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By Philip D. Pearson

CALAND Ore Co.'s dredging program has reached the halfway mark. The goal is to remove more than 160 million cu yd of lake bottom material overlying the iron orebody in Falls Bay of Steep Rock Lake, Ont. Two large dredges were constructed to remove this overburden, a layer of soil about 90 pct varved clay some 400 ft thick. Moisture content of the clay is high-approximately at the liquid limit of the soil. A few years ago, searching for additional reserves of ore, Inland Steel Co. looked toward Canada and the newly developed Steep Rock iron range. To this end an option agreement was signed Dec. 30, 1949, between Steep Rock Mines Ltd. and Caland Ore Co. Ltd., a wholly owned subsidiary of Inland Steel. Several drilling programs were conducted between this date and Jan. 1, 1953, when a 99-year lease was signed covering an area of 1249 acres. This lease calls for ore production to begin in 1960, with full production of 3 million tons per year by 1969.

Jan 8, 1957

By A. Daugiala, D. Coleman

In 2014 the New York State Department of Transportation (NYSDOT) awarded a Design-Build contract to the joint venture team consisting of Skanska which is the managing partner; Kiewit Corporation and Ecco III (SKE) and HNTB as the lead design firm. The design build team employed methods to provide schedule savings and commercial benefits to the owner NYSDOT. This paper presents the design approach and construction methods used to accelerate drilled shaft installation supporting the main span foundation for the cable stayed bridge. The drilled foundation contractor Underpinning and Foundation Skanska, a division of Skanska was responsible for assisting with design, budgeting and construction of the drilled shaft foundation construction. As part of the Request for Proposal (RFP) bid process, Alternate Technical Concepts (ATC) was encouraged by the NYSDOT. The JV team proposed an ATC which involved a scaled production shaft testing method to simulate the ultimate load level, commonly referred to as the “Chicago Method” was used in the rock socket employing the Osterberg Cell (O-Cell) testing procedures. To the authors knowledge this was the first NYSDOT project where this approach was utilized on a Demonstration/Production installation to validate construction procedures and load bearing capacities. Construction procedures, challenges and solutions are discussed in detail. Upon completion of successful drilled shaft construction, concrete integrity is verified using two methods – Cross Hole Sonic Logging (CSL) and Thermal Integrity Profiling (TIP). A comparison from the contractor’s perspective is provided. Finally the drilled shaft rock socket is tested to ultimate end bearing values while rock socket side shear stresses are not exceeding the service level loading. Lateral load test was also a RFP requirement; the test arrangement and performance are discussed in detail as well. The Kosciuszko Bridge location poised unique geotechnical challenges on many levels shifting towards an extreme boundary of parameters encountered in New York City Area for drilled shaft construction. The JV team and specialty drilled shaft contractor actively participating in the shaft design and construction practice coordination and working with NYSDOT yielded a successful foundation solution which provided cost and schedule savings to the project.

