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By J. C. Wilhoit, J. B. Cheatham

Although an oil well is a long cylindrical hole with an irregular bottom, it appears likely that the nature of the stress concentration at the bottom of the hole can be ascertained from an analysis of the stresses around a short cylindrical cavity with rounded comers and smooth bottom. Such a cavity is studied primarily because it leads more readily to a solution to the problem by the use of stress functions. In this paper the stress distribution around a short cylindrical cavity subjected to bit loading, overburden and drilling-fluid pressures is determined by means of an analytical solution which approximately satisfies the boundary conditions of the problem. From this solution the stresses at the corner of the hole are calculated to be about 35 per cent lower than comparable results obtained by photoelastic and relaxation analyses. This difference is apparently due to the large radius of curvature at the corner of the cavity in the present analysis. Since good agreement is obtained between the results of this analysis and the stresses calculated for a similar loading on a semi-infinite elastic solid, it is concluded that the bit action in the region near the center of the hole is not appreciably affected by the presence of the sides of the hole. INTRODUCTION Much has been written concerning drilling "under down-hole conditions" and pertaining to the stress distribution in the rock at the bottom of an oil well.1-5 For example, it is known that identical rocks can be drilled more rapidly at the surface than under subsurface conditions of pressure and stress.6 Information on the behavior of rocks under loading can be obtained from triaxial test data.7-9 From such tests it is found that rocks exhibit brittle failure when the confining pressure and pore pressure are equal, but the mode of failure may change to ductile as the difference between the confining pressure and the pore pressure is increased. Brittle failure implies that there is very little permanent deformation before fracture, whereas ductile failure indicates that permanent deformation takes place before fracture. Some rocks are ductile at differential pressures of 5,000 psi , but other rocks are brittle even at differential pressures of more than 50,000 psi. Cuttings embedded in mud at the bottom of the hole may act as a plastic mass which the bit teeth must penetrate in order to attack the virgin rock below. 10 During drilling, the mud pressure acts as the confing pressure, and in many cases the difference between the mud pressure and the formation pore pressure IS sufficiently low so that the rock most likely falls in a brittle manner. Very little penetratiori by the bit teeth may be required for brittle failure of the rock. For these brittle materials, the elastic stress distribution is of practical interest in determicing the possible effects of a given loading on failure. The penetration of bit teeth into ductile or plastic rock has been analyzed previously ll,12 and will not be considered further in the present work. During actual drilling, many teeth act at various points ovel an irregular hole bottom. In the present analysis a solution is obtained for an idealized problem of determining the elastic stress distribution caused by only one tooth acting alone at the center of a smooth hole bottom. It obviously is not proposed that any driller should put a special one-tooth bit on bottom, but it is hoped that extensions of the simpler problems can eventually lead to a better understanding of more complex actual drilling phenomena. To obtain a solution by the use of stress functions, a short cylindrical cavity with rounded corners and smooth bottom is studied. The solution to this problem should give an insight into the nature of the stress concentration at the bottom of an oil well which is in reality a long cylindrical hole with an irregularr bottom. After selection of the proper curvilinear co-ordinate system, the stress functions are expressed as series of Legendre polynomials. The coefficient of each term in the series is then determined to satisfy the boundary conditions at a finite number of discrete points on the boundary of the cavity with least square error. Results are compared with bottom-hole stresses obtained by photoelastic and relaxation analyses.

By Ellis H. Gates

BENEFICIATION of the manganese oxide ores at Three Kids Mine near Henderson, Nev., has evolved over a period of years. Commercial application of the process is on a secure basis, and an effective working hypothesis has been formulated to help solve the problems encountered in flotation. An outstanding difference between the Manganese Inc. operation and other flotation processes is the necessity for high recovery of slime mineral values. Part of the manganese minerals in the Three Kids ore deposit occurs as primary slime or as minerals that are easily slimed by grinding. High recovery of this slime portion is essential to an economic overall manganese recovery. The aim of much of the research at Manganese Inc. was to recover these slimed and finely ground minerals by flotation. Early Agglomeration Processes: Tunbridge' in 1880, followed by Everson2 in 1886 and Cattermole" in 1904, discovered that under certain pulp conditions precious metals and sulfide minerals could be agglomerated if enough soap or oil were mixed into the pulp slurry. These agglomerated particles, Which under ideal conditions ranged from 1/16 to 1/4 in. diam, were recovered by screening, tabling, and hydraulic classification. Even flotation by air occluded in the agglomkrate particles by vigorous stirring was suggested. Cattermole nearly succeeded in getting his process into commercial operation, but it was finally abandoned, presumably because he could nat compete with the low reagent requirements of the new agitation froth process. Since Cattermole's time Christensen in 1921, Borcherdt4 in 1926, and Chapman and Littleford- in 1934 have attempted to apply the principles of agglomeration. In the light of current research at Manganese Inc. on conditioning of pulp at high energy input levels, it is interesting to note the variety of equipment such as tumbling mills, mixing drums, and shakers described by Borcherdt in his patent to produce the agglomerated particles. Ore Characteristics: Manganese in the Three Kids ore occurs as a 'complex mixture of manganese oxides commonly Known as wad. The manganese minerals have Been identified as psilorflelane, hollandite, coronadite, and pyrolusite with minor amdunts of rhodonite and rhodochrbsite. The ppincipal gangue miherals are quaktz, opal, kaolinite, montmoril-lonite, calcite, gypsum, fcelesitite, and barite. Most of the mangandse ore minerals are either porous and sponge-like in structure or soft clay-like masses composed of fine grained mineral particles approaching the size range of colloids. Consequently the ore minerals from the Three Kids mine have relatively large surface areas and high void content per unit weight. The specific area of various manganese ore samples measured at the California Research Corp. laboratories ranged from 32 to 58 sq m per g. The great surface area and high void content explains in part the unusually high reagent consumption required in the treatment of this ore. Current Milling Practice: At Henderson6 he ore is crushed to —1 in. and ground with rod mills in a closed circuit with spiral rake classifiers and cyclones. Sulfur dioxide solution is added to the pulp as it leaves the grinding circuit, immediately followed by addition of a reagent mixture of soap, oil, and wetting agent in the form of an emulsion. The reagents are initially well mixed into the pulp during their passage through centrifugal pumps to the conditioning equipment. All reagents are added prior to conditioning. The intense conditioning or mechanical stirring of the pulp, in which large amounts of power are consumed, is the principal difference between the flowsheet used at Manganese Inc. and a standard flotation circuit. As much as 38 kw-hr of energy per dry short ton of ore in a pulp containing 18 to 23 pct solids is used for conditioning. After conditioning, the pulp passes to a flotation circuit that consists of one roughing stage, a scav-

