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By R. A. Wattenbarger, H. J. Ramey

The trend in gas well testing has been to rely more on the early-time flow data of drawdown and buildup tests than on stabilized flow tests. The stabilized testing methods often are not adequate for complicated modern gas reservoirs because many of these reservoirs have extremely low permeabilities and the transient flow period of a well test can last for days or weeks. It has become common to augment the flow capacity of gas wells by hydraulically fracturing their producing formations. In deep reservoirs the induced fractures are generally vertical and tend to follow a single plane of weakness. The presence of a vertical fracture at the wellbore complicates the transient flow behavior of a low permeability gas well. The flow is further complicated when turbulence occurs near the wellbore. Russell and Truittl published transient drawdown solutions for vertically-fractured liquid wells. They developed methods of drawdown and buildup testing utilizing these solutions, which were based on numerical simulation. Clark2 applied the basic Russell-Truitt solutions to analyze fractured water injection wells by falloff tests. Field examples were given to substantiate the method of analysis. The Russell-Truitt solutions and well testing methods can be extended to gas well flow. Millheim and Cichowicz3 presented the drawdown equations for *Now with Scientific Software Corp., Denver. ideal gas flow, including the effects of formation damage and turbulence. Two actual field cases were presented, both of which exhibited extensive fractures and turbulent flow. For the pressure range of these well tests, the use of ideal gas equations was acceptable. These field cases were significant in being the first published data showing the occurrence of turbulence in fractured gas wells. We, also, observed such cases in practice. Our purpose here is to extend the theory of fractured gas well testing to the flow of real gases. To determine the effects of wellbore storage and turbulence on well test interpretation, we developed a finite-difference model to simulate well test conditions. Since gas wells usually are widely spaced and have high compressibility, the emphasis has been put on the early transient behavior before the effects of the outer boundary are noticeable at the wellbore. The Mathematical Model The geometry of our mathematical model is similar to that used by Prats.4 The well is centered in a circular uniform reservoir. A vertical fracture of infinite flow capacity penetrates the formation and passes through the wellbore. The wellbore itself is not important. Fig. 1 shows a sketch of the problem. Fig. 2 shows the idealized model. The equation for the flow of real gas in a porous medium with uniform porosity, permeability and thickness is5-7

Jan 1, 1970

By C. C. Miller, A. R. Belyea, C. H. Herty, E. B. Burkart

The removal of sulfur from steel has been studied by many investigators, but the quantitative relationships between the factors involved have not been determined. This is undoubtedly due to the number of variables encountered in practice, an attempt to study any one reaction or relation being affected by other reactions the relative importance of which cannot be estimated. The factors that affect the extent of removal of sulfur in the basic open-hearth process are generally considered to be: 1. Basicity of the slag: (a) Iron oxides in the slag; (b) carbon in the metal or carbide in the slag; (c) temperature. These factors affect the reactions: FeS + CaO = CaS + FeO FeS + CaO + C = CaS + CO + Fe. 2. Amount of sulfur in the gas. 3. Amount of manganese in the metal. The factors which affect the rate of removal of sulfur include, in addition to these: 4. Fluidity of the slag: (a) Normal changes in fluidity with temperature; (b) changes in fluidity because of the addition of special reagents such as fluorspar. 5. Agitation of the bath. Within the past two years, two papers have been presented dealing with the influence of sulfur in the gas on slag and metal. Nead' has shown that if a gas high in sulfur be used for heating, the metal and slag will absorb sulfur. Whitely2 concluded from his experiments that if a