Jan 1, 2016

By J. E. Dorn, L. A. Shepard

LOW and Gensamer' demonstrated a number of years ago that the yield point phenomenon in mild steels was associated with the presence of fer-rite soluble carbon or nitrogen. More recently the yield phenomenon in body-centered-cubic metals containing interstitials was rationalized by Cottrell' in terms of a simple dislocation model. Interstitial atoms interact with dislocations in two ways; they cause not only local expansions but induce local tetragonality in the lattice. Consequently, interstitial~ interact with the hydrostatic tension and shear components of the stress about dislocations. They tend to migrate toward the expanded regions of edge dislocations and to assume sites that relieve the shear stresses of screw dislocations. Thus, a dislocation saturated with solute atoms constitutes a lower free energy state than that obtained when the dislocation threads through the average composition regions of the matrix. A greater stress will be required to separate the dislocation from its atmosphere than to move the dislocation through the matrix. This factor gives rise to the upper yield stress, which is required to unleash a series of dislocations in a localized region. This local yielding is propagated across the specimen to form a thin band of plastically deformed material known as a Lueder's band, making an angle of about 45" to the stress axis. Once the band has formed, deformation continues at the lower yield stress by the spreading of the Lueder's band in the direction of the applied stress. Undoubtedly the spreading of Lueder's bands at the lower yield stress is accomplished by the high stress concentrations at the band fronts, which serve to induce continued unlocking of new dislocations in advance of the migrating band fronts. Cottrell and Bilby have shown that the dependence of the yield point on temperature can be deduced by assuming that thermal fluctuations aid the stress in unlocking small dislocation loops from their solute atmospheres. Once a loop that exceeds a critical size has been nucleated, the entire locked diqlocation is released and can migrate. Fisher4 simplified this analysis by assuming that the locking forces were short range, so that if the dislocation loop were displaced only one Burgers vector from its atmosphere, it would be unlocked. Applying his model to the special case of delayed yielding under a constant stress of the order of the upper yield strength, he demonstrated that the delay time 7 for yielding should depend on stress and temperature according to where A and B are constants, G is the shear modulus, and u and T are the resolved shear stress and absolute temperature, respectively. Cottrell and Bilby, Fisher, and Fisher and Rogers' have shown that the above deductions are at least in qualitative harmony with the experimental facts. A number of investigators5-0 have shown that the yield point phenomenon can also be induced in sub-stitutional alloys of face-centered-cubic metals. In general such yield points are not as pronounced as those encountered in body-centered-cubic metals containing interstitials. The yield point phenomenon in these materials is usually enhanced by prestrain-ing at low temperatures and aging at intermediate temperatures. Undoubtedly the yield point phenomenon induced by strain aging substitutional alloys also results from locking of dislocations. But the locking of dislocations in substitutional alloys of face-centered-cubic metals differs somewhat from interstitial locking of dislocations in body-centered-cubic metals. Substitutional elements in face-centered-cubic lattices cause only radial displacements of the adjacent lattice points. Consequently, only the edge components of dislocations can be locked by the mechanism suggested by Cottrell. Additional locking, however, can be obtained by the Suzuki mcchanism.10 In face-centered-cubic metals, dislocations exhibit lower energies when they are present in the

Jan 1, 1957

By Wieland D. R., Hurst R. E., A. R. Hendrickson

Analysis of acidizing techniques, in correlation with reservoir data and a backlog of past treatments, has resulted in the development of a valuable engineering guide for planning acidizing treatments. Such treatments fall into three categories: (I) acid injection into the pores of the matrix; (2) acid injection into natural formation fractures at less than parting pressure; and (3) combination acidizing-fracturing treatments in which acid solutions (without propping agents) are injected at treating pressures sufficient to open and extend fractures through which the acid flows. Because the spending tirnze of acid during a specific well treatment does not change appreciably, maximunl penetration is attained when the first increment of injected acid is completely spent. Additional acid injection cannot be expected to further extend the benefits of the treatment. Depth of penetration will depend upon the reaction rate of the acid under treatment conditions, the injection rate of the acid into the matrix or fractures and the area-volume relationship existing in the flow channels. Based on Darcy's flow formula, extremely low injection rates must be used in order to keep bottom-hole injection pressures below formation fracturing pressure. As a result, only limited penetration of unspent acid will occur. Treatment records indicate that, in most acidizing treatments, formation parting pressures are exceeded, greatly extending acid penetration. Under these conditions, stimulation benefits are limited to the fracture area produced during the spending time of the first increment of acid injected into the formation. This area may be calculated from laboratory and well data to estimate depth of penetration. This, in turn, may be correlated with productivity data to assist The art of gas and oil well acidizing has been characterized by many changes in treating materials and techniques since its inception. These developments have been designed to provide greater production increases. prolong production declines and shorten payout time. Such improvements have been based primarily on data derived from laboratory research and field experience. As more of the variables influencing these treatments have been recognized and evaluated, acidizing has become less of an art and more of a science. Recent studies of fracturing treatments,' in light of individual well conditions and the results of thousands of fracturing treatments, made possible the formulation of an engineering guide that is now being used to select optimum treating techniques and to forecast probable results of such treatments. A similar analysis of the factors controlling acidizing treatments has been made and is the basis for this paper, The findings herein can be used as a guide in the selection of acidizing solutions and techniques, tailored to fit specific well conditions and to provide optimum stimulation per dollar cost. Acidizing treatments may be classified into three basic categories—(1) treatments in which the acid is injected uniformly into the pores and flow channels of the matrix, (2) treatments in which the acid enters natural fissures and fractures in the formation at less than fracturing pressures and (3) injection of acid into the formation at a pressure sufficient to open and extend fractures into the rock through which the acid penetrates (without the inclusion of a propping agent). TYPE 1 — MATRIX ACIDIZING This category consists of treatments in which acid solutions are injected into a homogeneous carbonate

