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By J Arola, B Kettunen, W M. Bond

Separation of fine and coarse particles as separate streams prior to flotation, and their differential treatments, could substantially increase recovery and reduce collector consumption. A series of simulations was carried out to quantify the benefits from plant scale implementation of such a treatment at the Ispat Inland flotation circuit (Minorca mine). The study involved detailed plant sampling and sample analysis to generate the data needed for simulation studies. Samples were taken from all the streams from flotation feed to final concentrate, as well as cell lips and cell pulp. Sampling was repeated for the second ore blend that was processed by the plant during different periods of a year. Sample analysis involved sieve analysis and size by size chemical analysis for total iron, magnetic iron and silica. Raw data were mass balanced, and flow rates within the circuit were calculated. Mass balanced data were used to calculate model parameters representing the existing conditions. Usim Pac was used for both mass balancing and simulation studies. Available pilot and bench scale data were utilised to define the modified flow sheet and expected performance. The simulated flow sheet included classification of flotation circuit feed by hydrocyclones. While the coarse fraction followed the existing flow sheet, the fine fraction was sent to a magnetic separator. Magnetic and rougher flotation concentrates were combined to form the final concentrate. Two simulations for each blend were carried out. The first one assumed that there would be no change in size by size kinetic behaviours of both gangue and magnetite due to this modification. This was named as the worst case scenario, in which particles had a longer residence time due to reduced flow rates within the flotation cells. The second scenario anticipated that there would be an increase in recoveries of gangue and magnetite going into froth launder due to more efficient use of collector in absence of fines. Results showed that pre-classification of flotation feed could improve the flotation recovery by 0.3 - 0.5 per cent, with the major benefit being significant reduction (40 - 60 per cent) in the amount of collector used.

Jan 1, 2005

By M. J. Sinnott, J. C. Tobin

A technique for collecting and analyzing the small quantities of gases desorbed on heating metal powders has been developed. The gases collected from copper powders of various types of manufacture have been analyzed and found to contain H2O, CO2, and SO2. The quantities of gas evolved vary from 0.05 to 1.0 milliliter, under standard con-ditons of temperature and pressure, per gram of powder. The source of the gas has been traced to surface films on the powder. IN the powder-metal industry it is recognized that the presence of gases in the powders affects the processing and resulting properties of the sintered product. Several sources for these gases are possible; namely, chemical reactions with the sintering atmospheres, inclusion. of gases during compacting, dissolved gases in the metal powder, and sorbed gases on the metal-,powder surfaces.1, 2, 3 Since the quantities of the gases are relatively small, no direct attempt has been made to collect and analyze them nor to isolate their source. The work that has been done has been concerned with the dimensional stability of the product., its porosity, density, or mechanical and physical properties. Conclusions are drawn on the basis of studies of this type as to the probable effect of gases on these properties. 4, 5, 0 In the present investigation an attempt has been made to obtain quantitative data on the analysis and quantities of gases associated with copper powders manufactured in different ways. Apparatus was constructed to collect gases evolved from loosely packed powder samples on heating to various temperatures under a reduced pressure of helium. The powders were outgassed thoroughly in a vacuum prior to starting the collection operation in order to eliminate mechanically entrapped gases and to prevent the possible reaction of the powder with these gases. Experimental Procedures Materials—The major portion of the investigation was carried out on a single heat of water-atomized, oxygen-free, high-conductivity copper. A relatively coarse fraction of this powder, 20 to 100 mesh, was tested as a mixed powder and as various graded size fractions. The microstructure of this powder showed equiaxed grains and was clean and free from oxides and oxide eutectic. The particle size, as determined

Jan 1, 1959

By K. Schimann, H. D. Knipping, J. P. Fouques

"The Cigar Lake uranium deposit, in the Athabasca area of north Saskatchewan, was discovered in May 1981 by Cogema (Canada) Ltd., as operator of the Waterbury Lake Joint Venture.The deposit, which is devoid of direct surface expression, is located at the unconformity between the Lower Proterozoic (Aphebian) Wollaston Group meta-sediments and the Middle Proterozoic (Helikian) Athabasca sandstone, at a depth of 410 m to 450 m below surface.At the end of 1984, the deposit had been defined by 105 mineralized holes over a length of 1800 m and a width of between 25m and 105m. It has the shape of flat-lying /em 5 m to 30m thick and shows a remarkable longitudinal and lateral continuity from the information obtained to date. A well defined bulge of the unconformity underlies the mineralized zone over its entire length.The deposit appears to be controlled by an east-west trending structure within which a particular rock facies (""augen gneiss'') has developed in the graphitic metapelites of the Wollaston Group. It is surrounded by a strong alteration halo affecting both sandstone and basement, characterized by extensive development of Mg-AI rich clay minerals (illite-chlorite). The mineralization is hosted principally by the Athabasca sandstone, and consists mainly of uraninite and sulfa-arsenides of nickel and cobalt. Numerous geochemical and geophysical surveys have been carried out in the Cigar Lake area, and although no specific geochemical or geophysical anomaly can be directly associated with the mineralized body, there may be locally a good coincidence.However, a combination of several methods proved extremely useful in interpreting the regional geology and defining the drilling targets within the favourable areas.Although the Cigar Lake deposit shows many similarities with several classical deposits of the Athabasca region, it is characterized by the intensity of its alteration processes and the presence of massive, extremely high grade mineralization.At the end of 1984, the global geological reserves estimated by geostatistics on the main mineralized ""pod"" of the deposit amounted to 110,000 metric tonnes of uranium at an average grade of 12.2% U, with a precision of 25% at the 95% confidence level."

