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By E. L. Anders

The Mile Six pool is located on the La Brea-Parinas Cullcession of International Petroleum Co., Ltd., in northwestern Peru on the west coast of South America. The reservoir pressure in this pool has been maintained within 200 psi of its initial value throughout its history, and gravity drainage has played an important role in the production behavior. It has now produced 95 per cent of its estimated ultimate recovery. It is estimated that this interesting oil pool will ultimately produce 67 per cent of the initial oil in place and that the resulting residual oil saturation may be as low as 19 per cent of the pore volume (29 per cent of the hydrocarbon pore volume). An evaluation of reservoir rock and fluid characteristics and ultimate oil recovery is presented. INTRODUCTION This study of Mile Six pool was made to evaluate its performance according to latest available information. The production performance of this pool has been discussed in various articles in the past. and the reported behavior has been used as an example for application of computation procedures for gravity drainage depletion' and as an illustration of field behavior under gravity drainage or expanding gas cap drive. There have been wide variations in reported values of initial oil in place, reservoir oil volume factor. connate-water sauration, volume of effective sand, and ultimate recovery because of the paucity of reliable basic data. These various factors have been determined as accurately as practicable with the latest available information, and this evaluation is presented herein. The production history of Mile Six is an excellent example of gravity drainage depleion with effective pressure maintenance by gar injection. GENERAL Mile Six pool was discovered by cable-tool drilling in November. 1927. when well 1996 was completed in the Parinas sand. After slow development with cable tools and sporadic production. the pool was opened to continuous pro(iuction in November. 1933. and develpment was completed with rotary rigs. Pressure maintenance was started in December, 1933. by returning gas to upstructure wells. Most of the development was 'completed by 1937, but some additional wells were drilled in the period 1939-1947. and several old wells were deepened. A total of 46 oil and gas wells and 4 dry holes were drilled on approximately 7-acre spacing. Of the producers. 21 are now flowing. 2 are pumping. 44 are gas input wells. 3 are abandoned. I is a gas well shut in. and 15 are shut in because of non-commercial production or high gas-oil ratio. The locations of all wells are shown on the map of Fig. I. Total oil production on Dec. 31. 1952, was 30,867,373 bbl: cumulative gas production was 22,023,777 Mcf; and 26,410,946 Mcf of gas had been returned to the reservoir. These figures do not include oil and gas lost ill a blowout in January. 1940. GEOLOGICAL DESCRIPTION Mile Six pool is located on the northern end of a structural spur projecting from the La Brea-Negritos uplift.' The spur is probably a reflection of a basement structure. It plunges gently to the north, i broken into a complex series of fault blocks. and contain.; the Verdun Alto. Section Sixteen. and Mile Six pools. The Parinas handstone (lower Eocene!. which is the producing formation in Mile Six. occurs at an average depth of 2,200 ft in the pool and dips north and east at from 15' to 20". The pool covers an area of approximately 350 acres. Mile Six is downfaulted about 600 ft from Section Sixteen pool to the soutb. and a major fault forms its western boundarv. The north and east boundaries are formed by the intersection of the sand top with the water-oil contact which occurs at approximately 2.440 ft subsea. An original gas-oil contact probably existed at about 1.875 ft subsea. Fig. 1 presents the latest structural interpretation of the pool. and Fig. 2 is an isopach map showing thickness of the total Parinas formation above the original water-oil contact. The heavy lints of Fig. 1 are contours on the sand top, and the fine lines are contours on the fault planes. This type of straight-line structural map was developed 1)) International's geologists to reflect structural conditions where the bedding planes dip and have no curvature. 'The La Brea-Parinas Concession is highly faulted by normal fault.. The beds are flat wherever exposed. The Parinas formation is approximately 635 ft thick. and it is etimated that 62 percent II the formation is effective sand. The original oil zone was about 565 ft thick. Fig. 211 presents an electric log showing typical Parinas sand development ill Mile Six pool. The Parinas band in Mile Six i. a well-sorted. medium- to-coal-se-grained. cross-bedded sand with minor lenses of shale and small lenses and pockets of pebble conglomerate. The sand grains are subangular to rounded and consist chiefly of quartz with feldspars. biotite hornblende, and augite as accessory minerals. Because of faulting of the Parinas- formation to the east and north of the pool. there is probably l possibility of a significant, natural water drive in Mile six. The faults within the pool. as indicated in Figs. 1 and 2. are of smaller disI,laceInent and seem to act only , partial barriers to fluid movement within the reservoir. RESERVOIR CHARACTERISTICS Core analysis data are available from five wells. The data were obtained from three wells (Nos. 3401. 3586 and 3719) at the time of their completion and from well 1996 when the original liner was sidetracked and the well was deepened ill 1946. Data from well 2779 were obtained in 1943 from old cores taken when the well was deepened in 1934. From these core analyses. the average porosity was estimated to be 22.6 per cent. and the average permeability to dry ail. was estimated to be 780 and Measured productility indices varied from 3.1 to 71.4 B/D per psi differential. Specific productivity indices varied approximately from 0.1 to 0.3 B./D per psi per ft of sand.

Jan 1, 1953

Very little has been written about the discovery and early history of prospecting of ironstone deposits in the Iron Knob-Middleback Range area of South Australia, the main source of supply of ironstone (hematite) for the Australian steel industry, and scanty though the record is, much of it is wrong or misleading.With the help of several people, to whom acknowledgement is made at the end of the paper, the writer has unravelled some of the story, but not the whole story. It is hoped that publication of this paper will stimulate interest in others, who may be able to supplement this account from family records, from the files of newspapers, or from archives.