Jan 1, 1925

By A. Thielen

In experiments for the production of steel the principal problem has always been the introduction into, or removal from, the iron of a definite quantity of carbon. Although the solution of this problem had engaged a more or less blindly and empirically directed effort for a long time, it was only with the development of analytical chemistry that the rôle which carbon plays in the constitution of steel was disclosed in its full significance, and that many of the problems confronting the steel-maker were thereby more clearly defined than before. For a long time, and to a certain extent even to this day, the percentage of carbon in steel has been the controlling means of deter mining its quality and hardness. In many cases the percentage of carbon alone has been relied on for judging large consignments, such as the large deliveries of wire-billets and blooms to America during the past twenty years; and though to-day the influences of the other constituents are recognized and many other considerations are taken into account, and though, moreover, some special steels, such as the Hadfield manganese steel, apparently dispense to a certain extent with the assistance of the carbon, yet fix our Bessemer, Thomas, Siemens and crucible steel the carbon still remains one of the most important constituents, and I venture to say it is indispensable. In order to produce in the Bessemer converter a steel of a certain required percentage of carbon, two methods, as is well known, present themselves. By the one the decarburization process is stopped when the desired carbon limit in the bath has been reached ; by the other the bath is completely decarburizecl, and then a calculated amount of carbon added in the form of spiegel, ferro manganese, etc. In Germany the latter method only was adopted, and it was retained when the Thomas-Gilchrist process for the production of low-Phosphorus steel from high-phosphorus iron eame into use, since from the nature of the new process complete decarburization is neces-

Jan 1, 1891

By J. A. Berger, R. J. Towner

The pjPocesses of polygonization and subgrain growth were followed in slightly stretched 99.99 pct A1 single crystals heated for cumulative times at 400°, 500°, and 600°C. Changes in sub-grain size and boundary angle were determined from measurements on X-ray micrographs. CAHN1, 2 studied polygonization in stretched and heated aluminum single crystals of 99.99 pct A1 using various techniques, including etch pits, Laue patterns, and a slight restretching of heated specimens to reveal small changes in lattice orientation by the direction of new slip lines. He found that kink bands, which had well-defined boundaries lying in a plane perpendicular to the slip direction (at least in the early stages of deformation), formed during stretching. Also, he noted that bending occurred about the <211> axis lying in the (111) slip plane perpendicular to the <110> slip direction. His studies of polygonization were largely confined to the kink bands rather than to the weakly bent matrix regions between bands. The minimum temperature required to bring about polygonization was established at about 450 "C. Guinier and Tenevin used a focusing Laue X-ray diffraction technique to study polygonization in aluminum single-crystal specimens of 99.99 pct purity. Stretching the crystals 5 pct in tension caused a widening of the focused Laue spots. After heating for 4 hr at 450°C, their method revealed no change in the specimen. Above 450°C polygonization was detected and further annealing near 600°C led to subgrain growth. They concluded that the orientation of the crystal relative to the axis of deformation had no marked effect on polygonization. Gay and Honey-combe4 using an X-ray microbeam technique found that polygonization did not occur in a stretched aluminum crystal heated at 250oC, although some decrease in the yield strength occurred. When the temperature was raised to 475oC, the asterism became broken up as a result of polygonization. Honeycombe5 noted that reflections on X-ray mi-

Jan 1, 1961

By W. C. Lacy

This article is a progress report of a program designed to quantitatively evaluate geological parameters to predict stable slope angles. The author feels that, although there does not appear to be any one technique or characteristics that will suffice as an index to the determination of stable slope angles, a combination of seismic methods as outlined here, detailed geological mapping and logging of core, and a better understanding of the angle of internal friction in rock types will permit accurate estimation. A mining engineer friend once remarked, "If we don&apos;t have pit slope failures, the slopes are too flat." Since economic justification of an open pit project is based upon ultimate pit slopes, we cannot wait until the pit is developed and slopes established by trial and error. Some combination of geological and/or geophysical techniques must be developed to permit accurate prediction of stable slope angles and to suggest modification of pit outlines to result in slopes enabling maximum economic yield. This presentation is a progress report of a program designed to quantitatively evaluate geological parameters to predict stable slope angles. The principle of the weakest link applies to any slope stability evaluation and necessitates detailed geological examination and recording of observations. Physical failure will occur along any surface where shearing stress exceeds shearing resistance. Rock conditions are seldom uniform even within a single cut or pit face, so stability conditions will vary. Consequently, both plan outline and ultimate slope limits must be tailored to the specific situation. Economic advantages to be gained by slight increases in slope angles in deep pits are well recognized and stimulate efforts to obtain maximum permissible slope angles. MECHANICS OF SLOPE FAILURE Excellent reviews on the nature of slope failure have been presented+2,6,8,11 For the purpose of this study, the writer considers failure to take place by one or more of three processes: 1) ravelling, 2) translation or slope failure, and 3) rotation or base failure (Fig. 1). The effects of increased pore water pressure in soils were recognized in 1923 by Terzaghi in his development of the concept of effective stress, and he has applied the same principle to jointed rocks designating water pressure as cleft-water pressure.8 Cleft-water pressure in a rock slope reduces normal pressure on the plane of failure, thus reducing shearing resistance and resulting in slope failure. In the following discussion it is assumed that adequate drainage is supplied to prevent any build-up of water pressure. This is the first step in any program to establish maximum slopes. Ravelling: Ravelling occurs in rocks with low cohesion and entails the free movement of individual fragments ranging from single rock falls to running ground. Removal of lateral support wherever slope angle exceeds the angle of repose of the rock parti-