By J. A. Norden

MUCH has been written concerning the hoisting equipment of various mines -throughout the country, but most of the literature on the subject, if not all, describes equipment of extraordinary capacity and the small and medium-sized outfits have been neglected. The hoisting equipment here described is capable 'of handling about 1200 tons per day. It has not been worked, to capacity, but 600 to 700 tons, together with men and materials, have been handled in two shifts without any great effort. Only one shaft extends from the surface to the lowest levels of the mine; see Fig. 1. It is composed of three compartments, two of which are used for hoisting and the third for water, air, and electrical lines, and for sinking purposes. In 1909, this shaft was raised from the 1000-ft. (304-m.) level to the 700-ft. There a station was cut and an air hoist installed. At this time all mining operations were confined to the upper levels of the mine; that is; above the 700-ft. The. ore was-dropped to this level through chutes from the 300-, 400-, 500-, and 600-ft. levels and then lowered in cars on cages to the 1000-ft., which is the main haulage adit, the mill and railroad bins being about 3000 ft. from the shaft. Orebodies were, in due time, developed on the 1000-ft. level and below. Therefore, in 1913, the shaft was sunk to the 1150-ft. level and thereafter sinking was continued at regular intervals until the 1500-ft. level was reached. But, hoisting from the 1300-ft. level was a difficult, and slow process with the antiquated air hoist, so, in 1916, it was derided to raise the shaft to the surface and install a modern, electrically operated hoist. This was completed early in 1917 and the hoist was in operation in March of that year.

Jan 8, 1920

By J Jasper, L Koroznikova, J Begelhole

This paper outlines a comparison of heavy liquid and heavy suspension separation methods for iron ore characterisation.As part of Arrium MiningÆs program to improve its knowledge of ore types being processed in its two hematite ore beneficiation plants (OBP), metallurgical evaluation of hematite feed for the upgrading process has become a major focus.Iron ore deposits almost always contain more than one valuable (and relatively dense) iron mineral. Iron ores are found with a range of iron contents, giving rise to density variations.At present the industry relies on heavy liquid separation techniques for the evaluation of iron ore samples arising from exploration, mining or metallurgical processes. There are only a few high-density liquids available, which are both high-cost and highly toxic.Low-toxicity solutions that can be utilised at relative densities up to 3.0 have been developed. However, highly-toxic organic liquids are required presently to achieve relative densities beyond 3.0.Bromoform which has a specific gravity (SG) of 2.89, and diiodomethane (SG 3.31) are commonly used; both present significant health and safety hazards.æClericiÆs solutionÆ (mixtures of thallium formate and thallium malonate) provide liquids having specific gravities between 4.0 and 5.0. ClericiÆs solution is highly toxic and testing is currently conducted by a limited number of laboratories worldwide.An alternative to ClericiÆs solution and organic liquids is a low-toxicity, high-SG, suspension technique, which has been demonstrated successfully for heavy mineral sands. This heavy suspension method employs mixtures of fine tungsten carbide particles in lithium heteropolytungstate (LST) solutions.Heavy liquid separation is not commonly used as part of routine iron ore feed characterisation due to the expensive, toxic and time consuming nature of the technique. Our results show that the heavy suspensions method can achieve low-cost, low-toxic separation and are outlined and results compared with results using ClericiÆs solution presented.Heavy suspension technique provides a theoretical/absolute separation of a particular sample. By placing an iron ore sample in a suspension of a particular density, it produces two products: a sink fraction and a float fraction. As iron oxides-hydroxyoxides are denser than the gangue minerals associated with the ore, particles that have high iron content will sink and the particles with gangue association will float. This process is repeated over several densities, thus also enabling separation of the constituent iron ore phases. A theoretical grade and a theoretical mass recovery can be calculated. With these calculations, a grade-mass recovery curve can be constructed for any element that is of interest.This type of characterisation can be used to either prepare OBP feed-piles of known processing characteristics, or to benchmark the performance of the process against what can be theoretically achieved on a continuous basis.CITATION:Begelhole, J, Jasper, J and Koroznikova, L, 2013. Application of the heavy suspension technique for ore beneficiation plant feed characterisation from the Southern Middleback Ranges, in Proceedings The Second AusIMM International Geometallurgy Conference (GeoMet) 2013 , pp 269-278 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 29, 2013