Jan 1, 1969

By K. F. G. Hosking

I read Professor Gaudin&apos;s paper with great interest and pleasure because for some time I have held that the chemical aspect of comminution is a subject of considerable importance to the mineral dresser and deserves to be thoroughly investigated. It does seem appropriate, however, to emphasize the fact that "fresh" edges and corners produced by the grinding of solids display enhanced reactivity has been recognized and utilized in the development of certain mineral identification techniques. Some of these techniques are worth noting, not only because they might facilitate research in the aspect of mineral dressing under discussion, but also because they emphasize the fact that many mineral species commonly regarded as being very inert can display a surprising reactivity when in the freshly ground state. In 1951 Isakov6 published a number of tests for the components of certain mineral species which depend essentially on grinding in a mortar a mixture of the material under investigation with a solid reagent. Thus when stibnite, 4(Sb2S3), is ground with sodium or potassium hydroxide. the antimony is revealed by a momentary development of a yellow color which changes in air to orange-red. Other antimony minerals need a preliminary treatment before the test can be carried out. This consists of grinding with aluminium sulfate, ferric sulfate or potassium bisulfate, and breathing upon the resultant mixture. I have employed a similar technique to determine the approximate magnesia content of certain limestones.&apos; The method depends essentially on the fact that when a sample of limestone is ground under controlled conditions with solutions of sodium hydroxide and Titan yellow the color of the final product is, within limits, a function of the amount of magnesia present. I have also shown that the components of a wide range of minerals can be identified by applying chemicals to their streaks on portions of vitrified, unglazed floor tiles, etc. Under such circumstances the diversity of the reactions which take place in the cold (because of the reactivity of fresh corners and edges) is surprising. Thus, for example, if a garnierite, (Ni,Mg)3Si2O5(OH)1, streak is treated first with a drop of 0.880 ammonia and then with a drop of a 1 pct alcoholic dimethyl-glyoxime it immediately becomes red, indicating the presence of nickel.&apos; Stevens and Carron9 have evolved a simple field test for distinguishing minerals by "abrasion pH." A soft nonabsorbent mineral is scratched in a drop of water on a streak plate until a milky suspension is formed. A piece of pH indicator paper is dipped into the suspension, after which it is removed and the maximum deviation from neutrality noted. When a hard mineral or one which absorbs water is being tested, fragments are first ground for 1 min with a few drops of water in a mortar to make a heavy suspension. The importance of the findings of such tests to mineral dressing may be judged by the abrasion pH values, Table 11, recorded by Stevens and Carron for certain species usually regarded as comparatively inert. The combined results Of the above researches clearly indicate that comminution is capable of altering the pH of a pulp and of causing the chemical nature of the surfaces of some of the components to be profoundly changed&apos; Depending On circumstances such surface alterations may have a beneficial or an adverse effect if these products are subsequently subjected to flotation. The tests also suggest that by grinding "inert" minerals with appropriate solid or liquid reagents "reactive" surfaces may be developed which might facilitate separations by flotation. It is an interesting and instructive problem to determine the reactions that are likely to take place when dry solid substances are subjected to comminution and to the unavoidable heat liberated during the process. To do this it is theoretically necessary to know the free energy values of the reactants and possible resultants, but unfortunately there is a dearth of such data! However, the heats of formation of many substances are known, and generally speaking, if in a reaction of the type AB + CD = AD + CB the sum of the heats of formation of AB and CD is less than that of AD and CB the reaction will probably proceed to the right. Thus, according to a note I have (the author of which I cannot name) if PbS (black) is warmed with CdSO, (white), PbSO., (white) and CdS (yellow) are formed, and that the reaction does, in fact, take place is indicated by the change in color of the mixture. The reaction is expected, as the sum of the heats of formation of PbS and CdSO, is less than that of PbSO, and CdS (as shown below). PbS + CdSO4 = PbSO4 + CdS 22.2 + 218.0 < 216.2 + 33.9 Finally, certain other aspects of the chemistry of comminution, which are neither mentioned by Professor Gaudin nor referred to by me are to be found in a paper by Welsh" and in the printed discussion thereof. A. M. Gaudin (author&apos;s reply)—The observations contributed by Dr. Hosking are indeed welcome, as they add to our experimental knowledge of a topic which is believed to be of the first importance. In connection with the experiments cited it should be kept in mind that oxidation, hydration, and carbonation at various rates should always be deemed to be possibilities when grinding is done in water or in air, even in "industrially dry" air. Special precautions might lead to sufficient minimizing of these reactions and to the assertion, instead, of deliberately-created reactions. The author wishes to thank Dr. Hosking for his contribution.

Jan 1, 1957

By R. D. Lord

The search for the best practical means of storing sulfide bearing tailings, where there is no residual excess of carbonate material is discussed in this paper• Usually the sulfide content decomposes, with the aid of bacterial action, and the resulting sulfuric acid escapes, along with any heavy-metal solutes, through embankments that are usually porous to some degree• The problem is typified in the tailings of the uranium operations of Elliot Lake, Ont., where mining started some 20 years ago• The approach to tailings disposal paralleled the practice for other hydrometallurgical plants treating gold and base-metal ores• Impoundment areas were designed to retain solids, and a clear and neutral overflow was considered satisfactory practice• Now experience has shown that these areas, some of which have been idle for over a dozen years, release acids in seepage and overflows to an unacceptable degree• To protect natural water courses, neutralizing plants are operated wherever required• Lime slurry is fed continuously into the tailings outflows in a quantity sufficient to raise the pH to 8•5 and precipitate heavy metals that may be in solution• The objection to this procedure is that the plants will require servicing indefinitely, unless a better remedy is found• The problem differs only slightly from that common to base-metal concentrators in that here the ore has been leached with sulfuric acid for the recovery of uranium• Any native content of calcareous material has been digested, and only that added for final neutralization is available to maintain a pH unfavorable to bacterial activity• Chemical oxidation slowly lowers the pH and when this reaches a level of 4•5 or less, bacteria become active and greatly accelerate the formation of acid. The bacterial process is probably at least ten times as fast as the chemical oxidation• Location and Processing The operations referred to, uranium and one copper mine, are located at approximately 46°N and 82°W longitude• This is typical Canadian Shield country, a land of lakes, deeply glaciated and rocky, with sparse soil which supports mixed forest cover• Drainage is to Lake Huron, 25 miles to the south• Average temperature is 45°F, ranging from -40° to +95°F• Annual precipitation is 38 in•, about half of which is snow• The ore is Precambrian, quartz-pebble conglomerate, with mineralization in the matrix• From 5 to 10% pyrite is present• All known means of pre-concentration have been tested, but a bulk sulfuric acid leach has proved the most efficient. Tailings have from the outset been neutralized before release• Current practice is to add ground limestone to bring the pH to 4•5, and then lime to raise the value to 10•5• Environmental regulations have recently been increased and the foregoing meets the new standards• Separate measures are taken to precipitate radium• Remedial Measures Since the outstanding environmental problem is the oxidation of pyrite by bacterial action, the solution is to contain the products, or arrest the process• Given the ambient temperature, favorable half of the time, four items are essential to the activity• 1) Pyrite• 2) Moisture pH < 4•5. 3) Oxygen• 4) Bacteria• Removing any one of these out of the range of tolerance will bring the reactions under control• A variety of proposals considered, and a number tested for the arrest of the process, are: (a) render embankments impermeable, (b) provide an impermeable cover, (c) cover with an oxygen absorbing layer, (d) provide a vegetative cover, (e) flood the site, (f) remove pyrite from current tailings, (g) add excess limestone to current tailings, (h) poison the bacteria• Bank Seal-On existing impoundment areas, where the embankments are several thousand yards in length, it is believed that any program of injecting sealants can have small chance of success• However, a moisture barrier is an indicated specification for future construction, and this can be highly expensive• Surface Seal-Depending on the configuration of the deposit, the downward travel of water should be prevented, and oxygen excluded• Burying a plastic membrane just below the surface has been considered, as has the application of a liquid sealant that would penetrate the surface. The objection to these remedies is the excessive cost of dealing with large areas and the expectation of only temporary benefit as a result• Frost penetration is over 4 ft, and frost action breaks up asphalt paving and all but heavy concrete in a few years• Organic Layer-An oxygen-absorbing layer, such as bark fines from paper mills has been proposed as a surface treatment• Cultivated into the tailings such material might be expected to arrest subsurface oxidation for some years• Estimates are 100 tons per acre of bark fines, or 35 tons per acre of sawdust, and these enormous quantities do not so far give assurance of providing a long-term remedy• Vegatative Cover-Several obvious benefits would result from a good growth of grass or other vegetation on abandoned tailings• While restoring the natural green of the tract the growth would prevent wind-blown dust and reduce erosion• Subsurface oxidation should be reduced, as well as the upward movement of ground moisture as occurs in dry weather. To this end, considerable research and field testing has been carried out to arrive at a formula - a prescription which will provide a self-sustaining growth on the tailings surface, or at least one that would survive with reasonable maintenance attention. Many test plots have been run with different combinations of surface treatment and seed mixtures. Generally, by addition and close cultivation of limestone, lime, and fertilizers, technical success has been demonstrated• Plants with a high tolerance for acid soil seem the more hardy, and a pH above 3 is indicated so that nutrients can be absorbed• Recommendations are for 12 to 15 tons of