Jan 1, 1963

By H. S. COOLEY

TO talk about a "factor of economy" in connection with the wire ropes used in mining practice may be coining a new phrase. If such be the case it needs no other apology than that economy in wire rope usage is so bound up with the factor of safety as to justify a similarity in naming that makes one suggest the other. At the latest Wire Rope Research Conference, held under the auspices of the A. S. M. E., Re- search Committee and Engineering Foundation, data were introduced covering mine ropes working under

Jan 1, 1930

By John D. Sullivan

A DISCUSSION of the developments of the past decade in the field of refractories, and the effect of these developments on the performance and life of open-hearth furnaces, is perhaps best introduced by reference to Fig. 1. This shows the most important types of refractories doing into various parts of the furnace, together with their cost, and remarks on their use. It has been estimated that half of the refractory makers market comes from the iron and steel industry, and something like one-third of all refractory brick used finds its way into open-hearth furnaces. To melt steel, a furnace capable of withstanding temperatures close to 3000 deg. F. is required. The furnace is assembled of various refractory materials selected by experience and economic considerations. Furnace structures today, from a refractory standpoint, are considerably different from a quarter century ago. Furnace operators, interested in the performance of refractories, care little how they are made or where they come from so long as they give reasonable service. But, the entire history of the refractory is important and technical knowledge of the ceramic materials used in making the refractory is necessary to the manufacturers, so that the proper article gets into the hands of the ultimate consumer.

Jan 1, 1936

By Earl L. H. Sackett

Barite, although not a glamour mineral and probably little known to many of those in the mining business, is produced in the US. in very respectable quantities and is an important factor in the mineral economy of this country. The scope of the various uses of this mineral is perhaps one of the most misunderstood things about it, for it is surprising how many of those with a casual knowledge of the subject speak of its primary utilization in terms of paint pigments and barium chemicals. The life blood of barite utilization, however, is the petroleum industry by virtue of its dependence on this high specific gravity, inert mineral as the weighting material in its drilling fluids. From 80 to 85% of barite usage in the U. S. is accounted for in this manner-in fact, its role in oil well drilling is the only excuse for some of the large-scale barite production facilities which have been developed. Other uses are important and widely varied, but comparatively they are of small scale. They include fine crushed barite of low iron content as a component in glass manufacture, barite as a source material for manufacture of barium chemicals, finely divided barite as a component of some rubbers and ground, bleached barite in pigments. The formerly substantial use of barite in lithopone, however, has declined markedly because of replacement of lithopone by titanium pigments