Jan 1, 1958

By R. E. Pray, M. C. Fuerstenau, J. D. Miller, B. F. Perinne

Quartz cannot be floated with potassium amyl xanthate as collector at any pH. Complete flotation is achieved with certain minimal additions of amyl xanthate and Pb from pH 5.8 to 8.5 and with amyl xanthate and Zn from pH 7.5 to 8.1. The active species of these metal ions responsible for activation are shown to be PbOH and ZnOH No flotation was effected when Cu and Mg were added as activators. Previous work1,2,3,4,5,6,7 has shown that quartz may be floated in the presence of sulfonate or soap when the following two criteria are met: 1) some portion of the added metal ion has hydro-lyzed to its first hydroxy complex, and 2) a precipitate of the metal-collector has formed. Since sulfonate and soap, two substantially different anionic collectors, conform to these criteria, this phenomenon may well be expected to be of a general nature. In this view then, it would appear that any anionic collector, including xanthates, should also float quartz when these two criteria are met. The object of this paper is to demonstrate the conditions under which quartz will respond to flotation in the presence of potassium amyl xanthate. EXPERIMENTAL MATERIALS AND TECHNIQUES Quartz was prepared by leaching the sized sample (48 x 150 mesh) with HC1 until no iron could be detected in the leach liquor. Conductivity water, made by passing distilled water through an ion exchange column, was used in the investigation. Potassium amyl xanthate was selected as collector. Purification of this reagent was effected with the following method: 1) a laboratory sample of potassium amyl xanthate was dissolved to saturation in hot ethyl alcohol and filtered several times while still hot, 2) the filtrate was cooled, during which time potassium amyl xanthate crystallized out of solution; the xanthate was then separated from the solution by filtration, 3) the amyl xanthate was dissolved in fresh hot ethyl alcohol and filtered, 4) step (2) was repeated, 5) the crystalline potassium amyl xanthate was washed with ethyl ether five times, and 6) the xanthate was dried for 24 hours under continuous vacuum. Inorganic impurities are insoluble in hot ethyl alcohol whereas dixanthogen is soluble in ethyl ether. Five-gram charges of 48x150 mesh quartz were conditioned for 3 min in 125 cc of solution containing the reagents and were floated with the small-scale cell and procedure described in previous papers.478 Reagent grade HCl and NaOH were used for pH adjustment, while reagent n-amyl alcohol was added as frother. Frother addition is necessary with this technique, but it should be mentioned that data, similar to those shown in Fig. 2 for 1 x10-4 mole per liter Pb++ and 1 X 10-4 mole per liter AX-, were also obtained with a Hallimond cell in the absence of frother. EXPERIMENTAL RESULTS Clean quartz could not be floated with amyl xanthate at any pH. Flotation was achieved under certain conditions, however, when various metal ions were added to the system. The activating ions used were lead, zinc, copper and magnesium, added as reagent grade chloride salts. Lead as Activator: The effect of Pb++ as activator was examined at five different additions of amyl xanthate (Fig. 1, 2 and 3). When 1x 10-5 mole per liter xanthate was added in the presence of 1 x 10"4 mole per liter Pb ++, no flotation was obtained at pH 6.5 or below, while a recovery of about 60% was effected at pH 7. When the pH was increased to 7.8, no flotation was again obtained, Complete depression was noted from pH 7.8 to about 10, at which point a recovery of about 30% was achieved. At about pH 11, the precipitate of lead amyl xanthate which was present below this pH dissolved and complete depression occurred. When 2.5x 10-5 mole per liter xanthate was added with the same lead addition, no flotation was obtained at pH 5.8 and below, whereas a recovery of

Jan 1, 1965

By George B. Clark

Many basic engineering principles of all four phases of mining operations, namely, prospecting, exploration, development, and exploitation, can be analyzed better in terms of quantitative geology. Geological data from both field and laboratory will also complement scientific methods now being developed. THE geological data emphasized so successfully in prospecting for new deposits, that is, structural controls, strength of solutions, and type of mineralization, are basically those required for successful exploitation. In the mining of newly discovered deposits the most economical methods should be employed as early as possible to keep the overall cost per unit produced at a minimum and to permit maximum extraction of valuable minerals. A crucial question is: How can geological data be translated into useful quantitative results which will aid in achieving this end? H. E. McKinistry' has suggested that a solution may be reached in one of two ways: 1—the usual approach, use of judgment based on experience; or 2—mathematical calculations and tests on models, both subject to certain limitations. He also suggests that in addition to better use of geology more case data and theoretical data are needed on which to base sound judgment. Further research, therefore, is necessary. Perhaps in this field the emphasis should be on more specialization in mining methods and ground movement by men with thorough training in physics, engineering, geology, and underground mining. These specialists would be equipped to point out the most economical and scientific methods of exploitation. Selection of a stoping method is governed by the amount and type of support a deposit will require in the process of being mined, or by the possibility of employing the structure of the deposit to advantage in mining the ore by a caving method. In addition to these factors there are others which almost invariably influence the choice of an economical method of mining:' 1—strength of ore and wall rocks; 2—shape, horizontal area, volume, and regularity of the boundaries of the orebody, and thickness, dip and/or pitch of the deposit and individual ore shoots; 3—grade, distribution of minerals, and continuity of the ore within the boundaries of the deposit; 4—depth below surface and nature of the capping or overburden: and 5—position of the de- posit relative to surface improvements, drainage, and other mine openings. In the final analysis it is usually necessary to disregard the less important of these factors to satisfy the requirements of the more important. Because of the variation of geological conditions throughout and surrounding the deposit, no mining method will be everywhere ideally applicable to the conditions encountered in one ore deposit. The immediate problem is to interpret the above physical characteristics of deposits in terms of geological characteristics. Very few quantitative geological data are available on the factors related to a choice of mining methods. However, there are many descriptive data in mining and geological literature which collectively show how important an effect details of geology have upon all phases of mining operations. The following categories of basic mining methods were investigated to establish the geological factors that have affected their successful application: 1— open stopes with pillars; 2—sublevel stoping; 3— shrinkage stoping; 4—cut-and-fill stoping; 5— square-set mining; 6—top slicing and sublevel caving; and 7—block caving. It should be noted that the first five of these methods are listed in the order of increasing support requirements. Mines were selected as examples only where geological descriptions were complete enough to warrant their use. A study of the geological factors involved in mining operations led to a choice of the following classifications, employed in Table I: 1—structural type of orebody; 2—dimensions (geometry); 3— country rock (type); 4—faulting, folding, and fracturing; 5—alteration of ore and rock; 6—type of mineralization; and 7—geological factors determining mining method (summary). Of these factors only one yielded results that can be defined from available data in a quantitative manner, i.e., dimensions of the deposit. These are the most reliable guides that can be used in selection of suitable mining methods. They are, in general, the properties of geologic structure most difficult to evaluate by studies of models, pho-toelastic studies, and other laboratory methods, all of which are at present more limited in their applications than the geologic method. Application of geology has proved a reliable guide in other phases