Jan 1, 1977

By Jacques Steininger, Robert E. England

Single crystals of ZnTe and ZnTe1-,Sex with x up to 0.13 have been grown from the elements by temperature gradient solution zoning using excess tellurium as a solvent. Best results have been obtained with charges with the compositions 45/55 at. pct Zn, Te, for ZnTe and increasing amounts of selenium for ZnTe1-xSex. The temperature in the molten zone was maintained at about 1070°C with a gradient of about 10°C per cm. Chemical analyses of quenched ZnTe ingots show tellurium concentrations in the molten zone as high as 70 pct with concentration differences across the zone of 1 to 2 at. pct Dark dots which are observed by transmitted light microscopy in as-grown crystals can be removed by annealing in zinc vapor at 900 C. INTEREST in wide band gap semiconductors has led to a new study of ZnTe and ZnTel-xSex crystal growth. ZnTe is the only II-VI compound with a wide band gap (2.3 ev) that can be made p type with low resistivity. Attempts to make it n type with low enough resistivity to be useful for p-n junctions have so far been unsuccessful.1 ZnSe has a band gap of 2.65 ev but can be made n type only. However, ZnTel-xSex solid solutions with x as low as 0.36 have been made both highly n and p type2 with a minimum band gap around 2.12 ev3 at room temperature and appear to hold the best promise for efficient injection electroluminescence in the visible. ZnTe has the lowest melting point of the zinc chal-cogenides (1295°C) and consequently attempts have been made to grow crystals from both the liquid and the vapor phase.4 Complicated apparatus is required for growth from stoichiometric melts because of the high vapor pressures of the elements at the melting point of ZnTe and because of the problem of quartz devitrification. Small crystals have thus been grown in high-pressure equipment by Fischer5 and by Narita et a1.6 with pressures of the order of 50 atm of argon to prevent excessive evaporation from the melt. Large crystals of ZnTe can be obtained by growth from the vapor phase4 but they often present numerous dislocations and inclusions. An improvement in the quality of vapor- grown ZnTe crystals was reported by Albers and Aten7 by equilibration of mixtures of small crystals with compositions lying on either side of the solid single-phase field at fixed temperature. The same technique was later applied by Aten8 to the growth of ZnTe1-xSex crystals with less than 1 pct inhomo-geneity. Because of the higher liquidus temperatures of the solid solutions and the high vapor pressure of selenium, previous attempts to grow ZnTel-xSex from the melt have been limited and unsuccessful.9 The phase diagram of the Zn- Te system is reproduced in Fig. 1, based on data from Kobayashi10 and Kulwicki.11 Carides and Fishher12 have reported lower liquidus temperatures on the tellurium-rich side, but their data would require confirmation. The liquidus temperature on the tellurium-rich side decreases rapidly with increasing tellurium concentration and the Te2 vapor pressure over the liquidus also decreases accordingly.&apos;3 The decrease in liquidus temperature and vapor pressure therefore makes it possible to use conventional apparatus if there is a sufficient excess of tellurium in the melt. Single crystals of ZnTe have thus been grown by Kucza,14 in a modified Bridgman technique, from solutions containing up to 60 at. pct of Te by lowering unsupported quartz ampoules through a temperature gradient at about 1200°C. Under these conditions, the phase diagram indicates that the entire charge is initially molten. Crystal growth can therefore proceed by normal freezing and rejection of excess tellurium into the melt. The modified Bridgman technique has several major limitations. Because of the rejection of excess tellurium into the melt during freezing, the melt composition and the temperature at the growth interface vary continuously. They tend to follow the liquidus until the eutectic which is very close to pure tellurium (447°C, >99 pct Te). Since the solidus composition also varies with temperature,15 crystals grown by this method are inhomogeneous. They present small variations from stoichiometry which may affect their structure and physical properties. The simultaneous increase in tellurium content and decrease in melt temperature also combine to reduce the rate of diffusion of tellurium away from the growth interface, thereby causing constitutional supercooling and possibly dendritic growth. To minimize these effects, the initial melt composition is in practice kept relatively close to stoichiometry (less than 60 pct Te). This however limits the possibilities of operating at low temperatures and pressures. This paper describes a modified method of crystal growth by temperature gradient solution zoning (TGSZ) which is an adaptation of the temperature gradient zone-melting technique developed by pfann16 and of the traveling solvent method of Mlavsky and weinstein.I7 The TGSZ method now applied to the growth of ZnTe and ZnTel-xSex crystals is characterized by its very simple experimental arrangement and sample preparation technique. Unlike the modified Bridgman technique, there is no increase in the tellurium concentration in the melt and therefore it is possible to operate at lower temperatures and pressures. This method is also suitable for maintaining a constant temperature at the growth interface.

Jan 1, 1969

By J. A. Procarione

Introduction Ventilation engineers have always felt somewhat uneasy when, in the normal course of their work, they are forced to consider the flow in any mine entry as "wholly rough." Under such conditions in pipes, friction factor is no longer a function of Reynolds&apos; number. This assumption serves to simplify the problem and overcome an evident lack of information concerning the variability of K-factors. During the summer of 1985, a unique opportunity to investigate the reaction of friction factor to changing flow rates presented itself. A Utah coal mine was temporarily directing a substantial airflow through a single entry. By regulating the quantity, it would be possible to obtain a wide range of Reynolds&apos; numbers in an entry composed of two distinct coal thicknesses. This paper describes the conducted study. The theory covering basic fluid flows in pipes that led to the questioning of constant K-factor is presented first. Experimental procedures are then described, including the use of a view camera in determining entry area and perimeter. Finally, a discussion of the observed variations in friction factor versus Reynolds&apos; number is given. Changes of 34% and 45% in K-factor were seen in this study for the two entry heights. The implications of ignoring the relationship in this case and a call for an expanded research effort are also presented. Theory Since a gas is a fluid, the general principles of fluid mechanics are equally applicable to airflow in a mine entry as they are to the flow of water in a pipe. The Darcy equation is, therefore, equally applicable to pressure drop calculations in both situations. This famous formula was originally derived for use in circular cross sections and may be modified for any other shape by the use of hydraulic radius (defined as the area divided by the "wetted" perimeter). The Darcy equation relates the physical parameters of the pipe (diameter, length, and resistance to flow) and fluid velocity to the energy needed to create the flow. A study by Nikuradse (1933) brought to light the suspected dependency of the friction factor term in the Darcy equation upon Reynolds&apos; number. By performing a large number of laboratory experiments, he established some basic concepts upon which other researchers built. Chief among these contributors was Colebrook (1939). Using commercial-grade pipe, he established an empirical relationship between Darcy&apos;s friction factor and Reynolds&apos; number that is still in use today. If Colebrook&apos;s system of equations is plotted on log-log paper (as was done by Moody, 1944), the curves have some very distinct features. Of importance to the subject at hand is the relative response of friction factor to Reynolds&apos; number at a given relative roughness. Simply stated, a pipe can appear to be either smooth or rough, depending on the flow conditions. As the Reynolds&apos; number increases, the friction factor decreases at a decreasing rate until it becomes nearly a constant value. At this point, the flow is said to be "wholly rough." It is upon this condition that many traditional ventilation solution methods depend. For calculating pressure drops in a mine entry, Atkinson&apos;s equation is much more popular than the Darcy equation. Both of these formulas describe the same phenomenon and so have similar forms and functions. The K-factor in the Atkinson equation is equivalent to the Darcy friction factor and can be related by simply changing units, equating the pressure drops, and canceling like terms. A generalized form of this relationship is presented below as equation 1. K = yf/28,800g = [q]f/28,800 (1) where K is Atkinson&apos;s friction factor in kg/m3 (lb - min2/ft4), Y is the specific volume in Nt/m3 (lb per cu ft), p is the fluid density in kg/m3 (slugs per cu ft), g is the acceleration due to gravity (9.8 m/s2 or 32.2 fpsz), and f is Darcy&apos;s friction factor, which is unitless. The objective behind Eq. 1 is to show the strong correlation between K and f. Since these two quantities differ only by a constant (q]/28,000), the assumed independence of K-factor and Reynolds&apos; number is suspect in light of the known response of f. Unfortunately, the formulation does not provide information concerning the actual relationship, as entry geometry likely plays a major role in the response curve. Only an indirect indication that such a relationship might exist is provided. Previous work on this subject was performed by Thomas Falkie (1958). The study involved a laboratory model composed of sections of straight, steel duct and a range of Reynolds&apos; numbers from 15,610 to 137,200. Results from this work show a variation in friction factor with Reynolds&apos; number in much the same way as it was found to vary in pipes. Over the range of velocities examined, a decrease of approximately 38% in K-factor was observed. This previous experience provides additional incentives to investigate the phenomenon under field conditions. Past model studies done at the University of Utah by Ozukurt (1945), Biron (1944), and Greenhalgh (1942) present similar evidence of this variability. More recently, Kharkar&apos;s (1974) study of coal mine friction factors contains similar indications that K-factor is not constant in all situations. Taken together, the data