Jan 5, 1962

By H. J. Modi, D. W. Fuerstenau

pH effectively regulates the flotation of corundum through its control of the surface charge. Since collector ions function as counter ions in the double layer, the collector must be anionic when corundum is positively charged and cationic when it is negatively charged. Even a long-chained xanthate will float positively charged corundum. Association of the hydrocarbon chains of the adsorbed collector ions gives rise to increased adsorption and increased rate of flotation. Because of this phenomenon, neutral organic molecules aid flotation although the unionized collector cannot promote flotation independently. Na2SO4 was found to depress the flotation of corundum with animic collectors through ionic competition whereas the same salt activates corundum for flotation with cationic collectors. A close connection between ion association, electrokinetic behavior, and floatability is suggested. As part of the increasing interest in the physical chemistry of flotation, strong experimental evidence has been obtained which indicates that organic ions are adsorbed as counter ions by non-metallic oxide minerals mainly because of the electrical forces which exist at interfaces(1-3). Thus flotation of such minerals without activation should then be appreciable only when the mineral and the collector are oppositely charged. Since only limited work has been reported on the effect of surface charge upon flotation behavior(3) this general hypothesis on the flotation of insoluble oxide minerals has not been tested thoroughly. Investigating the flotation of a mineral for which extensive electrokinetic data exist yields a way to test this hypothesis more extensively. The electrokinetic behavior of corundum in solutions containing both organic and inorganic electrolytes has been studied quite thoroughly&apos; Electro-kinetic measurements have shown that in the case of corundum, simply by controlling pH the solid surface can be charged either positively or negatively. For this reason corundum was chosen for study. The object of this paper is to present a study of the flotation of corundum and to explain the results in the light of our knowledge of adsorption mechanisms gained from streaming potential investigations. This paper is concerned mainly with the effect of surface charge on flotation with anionic and cationic collectors, but in addition such flotation phenomena as collector hydrolysis, Van der Waals attraction between hydrocarbon chains, ionic competition, and activation are discussed. While the present paper is confined to a reporting of the data obtained and some of the more immediate conclusions, it will be seen that the results, in combination with those on other systems, do promise a

Jan 1, 1961

By J. C. Detweiler

THE Jacksonville plant of Humphreys Gold Corp., operating under leases from National Lead Co. and Rutile Mining Co. of Florida, recovers rutile, ilmenite, zircon, and monazite from an ancient beach deposit containing about 2.5 pct of these minerals in combination. The location of the plant is unique, being situated about ten miles from the center of Jacksonville, Fla., surrounded by a thriving suburb. Two four-lane highways pass within 1 ½ miles of the plant. An expressway now under construction will pass over the tailing pile, within 500 ft of the mine office. The elevation at the plant site is approximately 50 ft above sea level. This is one of two mines producing titanium minerals and zircon in Florida. Another mine is also operated by Humphreys at Starke, about 50 miles southeast of Jacksonville. The Jacksonville deposit extends about six miles in a north-south direction, averaging approximately one half mile wide and twenty ft thick. Values begin at the surface over most of its length and ter-

Jan 1, 1952

By Lucien Eaton

THE use of filling in mines is less common in the United States than it is in Europe, where in some places it is required by law. In most cases the filling is placed by hand, and the material used for this purpose is the waste from development and that sorted from the ore in the stopes. When these sources provide an insufficient supply, additional rock must be quarried on surface or mined underground, or some other material :st be used. In different parts of the world sand, clay, gravel, culm, nulated slag, mill tailings, and even ashes, have been used as filling material in place of or in conjunction with rock. Of all the materials used for filling the best is probably clean sand. Gravel, finely crushed rock, granulated slag and coarse mill tailings are almost as good, and the methods used for placing and confining sand will apply equally well to any of these. The advantages of sand are its comparatively slight compressibility and the ease, economy and convenience with which it may be placed. On account of the voids, which necessarily occur in fills made of coarse waste, such fills will compress 10 to 25 per cent before they develop their full strength, whereas clean sand seldom compresses more than 5 per cent. Sand is especially desirable when practically complete filling is required, and sand derived from mill tailings, especially when not thoroughly cleaned, has been remarkably effective in sealing off and extinguishing mine fires. When natural sand is available on surface, or when it can be obtained from the mill tailings or tailings dump, the cost of using this material for fill will probably be considerably less than that of rock. In this paper it is not proposed to discuss all methods of filling, butt to describe some of the methods used for placing sand fill and to make a brief comparison of their merits.