Jan 1, 1955

By H. L. Sauder, J. L. Newman, C. Waddell

An instrument capable of measuring subsurface flow rates is described. The instrument is self-contained and may be run on piano wire line. It detects flow by means of an impeller suspended between two torsion wires. The force of the well fluid striking the impeller causes the impeller to rotate, exerting a torsional force to the wires. This force is determined by recording the angle of rotation of the impeller on film using a battery-driven, clock-controlled camera. The impeller is sensitive to flow in either direction and the same instrument can be used to measure injection as well as production rates. Changes in direction of flow, such as might be caused by thief zones, are measurable and are indicated by a reversal of the direction of rotation of the impeller. Adjustment of the sensitivity of the instrument to measure a wide range of flow rates is accomplished by the use of different size torsion wires. An umbrella-type fluid trap, which contacts the casing or wellbore, diverts the flow through the flow tube. The fluid trap remains closed while rurrning in the hole and can be opened at any point in the well. The instrument with the trap closed is 1 3/4 in. OD and may be run through 2 in. tubing. Successful flow profiles have been made on wells with flow rates ranging from 60 BID to 4,000 B/D at surface pressure up to 4,500 psi. The instrument is designed for high-pressure and high-temperature operations. INTRODUCTION The need for an instrument that will measure flow rates in a wellbore is as old as reservoir engineering itself. A number of such devices have been designed within the past few years. Some of these have been described in the literature1,2,3,4,5,6 One of these devices1. a hot-wire anemometer type, made many successful jobs in dry gas producing areas. Difficulty was encountered when this flowmeter was used in wells producing liquids and development work was initiated to overcome this difficulty. The present flowmeter is the result of this work and provides an instrument with a wider range of application in the industry. Consider- able development work has been done on the tool involving many changes in the mechanics of recording the flow readings, however, the basic principle of flow detection has remained the same since its inception. PRINCIPLE OF OPERATION The basic principle of operation of the flowmeter is the utilization of a portion of the kinetic energy of flowing well fluid to impart a rotational force to an impeller suspended between two torsion wires. This impeller, Fig. 1, is a screw-type propeller with the blades having a pitch angle of 40. As the well fluid passes through the instrument it strikes the impeller. The resulting force rotates the impeller and twist\ the torsion wires to a balance point at which the torque imparted to the wires balances the force required for the impeller to deflect the well fluid. The impeller is symmetrical and a change in direction of the flow results in a reversal of the direction in which the impeller is deflected. Thus the direction of the flow as well as its velocity is indicated. A dial is attached to the impeller by means of a hollow tube through which the upper torsion wire extends. The angular position of the impeller is recorded by photographing the dial at frequent intervals. In any given set of torsion wires the modulus of shear and polar moment of inertia remain constant and the deflection varies directly with the applied torque. The deflection imparted to the torsion wires also varies as the square of the fluid velocity. Further, the velocities represented by any two given deflections hear the following relationships: and Where V = velocity of flow stream and O ;= deflection. This is the basic relationship which is used to calculate the relative velocities represented by a given set of deflections where the velocity represented by any one of the deflection readings is a known quantity. For a given torque, the deflection of the torsion wire varies inversely as the fourth power of the wire diameter. Wire sizes ranging from ,005 in. in diameter to .022 in. in diameter can be used. The range of

Jan 1, 1957

ARTICLE I Name and Object Sec. 1. This Division shall be known as the Institute of Metals Division of the American Institute of Mining and Metallurgical Engineers. Sec. 2. The object of the Division shall be to furnish a medium of cooperation between those interested in the field of physical metallurgy; that is, the nature, structure, alloying, fabrication, heat treatment, properties and uses of metals; to represent the AIME insofar as physical metallurgy is concerned, within the rights given in AIME Bylaw, Article XI, Sec. 2, and not inconsistent with the Constitution and Bylaws of the AIME; to hold meetings for the discussion of physical metallurgy; to stimulate the writing, publication, presentation and discussion of papers of high quality on physical metallurgy; to accept or reject papers for presentation before meetings of the Division. ARTICLE II Members Sec. 1. Any member of the AIME of any class and in good standing may become a member of this Division upon registering in writing a desire to do so, but without additional dues. Sec. 2. Any member not in good standing in the AIME shall forfeit his privileges in the Division. ARTICLE III Funds Sec. 1. The expenditure of the funds received by the Division shall be authorized by the Executive Committee of the Division. ARTICLE IV Meetings Sec. 1. The Division shall meet at the same time and place as the annual meeting of the AIME, and at such other times and places as may be determined by the Executive Committee subject to the approval of the Board of Directors of the AIME. Sec. 2. The annual business meeting shall be held within a few days before or after the annual business meeting of the AIME. Sec. 3. At a meeting of the Division, for which notice has been sent to the members of the Division through the regular mail or by publication in the Journal of Metals at least one month in advance, a business meeting may be convened by order of the Executive Committee and any routine business transacted not inconsistent with these Bylaws or with the Constitution or Bylaws of the AIME. Sec. 4. For the transaction of business, the presence of a quorum of not less than 25 members of the Division shall be necessary. ARTICLE V Officers and Government Sec. 1. The officers of the Division shall consist of a Chairman, a Senior Vice-Chairman, a Vice-Chair -man, a Secretary and a Treasurer. The office of Secretary and Treasurer may be combined in one person, if desired by the Executive Committee. Sec. 2. The government of the affairs of the Division shall rest in an Executive Committee, insofar as is consistent with the Bylaws of the Division and the Constitution and Bylaws of the AIME. Sec. 3. The Executive Committee shall consist of the Chairman, Senior Vice-Chairman, Vice-Chairman, past Chairman, Secretary, and nine members, all of whom shall be nominated and elected as provided hereafter in Article VII. Sec. 4. The Chairman, Senior Vice-Chairman and Vice-Chairman shall serve for one year each, or until their successors are elected. Each member of the Executive Committee shall serve three years. The Chairman shall remain a voting member of the Executive Committee for one year after his term as Chairman. Sec. 5. The Treasurer of the Division shall be invited to meet with the Executive Committee, but without ex-officio right to vote. He shall be appointed annually by the Executive Committee, from the membership of the Executive Committee or otherwise. Sec. 6. The annual term of office for officers of the Division shall start at the close of the Annual Meeting of the Institute and shall terminate at the close of the next Annual Meeting. ARTICLE VI Committees Sec. 1. There shall be standing committees as follows: Programs Committee. Finance Committee, Membership Committee, Annual Lecture Committee, Technical Publications Committee, Mathewson Gold Medal Committee, Nominating Committee, Education Committee and such other Committees as the Executive Committee may authorize. Sec. 2. It shall be the duty of the Programs Committee to secure the presentation of papers of appropriate character at meetings of the Division. Sec. 3. It shall be the duty of the Finance Committee to inquire into and examine the financial condition of the Division and to consider proper means of increasing its revenue and limiting its expenses. The Finance Committee shall audit the accounts of the Division and report to the Executive Committee prior to the Annual Meeting of the Division. It shall render a budget to the Executive Committee estimating receipts and expenses for the ensuing year so that action can be taken on same at the first meeting following the Annual Meeting.