Jan 1, 1988

By R. Lincoln

A surplus of vanadium has resulted from the tremendous output of uranium, of which vanadium is a byproduct. More knowledge of the properties and characteristics of vanadium as a pure metal has been needed so that industry can take advantage of its availability and use it according to the best principles of conservation. Besides the problem of utilization, there is the ever present need of finding new materials which will satisfy specific and demanding requirements of science and engineering; vanadium or one of its alloys may possess the latent potential for filling some of these needs. Thus this? research project was undertaken by the Bureau of Mines to determine and examine some of the physical and mechanical properties of pure vanadium. A corallary investigation was the determination of the effect of the interstitial contaminants, oxygen, nitrogen, and carbon, on unalloyed vanadium. The vanadium used for these determinations was principally the product of experimental production of the Bureau. The evaluation of this material, as small quantities became available, provided most of the data that have been obtained.

Jan 1, 1962

Study on Iron Ore Sintering with High Proportion of Specularite Concentrates Specularite iron ores have poor hydrophilicity and ballability as well as bad high-temperature reactivity, which restricts the large-scale application in the sinter and pellet production. In order to make the best use of the specularite concentrates, a novel process of Composite Agglomeration Process (CAP) is applied in this research. By the CAP, the percentage of specularite concentrates in the sinter mixture is increased to no less than 40 per cent as well as the coke breeze dosage is reduced by 0.3 per cent~0.6 per cent. Moreover, all the sintering indexes are obviously improved compared with the traditional sintering process. Researches on the metallurgical performances shows the finished composite agglomerates are high-quality burdens for blast furnaces. This investigation can provide an effective route to utilize the inferior specularite concentrates in the sintering.

Sep 13, 2010

By D. A. Bruce

To check the feasibility and the quality of a very deep diaphragm wall, a tern of adjacent panels has been constructed, up to a depth of 100 m, in the Milan sandy gravelly soil by means of a hydraulically driven milling machine with reverse circulation of mud. The results of the monitoring carried out in order to evaluate the panel geometry and the quality of both joints and concrete are described. The use of instrumentation based on different physical principles allowed the cross checking of measurements as well as the evaluation of the accuracy and sensitivity of each single procedure to be performed. The trial field was successfully carried out and the feasibility of a diaphragm wall of such a depth was demonstrated and corroborated by a reasonable level of quality control.

Jan 1, 1989

By Kazuo Magata, Hirohiko Fujita, Masazumi Tashiro, Yoshiaki Sato, Yoshio Shinoda

The Shinkansen of the Japanese National Railways (JNR) presently in operation between Tokyo and Hakata, approximately 1,000 km, was constructed at three stages: between Tokyo and Shin-Osaka opening in 1964, Shin-Osaka and Okayama in 1972 and Okayama and Hakata in 1975. Works on the extension of 400 km from Okayama to Hakata were launched in March 1970. As construction of the undersea Shin-Kanmon Tunnel on the extension was the toughest task, requiring many, many months, the work was undertaken earlier than any other works on the extension. It was brought to its completion in June 1974 at the cost of about 22, 000 million yen. The Shin-Kanmon Tunnel is the third undersea tunnel constructed to connect the main island of Honshu and Kyushu across the Kanmon Straits. (Table 1)

Jan 1, 1976

By M. R. Geer

Efforts to evaluate washery performance are nearly as old as the art of coal cleaning, and in former years there was no uniformity in either the methods used or the basic terminology employed. This situation, and the con- fusion it bred, prevailed until about the time of the First International Coal Preparation Conference in 1950, Largely because of the interest of the French coal research organization Cerchar, the subject of evaluating performance was given prime importance at this conference. As a result, there is now a much greater degree of uniformity in terminology and methods of expressing cleaning results. The International Standards Organization currently is working toward international standards of terminology and methods that should improve the situation even more. The various criteria that are used to evaluate the performance of coal- cleaning equipment fall into three categories, depending upon the extent to which they are dependent on the density composition of the raw coal; some are directly dependent, others are only indirectly dependent, and still others are substantially indenpendent.

Jan 1, 1962

By Geo. A. Brown

This paper was written from conditions existing in the Cumberland Mines of The Union Pacific Coal Co., with my impressions and cause of same. These mines are located on a branch line of the 0. S. L. R. R., about sixteen miles south of Kemmerer, Wyoming. The mines are the property of The Union Pacific Coal Company with headquarters at Rock Springs, Wyo. There are two mines, No. 1 and No. 2, located about two miles apart in a seam of coal averaging about 16 feet thick which is split with a parting of rock from 4 feet to G feet from the bottom, averaging from 4 inches to 5 feet thick, mostly of soft fire clay, with an average pitch of 21 degrees to the west. The lower coal below the fire clay band being of a soft nature, much softer than the coal above it. The cover or roof next to the coal is of a massive sandstone formation from 4 feet to 60 feet thick while the floor of the seam is of a soft fire clay as far down as has been prospected. These mines are also of a dry and dusty nature, they generate lots of fire damp and are subject to fires from numerous causes. These mines were opened up in 1900 with five openings each, Main Slope, North and South Manways and North and South Aircourses driven parallel directly on the pitch of the seam which is fairly regular, entries being turned North and South at right angles off the slope about 300 feet apart. The entries were driven double 10 feet wide with chain pillars from 20 feet to 30 feet thick. The main entries are used for haulage and main intake, back entries for return aircourses. Rooms are turned on 50-foot centers and 20 feet wide, directly on the pitch Galvanized sheet iron chutes are used in the rooms. The slopes were driven down in No. 1 mine about 4,000 feet, in No. 2 mine about 5,000 feet where the cover would be about 1,400 feet in No. 1 mine and about 1,700 feet in No. 2 mine approximately. When a point about 2,000 feet on the slope had been reached with cover, 600 or 700 feet squeezing became noticeable on the entries and room pillars with slight bumps which extended clear to the upper entries of the mines and continued to get worse as the mines got deeper and with more cover as no pillars had been removed at this time. It should be stated here that faulty ground was found on the south side of No. 2 mine and on the north side of No. 1 mine, in some places looking like a fault of a down throw, some places like a want and other places like an erosion. The latter theory seems to be correct from prospecting and drilling. The mining of entries and rooms became more difficult with more squeeze and bumps of increased pressure which closed entries and aircourses and finally caused the lower workings of the mine to be abandoned. It was then decided to open up the entries from the 6th up, or approximately half way up the slope. The opening of these entries was started with the intention of extracting pillar coal but considerable danger was met with for the entries .were squeezed tight and when they were disturbed, bumps and squeeze conditions had to be contended with, also gas and fires due to the continued grind on the pillars due to overhead pressure bringing on friction which caused innumerable fires to start in the bottom coal. Bumps being so extensive at times as to close up aircourses for 500 or 600 feet, also discharge timber, coal and rock from the rib of slope and entries for hundreds of feet, dislodging tracks and throwing mine cars around just as if an explosion had taken place, filling places with gas and throwing dust into the atmosphere so that you could cut it with a knife. It became necessary to discontinue the use of black powder and naked lights entirely as a safety precaution. Permissible Powder Monobel No. 5 and Wico Electric Safety Lamps were installed. All coal was shot off the solid whenever it was necessary to shoot. &apos; In opening up these entries it was found that practically all of the weight was centered on room pillars up to the first crosscut chain pillars and barrier pillars between top of rooms of entry below and the one above which averaged 30 feet thick and in all practically 120 feet of ground. The rooms were mostly open from the first crosscut up to the face which showed that the points mentioned had taken the weight. It was found impossible in some of these entries to drive double entries again as the bumps and squeeze made it impossible to maintain them. This made ventilation quite a problem, small booster fans and flexoid tubing being used for ventilation until one entry could be connected with the one above which has solved most of our ventilation problems so far.