Jan 1, 1932

By G. H. Wigton

THE oxidized ores of the Chief Consolidated Mining Co., in the Tintic mining district, have never yielded to metallurgical treatment by any standard method except smelting. These ores occur in comparatively small, irregular, siliceous replacements in limestone, and carry silver, lead, and gold as the principal valuable metals. But smelting methods. are unsatisfactory, as the lead smelters heavily penalize the large amount of silica in the ore, and the copper smelters do not pay for the lead in the ore; therefore, an alternative treatment method was sought. In 1916, a metallurgical research department was organized under; the direction of the writer. Many standard metallurgical methods were investigated but, on oxidized ore, none yielded extraction great enough to make its use feasible. During some chloridizing-roasting experiments in 1916, made in an assay muffle, the writer noted that large proportions of the silver, lead, and gold were volatilized from certain oxidized ores. The possibility of volatilizing the metals and collecting them in the form of a concentrated fume was then studied. But, as it was shown that satisfactory chloride-volatilization would be limited to a furnace feed with a high fusing temperature of the non-volatilizable constituents, and as the results obtained in a rotary kiln with the admixture of common salt with the ground ore were inferior to muffle tests, further volatilization experiments were discontinued until 1920. During this interval, an experimental sulfidizing and flotation plant was operated. After the cessation of attempts to float the oxidized minerals, chloride-volatilization experiments were resumed. The highly infusible flotation. tailing from the operation of the flotation mill was used for furnace feed. The first tests made on flotation tailing mixed with 10 per cent. common salt showed -nearly complete lead and gold volatilization but considerably less volatilization of silver than in muffle tests. In attempts to increase the silver extraction, the percentage salt in the furnace feed was varied, and it soon became apparent that the use of smaller percentages of salt raised the fusion temperature of the furnace charge and thus,

Jan 8, 1925

By Henry Rawdon

THE following note describing the behavior of copper under rather unusual conditions is offered here for its suggestiveness rather than as a complete study of the question. The examinations described were made because of a request for information concerning the embrittlement of copper when heated in a molten salt bath (sodium chloride) for the purpose of cleaning it, as well as for softening it. It is well known that metals may be rendered very brittle by electrolytic action, in aqueous solutions, in all probability by the action of nascent hydrogen, and it was suspected that perhaps the embrittlement of the copper by means of the molten salt was produced in a somewhat similar manner. It was found that the copper was in contact with iron (or steel) while immersed in the molten salt, through the occasional use of an iron stirring rod, iron forceps for handling the pieces, and an iron pot for containing the fused salt. In the electrochemical potential series iron and copper stand some distance apart, copper being electronegative to the iron. Assuming that the two metals bear a similar relation to each other, when immersed in a bath of molten salt, as they do in an aqueous solution, they would be expected to form a galvanic couple, the copper being the cathode. By this means the copper might be embrittled, as often happens in aqueous solutions, by the action of the liberated hydrogen. In this case, however, the embrittlement must be attributed to some other agent than hydrogen. To demonstrate how copper may be embrittled under these conditions, two small rods, one of copper and the other of mild steel, approximately 3/8 in. (9.5 mm.) in diameter, were immersed in a bath of sodium chloride that had been heated somewhat above its fusion point (melting point of sodium chloride is 819° C.) and an electromotive force of approximately 6 volts from a storage battery was impressed upon the system. The copper was made the cathode and the action was continued for