Jan 1, 1986

By Çisem Tuba, Ismail Girgin

"Comparative atmospheric H2SO4 and H2SO4 + H2O2 leaching experiments were performed using a Turkish lateritic nickel ore containing 1.52% Ni. In the leaching tests; acid concentration, temperature and leaching time were considered as main variables and the dissolution behaviour of the ore were studied by Ni, Fe, Co and Mg analyses. The dissolution values in the H2SO4 leaching experiments were observed to increase up to 60% (w/w) acid concentration and a temperature of 95°C. Increase of H2SO4 concentration above 60% at 95°C caused decreases in all the dissolution values and this was attributed to the formation of Fe2(SO4)3 and Fe2(SO4)3•HSO4•H2O type compounds. At the considered optimum H2SO4 leaching conditions ([H2SO4]= 45%, T=95°C, t=45 min); dissolution values reached for Mg, Ni, Fe and Co were determined as 99.12%, 95.58%, 92.63% and 78.37%, respectively. An activation energy of 30.57 kJ/mole was determined for nickel dissolution from the evaluation of the kinetic data obtained for H2SO4 leaching showing that the dissolution was controlled by external diffusion and chemical reactions. Caro’s acid solutions were prepared at H2SO4/H2O2 mole ratio of 3 for varying initial H2SO4 concentrations and leaching tests were performed at 25°C. Co behaved very differently and a dissolution value of 83.94% was obtained even at 5% acid concentration in 60 minutes reaction time. Complete dissolution of Co was obtained after 30 minutes of reaction time at 60% initial H2SO4 concentration and the corresponding dissolution values for Mg, Ni, and Fe were reached to 96.16%, 94.27% and 89.46%, respectively. Although the H2SO4 + H2O2 leaching experiments were conducted at 25°C, it was realized that the temperature of the leaching medium was increased to 122°C after four minutes of reaction time due to probable exothermic reaction(s) at 60% initial acid concentration."

Jan 1, 2014

By Z. Smieszek, Z. Szota, P. Gajowski

"A new method for measuring the total quantity of suspensions contained in the electrolyte during copper electrorefining has been developed by the Institute of Non-Ferrous Metals, Gliwice. Electrolyte purity, and the suspensions content in particular, is of essential importance for the quality of electrolytic copper produced during the process. Suspensions contribute to the increase of slime impurities in electrolytic copper and increase losses in noble metals. Up to now, the laboratory method has been used for the· process control which provides analytical results with considerable time delay. Therefore, it is useless for the process control.The measuring method developed enables continuous measurements of a suspensions content directly within the stream of the flowing electrolyte. The principle of measurement is based on changes in optic properties of the electrolyte with varying suspensions content.1. IntroductionIn copper electrorefining, it is essential that the final product quality be maintained. If suspensions are present in electrolyte, the content of slime impurities in electrolytic copper, (such as Pb, Ag, Bi, Sn, Sb, and As), increases which can result in the occurrence of the so-called slime dendrites on cathodes, thus debasing the quality of the product. As a normal suspension content in electrolyte, the contents in the order of several mg/l are considered, usually not more than 5-6 mg/l, while contents exceeding 20 mg/l are referred to as an alarm state. Suspensions present in an electrolyte are usually a light fraction of anodic slime whose main components are metallic silver, lead sulphate, complexes of antimony and arsenic, and sometimes even calcium sulphate; they are mainly composed of very small particles of lead sulphate and antimony complexes. Sometimes - - in case of alarm states - - the so-called ""wandering slims"" can be observed, which are composed mainly of arsenic and antimony complexes. These ""wandering slims"" make a serious technological disturbance. In normal operating conditions of the electrorefining plant, an increase of suspension content can result from such reasons as crack of filter cloth in a slime filter, turbulent flow of the electrolyte, anodic passivation, aeration of the electrolyte, and improper servicing of an electrolytic tank. It is necessary to continuously control the suspensions content in electrolyte for proper electrorefining."