Jan 1, 1953

By Morris Muskat, James M. McDowell

A report is given of electrical analog experiments on the effect of casing perforation completions on well productivity for ideal uniform reservoirs under steady state homogeneous fluid flow conditions. Graphical results are presented for the effect of perforations of different densities and various degrees of penetration into the formation surrounding the casing or cement sheath. The latter range from perforations which terminate flush with the casing to those which extend into the formation a distance equal to the casing diameter. For the special case where the perforations terminate immediately beyond the casing with semi-spherical protrusions the electrical analog results correct and supersede previously reported theoretical calculations. It is found that the penetration of the perforations into the surrounding productive section may so increase the resultant productivity as to approach or even exceed that for open hole completions. INTRODUCTION A theoretical analysis of the effect of casing perforation on well completion, without additional penetration of the formation, was reported several years ago.' In this treatment the perforations were represented by mathematical sinks and no explicit account was taken of the imperviousness to flow of the casing in which the perforations were made. This procedure was justified by consideration of the mutual interference between the sinks, which seemed to indicate that the analytical method automatically eliminated flow across the casing radius between the perforations. However, both further theoretical investigation and the electrical analog measurements described below showed that this assumption in the previously developed theory was not valid and that the calculated well capacities would considerably exceed the correct values. § Even if this simplifying assumption were retained the generalization of the theory to situations where the perforations extend into the producing formation would be extremely difficult. On the other hand, the electrolytic model analog of this type of steady-state homogenous fluid system is, in principle, so simple that it has not been worthwhile to attempt to develop the exact analysis even for the more limited case of strict casing perforation. The whole problem has therefore been studied anew by electrical models to cover the range from holes in the casing which terminate flush with the casing surface to such as extend into the surrounding rock to a depth equal to the casing diameter. It is to be expected that the productivity of a well completed with casing radius* rw, perforated with holes of radius Tp, and extending to depths d into the surrounding rock, would depend primarily on the dimensionless geometrical parameters Te Tw TP d of the system, such as —, —, — , and —, where re is the Tw, a a rw external boundary radius and a is the mean vertical separation between the perforations. It will also depend, of course, on the number of perforation lines or total perforation density. While such parameters automatically control analytical treatments, it would require a very extensive series of model experiments to fully disclose their functional effect on the well productivity. For practical purposes, however, a rather limited range of the variables suffices to give the effects of interest. Accordingly, in the experimental study reported here the casing diameter has been kept fixed? at six in., the perforation diameter has been taken as either one-fourth in. or one-half in., their density has been varied from one to eight per ft, and their penetration into the formation has ranged from zero to six in., with a primary scaling factor of one to six. EXPERIMENTAL METHOD AND RESEARCH Although electrolytic model experiments have been made for years, there still seems to be some question as to what set of conditions will give the best results. For this reason a number of electrode materials and electrolytes were tried. Those chosen for the final experiments were copper electrodes and a weak copper sulphate solution as the electrolyte. This 'Assuming that the casing has been cemented with a uniform and impermeable cement sheath, the effective casing radius is to be interpreted as the external radius of the cement annulus. It is also assumed here that the casing is set through the complete producing section that it is perforated throughout with a uniform density and that tie formation has a a and isotropic permeability. As noted later, however, a uniform scale change permits reinterpreting the primary measurements to correspond to a casing diameter of 12 in.

Jan 1, 1950

By W. A. Sheaffer

Electric furnace installation and tough-pitch copper-casting operation at Canadian Copper Refiners Ltd. are described. General layout, power supply and control, refractories, induction pour hearth, casting equipment, metal temperature control, and oxygen-content control are discussed. E LECTRIC furnace at Canadian Copper Refiners Ltd. was put into operation in August 1949. The installation was designed primarily to melt electrolytic copper cathodes and to produce vertically cast tough-pitch shapes. To meet emergencies during reverberatory furnace shut-downs, provision was made for casting horizontal tough-pitch wire bars. Due to the ease with which metal temperatures, melting rates, oxygen content, and casting hours can be altered to suit production demands, the electric furnace has met the requirements admirably. General Layout The plant has been described by H. S. McKnight' and by J. H. Schloen and E. M. Elkin.2 Since these papers were written, the electric furnace has been installed in a 140 ft extension of the original casting building. The extension, 264 ft in width, is divided into one 24 ft and four 60 ft bays, each a continuation of similar bays in the older building. Fig. l, a partial floor plan of the extension, shows the location of major equipment. Power Supply and Control Power for the electric furnace and its auxiliary equipment is delivered by Quebec Hydro over two 12 kv three-phase, 60 cycle pole lines. One line is in use, while the second serves as an emergency standby. Connections for the electric furnace from these lines, which also feed the refinery power house, are made to a substation on Canadian Copper Refiners Ltd. property. Connections between this substation and the furnace transformer room are in underground conduit. The are-furnace transformer, 550 v auxiliary-equipment transformer, and 12 kv switchgear are in the transformer room near the furnace. Stepdown from 12 kv to are-furnace voltages is done by a 7500 kva oil-immersed water-cooled three-phase 60 cycle transformer. This unusually high kva rating is available because the arc furnace and transformer were adapted from a steel-melting unit. Secondary voltages are available in 16 steps between 251 and 95 v (across phases). A four-position tap changer, operated from the furnace switchboard is connected to taps yielding 199, 164.5, 127, and 115 v, respectively. The transformer primary feeder is equipped with a General Electric Magne-Blast circuit breaker. The pour hearth, casting wheel, bosh conveyor, and other auxiliary equipment are fed by a 12 kv to 550 v, 750 kva three-phase 60 cycle transformer. Graphite electrodes, 14 in., with tapered nipples are held in the are-furnace electrode arms with steel wedges. The arms, fed from the transformer secondary busbars by flexible cable bundles, are actuated by winch cables. Direct-current motors operating the winches are supplied by a 250 v motor-generator set. Power input to the furnace is controlled from the furnace switchboard. The board has the usual combination of remote controls for the 12 kv breaker and transformer tap changer, meters, overload relays, automatic and manual electrode motion controls, switchgear, etc. When power is on the furnace, automatic electrode feed is used; the power draw on a given voltage tap then is governed by current-limiting rheostats. The rheostats work through a Westinghouse bal-lanced-beam control circuit which, in turn, operates reversing contactors in the winch motor circuits. Arc Furnace The arc furnace is a standard-type NT Pittsburgh 'Lectromelt with inside shell diameter 12 ft 3¾ in. By hydraulic-lift cylinders, the furnace can be tilted forward to 39 ½ ° and backward to 5" from horizontal. A roof-swing cylinder is also installed but not used. Fig. 2 shows refractory-lining details. With the lining as shown, operating life between repairs is 6 to 9 months of two-shift casting operation. Charge slot, skim door, and roof refractories then are replaced completely, and any necessary repairs made to side walls. Bottom life is longer and the original under courses are still in service, while the top course was replaced in October 1951. The launder is lined with high chrome-magnesite trough brick backed by fireclay and insulating brick. The trough is covered with fireclay brick and asbestos sheet. At intervals, openings 12x6 in. are left in the permanent covering. These are covered with