By P. J. Petit

Changing mining conditions, legislative issues, rising costs, and the constrained supply of skills and electricity in South Africa, have been some of the criteria forcing mine operators and engineers to review the traditional use of compressed air to power underground hand-held drilling operations in narrow-stope, hard-rock mines. In answering the question, 'What is the best way to power this mine?', they have resorted to trade-off studies with results of limited usefulness, because they are site-specific, and use only a few comparative parameters, based on a small data set. This paper aims to provide definitive, objective and quantitative decision-making approaches to evaluate alternative energy delivery systems for stoping in narrow-reef hard rock mines, based on primary and relevant criteria. These include quantifiable criteria such as capital expenditure, operating costs, resulting financial returns, energy and water usage, production performance, life-of-mine, etc. Other important criteria with qualitative aspects are integrated. These are safety, health and environmental concerns, legislative requirements, business risk, technology change, etc. The paper begins with a review of decision making processes available, and proposes the use of the analytical hierarchy process (AHP) method as a multiple-criteria decision making tool and the decision tree-based on milestone-driven uncertainties, to validate the selection of the energy delivery system. The results of a multiple-criteria questionnaire completed by a variety of industry experts and professionals, such as production and operation managers, consultants, suppliers and techno-financial analysts, are discussed. The evaluation indicates the potential for hydropower to be the best solution for narrow reef hard rock mines, based on current information, mine design layout and production.

Jan 1, 2012

By Tracy C. Jarrett

The alloying of various elements with gold to produce different colors is an old art, but the effect of each element or combination of elements is not generally known. Investigations have been made in this field by F. E. Carter,' E. M. Wise,2 and L. Sterner-Rainer.3 It is the intent of this paper to present in schematic diagrams the effects of various elements upon the color and the melting point of gold alloys. In this investigation 2-oz. melts made in a high-frequency inductiorl furnace were used for all tests. The melting points were determined by the use of a calibrated chromel-alumel thermocouple sealed in a thin-walled alundum tube. The couple was then submerged in the molten metal and the freezing and melting points were determined by using a Leeds and Northrup potentiometer. By checking several melts using copper, zinc and gold, it was found that the accuracy of determining the melting point was within ± 5" F. To determine the differences in colors and to classify them, a wet tissue paper was placed over the polished samples and examined under a uniform white light. The 2 by 45 by 0.045-in. samples were cut from 50 per cent cold-worked strip. The original ingot was 3 by % by % in. This visual comparison, although not scientifically exact, served as a quite satisfactory means for determining and classifying these differences. The base alloys, or "hardeners," were first made up and analyzed, after which the gold and silver were added. This method of making up the alloys was chosen so that the same base alloy could be used for the gold alloys of different karats. The first series consisted of gold, copper, zinc and nickel, the copper comprising from 85 to 65 per cent of the content of the "hardener" and the zinc and nickel replacing copper. The second scries consisted of gold, silver, copper, zinc and nickel, the copper comprising from 85 to 65 per cent of the content of the hardener with the silver, nickel and zinc replacing copper. There was no particular reason for choosing these combinations, except that they gave a color range from red-pink to white.