Jan 1, 1989

The present paper records a chapter in the history of the development of an electrolytic manganese industry in the United States.l A relatively large pilot plant at Boulder City, Nev., for the production of electrolytic manganese was conceived as part of the national defense program to make this country independent of foreign importation of manganese ore. The first authorization, in November 1940, however, was for a plant of only several hundred pounds capacity. Improvements in process and subsequent additions to some sections authorized by subsequent appropriations have enabled the plant to produce as much as 2000 lb. of manganese a day for substantial periods of time. Since in pilot-plant operation frequent changes are made in the flowsheet, the electrolytic manganese plant was constructed so that equipment could be rearranged with a minimum of expense and time. Several different kinds of the various ¦ types of equipment were installed for many of the operations, so that the one most suitable for the process might be chosen. Pilot-plant operations were begun in November 1941 and have been continuous except for brief shutdowns for installation of new equipment. More than one-half million pounds of manganese have been produced. The improvements in process, selection of equipment, and operational procedure developed are recorded here. Development of the Process The work of early investigators on the electrolysis of manganese solutions has been reviewed by Shelton. Until the method of continuous production of manganese by electrolysis was perfected by SheltonZm4 and coworkers, no practical procedure had been devised. The method as originally developed by the Bureau of Mines will be described here only briefly. The manganese dioxide ore was reduced in a rotating kiln-type furnace at 650°C. in an atmosphere of "city" gas. The calcine was leached with spent electrolyte and enough sulphuric acid to produce a neutral leach solutioil containing 2 5 grams manganese per liter as manganese sulphate and zoo grams per liter of ammonium sulphate. Arsenic was removed from the solution by the addition of 0.0j gram iron per liter as ferrous sulphate followed by oxidation with manganese dioxide, and aeration, which precipitated the iron as ferric hydroxide. The solution, free from iron

Jan 1, 1944

By E. C. Harder

CERTAIN typical characteristics of bauxites, ores of aluminum permit their classification into a number of categories, each of which has special features that enable grouping within it deposits from widely scattered bauxite fields. There may be local variations in conditions that have introduced differences of lesser importance between similar deposits in different fields, resulting perhaps from the influence of factors other than those that have produced the distinguishing features of each of the different groups, but in general the similarity between the deposits in each group is sufficiently striking to make the grouping indicated of interest. Several factors and conditions enter into the decomposition process that results in the development of bauxites and laterites and affects their character. The principal conditions favoring optimum bauxitization are: (1) pres- ence of rocks with easily soluble minerals yielding residues rich in alumina; (2) effective rock porosity, enabling easy access and free circulation of water; (3) normal to abundant rainfall alternating with dry periods; (4) vegetation, including bacteria, advantageously distributed; (5) available sources of appropriate solution and precipitation agencies; (6) a tropical or at least a warm climate; (7) low to moderate topographic relief, allowing free movement of the water table but a minimum of erosion; and (8) long quiet periods in earth history. Assuming that all of these factors are operating to a greater or less degree, it seems that the one that affects most the physical character of the resulting bauxite or laterite deposits is the chemical and mineralogical composition of the rocks that by their decay yield the bauxite deposits. This factor is used in the classification followed in this paper. In the process of weathering of ordinary silicate and carbonate rocks of the earth's crust, whether of igneous or sedimentary origin, it is the minerals of the alkalies that yield most readily to attack. The alkaline earth compounds also offer no great resistance, particularly the carbonates. Thus, among the rock-forming elements, those producing the most insoluble compounds are aluminum, iron, and silicon, together with certain ones of less importance, such as titanium, manganese, and phosphorus. In normal weathering processes such as prevail in temperate regions, the common minerals of rock decay are