Jan 1, 1924

By Francis S. Kendorski

INTRODUCTION Caving offers the lowest cost per ton of any large-scale mining method, but its successful application demands an ore body that conforms to several rigid requirements. The deposit must be of wide areal extent, massive and not spotty in ore values, and insensitive to ore dilution. It must also be a rock mass that breaks up readily. There are only three active caving operations in the US-Climax, Henderson, and San Manuel-but caving methods have recently taken on new importance as deeper lower grade mineral occurrences and ore bodies are found. These deposits are too deep for surface min¬ing methods, and too low grade to support any type of underground mining except a bulk method such as caving. Announced discoveries or indications that may be amenable to caving include: Climax&apos;s Mt. Emmons molybdenum discovery in Colorado; Molycorp&apos;s Goat Hill molybdenum prospect in New Mexico; the Phelps Dodge molybdenum deposit in Beaver County, UT; Arizona copper occurrences such as Asarco&apos;s Sacaton, Hanna&apos;s Casa Grande, Noranda&apos;s Lakeshore, and Ken¬necott&apos;s Safford; Anaconda&apos;s suspected deep copper de¬posit in Butte, MT; Anaconda&apos;s Carr Fork, UT, deposit; and perhaps others. CAVABILITY&apos;S ROLE IN FEASIBILITY STUDIES Caving is a system of underground mining which removes support from underneath an ore body. As a result, the rock mass fractures, fails, and flows vertically downward by gravity to be collected in previously ex¬cavated funnels. Types of ore that have been mined by caving include molybdenum, copper, iron, nickel, as¬bestos, and diamonds (Julin and Tobie, 1973). It is primarily a large-scale method, with production rates of more than 45 300 t/d (50,000 stpd) having been achieved. However, the initial capital investment before return is very high, often in the hundreds of millions of dollars. The cavability of an ore body or mineral occurrence is a critical item in the feasibility study of a proposed mine, not only from the point of overall minability, but from the point of impact on other costs such as blasting, loading, hauling, crushing, and grade recovered. Aside from the often-asked question of, "Will it or will it not cave?" the real questions are, "Can we afford to make it cave, carry the rock away, and extract the mineral?" The last is not a topic of this chapter, but the first two are. The cavability of an ore deposit or mineral occur¬rence is based on many things, but clearly, if a large enough area is undermined, any rock mass will cave. The result could be a violent collapse as occurred at Urad, CO (Kendrick, 1970), or perhaps the rock mass will cave beyond the ore boundary. Another unfavor¬able result could be ore blocks that are too large for the equipment and orepasses to handle without considerable secondary blasting. Weak rock with numerous fractures may produce a very fine ore when it caves, resulting in dilution and ground control problems. DETERMINING STRUCTURAL DOMAINS It has long been recognized that the geologic nature of an ore body is important to cavability (King, 1946). Such items as weak rock material, intensity of fractur¬ing, and severity of faulting all contribute to the success of a caving operation, and information regarding these is required as a minimum for the cavability determination. In practice, the rock mass-defined as the blocks of intact rock together with the intervening fractures, joints, faults, bedding planes, and other discontinuities-that contains the ore body, as well as the surrounding and overlying rock, must be examined in a systematic and detailed fashion. Surficial geology maps must be pre¬pared, exploration holes drilled, and core logged for en¬gineering information. The fracturing of the rock mass must be studied to ascertain the three-dimensional dis¬tribution of fractures and their characteristics, and faults must be located and described. The strength and other mechanical properties of the rock material, the fracture surfaces, and the fault filling materials must be tested and reported for later use by designers and planners. With this basic information and an understanding of the geologic setting, the rock mass can be divided into one or more structural domains which tend to behave similarly in response to engineering activities (Robertson and Piteau, 1970). One must keep in mind that the determination of the structural domains goes beyond the geologic units present. Several lithologic units may be lumped together, while a single lithologic unit can be divided into multiple domains. Major faults often form their own domain, and the direction of engineering ac¬tivity-for example, cave advance to the north rather than the south-may alter rock mass behavior, resulting in different domains. As an example of the detailed rock fracture map¬ping required for such studies, the structural domain determinations at the Climax mine (Kendorski, 1973) are shown in Fig. 1. The circles are Schmidt equal area projections of the three-dimensional attitudes of frac¬tures (Ramsay, 1967) mapped in detail at various rock exposures. The attitude of fractures is important to the cavability determination since it dictates the directional behavior of the rock mass as it fails, and determines the effectiveness of arching, keying, and rock block inter¬locking. Low-angle fractures must be present to allow movement of the rock in the vertical direction during undermining (Mahtab and Dixon, 1976); if low-angle planes of weakness are absent, the rock mass may arch with a keystone effect, rather than moving vertically downward.

