however, after World War II. Today over 700 private venture capital sources (and more government related ones) exist providing funding to a range of industries, but typically related to the "high-tech" business. The table to the right shows industries financed by venture capital funds between 1980 and 1982. Conspicuous by its absence is, of course, mining which is not a great surprise since typical mining projects up until recently were large, required long lead times and were undertaken by established mining houses. Also, as most of you can appreciate, successful investing in the mining business requires a great deal of experience and understanding of the industry which most venture capital funds do not have. TARGETED RETURN The rate of return objective of most venture capital funds is, according to the legalized version, to realize a return which is "substantially in excess of that obtainable through conventional equity investments." One may say that a common objective of venture capital funds is to realize a 20% to 30% compound annual internal rate of return on capital over their life which is often ten years. [ ] This objective has been recently more feasible in the mining sector due to the development of smaller precious metal projects which cost less and are quicker to generate cash flow than the behemoth projects of the 1970s. In addition to precious metals, venture capitalists have invested or looked at a variety of other mining projects including ones as prosaic as a high margin "kitty litter" operation, sand-gravel and other industrial minerals operations, the leveraged buyouts of coal mines or mineral divisions of major energy or mining corporations, and even in in-situ copper or gold leaching ventures. A typical investment is in the range of $100,000 to $3,000,000 although some can go as high as 5,000,000. This implies a project size up to, say, 30,000,000 if a venture capital fund does it on its own. Larger projects can be done if the equity portion of the investment is syndicated with other investors.

Jan 1, 1985

By W. R. Ingalls

During the last two years, and especially during the last six months, a number of important articles upon the new methods for the desulphurization of galena have been published in the technical periodicals, particularly in the Engineering and Mining Journal and in Metallurgie. I proposed for these methods the type-name of " lime-roasting " of galena as a convenient metallurgical classification,'and this term has found some acceptance. The articles referred to have shown the great practical importance of these new processes, and the general recognition of their metallurgical and commercial value which has already been accorded to them. It is my present purpose to review broadly the changes developed by them in the metallurgy of lead, in which connection it is necessary to refer briefly to the previous state of the art.

Jan 1, 1907

By E. A. Hartsook

Since World War II we have all been treated to repeated public and private evaluation of the population explosion. Improvements in medicine and agriculture have brought greater life expectancy and a decreasing infant mortality rate. This has brought boom conditions to some areas-and pressures everywhere for the amenities of life: a little elbow room, a tree or a pot of flowers, space to hike or swim with some freedom have all become premium commodities. With this wish for privacy and attractive surroundings, people are being forced to move to once sparsely populated areas. Suburbs expand or sprawl, depending on the point of view. New country is touched by a freeway and houses and stores spring up overnight. Entire new cities are born to meet the increasing mobility of industry and its work force.

Jan 6, 1965