Jan 1, 1952

By E. E. Mason

The Canadian National Railway system is completing construction of the first major tunnel in North America to use the coarse-aggregate (+ ½ -in.) shotcrete technique of primary support and lining. The tunnel was driven 29 X 20 ft in cross section, 10,760 ft long, passing under an industrial and residential area of Vancouver. With the exception of 1400 ft in soil (mainly impervious tills), the tunnel passes through a very young series of flat-lying sedimentary beds, to a large extent so incompletely consolidated as to fall apart on exposure. For a conventional steel-support system, it was estimated 8 WF, 28-lb steel should be used at 5-ft centers. Cost of erection including lagging would amount to $120 per linear ft of tunnel at 304 per lb. With 8-in. concrete cover on the steel, lining would cost $280 per linear ft at $82 per cu yd. Thus, minimum cost of support and lining would have amounted to $400 per linear ft. It is now estimated that at least 90% of the rock tunnel would have required support and 100 % required lining. Specified shotcrete support was 6 in. in the arch and 4 in. in the walls. Cost of this work was roughly $100 per linear ft. In addition, another $10-$12 per ft has been incurred in drainage and waterproofing wet areas and in measures to correct experimental and operational errors, thus finalizing the support as a lining (Fig. 1). The tunnel was driven three shifts five days weekly, no blasting being allowed by city ordinance between midnight and 7.00 a.m. A six-drill jumbo was used to drill a 10-ft round, 110 holes blasted three times daily. A 2-in. layer of shotcrete was applied to the newly blasted arch commencing within 45 min of blasting. This was done from a flying deck that extended over the muck pile from the jumbo (Fig. 2). Also during the mucking cycle, the preceding incomplete arch support was brought up to 6-in. specification. The walls were sprayed 4 in. thick during the drilling cycle (Fig. 3). The shotcrete machines and aggregate bins were mounted on the jumbo.

Jan 1, 1970

By Alden B. Greninger

In a recent study1 of the deformation of metastable beta copper-zinc and beta copper-tin crystals, it was established that the parallel markings that appear on the surface of these crystals after slight deformation represent lattice transformations. Stereographic analysis showed that in beta copper-tin, the poles of these markings deviated one or two degrees from poles of the form {133}. In beta copper-zinc the plane delineated by the markings corresponded to no one crystallographic plane of the matrix lattice; the marking poles deviated from 5° to 11° from poles of the form (110), and were displaced only a few degrees from (110) zones. Inasmuch as the new transition lattices produced as the result of the deformation are different for each alloy, it was thought at first that the habits of the markings might be the result of a cooperation between matrix lattice and transformed lattice. (Both matrix lattices arc body-centered cubic.) Such a cooperative effect is thought to obtain in the formation of Widmanstätten figures, for it is well known that the matrix lattice alone does not determine the habit assumed by a segregate phase. On the other hand, it is not inconceivable that the dissimilarity in marking habit might be due to differences in deformation characteristics of beta brass and beta bronze. Barrett and co-workers2 have shown that slip-line habits in pure iron are markedly different from those in high-silicon ferrite; and although it has been repeatedly demonstrated that plates of a segregate phase do not necessarily form on the slip planes of the matrix lattice, deformation characteristics may be more intimately allied with plate formation when a transformation is made to occur as a result of plastic deformation. However, there is the fact that the orientation of markings in metastable beta copper-zinc is the same whether the markings be produced by quenching, deformation, or liquid-air cooling; likewise, quenching and deformation markings in metastable beta copper-tin are oriented in the same way with respect to the beta matrix. During the course of the investigation referred to above, samples of 52:48 (stable beta) copper-zinc were deformed and examined micro-

Jan 1, 1938

By Alden B. Greninger

In a recent study1 of the deformation of metastable beta copper-zinc and beta copper-tin crystals, it was established that the parallel markings that appear on the surface of these crystals after slight deformation represent lattice transformations. Stereographic analysis showed that in beta copper-tin, the poles of these markings deviated one or two degrees from poles of the form {133}. In beta copper-zinc the plane delineated by the markings corresponded to no one crystallographic plane of the matrix lattice; the marking poles deviated from 5° to 11° from poles of the form (110), and were displaced only a few degrees from (110) zones. Inasmuch as the new transition lattices produced as the result of the deformation are different for each alloy, it was thought at first that the habits of the markings might be the result of a cooperation between matrix lattice and transformed lattice. (Both matrix lattices arc body-centered cubic.) Such a cooperative effect is thought to obtain in the formation of Widmanstätten figures, for it is well known that the matrix lattice alone does not determine the habit assumed by a segregate phase. On the other hand, it is not inconceivable that the dissimilarity in marking habit might be due to differences in deformation characteristics of beta brass and beta bronze. Barrett and co-workers2 have shown that slip-line habits in pure iron are markedly different from those in high-silicon ferrite; and although it has been repeatedly demonstrated that plates of a segregate phase do not necessarily form on the slip planes of the matrix lattice, deformation characteristics may be more intimately allied with plate formation when a transformation is made to occur as a result of plastic deformation. However, there is the fact that the orientation of markings in metastable beta copper-zinc is the same whether the markings be produced by quenching, deformation, or liquid-air cooling; likewise, quenching and deformation markings in metastable beta copper-tin are oriented in the same way with respect to the beta matrix. During the course of the investigation referred to above, samples of 52:48 (stable beta) copper-zinc were deformed and examined micro-