Jan 2, 1920

By J. Mason

In order that the use of continuous mining machines be most economic, the equipment must be operated at rated capacity as much of the time as possible and downtime for repairs kept at a minimum. A large factor influencing any proposals for maintenance programs is the complexity of the machines—often the intricacies of the precision-built equipment requires engineering knowledge byond the scope of mine maintenance personnel. A training program is part of the answer; there are others. In addition to investigating a joint company-manufacturer equipment-design improvement program, Mountaineer Coal Co. has launched a detailed preventive maintenance program to combat extensive downtime often encountered when continuous miners are used. Everyone in the coal industry knows the importance of operating continuous miners at their rated capacity as much of the time as possible. The potential capacity of the miner is such that l-min. delay means from 3 to 8 tons of coal lost. Today&apos;s miner is a well designed, precision-built machine, complicated with hydraulic and electrical equipment and complex mechanical assemblies with all possible combinations of gears, levers, and rods. As the machine becomes more complex, the problem of maintenance becomes more acute. The transition from conventional to continuous mining has taken place so rapidly that the mine mechanic has been unable to keep pace. When there is a machine failure, the maintenance man usually employs either the trial-and-error or the play -it-by-ear method of repair, but this represents an expensive proposition. Mine mechanics, maintenance superintendents, and mine managers complain about the complexity of the machines and how difficult it is to keep them running. There is no shortage of skill in the design offices. Machines are designed by good engineers, but for engineers. What&apos;s the answer? Better trained maintenance men—better maintenance instruction—better field service. These are all good corrective measures, but the real cure is to provide for the men at the do it level. Designers have got to cooperate with men to simplify maintenance on the increasingly complex machine. The designer must be aware of how his machine can be maintained. Sometimes a simple design change in a bearing, oil seal, or a bushing will correct recurrent breakdowns. Parts such as filters, hydraulic cylinders, hoses, and pivot pins can be better arranged on the machine to make them accessible so that the mine mechanic can easily keep the machine producing. Therefore, to get our maintenance off on the right foot, the program must begin back in the design office, before the machine is delivered to the mine property. Realizing the importance of this aspect, our company set out on a joint company-manufacturer equipment-design improvement program. This is a formal program in which quarterly reports are written covering machine defects, design weakness, and general problems relating to machine maintenance and performance. The manufacturer in turn comments on each and every item appearing in these reports, indicating the action he proposes to take. In addition to the written reply, meetings are scheduled with manufacturer representatives to discuss further the outlined problems. The benefits realized from this organized program are both substantial and far reaching. Good maintenance begins with a good design. The subject of preventive maintenance is not a new one. Volumes have already been written on the subject. It consists primarily of doing the maintenance work when you want to rather than when the machine demands it. However, a preventive maintenance program can be overdone. Theoretically, if carried to ridiculous extreme, all working time could be spent inspecting and repairing the machines and no time would be left for production output. Our goal in maintenance is to maximize running time for all machines, and to prevent or at least minimize production downtime. INSPECTION AND REPORTING The center of any preventive maintenance schedule is regular and systematic inspection, but it must be

Jan 1, 1961

By S. S. PRENZ

Discussion held at the sessions of the Canal Zone meeting, November, 1910. [SECRETARY&apos;S NOTE.-As a result of this discussion, an expression of opinion covering all points of unanimous agreement was adopted and signed by members of the Institute party as individuals. This statement, which was printed in the Bulletin of December, 1910, is reprinted below, as an introduction to ,the detailed report of the discussion. It should be clearly understood that it does not constitute an official declaration by the Institute, the rules of which forbid any such action. At the New Haven meeting of February, 1909, a free discussion of the plan of the canal took place, in which the critics of the present scheme had full opportunity to state their views. That discussion was continued at considerable length in the newspapers ; but the Institute was never committed to either side. In this latter instance, the situation is precisely similar. Certain members and guests of the Institute, having had the advantage of later developments and of personal inspection, have expressed the conclusions at which they have arrived on their own responsibility ; and these conclusions are published by the Institute, subject to further discussion, like any other part of its proceedings. The general statement referred to contained a promise that the full discussion, of which it was a result, would be subsequently printed ; and this promise is fulfilled, so far as practicable, in the subjoined report, concerning which I beg to offer the following explanations: 1. It is made up of the remarks of the members of our party, as subsequently written out and mailed to the Secretary. 2. In editing these contributions, I have condensed the material considerably by cutting out much repetition of opinions and phrases ; and in such excision, I, frankly confess, the authors whose communications were delayed, have been the sufferers. Nevertheless, there is a good deal of repetition left, as there must be, when so many persons had talked so much, and come to agreement upon so many points. 3. It has been found impossible to preserve in this report the actual order of the free and informal oral discussion represented. Consequently the several contributions have been arranged in alphabetical order under the names of their authors. References to "preceding speakers," etc., must therefore be construed as referring to other speakers, preceding or following.-R. W. R.] S. S. PRINZ AUGUST WILHELM, AT SEA, November 14, 1910. We, the undersigned, members and guests of the American Institute of Mining Engineers, after a visit to the Isthmus of Panama, and inspection of the work of the United States Isthmian Canal Commission, and after full discussion of our individual impressions, find ourselves in unanimous agreement as to the following conclusions:

Jan 1, 1911

By H. H. Bein

Most western industrial sands are feldspathic and contain feldspars in variahle amounts. A few deposits will show alumina contents of less than one per cent while others will contain over twelve per cent alumina. Some industries are not especially concerned with the amount of contained feldspar but are interested only in color and in size gradations. Other industries, however, are very much interested in alumina content and set up quite rigid specifications as to this content. Several western plants are beneficiating crude sands both to reduce the iron content and to control the feldspar content. At present, most of these plants are using flotation processes for both controls although a few plants do use magnetic separators for the reduction of iron content. Some work has also been done with electrostatic separators for the feldspar-quartz splits. The Sand Division of Del Monte Properties Company, Pebble Beach, California, has been operating a flotation plant for iron removal and the separation of quartz and feldspar since 1950. As mentioned before, control of alumina, especially in the quartz product, is very important. Some users require an alumina content of less than two per cent and others will demand a product of less than one per cent Al2,O3,. The standard Quinolate method for determining alumina is lengthy and time consuming. The HF —H2SO4 evaporation step is especially critical and, if forced, spattering is apt to occur with a resulting loss of sample. Standard procedure is to evaporate to dry ness overnight at a low temperature. Even with forcing under optimum conditions the procerlnre will take two or three hours to complete. A new method has been developed which allows an alumina determination to be made, in 10 or 15 minutes, on a sample of the quartz product as it is being produced. This pocedure, although simply an approximation, has been most helpful in showing whether or not the product being formed is meeting specifications. This procedure has also allowed a great reduction in samples formerly taken and analysed by the orthodox method. Whether a sample assays 1.75% A12O3, or 1.95% .41,0, is not so important as the fact that it is below the allowable maximum of 2.00%. Similar thinking would apply to products required to contain less than 1.00% A1,0,. The step by step procedures for the method are as follows: 1. Collect sample, drain or filter, and dry. 2. Split sample through standard Jones sample cutter to quarter or eighth sample, depending on size of original sample taken, then split down to about 5 grams through Sepor micro-splitter. 3. Screen through 50 mesh sieve (U.S. Series) and reject undersize. 4. Weigh out 2 grams and place in platinum crucible. 5. Etch with 3-5 mls 38.40% HF for 1 minute at room temp. 6. Decant and wash. 7. Stain for two minutes in same crucible with a 1% water solution of Nigrosene dye. 8. Decant and wash well. 9. Dry.

Jan 1, 1961

By Grafton Thrasher

For the proper discussion of this subject it is necessary to incorporate in this paper the substance of part of an article published by rite in Metallurgical and Chemical Engineering for January, 1915. CAST-IRON TEST PLUG, SHOWING LOW MOTTLED AND MOTTLED FRACTURES. Several years ago, while I was thief chemist for a malleable-iron company, it occurred to me that the mere regulation of silicon was not sufficient to produce the desired degree of chill. But only after repeatedly encountering this fact in befit open-hearth and air-furnace, as well as cupola metal, was I led to regard variation in the total carbon as the principal cause of disturbance; and I resolved to determine the total carbon as well as the sill on ill every heat, and by a carefully prepared test piece to gauge the approximate degree of chill. The test piece selected was a round plug, 2 in. in diameter and 12 in. long, cast on end