Jan 1, 1956

By V. T. Stringfield, H. H. Cooper

ONE of the earliest investigations of ground water in Florida was made in 1513 when Ponce de Leon arrived at St. Augustine in search of the Fountain of Youth. The history of the development of the water resources of the State shows that the large artesian reservoir that underlies Florida was discovered in the latter part of the last century. Part of that history is given by L. C. Johnson' who states that the first successful artesian well in Florida was drilled at St. Augustine between 1880 and 1882. After the City of Jacksonville failed to obtain a flow of artesian water at a depth of nearly 400 ft and abandoned the drilling, R. N. Ellis, City Engineer, and L. C. Johnson, using Johnson's knowledge of the geologic structure of the artesian reservoir, estimated correctly that artesian water could be obtained at a depth of about 500 ft. Thus began the development of the large artesian system in the northeastern part of the State. In 1908, the Florida Geological Survey issued its first report on the ground water of central Florida. More recent reports give the results of investigations that have been in progress in cooperation between the Florida Geological Survey and the U. S. Geological Survey since 1930. As a result, the ground-water geology and hydrology of Florida are now so well known that ground-water problems such as confronted early investigators no longer exist. However, new problems arise and new discoveries are made as the demand for more ground water increases with the development of the State. Water-Bearing Formations Descriptions of the geologic formations and a map showing their distribution at the surface are given in a recent report by Cooke.' The geologic formations that yield the ground-water supplies in Florida represent only a small part (about 1000 ft) of the total thickness of the sedimentary rocks (more than 15,000 ft) that underlie Florida. The water-bearing rocks that yield fresh water include more than two dozen formations that range in age from Eocene to recent. The Eocene formations, which consist chiefly of limestone, are the oldest and the deepest of the formations in Florida that yield fresh water to wells. In 1944, on the basis of a study of the foraminifera, the Applins -ivided the limestone of Eocene age into six formations, which are as follows, from top to bottom: Age or group Formation Jackson Ocala limestone icAvon Park limestone Claiborne -j Tallahassee limestone aLake City limestone Wilcox Salt Mountain limestone xOldsmar limestone The Ocala limestone and the underlying formations of Claiborne age, along with some of the overlying limestone of Oligocene and Miocene age, constitute the principal source of water in Florida and southeastern Georgia and generally may be regarded as forming one artesian aquifer or water-bearing unit. This aquifer is referred to in this paper as the Floridan aquifer, a name proposed by Parker.' In part of Seminole County alone more than 1000 flowing wells yield water from the Floridan aquifer. A yield of 6500 gpm, or about 9.5 million gal per day, by natural flow from a well penetrating the aquifer at Jacksonville, Florida, was observed in 1942. A yield of about 7000 gpm by natural flow was reported for a well 1390 ft deep at St. Augustine in 1887. The largest yield reported from a pumped well penetrating the aquifer is 7500 gpm, or about 10.5 million gal per day. The aquifer is also the source of some of the large springs, such as Silver Springs, whose discharge, according to measurements made by the U. S. Geological Survey, has ranged from 526 to 1350 sec-ft, or from 340 to 872 million gal per day. The top of the Ocala limestone, as represented by contours in Fig. 1, indicates in a general way the geologic structure of the formations that comprise the Floridan aquifer. The Ocala is at or near the surface in the areas represented by shading in the northwestern part of the peninsula and also in an area in western Florida adjacent to Alabama and

Jan 1, 1952

By E. A. Slover

THE usual reverberatory system of smelting cop--1- per concentrate or calcine has for its component parts a furnace and one or two waste heat boilers. These parts are operated on a basis of compromise, since the furnace can send gas to the boilers at too high a temperature and the boilers by plugging, due to dust or slag, can place a definite limit on the amount of fuel the furnace can burn. Over the years the copper concentrate smelting furnace has had few advances in design. The simple rules of design such as the flame should wipe the bath and the speed of the gases should be reasonably low for dust carrying purposes seem to cover the main features. In the construction of the individual furnaces some innovations are always being introduced. Among these are charging so that the work of smelting is a complete bath process, the use of suspended brick arches in place of sprung arches, the use of basic brick, not only in the crucible, but also in roof and sidewalls, the use of various means to feed the charge, the use of magnetite or other heavy material to construct the hearth, water cooling of bridgewall and slag skimming bay, the smelting of raw charge instead of calcine, the use of preheated air, and possibly the use of oxygen-enriched air for combustion. But the general outlines of the furnaces have not changed much except as to size. Furnaces at Hurley As shown on Fig. 1, the furnace at Hurley is 126 ft long between the longitudinal buckstays and 32 ft wide at the skewback plates. The foundation is a concrete retaining wall with piers at intervals that go deeper into the earth. Purposely the wall at the burner end of the furnace is not backed-up as tightly as the other parts of the foundation so that movement due to expansion may take place here rather than into the boiler foundations. Within these foundation retaining walls of concrete, the earth has been removed to allow the placement of the crucible brick base inside of which a silica hearth is laid 4 ft 6 in. in depth. No expansion is left in the brick base and crucible where they are in contact with the hearth. The hearth itself is of quartzite crushed to 1 in. size with fines left in the product. An 8 in. layer is laid and tamped with paving tampers to about 6 in. in thickness. Then a layer of silica flour is spread and vibrated into the hearth. This operation is repeated until a depth of 4 ft 6 in. is occupied by the silica mass onion-skinned in layers of approximately 6 in. Before firing the entire hearth is covered with broken slag to a depth of 4 in. so that a seal may be formed on the hearth. The crucible is completely faced with magnesite chemically bonded brick while the outside, against the foundation, is made of silica brick. The side-walls are carried up with silica brick in which expansion joints are left at intervals. Above the crucible the sidewall is corbelled to form a shelf on which the charge may build up along the side-walls, see Fig. 2. The arch of the furnace is sprung 20 in. silica brick, with the longitudinal centerline horizontal the length of the furnace, and some 9 ft in the center above the bath. Both straight and wedge brick are used in the construction and a thin silica mortar is troweled for joints. After the arch under heat has assumed its permanent shape, a silica slurry is spread over the arch to fill any cracks that have formed, thus giving bearing surface to the brick and preventing dust from entering the body of the arch to act as a future fluxing agent. The uptake of the furnace slopes up to the boiler entrance where a brick pilaster divides the gas stream for the two boilers. Over this flared uptake is a suspended flat arch of firebrick. The pilaster and sidewalls are constructed of firebrick but the bottom of the uptake is lined with silica brick and fettled through holes in the roof with siliceous fettling. Close to the entrance of each boiler is a brick covered slot through which water-cooled dampers may be lowered in event of boiler trouble. These water-cooled dampers are hung permanently in position ready to be lowered when needed. Flexible hoses to follow the dampers as they are lowered are connected at all times and individual chain blocks are used to lower the dampers. A pump supplying water is started before the dampers enter the heat. Charging of the furnace along the sidewalls for some 80 ft from the bridgewall is accomplished by electric vibrating conveyors fed by belt from charge storage bins above the furnace. These conveyors