Jan 1, 1938

By M. W. Hayward

THE object of this paper is to record and tabulate the data and field observations obtained by the writers and their associates during 10 years of intensive study of lead-zinc deposits in the Cretaceous limestone areas in Northern Mexico, particularly in mines owned by the Cia. Minera de Peñoles, S. A. At the El Paso meeting of the Institute in October, 1930, J. G. Barry made the suggestion that a committee be appointed to correlate the ideas and theories which have been put forth by the numerous engineers and geologists who, during recent years, have made a study of the ore deposits in the area under consideration. The first step in a work of this nature would be to record all of the observed facts, and these facts would have to be accepted by those who attempted to formulate a theory of ore deposition. It is with this idea in mind that the writers have described as accurately as possible the actual, observed facts pertaining to the occurrence of the lead-zinc ores in dolomites or dolomitic horizons. By most writers the term "dolomite," is loosely used to designate magnesium-bearing limestones, regardless of the percentage of MgO present. Since no, true dolomites exist in the areas to be described, we shall also loosely use the terms dolomite or dolomitic limestone to designate those beds having a magnesia content of 5 per cent MgO or more. For a number of years the writers have been impressed with the fact that in areas where dolomite beds occur a large percentage of the lead ore is found in these beds, in spite of the fact that such beds may be overlain or underlain by pure limestone, which one would assume to be just as favorable for replacement as the dolomite, if not more 'so on account of its being more soluble in acid solutions than the dolomite. It is true that in some limestone areas, such as the Erupcion-Ahumada districts, Chihuahua, and Providencia, Zacatecas, where no dolomite beds have been found, large tonnages of lead ores have been mined. We venture to predict, however, that had there been dolomitic beds in these areas, the

Jan 1, 1931

By P. C. Benedict, J. A. Titcomb

"The following description of the geology of the Northern Empire mine is recorded by the senior author, but is based on the observations and complementary mapping of both authors.The Northern Empire mine is a mile east of Beardmore, and 100 miles northeast of Port Arthur, Thunder Bay district, Ontario.Production started in March, 1934, and terminated in September, 1941. The total number of tons milled was 425,866, from which 153,103 oz. gold was recovered. The lowest level from which ore was mined was the 1,900- foot, but considerable drifting was performed on the 2,075-foot and 800 feet on the 2,415-foot, the deepest level,The vein strikes about N.72°E. and has an average dip of80°S. In stoped sections the quartz averaged about 2 feet wide. On the 1,725-foot level it was drifted on continuously for more than 7,000 feet. In the mine as a whole, practically no commercial ore was found beyond points 1,400 feet west and 1,800 feet east of the shaft. Conspicuous wall-rock alteration extended throughout this section, but not much beyond these limits. There was some splitting of the vein and widening of the alteration halo on the lower levels. The average general colour of the quartz changes with depth .The vein is in 'Keewatin' lava flows, to which it is almost conformable. It lies only a short distance south from the 'Timiskaming-Keewatin' contact ( Fig. 2). This contact is marked by a black graphitic fault. So far as is known the fault is parallel, or nearly parallel, with the flows in strike, but dips at 72°S., which is 5° flatter than the average dip of the overturned flows. A post-vein, almost flat sheet of diabase about 550 feet thick cuts outthe vein from about the 650- to the 1,200-foot level ( Figs . 1 and 3). If it was intruded along a fault, the fault had only minor displacement. A strong alteration halo extends to 200 feet above and below the diabase. The vein was conspicuous but in large part it was below ore grade for 100 feet or more above and below the diabase.-"

Jan 1, 1954

By Brian T. Nereson, Michael C. Penney

"This technical paper was prepared to describe the general process by which the authors designed a column-supported load transfer platform (LTP) for support of asphalt-paved runway and taxiway embankment areas over soft organic soils, the major tasks and observations made during construction of the system, and the outcome of the system design and construction. The LTP system was designed using the method described in the BSI Standards Publication Code of Practice for Strengthened/Reinforced Soils and Other Fills, BS 8006-1:2010 (BS8006), to achieve performance objectives for the project. The design also used other criteria based upon the authors’ engineering experience and understanding of the expected interaction between soils, structures and geotextiles. The design and construction process described herein resulted in a successful project. The authors identified that full-time engineering oversight during construction was of high importance and provided a high degree of value to the project. The authors’ intimate knowledge of the potentially adverse effects of variations in construction practices promoted timely identification of construction deficiencies that could have affected the reliability of the system. This resulted in corrections and modifications being made, while materials were still accessible, to protect the intended performance of the finished system.INTRODUCTIONThe site is located in Marshfield, Massachusetts. The surrounding area is flat and contains significant wetlands with a water table at or near the ground surface. Prior to the improvements described herein, the site was already developed as an existing municipal airport with an approximately 3,000 linear-feet asphalt-paved runway oriented in a northeast/southwest direction (northeast end referred to as Runway 24 and southwest end referred to as Runway 6; Fig. 1). The site is located approximately 0.5 to 1 mile (Runway 24 and Runway 6, respectively) from the Atlantic Ocean.The improvement project included total extension of approximately 550 linear feet of asphalt-paved runways and taxiways over existing wetlands that consisted of soft, organic soils with varying but significant depths (Fig. 2). An initial geotechnical assessment performed for the project identified that the subsurface conditions were not suitable to support the proposed runways and taxiways in their existing state."