Jan 1, 1940

By Tracy C. Jarrett

The alloying of various elements with gold to produce different colors is an old art, but the effect of each element or combination of elements is not generally known. Investigations have been made in this field by F. E. Carter,' E. M. Wise,2 and L. Sterner-Rainer.3 It is the intent of this paper to present in schematic diagrams the effects of various elements upon the color and the melting point of gold alloys. In this investigation 2-oz. melts made in a high-frequency inductiorl furnace were used for all tests. The melting points were determined by the use of a calibrated chromel-alumel thermocouple sealed in a thin-walled alundum tube. The couple was then submerged in the molten metal and the freezing and melting points were determined by using a Leeds and Northrup potentiometer. By checking several melts using copper, zinc and gold, it was found that the accuracy of determining the melting point was within ± 5" F. To determine the differences in colors and to classify them, a wet tissue paper was placed over the polished samples and examined under a uniform white light. The 2 by 45 by 0.045-in. samples were cut from 50 per cent cold-worked strip. The original ingot was 3 by % by % in. This visual comparison, although not scientifically exact, served as a quite satisfactory means for determining and classifying these differences. The base alloys, or "hardeners," were first made up and analyzed, after which the gold and silver were added. This method of making up the alloys was chosen so that the same base alloy could be used for the gold alloys of different karats. The first series consisted of gold, copper, zinc and nickel, the copper comprising from 85 to 65 per cent of the content of the "hardener" and the zinc and nickel replacing copper. The second scries consisted of gold, silver, copper, zinc and nickel, the copper comprising from 85 to 65 per cent of the content of the hardener with the silver, nickel and zinc replacing copper. There was no particular reason for choosing these combinations, except that they gave a color range from red-pink to white.

Jan 1, 1940

By Theo Lehner

The Rönnskär smelter has wide experience in the field of recycling of non-ferrous metals. Today, smelting and refining from primary and secondary raw materials are integrated in a complex flow sheet. Boliden operates a lead flash smelter, a lead refinery, a copper smelter and refinery, a sulphuric acid and sulphur dioxide plant, a slag fuming plant and precious metal plant at its Rönnskär smelter in the North of Sweden. The Business Area Smelter Operations also incorporates Boliden Bergsöe, Sweden's lead-acid battery recycler, and Norzink, Norway, where recycled zinc accounts for more than 20% of it's production. Raw materials for the Rönnskär smelter include concentrates, ashes, slags, metallic scrap, electronic scrap and electric arc furnace (EAF) steel-making dust. This paper briefly describes current operations and illustrates current measures to improve productivity. Limits to recycling as experienced by a non-ferrous smelter are, illustrated. The deportment of critical elements contained in raw materials is also discussed.

Jan 1, 1998

C. M. GARLAND,§ Chicago, Ill. (written discussion 11).-7--In the development of Mr. Traer's method for the low-temperature coking of Illinois coals, after the principles laid clown by Prof. Parr, it was soon demon-strated that, in order to yield a product of uniform volatile content, the coal must be coked in thin layers, and consequently must be handled in comparatively small volumes. This means, in the first place, that great care must he observed in the design of the coking chambers to prevent the loss of the products of distillation, and, in the-second place, that labor-saving devices must be utilized to the greatest possible extent in order to reduce the cost of handling.

Jan 12, 1918

By Ligon Johnson

ONLY the acute phase of the smelter fume problem is new. The problem itself is older than the Christian era. While both lead and copper were mined and crudely smelted ; on: 3,000 years ago, it was not until the Roman occupation of the Iberian Peninsula and the British Isles, which occurred but shortly before the beginning of the Christian era, that there was any evidence of smelting operations on a scale sufficiently large to permit a fume problem. Around Huelva, Spain, are found more than 30 million tons of slag from lead smelting conducted there by the Romans. Pliny tells us that more than 20,000 slaves were employed in the Iberian mines. Extensive mining and smelting by the Romans in England and Wales were coincident with the Iberian proceedings. These metallurgical operations were upon a sufficiently large scale to produce marked results both upon the surrounding country and the smelter workers, but, as they were those of a conqueror upon conquered soil, conducted by slaves, imperial Rome failed to recognize that such a thing as a smoke problem did or could arise, and whatever fume question there may have been at that .time remained a question only to those who had no chance to answer any phase of it. The smelting operations of the Romans extended over about 400 years, and little is "recorded of lead and copper smelting from that time until the 16th century. From the revival beginning at this period up to the present century, the growth of smelting has been comparatively gradual. In Great Britain, smelting was conducted almost wholly in localities where metallurgical operations were of paramount importance, and the communities that grew up in smelter localities were due to and dependent upon the mines and smelters. This fact had much to do with the comparative freedom of British smelting from burdensome fume litigation and legislation. To a lesser degree these conditions applied to the German operations which, when coupled with the further fact that most of the early German operations were to a greater or less degree fiscal workings of Prussia, Saxony and Brunswick-Hanover; accounts for