Jan 1, 1982

By Ravindra M. Nadkarni, Milton E. Wadsworth

The kinetics of cetnentation of copper with iron were observed to follow first-order kinetics and increase with speed of agitation to a limiting value. Maximum rates agree closely with theoretical values based upon a model of aqueous solution diffusion through a litniting boundary film. Back reaction kinetics are shown both theoretically and experimentally to be independent of ferrous iron concentration in solution. The inlportance of attnospheres of air, oxygen, nitrogen, and hydrogen was studied and the results have been correlated with several impovtant oxidation processes involving metallic iron and copper. The kinetics of the reaction of ferric ion with metallic iron were found to be slow in the absence of metallic copper and essentially proportional to the surface area of metallic copper present in the system. THE precipitation of copper on iron is classic as an example of a relatively ancient art applied successfully for centuries with little fundamental understanding of the important parameters involved. There is some indication that the process has been a commercial means to produce copper since the sixteenth century.&apos; The amount of fundamental work on the cementation of copper with iron is not great. Wartman and Roberson2 carried out a series of detailed copper cementation experiments using natural and synthetic mine water. The following were presented as the three principal reactions: Reaction [I] is the desired cementation reaction and accordingly 0.88 lb of iron would produce 1 lb of copper. In actual practice iron consumption would more normally fall in the range of 1.5 to 2.5 lb per lb of copper. Wartman and Roberson attributed the excess consumption of iron to Reactions [2] and [3]. They found that Reactions [I] and [2] proceeded at approximately the same velocity while Reaction [3] was much slower and would be diminished by controlling the contact time. It was also pointed out that increased agitation is beneficial in removing hydrogen bubbles and barren layers of solution at the iron surface as well as removing contaminants resulting from the hydrolysis of iron. Episkoposyan3 and Episkoposyan and Kakovskii4 studied copper and silver cementation on rotating iron disks in chloride solutions. The kinetics based upon a diffusion model were first order and varied linearly with surface area and with angular velocity raised to the one-half power according to the Levich equation. The experimental activation energy for both copper and silver was approximately 3 kcal per per mole. Excess iron consumption was found to increase with temperature. The rate of cementation first increased with increasing acidity and then diminished at high acid concentrations. sutolov5 has presented an excellent review of the Leach-Precipitation-Flotation (LPF) process including a discussion of copper cementation from an electrochemical point of view although few experimental results were presented. From voltage considerations he predicted that cementation should not be influenced by the concentration of ferrous iron in solution. He considered several secondary reactions including Reactions [2] and [3] and pointed out the importance of oxidation of ferrous iron to ferric with oxygen. In addition it was suggested that Reaction [2] was enhanced by the dissolution of metallic copper by ferric iron which in turn consumed excess iron by the cementation reaction, Eq.[1]. Cementation of copper on metals other than iron has been studied by several investigators but, as in the case of iron, the amount of fundamental work is not extensive. Bashkova and kovalenko6 and Bashkova7 studied the cementation of copper on indium from copper and indium sulfate solutions. The rate was found to be first order and to increase with acidity. This was associated with a decrease in potential (EIn — ECu) and the simultaneous reduction of hydrogen ions at low pH. The rate of cementation also decreased with increasing indium concentrations which was postulated to be due to the decrease in the rate of diffusion of the ions in solution. Below 97°C the experimental activation energy was found to have the unusually low value of 2 kcal per mole and was attributed to diffusional control. Above 97°C the rate increased suddenly and was explained as a change in the rate-controlling step to a chemical reaction. In Part I of this study Nadkarni et a1 .1 have reported on preliminary results obtained in a laboratory study of the kinetics of the cementation process. The rate was found to be first order, proportional to the surface area of the iron, and to increase with speed of stirring until a maximum rate was observed. At low stirring speeds the deposit was spongy and adherent. At medium speeds the copper peeled off in bright strips and at high speeds finely divided copper was produced and continually removed from the surface. The amount of excess iron consumed increased with speed of stirring and with temperature. The average experimental activation energy combining results from several types of iron was 5.8 + 1.6 kcal per mole suggesting diffusional control through a limiting boundary film. Traditionally copper cementation has been carried out over the centuries in gravity-fed launders of various design containing scrap iron. More recently rotating drum precipitators and activated launders8&apos;10 have been used. In the latter, copper-bearing solutions are

Jan 1, 1968

By Zephaniah Varley, Dominic Parmantier, Paul Schmall, Richard Martin, Red Robinson

Sequential excavation of the Beacon Hill Station has been accomplished in highly complex and inter-layered sand, silt, and clay soils with multiple perched groundwater tables. Successful excavation has required the installation of a deep dewatering well system used to depressurize sand layers above and within the excavated tunnel faces. In addition, the presence of sand layers in the crown of the Platform Tunnels required jet grouting at depths of up to 49m (160-ft). This paper discusses the design intent, construction, and effectiveness of the dewatering system and the selective jet grouting of sand horizons and shaft break outs for the successful construction of the deepest SEM station excavation ever completed in the United States.

Jan 1, 2007

By Cheryl Lee Vatter

The Lupin mine consists of a gold mining, milling, and refining complex located 400 km (250 miles) northeast of Yellowknife and 100 km (60 miles) south of the Arctic Circle on the southwest shore of Contwoyto Lake in Canada&apos;s Northwest Territories. (Fig.1). The mine was commissioned in 1982 and is presently producing 6 t/a (193,000 oz per year) of gold from 612 kt (675,000 st) of ore. The success of this operation is due to such factors as the continuity and grade of the ore body, the competency of the host rock, low mining costs, efficient milling, the transportation of people and materials to and from the mine site, and the unique work schedule resulting in a stable workforce. Echo Bay Mines operated a silver mine at Port Radium on Great Bear Lake, Northwest Territories, from 1965 to 1982. The company first obtained an option on the Lupin property from Inco in February 1979 and completed an underground exploration program in 1979-1980. Topographic relief is low and vegetation is sparse in the continental subarctic climatic zone, consisting mainly of moss and lichens. Temperature extremes are from 24°C to - 45°C (+75°F to -49°F) with an annual mean of -12°C (+54°F). Permafrost extends from near surface to 500 m (1640 ft) in the ore zone. Its remote location and harsh climate presented some challenging design and logistical problems. Construction began at the Lupin mine site in 1980. Before construction, a 1.5-km (5000-ft) gravel landing strip was prepared, suitable for landing a C 130 Hercules. The entire Lupin project took 20 months to build. All of the men and materials were transported to the site with some 1100 Hercules flights and several hundred Convair 640 flights. The Convairs transported construction crews, which numbered 400 at their peak, and also carried 3.2 kt (3535 st) of supplies during construction. The facilities were constructed and commissioned for a total cost of C$135 million. The operation was originally designed to throughput 860 t/d (950 stpd). Expansion in 1983, circuit refinements, and some capital projects brought the daily throughput to 1.7 kt/d (1850 stpd). The underground mine delivers 612 kt/a (675,000 stpy) of ore to the mill at an average head grade of 8.46 g/t (0.3 oz per st). The ore is nonrefractory and is processed in a conventional cyanide leach using the Merrill Crowe process. Gold recovery is about 95.0%. The average production cost is $US5.85/g ($US 182 per oz) based on 1987 figures. Mining at Lupin – Geology The Lupin deposit occurs in amphibolite grade iron formation overlain by mudstones (phyllites) and underlain by graywacke (quartzites). Contacts between the wallrock units and the iron formation are well defined. It has been folded and tilted into a megascopic antiform-synform-antiform structure (Fig. 2). Gold occurs primarily within the sulfide rich iron formation, with some minor occurrences in sulfide poor iron formation. The distinction between sulfide rich and sulfide poor iron formation is based on a visual cutoff of 5% total sulfide content. Mining widths are determined by an assay cutoff of 4.2 g/t (0.15 oz per st) gold. There are few tons between 1.7 and 4.2 g/t (0.06 and 0.15 oz per st). The amphibolitic iron formation at the mine ranges from 1.5 to 20 m (5 to 65 ft) wide and has been followed over a strike length exceeding 1.7 km (5600 ft). The wider portions of the ore body tend to occur at its north extent and south nose. The gold-bearing iron formation appears on plan as a Z-shaped structure made up of three zones: the West, Center, and East. The West and Center zones dip steeply to the East (75° to 90°). Each of the ore zones plunge at an angle of about 65°. Total strike length of the three zones is more than 610 m (2000 ft). The zones are confirmed at a depth of 650 m (2130 ft) below surface. The Center zone is the widest and varies from 4.5 to 20 m (15 to 65 ft) while the West is the narrowest, averaging 1.5 m ( 5 ft). The footwall is comprised of quartzites that are strongly jointed and locally grades into phyllite, which comprises the hanging wall. The hanging wall and footwall are reversed in the West zone. Mineralogy of the ore at Lupin consists of amphibole minerals (hornblende, cummingtonite, and grunerite), feldspars, quartz, occasionally garnet, pyrrhotite, arsenopyrite, minor pyrite, and trace chalcopyrite. Also found in minor amounts are scheelite, apatite, epidote, calcite, tourmaline, and some arsenides (notable loellingite). Quartz