Jan 1, 2016

By G. M. Proudfoot

IN June, 1940, a .Longyear 3420 blast-hole drill rig was purchased by the Hudson Bay Mining and Smelting Company for the purpose of experimental blast-hole drilling. The early work showed that the diamond drill could compete successfully with the Leyners then in use and by January 1st, 1941, the diamond drills were placed on contract. They may be considered to have gone into production at that date. Prior to the introduction of the diamond drill, our bench drilling was clone with Ingersoll-Rand S-70 and S-83 Leyner drills. Holes up to 40 feet were drilled, using coupled 1 1/4-inch round steel. Bits with a gauge of 3 inches were used in starting these holes, with a drop of 1/8-inch gauge for each 2-foot run. The holes normally ended with 1 1/2 or 1 5/8-inch gauge, though bits down to 1 1/8 inches were available for holes longer than normal or for holes drilled in abnormally abrasive ground. This method of drilling was very satisfactory and was a distinct improvement on the three-bench method of drilling used previously (1). During 1937 and 1938, when experienced miners were available, Leyner long-hole drilling was carried on with great efficiency. However, in 1939 and 1940, labour turnover in-creased and experienced long-hole drillers became scarcer. This resulted in a marked decline in the efficiency of the drilling operations, with increased costs and less tonnage drilled per shift. EXPERIMENTAL WORK The initial diamond drilling test work was carried on with a Longyear 3420 air-driven blast-hole drill rig, drilling a 1 3/16-inch hole. Tests indicated that diamond drilling could be successfully adapted to local conditions and that overall costs compared favourably with costs obtained with the Leyner long-hole method. Leyner long holes on stope benches usually bottomed at a larger diameter than the 1 3/16 -inch diamond-drill holes. More powder could therefore be introduced into the Leyner holes than could be placed in the diamond-drill holes. To achieve the same fragmentation, additional footage was required when diamond drilling. Diamond-drill holes of larger diameter were indicated.

Jan 1, 1943

By G. F. Loughlin

MORE than 20 years have passed since the publication of Lindgren and Ransome's report on the Cripple Creek District,1 which was made when the district was much more active and prosperous than in recent years. They concluded that the output of the district, which had reached its maximum of about $18,200,000 in 1900, would gradually decline, and their conclusion has been confirmed by records of subsequent production. Between 1905 and 1915 annual declines were partly compensated by encouraging increases, especially in 1908, 1914, and 1915; but a rapid continuous decline followed, due largely to the war and its after effects, until in 1922 the value of output was only slightly more than $4,000,000. It rose to nearly $5,000,000 in 1924, but declined again in 1925 and in 1926, when it was about $4,400,000, and represented ores mainly from deep levels in the Cresson and Portland mines. The profitable operations of these two mines were the principal factors leading to a study of ore at deep levels in the district by the U. S. Geological Survey in coopera-tion with the Colorado Metal Mining Fund. The principal reasons for Lindgren and Ransome's conclusion were three: (1) the greater cost and difficulty of deep exploration; (2) the decrease in the number of veins with increasing depth, and (3) the probability that the telluride ore, believed to have formed at comparatively low temperature, would not persist indefinitely -in depth. Operations since they made these statements have verified the first two, but call for some modification of the third. Rich telluride, ore is being mined at a profit in the Portland mine 3000 ft. below the present surface, or 5000 ft. below the original surface of the Cripple Creek volcano, as interpreted by Lindgren and Ransome. The question arises, therefore, as to the number and extent of such deep deposits, and the answer, so far as it can now be given, must be based mainly on structural data, supplemented by data on the processes of ore deposition. Presentation of these data based on a study of deep mines in the southeast quarter of the district, the only