Jan 5, 1917

By Andre Zingano, Anderson Weiss

"The objective of this paper is to study the behavior of a low thick and low depth coal seam and the overburden rock mass. The mining method is room and pillar in retreat and partial pillar recovery. The excavation method is conventional drill and blast because of the small production. The partial pillar recovery is about 30% of the previous pillar size, 7m x 7m. The roof displacement was monitored during retreat operation; the surface movement was also monitored. The effect of the blasting vibration on the final pillar strength had been considered. Due to blasting, the pillar reduced about 20%. The consequence is more pillar deformation and roof vertical displacement. The pillar retreat and ground movement were simulated in a three-dimensional numerical model. This model was created to predict the surface subsidence and compare to the subsidence measured. This study showed that the remaining pillar and low seam reduce the subsidence that was predicted with conventional methods.INTRODUCTIONMining low coal seams (less than 1 meter) with the room-and-pillar mining method implies the excavation of part of the roof or floor to maintain adequate entry height for equipment and compliance with labor regulations with respect to ergonomic issues. The direct consequence of this is an increase in coal contamination and the increased cost of transportation and preparation at the plant. Traditional room-and-pillar mining only uses a development operation, even with a continuous miner or drill-and-blasting. Therefore, all the run of mine (ROM) coal is carried out with part of the roof or floor. However, if the retreat mining method were to be considered, the quality of the coal could increase because only the coal seam is excavated from the pillar recovery. The quality of ROM coal will increase, and the cost will decrease proportionally to the amount of coal produced from the retreat operation.However, the retreat mining with partial or total pillar recovery implies greater subsidence. This is particularly true if the coal seam depth is shallow, under 100m depth. Subsidence prediction in room-and-pillar retreat mining is complex; it depends upon the remaining pillar (stump pillar) size, as well as the pillar height, and the behavior of the pillar after retreat. This is related to the strength of the remaining pillar, which is related to the excavated method (drill-and-blast or mechanical)."

Jan 1, 2018

By J. A. Sutton

In connection with, a stream-pollution-research project, a colorimetric method was developed for concentrations of 1 to 30 p.p.m, of pine oil in water. The method depends on the stable bluish-green color produced when pine oil is reacted with vanillin-hydrochloric acid reagent at 60° C. for 25 minutes. The colorimetric reaction actually determines alphaterpineol, the principal component in pine oil. Some waters may require a special preparation in which ion-exchange resins are used to remove interference. ACKNOWLEDGEMENTS The author wishes to express his appreciation to A. A. Berk, chief of the Industrial Water Branch, Bureau of Mines, for helpful advice and encouragement during the progress of this work and to B. P. Martinez, chief of the Stream Pollution Research Section, for assistance with the manuscript. Thanks are also due to W. D. Bishop of the Hercules Powder Co., Wilmington, Del., for supplying the principal components of pine oil.