Jan 1, 1989

By Hugh H. Waesche

Of all the military services, the Signal Corps is the most concerned with piezoelectric minerals because of its function as a supply service to the strategic and tactical military forces. Consequently this paper is written from the point of view of one associated with that organization. The Signal Corps is responsible for the research, development, and supply of communications, radar, and components to the using services of the Department of the Army and to some extent the Other branches of the National Defense Department. Their work therefore includes the study of the sources* characteristics, and application of quartz and other piezoelectric materials. These materials have become a vital consideration in strategic planning and are essential for efficient tactical operation by all the Armed Forces. The Signal Corps at the beginning of world War 11 Was respon-sible for both Army Ground and Air Force electronic equipment. Since that time this Army service organization has probably done more in the development of frequency control devices using piezoelectric materials than any other group. The U.S. Department of the Interior, Bureau of Mines, Minerals yearbook of 1945, shows that during the four war years, 1942 through 1945, 9,598,-410 Ib of quartz crystal were imported for all uses and of this total, 5,168,000 lb were consumed to produce 78,320,-000 crystal units for electronic application. Other government records confirm these data which conclusively show that approximately 53 pct of the crystalline quartz imported was consumed in the production of electronically applied quartz crystal units. It may be assumed that some effort was made to maintain a stockpile over demands for all purposes. and this would mean that the actual percentage of quartz used electronically was considerably over the 53 pct figure. These data only emphasize that electronic application of crystalline quartz was the greatest requirement, and per- haps the actual value in this application to national defense is many times greater in importance than is apparent on first inspection. Current electronic research and development programs of the Armed Forces are planned around the fundamental use of piezoelectric minerals for frequency control and this at present, at least, means quartz. Definition and Early Development The word piezoelectricity is formed from combination of the Greek word "piezein". meaning "to press," and "electricity." It is that property shown by numerous crystalline substances whereby electrical charges of equal and opposite value are produced on certain surfaces when the crystal is subjected to mechanical stress. It appears to be intimately associated with the better known property, pyro-electricity and in fact, the two may be manifestations of the same phenomeuon. This property was discovered by Pierre and Jacques Curie in quartz, tourmaline, and other minerals in 1880 while studying the symmetry of crystals. The converse effect, that is, mechanical strain in the crystal when placed in an electrical field, was predicted by the French physicist, G. Lippman, in 1881, and verified by the Curies almost immediately. As has been the case with many discoveries of similar character in the basic sciences, not much attention was paid to this property for man)- years except as an entertaining curiosity. Between 1890 and 1892 a series of papers was published by W. voigt in which the theoretical physical properties were put into mathematical form. The first practical application of piezoelectricity occurred during World War I when professor P. Langevin of France used quartz mosaics to produce underwater sound waves. The same mosaics were used to pick up the sound reflections from submerged objects which were in turn, amplified by electronic means and used to determine the distances to such objects. This device was intended for use as a submarine detector but development was not completed in time for war service although it was used later for determining ocean depths. About the same time, A. M. Nicholson, of Bell Telephone Laboratories, developed microphones and phonograph pickups using Rochelle salt crystals. A major step in the application of piezoelectric quartz came in 1921, when professor W. G. Cady, of wesleyan university, showed that a radio oscillator could be controlled by a quartz crystal; from that date, this use of quartz has increased steadily, reaching its peak in world war 11 as is shown by the figures previously given. Essentially all American electronic equipment for communication, navigation, and radar, utilized quartz crystals for the exacting frequency control required. Crystalline Minerals with piezoelectric Properties QUARTZ Hundreds of piezoelectric crystalline materials are known, most of which are water soluble. Of these, quartz appears to be without a peer for electronic frequency control. Unfortunately, the quartz must be of superior quality. It must be free of mechanical flaws, essen-tially optically clear, free of both Brazil and Dauphiné twinning and must be, for average uses, over 100 g in weight. Because of these stringent requirements, raw quartz of the quality desired is of rare occurrence. In addition to quartz, several other naturally occurring crystalline materials are known to have the piezoelectric property and could perhaps be substituted for quartz in some applications. These

Jan 1, 1950

By Christopher M. St. John, Michael P. Hardy

INTRODUCTION Geotechnical models, particularly those based on the finite element method, have been available to aid in en¬gineering design of underground mining excavations for over ten years. Despite this fact there are remarkably few cases of their use in mine design documented in the literature. It is therefore to be anticipated that many potential users of these models are relatively unaware of their capabilities and limitations and also of the form and detail of geotechnical data needed for their success¬ful application. This subsection attempts to address this problem by discussing the different types of numerical models now available and by noting how some of them have been used to study a variety of problems associated with underground mining. The subsection concludes by discussing the applica¬tion of the displacement discontinuity method to the de¬sign of possible mining systems for the copper-nickel deposits of northeastern Minnesota. The object of the analyses, which were nonsite specific, was to determine the significance of geotechnical parameters, such as ini¬tial stress and rock structure, on the stability of under¬ground excavations and hence to provide guidance for future geotechnical investigation. NUMERICAL MODELS In the geomechanical design of underground mining openings, use is made of numerical models that repre¬sent or simulate the large-scale mechanical behavior of rock. There has been less interest in analysis involving fluid flow and heat transfer, but with increasing interest in such areas as in-situ retorting and solution mining it is likely that there will be a growing need for numerical models embracing and coupling all three physical proc¬esses. However, the emphasis in this subsection will be on mechanical behavior. Models which simulate such behavior will be divided into two groups: continuum models and discontinuum models. These will be dis¬cussed in turn in order that some insight into alternative solution strategies and their merits may be gained. Continuum Models Almost all geomechanical numerical models must be classed as continuum models even though particular computer codes incorporate special provisions for rep¬resenting discontinuities such as faults, bedding planes, or joints. They are continuum models because they pro¬vide solutions for cases where material behavior is governed by the differential equations of continuum mechanics. Two basic solution strategies for such equa¬tions may be identified immediately: the differential ap¬proach and the integral approach. In the differential approach a means of approximating the differential equations over the entire region of interest is sought. In the integral approach use is made of fundamental solutions from continuum mechanics, and these are used to construct a solution to the whole problem, making approximations only on the boundaries of the region of interest. The several differential and integral methods are identified in Table 1. Differential Methods: Problems in continuum me¬chanics involve the solution of three types of partial differential equations. Two of these govern the behavior in so-called initial value problems, in which variables change both in time and in space. Examples of such problems include nonsteady heat transfer and fluid flow, and stress wave propagation. The last type of partial differential equation governs the behavior in boundary value problems. In these, variation is in space but not in time. Solution of initial value problems may be achieved in two significantly different methods: implicit and ex¬plicit. The differences between these two methods will be illustrated by considering a very simple initial value problem, that of one-dimensional heat diffusion. The equation governing this process may be written as: [ ] where T is temperature, K is thermal conductivity, p is density, c is specific heat, t is time, and x is the spacial coordinate. In finite difference form this equation might be written as: [ ] where the superscripts refer to the time and the sub¬scripts to the spacial location. Several solution strategies for this equation have been used. Two important ones may be illustrated very simply by discussing the signifi¬cance of the superscript * on the right-hand side of the equation. If i + 1 is substituted for * then the second derivative of the temperature with respect to distance is evaluated at the end of the next time step (using tem¬peratures not already known). Such an approach leads to a set of equations involving unknown temperatures and a solution procedure which is referred to as being implicit. The important characteristic of the implicit procedure is that it leads to a set of equations that must be solved for each time at which the temperature dis¬tribution is required. If instead of substituting (i + 1) for the superscript *, i is substituted, the following equation is obtained: [ ] {In this case the new temperature is defined in terms of an already known temperature distribution. The solution procedure is now known as explicit and has the impor¬tant characteristic that there are no equations that have to be stored or solved. A practical advantage of this