Jan 1, 1927

By Judy Che, Steven Ceccio, Gretar Tryggvason

"A detailed understanding of drop deformation and solidification is essential for quantitative predictions of geometry and properties of parts made by ballistic particle manufacturing processes. Axisymmetric results of towers built by the deposition of multiple drops are presented. In these computations, fluid flow, heat transfer, and phase change are fully coupled throughout the drop deformation and solidification process. The solution method is based on a single set of conservation equations written for all phases simultaneously. The phase boundaries are treated as imbedded interfaces by adding the appropriate source terms to the conservation equations. A parameter study demonstrates the effect of Weber number, Peclet number, and drop deposition frequency on the final shape of the structure. Numerical results show that most successful method of building uniform towers is when two conditions are met. First, each drop solidifies at the same time in the deformation cycle. Second, the cooling rate is slow enough to allow for good adhesion between drops. The ability to predict the final shape of deposited drops will greatly reduce the amount of experimentation necessary to select the desired set parameters for any artifact.IntroductionMaterial deposition techniques based on layering small molten metal drops holds the promise to build small, coherent three-dimensional structures. In this process, numerous 20-100 µm diameter drops are deposited onto a substrate with a low impact velocity (on the order of 5 mis). Both the momentum and trajectory of each drop are controlled precisely to build up complex parts. This material layering technique differs from processes such as spray forming and splat quenching which involve high impact velocities (on the order of 100 mis) and high frequency deposition with little control over individual drop placement. This process of layering small drops can produce parts with near net shape accuracy, while spray forming produces slabs of material which must then be worked to make a usable part. The ability to computationally predict final drop shape and cooling rate has the advantage of allowing the user to select parameters to produce parts with desired shape and mechanical properties. However, the development of such predictive methods requires a thorough understanding of both drop deformation and solidification"

Jan 1, 1999

By Daniel N. Beshers, Ralph H. Condit

An expression for the diffusion coefficient of inlerstitial atoms in the bcc lattice is derived which takes account of jumps between tetrahedral and octahedral interstices. The results are applied to the case of carbon diffusion in iron to find values for jump frequencies of the types, octahedral-tetrahedral, tetra-hedral-tetrahedral, and tetrahedral-octahedral. ThERE has been renewed discussion1-5 concerning the mechanism of diffusion of interstitial atoms in the bcc lattice. This lattice has both octahedral and tetrahedral interstices, and it has been pointed out that the interstitial atoms may make jumps of the octahedral-octahedral type, hereafter called o-o, octahedral-tetrahedral, o-t, and tetrahedral-tetrahedral, t-t. Doubt has been cast on the applicability of snoek's8 model for internal friction to some interstitial migrations in bcc lattices wherein he proposed 0-0 jumps as the principal mechanism of migration. Also, the anomalous excess diffusion of carbon in iron at high temperatures4 has been ascribed to t-t jumps,2,3 even though the sites normally occupied are octahedral. In this paper an expression is derived for the diffusion coefficient, taking account of the three types of jump mentioned. The relationship of the coefficient to the corresponding internal friction relaxation time is found for the limiting case corresponding to Snoek's model. These results are applied to the case of carbon in iron. In Fig. 1, the structure is shown with the locations of the interstices. Each o site has four t site neighbors coplanar with it, and each t site lies halfway between two o sites. There are four t site first neighbors of a t site lying at the vertices of a tetrahedron. Let the average fraction of atoms occupying o sites and t sites be No and Nt, respectively. Since there are six o and twelve t interstices per unit cell having a lattice parameter, a, the concentration of interstitial atoms will It will be assumed that three types of jumps may be distinguished, t — t, t — o, o — t,for which the jump

Jan 1, 1968

By E. J. Walls, W. C. Joughin

"In selecting a suitable method to sink a vertical shaft for underground access, a number of constraints influence the ultimate decision of where and how to develop the shaft, not least among these being safety, development and construction time, and cost. Two additional considerations stand out: these being geotechnical conditions and technology, the latter taking into account existing underground access. Assuming a project for which an existing underground excavation is available, it is tempting to build a shaft sinking project from the outset based on the raiseboring method, which has the potential to be the safest, fastest, and least expensive method provided that geotechnical conditions permit.And therein lies the rub: regardless of project time or cost constraints, when it comes to raiseboring a long (say, greater than 500 m), large diameter (greater than 4.5 m) shaft, the rock mass conditions ultimately dictate what method of shaft sinking will be feasible. Over the course of several studies for a particular project, several geotechnical analyses were carried out specifically for the purpose of developing a shaft by raiseboring. Risk analysis and experience showed that where the rock mass conditions indicated an unacceptably high risk potential, an alternative method needed to be considered, even if this meant increasing both the time and financial requirements.In this paper we present an overview of geotechnical investigation practices for shaft sinking. Decisionmaking thresholds for raiseboring or other methods of shaft sinking are discussed, including probabilities of failure, empirical rock mass classification, basic wedge failure, and back-analysis of a failed case. The design of appropriate support, and analysis of relative safety benefits for various shaft sinking methods, falls outside the scope of this work, and will be presented in a separate paper.IntroductionThe potential for rock mass failure during shaft development and construction affects whether a project is completed safely, on time and on budget. During the course of several studies to establish ventilation shafts for an underground expansion project, geotechnical analyses of rock mass stability and failure potential were carried out. The investigations were guided by the method of McCracken and Stacey (1989) together with insights from Peck and associates (Peck, 2000; Peck and Lee, 2007, 2008; Peck, Coombes, and Lee, 2011)."