Jan 1, 1953

By Harold M. Mooney

WHEN apparent resistivity data are taken with the symmetrical Wenner 4-electrode spread, a fixed center position is used and readings are taken for values of electrode separation. Basic data consist of apparent resistivity plotted against separation of adjacent electrodes . The interpreter attempts to infer geologic structure, such as the depth to discontinuities and the nature of subsurface earth materials. An earlier paper' described methods for interpreting resistivity data. All of these involve a severe assumption, namely that the earth in the region of interest consists of horizontal layers, electrically homogeneous and isotropic. The actual earth never satisfies this assumption exactly and may deviate from it so much that none of the above methods can be applied. Attempts in three directions have been made to modify the assumption so that it approaches known geologic complexity more closely. First, curves have been calculated for dipping discontinuities. Stern' and Aldredge hose a few widely separated dip angles. Trudu confined his attention to small dips. Berel'kovskiy and Zubanov computed gradient curves for widely separated dip angles. Unz as given the most complete solution, with a brief attempt to treat the three-layer case. Second, anisotropy can be taken into account. It seems geologically probable that layered materials have different vertical and horizontal conductivities. Cagniard, Maillet, and Pirson et up methods for finding an equivalent hypothetical isotropic medium. Standard interpretation methods can be applied to this, and the actual medium can then be deduced. Belluigi'" discounts the practical importance of anisotropy; Geneslay and Rouget do not agree with his conclusion. Third, the effect of variable resistivity in a layer can be considered. Keck and Colby" examine the mathematics of an exponential increase in a surface layer. Several authors, for example, Stevenson,'' consider a continuous variation of resistivity with depth. The present paper deals with a linear variation of resistivity in a surface layer. Geologically, surface variations should be expected. Unconsolidated materials such as glacial drift show marked irregularities over short distances. The effects of weathering change with depth. The moisture content of material above the water table may vary from a dry sand to a saturated clay, and both of these will be changed by rainfall. Figs. 1 to 6 present apparent resistivity curves to show the effect of a variable surface layer. In all cases resistivity varies linearly with depth down to a depth of one unit. Material beneath this depth has constant resistivity. Electrode separation is plotted in depth units. Insets on each figure show the corresponding cross-sections, plotting true resistivity against depth. To illustrate, consider curve A of Fig. 1. The true resistivity of material at the surface of the earth is taken as 0.4 units. Resistivity increases with depth, reaching a value of 1.0 at 0.5 depth units and 1.6 at 1.0 depth units. Below a depth of 1.0, all the material has very low resistivity (zero, for purposes of calculation). Curve A in the main part of Fig. 1 shows how apparent resistivity varies for this case as the electrode spacing is increased. To illustrate further, curve E of Fig. 4 corresponds to true resistivity of 1.2 units at the surface, 1.0 at a depth of 0.5 units, 0.8 at a depth of 1.0, and 1.5 for all depths greater than 1.0. Apparent resistivity curves have been plotted logarithmically so that the shape of the curves becomes independent of the units, giving the curves wide validity. A certain drift-covered area, for example, shows a gradual decrease of resistivity from 230 ohm-meters at the surface to 150 ohm-meters at the bedrock surface, 275 ft down; bedrock resistivity is 800 ohm-meters. Curve E of Fig. 5 indicates that true resistivity decreases from 1.2 units at the surface to 0.8 at a depth of 1.0, then increases abruptly to a constant value of 4.0 units. Since resistivity and depth ratios are the same, this can be used to predict the field curve. For Fig. 5, multiply true resistivity and apparent resistivity by

Jan 1, 1955

By I. S. Parrish

The Sullivan Mining Group has been exploring the mineral deposits of the Mount Pleasant area of southern New Brunswick since 1967. Sullivan geologists have logged over sixty miles of drill core, mapped over five miles of underground workings and studied numerous surface exposures. Several mineralized zones have been defined and two higher-grade porphyrytungsten bodies delineated. Within these bodies, wolframite occurs as disseminated specks and grains up to one-half inch in diameter. The host rock is a brecciated rhyolite porphyry that has intruded quartz-feldspar porphyry of Mississippian age . The host rock is itself intruded by an acid porphyry called transition rock and is underlain by a microgranite. The mineralized zones, host rock, transition rock and microgranite are all thought to be derived from the same magma, and to represent progressive fractions from that magma.

Jan 1, 1978

By Garrett R. Hyde

Sintered yttria and sintered dysprosia can be used to make single-component thermistors having typical thermistor geometries and operable in the temperature range 500° to 1,200° C. However, at these temperatures, instability of platinum electrodes can contribute a slight aging effect if the thermistor is kept near 1,200° C for long periods. Typical characteristics at 500° C for an yttria disk thermistor 12 mm in diameter and 2 mm thick are a resistance of 12.4 megohms and a temperature coefficient of resistance of -3.8 percent per degree C. Typical characteristics at 500° C for a dysprosia thermistor of the same dimensions are a resistance of 1.84 kohms and a temperature coefficient of resistance of -1.6 percent per degree C.

Jan 1, 1970

By Khamis Y. Haramy

The U.S. Bureau of Mines installed surface and subsurface instruments at a mine site in Colorado to monitor and characterize the overburden response to longwall mining. Instrumentation included a time-domain reflectometry (TDR) cable, a multiple-point borehole extensometer (MPBX), vertical-deformation surface monuments, and in-mine pressure cells for monitoring abutment, pillar, and gob loads with respect to the longwall face position. The purpose of the study was to collect data that could be analyzed to correlate the design of the mine layout to the resulting overburden failure response. Determining the extent of caving and fracturing is important in the arid and semiarid Western United States because of the possible consequences to the fragile underground hydrologic systems upon which the land owners and users depend. This study is unique from past investigations because the overburden contains massive sandstone members that impose special conditions on the caving process and pressure redistribution above the longwall panels. This paper presents the results of this correlation study.