Jan 1, 1954

By P. A. Witherspoon, R. W. Donovan, T. D. Mueller

The use of petroleum-barren aquifers for underground storage has become extremely important to the natural-gas industry. A critical problem in assessing the feasibility of a specific aquifer for such use is the permeability determination of the caprock over the proposed storage project. The approach used here is to conduct both static and dynamic field tests on the aquifer being analyzed. Valuable information on the possibility of communication between the storage aquifer and any other aquifers above can be obtained by measuring hydrostatic water levels and water analyses. Significant differences in such data give evidence of the lack of communication between the intended storage reservoir and other horizons. The dynamic approach requires that one well be pumped in the storage aquifer, and changes in fluid levels recorded in both the aquifer and its caprock. The interpretation of the data from such pumping tests involves the solution of nonsteady radial flow in an infinite aquifer and the influence on such flow of a leaky caprock. A finite-difference model has been used to investigate this problem, and the transient behavior has been solved numerically with a digital computer. It has been found that the pressure transients in the storage aquifer are not affected significantly by moderate caprock leakage. The pressure behavior of the caprock is a much better indicator of the degree of leakage, and generalized solutions for this behavior are included. Field data are presented to demonstrate both the static and dynamic approach. If is concluded that appropriate investigation of the groundwater hydrology in an aquifer-type gas-storage project can provide much valuable information for determining the effectiveness of the caprock to hold gas. INTRODUCTION Underground storage of natural gas in the United States has been developing at a rapid rate over the past few years. In 1955, the total gas-storage capacity was about 1.6 trillion cu ft; by 1961, this figure was almost 3.2 trillion cu ft, an increase of 100 per cent in six years.' This trend un- doubtedly will continue because the economics favor the development of gas storage, as opposed to the construction of new pipelines, to meet the inherent cyclic demand for fuel in the metropolitan areas of this country.' About 15 per cent of the current underground gas storage has been developed in petroleum-barren aquifers, i.e., geological domes or anticlines in which no commercial quantities of oil or gas had been produced prior to the storage operations. The necessity for using barren aquifers outside many metropolitan areas of this country has been due to the lack of depleted oil or gas fields that were near enough and large enough to meet the demands of such consuming areas. Pipeline companies have developed aquifer storage along their transmission lines to meet the fluctuating needs of their complex systems. Considerable thought has also been given to the problem of storing gas in a structureless aquifer, both in this country' and in the Soviet Union outside the city of Leningrad.'," Conditions such as these have led to the development of aquifer gas-storage projects in many parts of the U. S. Most of these developments have centered in the Mid-Continent area, and the greatest amount of activity has been concentrated in Illinois.6 Thus, the use of petroleum-barren aquifers for gas-storage purposes has become extremely important to the natural-gas industry. There are three basic problems in developing aquifer-type storage: (1) finding an adequate geologic structure, (2) finding a suitable storage reservoir within the structure and (3) determining the tightness of the caprock over the intended storage zone. The first two problems can be solved by applying conventional methods of exploration geology, but once these problems are solved, the question arises as to why no oil or gas is present in an otherwise favorable setting. Two situations are possible: (1) an adequate source bed was never present, or (2) a source bed was present but the petroleum seeped away because of a leaky caprock. Determining the tightness of the caprock is one of the most critical problems in assessing the feasibility of a specific aquifer for storage purposes. In attacking this problem, one usually takes cores of the caprock and subjects them to a rigorous investigation. Such core data are desirable, but they only detail the matrix properties and cannot be expected to reveal the gross characteristics of the caprock. Several gas-storage projects in the U. S. have had considerable leakage where

By John W. Vanderwilt

THE Climax district is in northeastern Lake County, Colorado, on Fremont Pass (elevation 11,320 ft.) where the continental divide runs east-west joining high peaks of the Mosquito Range with the Sawatch Range, prominent ranges of the central Rocky Mountains. The mining towns of Leadville and Kokomo are thirteen miles to the south and seven miles to the northwest, respectively. Good highways lead to Denver (100 miles) by way of Kokomo, Loveland Pass, and Idaho Springs, and to Colorado Springs by way of Leadville. The ore deposit and mine are about one mile east of Fremont Pass and 600 ft. higher on the southwestern flank of Bartlett Mountain, a peak in the Mosquito Range. The molybdenite deposit lies within the Colorado mineral belt extending across the Rocky Mountains from Jamestown and Boulder to Silverton and Telluride. (See Fig. 1.) Fremont Pass is a broad open divide separating the drainage of the Tenmile River (Pacific) from that of the East Fork of the Arkansas River (Atlantic) ; but the mountains to the east and west are extremely rugged; altitudes in the vicinity vary from 11,000 to 14,000 ft. A conspicuous topographic feature within the district is the Tenmile Amphitheater, which lies east of Climax and

Jan 1, 1946

By A. N. Holden

A DISLOCATION mechanism has been described by Cottrell' by which metals can yield locally, I. form Liiders bands, giving rise to a characteristic stress-strain curve with a sharp yield point and appreciable strain at constant or decreasing stress. It is undoubtedly the best mechanism that has been suggested to date." In its present development, however, the dislocation mechanism provides a more satisfying explanation for the sharp yield point than for the extensive localized flow occurring at the lower yield stress. The primary objective in this paper is to extend the dislocation mechanism to account for localized cataclysmic flow by a dislocation collision process and to give experimental evidence to support such a process. Only the yielding of iron containing carbon -will be discussed, although other metal-solute systems are known to behave similarly. Cottrell Mechanism In brief, Cottrell explains the yield point in the following way: The dislocations in iron which must propagate to produce slip usually lie at the center of local concentrations of carbon atoms, since segregation about these dislocatlons relieves some of the local stress resulting from them. A dislocation surrounded by a "cloud" of carbon atoms is thus anchored, and a higher stress is required to set it in motion than to move a free dislocation. Considering all available dislocatlons to be anchored in this fashion, the iron exhibits a yield point when the first dialocations break free and move through the lattice causing slip. This first breaking away of a dislocation enables other dislocations to break loose by "interaction" and the process becomes a cataclysm producing local deformation or Luders bands. The yield point in the stress-strain diagram for iron is absent in freshly deformed material, but returns gradually with time; the phenomenon is one aspect of what is called strain aging. The rate at which the yield point returns following straining depends on the temperature of aging. According to Cottrell the rate of return of the yield point in strained iron is limited by the rate of diffusion of carbon at the aging temperature, the mechanism is onr: of reforming the solute atmospheres around carbon-free dislocations that had stopped moving coincident with the removal of stress. If the specimen is retested immediately after straining and unloading, carbon will not have had time to diffuse to, and re-anchor, dislocations and the yield point will not occur. The carbon diffusion limitation for the rate of strain aging apparently applies if the criterion for strain aging is either the change in hardness" or the change in electrical resistance" of the strained speci- men with aging time. The possibility exists, however, that the yield point actually returns to strained iron at some rate other than that deduced from hardness or electrical resistance data. Therefore, as a preliminary experiment, the rate of yield point return in a rimmed sheet steel strained 6 pct in tension was measured at 27°, 77°, and 100°C. A plot of yield-point elongation for each of these temperatures against aging time appears in Fig. 1. The aging process is described by curves which rise to a plateau value of elongation that seems independent of temperature, but at a rate that depends on temperature. Very long times lead to a further rise in the yield-point elongation above the plateau value. However, if the later increase in yield-point elongation is ignored and the log of the time to reach half the plateau value of elongation is plotted against 1/T, a straight line results for which an activation energy of about 25 kcal pel- mol may be assigned. Within the accuracy of this sort of experiment this is approximately the activation energy for the diffusion of carbon in iron (20 kcal per mol), and the carbon diffusion limitation suggested for the yield-point return on strain aging is valid. The Cottrell mechanism thus explains in a qualitative manner the occurrence of a yield point in iron and its return with strain aging. It fails, however, to explain some of the other experimental observations that have been made of the yielding behavior of iron. For example, it is known that the yield point in iron becomes less pronounced with increasing grain size. Annealed single crystals of iron have very small yield-point elongations .if indeed they have any,' compared to a polycrystalline steel. If the only requirement for a yield point is that the dislocations in the lattice of the annealed. material be anchored by carbon atoms, the difference in the behavior of single crystals and polycrystals is not explained. That a dislocation mechanism may be entirely consistent with little or no yield point in an annealed single crystal will become apparent later when dislocation interaction is discussed. Strain aging produces a definite yield point even in single crystals. This accentuation of the yield-point phenomenon in single crystals after strain