Apr 1, 2019

By F. F. Aplan, D. W. Fuerstenau

This chapter presents modern concepts of the flotation of nonmetallic minerals. In particular, it is concerned with such minerals as metal oxides, silicates, sulfates, carbonates, and halides. It is not the purpose to duplicate the two excellent, recent treatises by Gaudin l and by Sutherland and ark,^ but rather to expand their correlations of data, particularly in view of recent, ex- citing new experiments which are shedding new light on the principles of non- metallic flotation. Outstanding success has been achieved over the past 50 years in separating one sulfide from another or from the gangue constituents using a minimal amount of reagents. In contrast to this, selective flotation in nonmetallic systems is often difficult in that one class of compounds alone, the carboxylic acids, may float nearly the entire array of nonmetallic minerals under the appropriate conditions. Moreover, while lime, cyanide, sodium sulfide, and copper sulfate constitute the principal modifiers used to obtain selectivity among the various sulfides, the nonmetallic mineral flotation engineer is con- fronted with a vast array of modifying reagents. In sulfide and native-metal flotation, a strong chemical adsorptive bond is generally formed between the mineral and sulfhydral collector. In fact, re- cent radio-tracer studies show that below monolayer coverage as much as 98% of added mercaptan is adsorbed on the surface of gold.? Contrasted to this, Bloecher's data indicate that for the quartz-amine system only about 60% of the collector is adsorbed.4 This greater residual concentration in the solution results from weaker forces of adsorption in the quartz-amine system. In the flotation of nonmetallic oxide minerals, collector adsorption results from rather weak physical forces, primarily electrostatic attraction and van der Waals attraction between hydrocarbon chains. In the flotation of salt-type minerals, carboxylic acids are adsorbed both by physical and chemical forces. It is a study of the effect of the electrical forces superimposed on the chemical adsorptive forces that contributes most toward understanding nonmetallic mineral flotation. The need for quantification of general rules for the flotation of nonmetallic minerals is so great that Taggart 5 with unmasked sarcasm suggested that after trying anionic and cationic collectors:

Jan 1, 1962

By Lawrence Tsang, Conrad Ng, Tim Leung, Peter Poon

"The Mined Tunnel is part of the Central-Wan Chai Bypass and Island Eastern Corridor Link (CWB), which is a strategic road along the northshore of Hong Kong Island targeted to alleviate the heavy traffic congestion. The objective of this paper is to review the performance of construction of a 160m long Mined Tunnel with a trinocular cross-section area of 460m3 and a combined span of about 50m underneath the southern approach ramp of the existing Cross Harbour Tunnel (CHT) by sequential excavation method (SEM) in hard rock. Mechanical excavation of tunnel by adopting the New Austrian Tunnelling Method (NATM), together with the application of NGI Q-system allows observational approach and optimization of temporary rock support design. Furthermore, the good performance of optimized rock support and the stability of CHT are confirmed through instrumentation monitoring review. This paper also discusses the selection of different tunnel plant and equipment to achieve the planned rates of tunnelling by mechanical excavation in competent granitic rock. INTRODUCTION The works of the CWB mainly comprise construction of 4.5km dual three-lane trunk road, of which a 3.7km section is a tunnel connecting Rumsey Street Flyover in Central to Island Eastern Corridor at North Point (see Figure 1). The Mined Tunnel runs from the ex-Public Cargo Working Area (ex-PCWA), underneath the existing CHT, and through the Causeway Bay Typhoon Shelter. The Mined Tunnel passes at an oblique angle beneath the southern portal of CHT over a length of some 50m. CHT is the primary sensitive receiver of this project and is utilized with a daily throughput of about 120,000 vehicles that far exceeds the design capacity of 78,000 vehicles per day. Hence it is necessary to choose the most appropriate tunnelling method to preserve the structural integrity of the CHT. Other sensitive receivers in the vicinity include structures to the north of the Mined Tunnel, the Royal Hong Kong Yacht Club (RHKYC); and to the south of the Mined Tunnel, the Police Officers’ Club (POC)."

Jan 1, 2016