Jan 1, 1992

By W. L. Remick

WHEN the senior author of this paper presented an article on Fine Coal Cleaning by the Hydrotator Process,1 at the February, 1927, meeting, that process had been developed in the anthracite region only to clean what is known as "slush" or "silt." This size constitutes all of the material passing the finest screen used at any particular colliery. Some use a 3/32 -in. round-hole screen and others use a 1/16 -in. round hole. Practice varies throughout the anthracite region, but the tendency seems to be toward the finer screen.

Jan 1, 1929

By J. L. Lehman, J. D. Sudbury, J. E. Landers, W. D. Greathouse

A full scale field experiment on cathodic protection of casing answers questions concerning (1) the proper criteria for determining current requirments, (2) the amount of protection provided by different currents, and (3) the transfer of current at the base of the surface pipe. Three dry holes in the Trico pool in Rooks County, Kans., were selected for cathodic protection tests. The three holes were in an area where casing failures opposite the Dakota water sand often accur in less than a year. Examination of the electric togs showed the wells to be similar to other wells in the field where casing in four of seven producing wells has failed. The three holes were cleaned out and cased with 75 joints of new 51/2-in. 14-tb J-55. Each joint was visually inspected and marked before it as run. The casing was bull plugged and floated in the hole 50 that the inside might remain dry and free of excessive attack. Also, if a leak occurred, a pressure increase could be observed on gawge at the surface. Extensive testing was done, including potential profiles, log current-potentid curves and electrode measurements from both surface and downhole connections. Based on these data, a current of 12 amps was applied to one well and 4 amps to mother. The third well was left to corrode. During the two-year period when the casing was in the ground, [he applied current was checked weekly, and reference electrode measurements were made about every two months. Three sets of casing potential profi1e.c were run. When the three strings were pulled, each joint was examined for type of scale formed, presence of sulfate-reducing bacteria, extent of corrosion nttnck and pit depth. Since the pipe was new when run, quantitative determination of the protection provided by current was possible. This is the first concrete field evidence to help resolve the many arguments about the proper method for selecting adequate current for cathodic protection of oilwell (-using. INTRODUCTION A casing string is run when a well is drilled. This pipe is supposed to protect this valuable "hole in the ground" for the life of the well. Often the casing does not last the life of the well; it is with these casing failures that this work is concerned. The cost of repairing a casing failure varies from field to field—from as much as a $30,000 per leak average in California to $5,000 per leak in Kansas. Additional costs other than actual repairs are also important. These include formation damage, lost production, etc. Casing damage caused by internal corrosion is important in some areas. Treatment normally consists of flushing inhibitor down the annulus, but further research is being done on control measures. The test described in this paper is concerned only with external corrosion. The problem of casing failure from external attack has appeared in several areas including western Kansas, California, Montana, Wyoming, Texas, Arkansas and Mississippi. Cathodic protection is currently being used in an attempt to control external corrosion. From reports in the NACE there are thousands of wells currently under cathodic protection. The quantity of current being applied ranges from 27 amps on some deep California wells to a few tenths of an amp being supplied from magnesium anodes on wells in Texas and Kansas. Considerable field and laboratory effort1,9,5,6 was exented on the problem of cathodic prctection of casing, and it became fairly obvious that this method could be used to protect wells. Early workers showed that current applied to a well distributed itself over the length of the casing and was not concentrated on the upper few hundred feet. Basic cathodic protection theory had shown that corrosion attack could be stopped by applying sufficient current. The problem resolved itself, then, into one of trying to decide just how much current was necessary. Various criteria were utilized in installing the many existing cathodic protection installations. These methods included the following. 1. Applying sufficient current to remove the anodic slope as shown by the potential profile." 7. Applying enough current to maintain all areas of the casing at a pipe-to-soil potential of .85 v.' 3. Applying the current indicated by a log current-potential (or E log I) curve." 4. Supplying the current necessary to shift the pipe to-soil potential .3 v." 5. Applying 2 or 3 milliamps of current per sq ft of casing."