Jan 1, 1953

By A. J. Mortlock

Cobalt and nickel have been diffused at tracer concentrations in gold at several temperatures in the range from approximately 700° to 950°C. The diffusion penetration profiles were determined by a serial sectioning technique in which the gold is first anodized and then the anodic layer is dissolved in acid. In this ulay sections as thin as 250A could be removed reproduci-bly. In all cases, the region close to the specimen surface was characterized by irregular behavior in the sense that the logarithm of concentration was not linear in the square of the penetration distance. In sotne cases, there zuas an indication of the operation of very slow dijfusion in this region, while in others the apparent diffusion coejj'icient was negative. Possible reasons for this anomalous behavior are briefly discussed. In recent years it has been found that the region close to the surface of a metal can sometimes exhibit anomalously slow diffusion characteristics relative to the interior of the metal. One of the best examples of this fact is the work of Styris and omizuka,' who showed that the apparent diffusion coefficient for zinc in the region withi: about 1 p of the free surface of copper was about ,,,, that at deeper penetrations. This result is particularly interesting, because it is free from the possibly complicating effects of low solubility of the diffusing tracer in the solvent metal. In the case of diffusion under conditions of low solubilitjr, interpretaticn of the results in terms of lattice diffusion is difficult because of the enhanced short-circuiting produced by segregation to dislocations.2'3 Measurements by Duhl et 1. suggest that cobalt diffusing in gold may also show a near-surface effect of this type. Once again the solubility is high, so that this result could be of great interest. However, the technique used for analyzing the diffusion penetration zones by Duhl, viz. the counting of residual gamma activity in the specimen following sectioning, appears to have indicated a near-surface effect in a parallel experiment on the self-diffusion of gold reported at the same time. The latter result is known to be spurious, since Kidson5 has demonstrated that self-diffusion in gold does not show this effect. Duhl et 01. also reported some measurements on the diffusion of nickel in gold, but failed to give any data for the near-surface region. As the solubility of nickel in gold is high, such data would also be of special interest. We, therefore, decided to conduct another set of experiments on the diffusion of nickel and cobalt in gold, using a sectioning technique that allows the individual sections to be assayed for solute content and thus gives direct determinations of penetration profiles. Also, by sectioning with an anodizing/stripping tech- nique, very thin layers can be removed and the region close to the surface studied in detail. MATERIALS The gold specimens were supplied as single crystal disks $ in. in diam by a in. high by Monocrystals Co. of Cleveland, Ohio. The gold itself was of spectro-scopic purity, i.e., better than 99.99 pct pure. METHOD Specimen Preparation. One flat end face of each gold crystal was spark planed with a Servomet spark erosion machine set for minimum spark energy. Following this treatment the crystals were preannealed for 2 to 4 days at temperatures of either 400" or 700°C. The three crystals preannealed at 700°C showed signs of recrystallization. The spark-planed end face of each crystal was then coated with the appropriate amount of 63i or 60 radioactive tracer. This deposit was laid down in a simple plating bath containing the as-supplied solution of the radioactive isotope as well as sufficient ammonium oxalate to saturate the solution. Some ammonium oxalate remained undissolved on the floor of the bath for this purpose. During plating further additions of ammonium oxalate were sometimes required to allow the plating to continue satisfactorily, perhaps due to passivation of the undissolved oxalate already present. The thickness of the deposited layer was determined by comparison of the apparent surface activity of the plated specimen with that of a similar specimen having a weighable deposit of the isotope on its end face. Correction for self-absorption of the radiation was made in this calculation. Annealing. The deposited crystals were annealed in a hydrogen atmosphere in sealed silica tubes. During this heat treatment they were supported, active face down, on optically flat silica plates. The temperature was measured with calibrated Pt vs Pt-10 pct Rh thermocouples, and the tabulated values can be taken to be correct to Z°C. All the crystals showed evidence of recrystallization following these heat treatments, suggesting that initially they may not have been good single crystals or had suffered strain during delivery. Concentration Profile Analysis. After annealing, the crystals were sectioned by the anodizing-stripping technique.6 The anodizing involved suspension of the specimen with its cylindrical axis horiz6ntal by a gold wire in a 200-ml beaker containing 1 M Hg304. A cathode in the form of a strip of gold sheet, 2 in. wide and positioned to be in contact with the curved side of the beaker, completely encircled the specimen. An anodizing current of 30 ma, corresponding to a current density of 5 ma per sq cm on the surface of the specimen, was passed for times ranging from 5 to 150 min depending on the thickness of gold to be removed; the solution was stirred continuously during this process. Following this treatment, the specimen