Jan 1, 1982

By J. J. Tietien, R. O. Clough

A technique previously used to prepare alloys of InAs1-xPx and GaAsl-x Px, miry: the gaseous hydrides arsine and phosphine, has been extended to grow single -crystalline GaAs 1-x Sb x by replacing the phos-phine with stibine. Procedures were developed for handling and storing stibine which now make this chemical useful for vapor phase growth. This represents the first time that this series of alloys has been grown from the vapor phase. Layers of P -type GaSb and GaSb-rich alloys have been grown with the carrier concentrations comparable to the lowest ever reported. In addition, a p-type alloy containing 4 pct GaSb exhibited a mobility of 400 sq cm per v-sec which is equivalent to the highest reported for GaAs. RECENTLY, interest has been shown in the preparation and properties of GaAs1-xSbx alloys, since it was predicted1 that for compositions in the range of 0.1 < x < 0.5, they might provide improved Gunn devices. However, preparation of these alloys presents fundamental difficulties. In the case of liquid phase growth, the large concentration difference between the liquidus and solidus in the phase diagram, at any given temperature, introduces constitutional supercooling problems. It is likely that, for this reason, virtually no description of the preparation of GaAs1-xSbx by this technique has been reported. In the case of vapor phase growth, problems are presented by the low vapor pressure of antimony, and the low melting point of GaSb and many of these alloys. In previous attempts1 at the vapor phase growth of these materials, using antimony pentachloride as the source of antimony vapor, alloy compositions were limited to those containing less than about 2 pct GaSb. This was in part due to the difficulty of avoiding condensation of antimony on introducing it to the growth zone. A growth technique has recently been described2 for the preparation of III-V compounds in which the hydrides of arsenic and phosphorous (AsH3 and pH3) are used as the source of the group V element. With this method, GaAs1-xPx and InAs1-xPx have been prepared2&apos;3 across both alloy series with very good electrical properties. Since the use of stibine (SbH3) affords the potential for effective introduction of antimony to the growth apparatus, in analogy with the other group V hydrides, this growth method has been explored for the preparation of GaAs1-xSbx alloys. In addition to GaSb, these alloys have now been prepared with values of x as high as 0.8. In the case of GaSb, undoped p-type layers were grown with carrier concentrations equivalent to the lowest reported in the literature. Thus it has been demonstrated that, with this growth technique, all of the alloys in this series can be prepared. EXPERIMENTAL PROCEDURE A) Growth Technique. The growth apparatus, shown schematically in Fig. 1, and procedure are virtually identical to that described2 for the growth of GaAs1-xPx alloys, with the exception that phosphine is replaced by stibine.* HCl is introduced over the gallium boat to *Purchased from Matheson Co., E. Rutherford,N+J. transport the gallium predominantly via its subchlo-ride to the reaction zone, where it reacts with arsenic and antimony on the substrate surface to form an alloy layer. The fundamental limiting factors to the growth of GaAs1-xSbx alloys from the vapor phase, especially GaSb-rich alloys, are the low melting point of GaSb (712°C) and the low vapor pressure of antimony at this temperature (<l mm). Thus, relatively low antimony pressures must be employed, which, however, imply low growth rates. To provide low antimony pressures, very dilute concentrations of arsine and stibine in a hydrogen carrier gas were used. Typical flow rates (referred to stp) were about 4 cm3 per min of HC1 (0.06 mole pct)+ from 0.1 to 1 cm3 per min of ASH, (0.002 to 0.02 mole pct), and from 1 to 10 cn13 per min of SbH3 (0.02 to 0.2 mole pct), with a total hydrogen carrier gas flow rate of about 6000 cm3 per min. Although no precise data on decomposition. kinetics exist, it is known4 that stibine decomposes extremely rapidly at elevated temperatures. However, the high linear velocities attendent with the high total flow rate (about 2000 cm per sec) delays cracking of the stibine until it reaches the reaction zone and prevents condensation of antimony in the system. To improve the growth rates of the GaSb-rich alloys, growth temperatures just below the alloy solidus are main-

Jan 1, 1970

By V. S. Gupt, W. H. Somerton

Rock may undergo great changes in physical properties when heated to high temperatures and then cooled, The temperature and intensity of reactions causing rock alterntiorl.s can he controlled by introducing certain chemicals during heat treament. Three typical outcrop sandsone samples were saturuted with common salt solutions, then heated to several maximum temperatures. After cooling, it was found that per-rrreuhilities had increased much more for salt-saturated samples than for samples not saturated with salt but heated to the same temperatures. This was only true, however. for samples heated above the melting points of the particrrlar salts. Potassium chloritle was particularly effective with Bandera sandstone. Samples saturated with potassium chlorirlr. solution and heated to 900C showed an 11-fold increa.se in perrrieahility. Samples without potasrirdrn chloritle but heated to the .same temperature. showed only a 2.7-jold incresre in permecahility. In application, it seems possible that injecting chemicalc into the formations from a wellbore followed by applying intensive borehole heating might promote reactions which would greatly improve permeability of the formalions. INTRODUCTION Great changes in the physical properties of rocks which have been heated to temperatures in the range of 600 to 800C have been reported earlier. Permeability increases of 50 per cent, and equivalent decreases in sonic velocity and breaking strength have been observed. Although there might be some suppression of certain reactions responsiblc for these changes when rock samples are heated under simulated reservoir pressure conditions. recent work has shown that these are more than offset by the increased importance of other reactions at high pressures. The reactions considered responsible for alteration of rock properties by heating include differential thermal expansion. dehydration, phase changes and dissociation of mineral constituents. Ceramists have long known that temperatures at which reactions occur can be raised or lowered by the presence of certain impurities in the sys. tcnl. Thus the possibility exists that by saturating rocks with appropriate solutions. desired reactions might be accelerated and unwanted reactions might be suppressec! when the rock is heated. Purpose of the present work was to investigate the effects of several common salts on the thermal alteration of sandstones. Types of reactions which might occur are reviewed and the findings of a number of thernlochemical alteration tests are presented. THEKMOCHEMICAL ALTERATION OF SANDSTONE MINERALS The most abundant constituent of most sandstones is quartz. Quartz can exist in at least eight different forms, but in the temperature range of immediate interest (to 900C) only two forms are of importance—alpha quartz below 573C and beta quartz above this temperature. Although the alpha-beta inversion of quartz is very important in thermal alteration of rocks. it is not particularly important here because the inversion temperature cannot be changed significantly by the presence of impurities or by the application of pressure. The transition temperature of quartz to tridynlite is 867C. but the transition is very sluggish. Tridymite is considered to be a metastable transition phase between two stable phases: quartz?cristobalite. However. the transition temperature for tridymite to cristobalite is 1,470C.4 The quartz-tridymite-cristobalite transition can be accelerated substantially by the presence of fluxing agents. Finely divided calcium, magnesium and titanium oxides accelerate the conversion while A12O may retard it. A Mixture of NaCl and CsCl reduces the temperature for cristobalite formation by as much as 420°C. Feldspars constitute the second most important mineral in sandstones. The melting points of feldspars range between 1,000 and 1,700°C depending upon the variety. No information available indicated any important role of fluxing agents in the thermal alteration of this group of minerals. Carbonate minerals are subject to dissociation at temperatures within the range of the present investigation. The dissociation of magnesite can start at a temperature of 373°C, but the reaction is sluggish and might not occur untll a temperature of 500°C or ,higher is reached. Dolomite dissociates in two stages at 500°C and 890°C, whereas calcite dissociation temperature is about 885C. Because CO2 is released in the dissociation of carbonates, ail such reactions are somewhat dependent upon CO2 partial pressure. In the absence of CO: in the surrounding atmosphere, the dissociation of calcite starts at 500°C.However. when 1 atm of CO surrounds the sample, the dissociation does not start until 900C. The other carbonates are apparently much less sensitive to a change in partial pressure of CO2. The aragonite calcite transformation can occur an!. where within the temperature range of 357 to 488°C. depending upon the presence of impurities such as barium. strontium. lead and perhaps zinc. The differential thermal analysis (DTA) curves of both magnesitc and dolomite vary with the presence of &apos;impuritics. The presence of iron in particular seems to affect the magnesitc DTA curve

Jan 1, 1966