Jan 10, 1915

Jan 1, 1882

By W. T. Swensen

The Anaconda Company, through its wholly owned subsidiary The Anaconda Company (Canada) I,td., has investigated, optioned and partially developed a large tonnage of iron ore amenable to magnetic concentration. The reserves are located in north central Ontario and occur within a twenty-two mile belt of layered and intricately folded metamorphic rocks, close to the north contact of an igneous complex of granite and pegmatite. There are two main ore deposits. Briarcliffe is a steep to vertically dipping body 100 to 500 feet wide and over a mile long. The Two Mile Lake deposit is flat dipping and confined to an open, asymmetrical syncline; it is approximately a mile wide and 250 to 4.00 feet thick. On its flat easterly plunge, it is saucer-like in shape and is amenable to open-pitting for an east-west distance of several thousand feet. Magnetometer surveys, confirmed by drilling, indicate at least a three mile eastward extension of this deposit. Location and Access: The property is located in the Kowkash Mining District of northwestern Ontario, roughly forty-four miles north of the community of Nakina on the main line of the Canadian National Railroad. (Fig. 1) Access during the exploration period was entirely by airplane from Nakina. (Fig. 2) More recently, Nakina itself was connected to the highway network in the Longlac-Geraldton area. The property is roughly 190 miles by existing and proposed railroad from the Great Lakes waterway system. The Trans-Canada pipeline moving natural gas from western to eastern Canada passes through Geraldton roughly 70 miles south of the property. Topography: The surface features in the area are typical of a sizeable portion of the Precambrian Canadian Shield area. Many lakes, low swampy areas, and glacial moraine-type features are common in this recently glaciated area. The average elevation is between 800 and 900 feet above sea level, and the relief is probably less than 100 feet with the greater portion of the area less than 50 feet. The topographic "highs" are formed either by outcrops that were resistant to glacial erosion or by moraines, eskers and drumlins which were deposited upon the recession of the glaciers. The better mineralized zones, composed chiefly of quartz and magnetite, commonly form ridges resistant to erosion and weathering. Previous Work: Because of the obvious outcrop of the magnetite-rich beds and knowledge gained from early developments within the general region, it is conceivable that the presence of the iron formations has been known for a good many years. The area is included in studies made by L. F. Kindle* in 1929 and 1930 for the Ontario Department *Kindle, L.F., Kowkash-Ogoki Gold Area, Ontario Department of Mines Annual Report, Vol. XL, Part IV, 1931; map 40f

Jan 1, 1961

By J. L. Green

Virginia-Carolina Chemical Corp., in its Florida Operations, uses both roller mills and ball mills in dry grinding of phosphate rock. The roller mills have been in operation since 1954 in the Mining Department. The ball mills were installed and put in operation in 1960 in the Harding Concentrate Superphosphate Plant. The roller mills in the Mining Department are used to grind phosphate rock for outside customers. This involves short notice orders of varying grinds, grades and types of rock. The ball mills grind phosphate rock for use in V-C&apos;s own phosphoric acid and concentrated superphosphate plant. These mills grind generally two types of rock for one grind. In 1959, a 10% speed-up was made in the operating speed of the roller mills. This paper deals with the operational gains and resulting cost savings which evolved from this speed-up, and compares the present operations of the roller mills and of the ball mills in grinding phosphate rock. In January 1959 it was decided to determine the feasibility of reducing the grinding cost on the roller mills, per ton of rock ground, by speeding up the mill 10%. At this time, one mill was completely reconditioned so it would be in top condition. Extensive tests were then run both before and after the mill speed-up to determine the resultant cost savings. These tests showed that, due to this 10% speed-up, the mill capacity was increased approximately 20% and grinding power costs per ton were reduced approximately 10%. Table I shows the results of these tests. Each test run actually represents several of the same type rock and grind, averaged together. Each test run lasted approximately three hours. The validity of these runs has been confirmed by subsequent operation in grinding over 1,000,000 tons of phosphate rock. Accurate records have been kept on power consumption and repairs as well as tons produced. Two types of feed rock were used for the tests — pebble phosphate rock and flotation phosphate concentrate rock. The pebble phosphate rock size is approximately 80-90% plus 20 mesh. The flotation concentrate phosphate rock size is approximately 15% plus 20 mesh and 95% plus 100 mesh. The same two types of rock feed were used for all the test runs. Test Nos. 1, 2, and 3 were run at the original mill speed of 85.5 rpm on the vertical shaft. Test No. 1 was run with a feed of high grade flotation concentrates; the desired fineness attempted was 70% passing 200 mesh. Tests No. 2 and No. 3 were run with a feed of low grade pebble. Test No. 2 was run for a desired fineness of 70% passing 200 mesh. Test No. 3 was run for desired fineness of 50% passing 200 mesh.

Jan 1, 1963