Jan 1, 1969

By K. W. Nelson, John N. Abersold

INDUSTRIAL hygiene has been defined by Patty' as "the science and art of recognizing, evaluating, and controlling potentially harmful factors in the industrial environment." This definition implies thorough study of operations, evaluation of potentially harmful factors through air sampling, micro-analyses and other means and finally, appropriate medical and engineering control wherever indicated. The prevention of industrial health injuries is a vital part of operations of American industry today. Progress and interest in this field has increased steadily for many years, the most rapid progress having been attained, perhaps, during the last three decades. It is significant to note that there are now official agencies in 46 states actively concerned with industrial health problems and that a western field station has been established recently in Salt Lake City by the U. S. Public Health Service to augment its industrial hygiene services directed from headquarters of the National Institute of Health, Bethesda, Md. Many of the larger industries have found it advantageous to establish their own industrial hygiene departments. The American Smelting and Refining Co. is a world-wide organization engaged in the mining, smelting, and refining of lead, copper, zinc, silver, gold, by-product metals, including cadmium, arsenic, and others. In the United States there are 13 smelters and refineries, 11 secondary smelters or foundries, and a number of mines. Approximately 9000 workers are normally employed. It has long been the established company policy to seek out occupational hazards and provide safeguards for employee health. Protective equipment has been supplied to individual workers and exhaust ventilation installations have been in use in some operations for more than 40 years. All of the major units have their own medical departments which provide employees with excellent medical and hospital care. In 1937 full scale industrial hygiene studies were undertaken at the Selby Plant and were extended to most of the other smelters during the next three years. In 1945 the Department of Hygiene was organized with Professor Philip Drinker of Harvard University as Director and with Dr. S. S. Pinto as Medical Director. The department is responsible for coordinating and maintaining a program for the good health of all employees from top management down to the lowest paid day worker. It is essentially a service organization serving all of the United States plants regardless of location or size. Full and part-time physicians employed in all of the company's American plants and working in close cooperation with the Medical Director are responsible for de- termining the state of health of all the employees and giving treatment when necessary. In general, medical care is confined to accidents or illnesses occurring while the men are on the job. Among the duties of the doctors is the making of careful physical examinations of new employees and routine check-ups of old employees. In addition to medical care a primary responsibility of the department is the prevention of occupational illnesses. In this the main concern is with the working environment in relation to its effect on the worker. Environmental factors may be dusts, fumes, gases, toxic materials, heat, humidity, radiation, or noise. The objectives are: (1) Immediate control of these factors through the education of the worker, through providing the wearing of respirators or other protective devices, and through careful medical examinations and regular analysis of urine specimens; (2) a long range control program which may be accomplished by local exhaust ventilation, wetting of materials, changes in metallurgy, changes in methods of handling, or by use of special devices and special equipment. To accomplish these objectives a fine industrial hygiene laboratory was built in Salt Lake City and equipped to do routine and experimental work. Trained and experienced industrial hygienists obtain the facts by making frequent hygiene surveys. These surveys include tests of the air, studies of all processes, and careful investigation of ventilation, lighting, and general working conditions. Except in emergencies, the air contaminants and often the substances handled by the worker are sent to the laboratory for analysis by chemists and technicians specially trained in industrial hygiene methods. The findings are evaluated in terms of limits recommended by various State and Federal agencies, and in light of all available medical data. The methods used for studying the working environment involve all of the usual chemical and physical procedures employed in industrial hygiene. The Impinger, electric precipitator, thermal pre-cipitator, and filter paper sampler have been used to collect atmospheric dust and fume samples. Of special interest here is the filter paper sampler, shown in Fig. 1, which was developed by Dr. Silver-man at Harvard University. The instrument has been improved and is used very extensively in field studies. A water manometer connected behind an orifice is used to determine the rate of air flow. Calibration is effected by use of a standard gas meter or rotameter. The dust or fume is collected on a filter paper clamped between two rings, as shown in Fig. 2. The filter paper, such as Whatman No. 52, collects both dust and fume with a very high efficiency. The instrument is very convenient and easily transported. The solids collected on the filter paper are analyzed in the laboratory usually by use of a polar-ographic procedure. By this procedure it is possible to measure quantitatively in a single analysis the

Jan 1, 1952

By James R. Stuart

AS the available remaining properties of iron ore reserves on the Mesabi Range are opened up for mining, the various properties located under lake beds are brought nearer an active status. The actual physical problems involved in stripping these properties do not act as a deterrent so much as the legal and political problems that are encountered. When it is proposed to destroy a natural lake that has been used by the public for many years, much local as well as state opposition may be encountered regarding its destruction. Public hearings must be held and some adverse publicity is likely to result. The ownership of the ore under the lake and the rights of the abutting property owners must be settled, and protection from damage caused by a disturbance in surface and subsurface drainage is likely to be demanded by property owners some distance from the proposed mine area. The Embarrass Mine, located near Biwabik, Minn., falls into this classification. A portion of the orebody lies under what was formerly Syracuse Lake, this body of water having been removed in the process of stripping the mine. An additional problem in the case of a meandered body of water is the establishment of a meander line that can be projected downward as mining progresses to form the basis for a satisfactory division between lake bed and upland ore shipments for royalty purposes. Fig. 1 illustrates the complications encountered in maintaining these divisions. A balance point was agreed upon in the center of the lake to make an equable division of lake bed ore to the abutting properties. The entire lake bed has since been adjudged the property of Minnesota. Lake Characteristics Lake bed stripping problems with which this paper is concerned necessarily are limited to a specific type of lake, namely the glacial lakes of the Lake Superior region. One characteristic common to these bodies of water is a deposit of fine black mud or silt on the bottom, frequently underlain by a layer of impervious blue clay. This is also true of the muskeg areas of the region, which present almost identical problems as lakes in stripping. The actual removal of the water and the lake bed material is a routine matter more or less standardized as to equipment, and the period of time required can be estimated easily on the basis of volume and capacity. More important than the foregoing is the execution of preliminary work, and above all, the timing involved. An account could be prepared based entirely on statistical and cost data which would give a very fair picture of the time required and cash outlay needed to effect the removal of a body of water preliminary to stripping the orebody. However, the real interest from the standpoint of the operator and the engineer who carry responsibility for completion of the job lies in the unexpected emergencies and the action of various materials involved in the stripping when the balance has been upset through diversion of water courses and the reduction of the lake level. Runoff and Drainage Lakes are located in natural basins that catch all the rain water and runoff water for a considerable area. Where a lake is involved having an inlet and outlet or a sizeable water course running through it, the drainage area may include a watershed covering many square miles. All available data then must be collected to supply a history extending over as many years for which information can be gathered on the flow of streams, annual rain and snowfall, and most important, the peak flows to be expected. Where the diversion of a stream around the stripping area is a part of the problem, this last factor is of great importance since it controls the cross-section to be selected for the diversion channel and the volume to be removed in its excavation, as well as affecting the hydraulic considerations to be met in the design of the completed channel. Characteristic material in the overburden found at the Embarrass Mine is illustrated in Fig. 2. Well Pumping Pumping from the well holes was started well in advance of the draining of the lake. Fig. 3 shows a gradual lowering of the water table with no noticeable fluctuations during the period in which the lake was being dewatered. Unfortunately, because of tight ground, a maximum flow to the wells was not maintained. This retarded the rate at which the water table was reduced so that in the course of stripping the excavation soon extended below the water table, and the great bulk of the pumping was handled from a system of sumps in the pit itself. Any dewatering program projected by prepumping from wells, a glorified well point system, would have to be started well in advance of the stripping to be of any great advantage. Preliminary drainage of the surface over the mine area is entirely apart from the actual elimination of the lake bed itself. Since the lake is what is called a perched water table because of the impervious character of the lake bottom, the adjoining surface may be dewatered below the surface of the existing lake and the flow will not be affected by the proximity of that body of water. This condition actually has been demonstrated through the establishment of a number of observation holes where a small churn drill was used to put down the holes and a 3-in. pipe was installed for taking water level

Jan 1, 1952

JOINT convention week has become a feature of the year with western mining men. The first was held at Denver in 1926, the second at Salt Lake City in 1927, and the third is about to take place at Los Angeles. Holding in the same city in one week the meetings of a number of related organizations has been .found to permit the development of a well-rounded program at the same time that it economizes time and money for the members, many of whom belong to several of the participating organizations. For the third time, the American Mining Congress, represented by its Western Division, of which Philip Wiseman is chairman, is taking the lead and assuming the main responsibility. The A.I.M.E. is furnishing the papers for two sessions and the California Development Association for one. The Southern California Section, Robert R. Boyd, chairman, is host for the Institute, but the Northern California, Nevada and Arizona sections have each contributed papers to the program, and will be represented at the meeting. It will be a real reunion of the mining men of the Southwest.

Sep 1, 1928

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Jan 1, 1947