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By George M. Potter

During the late 1960s the mining industry became interested in USBM's efforts to develop the gold-silver heap leaching process. In the last 10 years, successful commercial heap leaching applications increased. The process provides a relatively low-capital means of processing treatable ores with reasonable operating costs. Present gold and silver prices compared to capital and operating costs favor the continued growth of the heap leaching method. The heap leaching process is defined as the percolation leaching of relatively coarse gold-silver ore piled on an impervious surface. To date, dilute alkaline cyanide solution is the lixiviant of choice. Other leaching reagents such as thiourea and hypo have been suggested, although they are not currently used in heap leaching practice. Ores may he run-of-mine in size or crushed as fine as about 10 mm. Gold and silver-bearing so¬lutions draining from the heaps may be processed in activated carbon columns to recover the metals, followed by stripping and metal recovery, or treated by the Merrill-Crowe zinc dust precipitation method. All solutions are generally recycled to the leaching step, and the heap leaching system is relatively easy to control from an environmental viewpoint. One advantage of the heap leaching process is its flexibility. Plants may be very simple, comprising leaching of uncrushed ore on a locally-obtained clay-lined pad and shipment of gold and silver recovered in the form of loaded carbon or zinc precipitates; such a plant would have a relatively low capital investment. More elaborate plants for larger ore bodies may be planned to include crushing, cyclic pad use, and done gold-silver recovery. General planning is similar in each case. This article outlines some key factors in designing heap leaching operations and discusses how these factors are applied to develop a working operation. Since individual design requirements and local costs vary so much, only a few general comments on costs are included.

Jan 3, 1981

By C. K. Leith

(San Francisco Meeting, September, 1915) IN 1911, we published in the Transactions a brief account of the lateritic alterations of serpentine in eastern Cuba, producing the important iron-ore deposits of the Mayari and Moa districts.1 The special feature of that article was a quantitative treatment of the alterations, based on analyses, to show just what had happened in terms of chemical, mineralogical, and physical changes. Professor Kemp has published an interesting account of these ores, containing additional quantitative observations of his own, tending essentially to confirm the nature of the changes that bad been worked out.2 His discussion suggests to us the desirability of publishing certain additional quantitative data which were not included in our previous article. This seems especially desirable because quantitative measurements of lateritic alterations are rare, and this particular case may be regarded as typical of a fairly wide range of alterations of this kind. At the time of our examination of Mayari and Moa, deposits 29 complete analyses were supplied us by the chemists of the Spanish-American Iron Co., representing a graded series from the unaltered serpentine rock below to the ore at the surface. These analyses are given in the accompanying table. The mineralogical composition of the ores and rocks is comparatively simple and can be easily calculated from the analyses. In Fig. 1 the mineralogical changes from the serpentine rock below to the lateritic iron ore above, based on the analyses, are expressed in terms of weight, assuming alumina to have remained constant during the alteration. It is not easy by mere inspection of this table of analyses to determine what the progressive chemical changes have been in development of the ore. In order to bring this out clearly, the analyses have been graphically represented in Fig. 2 in such a fashion as to show the progressive additions and losses of substances during alteration.

Jan 7, 1915

By Robert S. Mayo

My subject today is "Drilling-and-Blasting of Small Rock Tunnels". I am sure that there will be some people, primarily college students but perhaps Consulting Engineers, who will say: "Drilling-and-Blasting? I thought that went out with the invention of the Tunnel Boring Machine". A few years ago I talked to Dick Robbins and he told me that they did not try to sell a "Robbin's Mole" unless the tunnel was two miles long. For easy calculations I use 10,000-ft. (3km) A Mole today costs about $2 million which works out to $200 per lineal foot. This is "front money". You cannot place the order unless you are the uncontested low-bidder and then you must wait at least six months for delivery. Of course you can use that time to square up your portals, ready to turn-the-eye, or maybe you have shafts to sink to reach the tunnel grade. But this is "dead-time" and it is a heck of a feeling to have a job and cannot get started on it until you get the Mole. I do not have the American figures on lengths of tunnels but there was an article a few years ago in a British magazine: NEW CIVIL ENGINEER, on this subject. This stated that 75% of the jobs in Britain which were bid were for small tunnels, 10-ft x 10-ft (3m x 3m) or smaller, and that they were less than 5,000-ft (1.5km) in length. This does not mean that 75% of the footage in Britain was less than 10-ft in diameter. It means that 75% of the jobs being bid were less than a mile in length. From my experience in the U.S. I believe that we have substantially the same ratio: so many of the jobs with which I was associated were short, 1,000-ft, 2,000-ft and sometimes 3,000-ft and the bores

Jan 1, 1981

By A. J. Strane

PRACTICALLY all explosives may be classified as those that burn and those that detonate. Explosives of the first class include .black blasting and smokeless powder and are broadly known as low explosives, or propellants, because they merely burn and do that comparatively slowly. Their action is as if some power behind the projectile or other material were exerting a shoving or pushing action. The high, or detonating, explosives produce a smashing or shattering effect upon the material. In using military- explosives of the propellant class, the breach of the gun, or cannon, is proportioned to hold the pressure, so that the projectile will be blown out. Exactly the reverse is true in commercial work; then it is the "breach" or the hole that is to be broken. The projectile and the explosive are therefore chosen that will rupture the largest "breach" possible. With the high explosives, it is easy to break out the breach (that is why they are of no value as military propellants)" but with some of the low explosives it is necessary to plug up the muzzle very carefully before shooting. Doubtless the use of incendiary mixtures of oxidizing and combustible materials led, in the 13th century, to the perfection of the mixture which is practically the same as our commercial black blasting powder of today. About 600 years later, nitroglycerine was discovered, but for years its great potential energy could not be used because, in its liquid state, it could not be safely handled or transported. The accidental discovery that kieselguhr would absorb, and hold fairly well, three times its own weight of nitroglycerine, however, made it available. "Guhr dynamite," as the combination nitroglycerine and kieselguhr was termed, was soon displaced by dynamites having an active base, such as wood pulp and other carbonaceous combustible materials. These do all that kieselguhr will do and, in addition, increase the chemical reaction and add materially to the effectiveness of the blast.

Jan 8, 1920

G. SHERBURNE LAGERS (written discussion*).-Mr. Kennedy's discussion1 of Mr. Matteson's paper takes the form of a criticism of my own comments2 on this paper. Mr. Kennedy is a respected authority on matters pertaining to the salt-dome oil fields, and his characterization of my "elaborate figures" as "worthless" is sufficiently candid and comprehensive to warrant a brief reply. My "voluminous figures" consisted simply of a rough estimate of the minimum quantity of salt in the Humble dome, and a calculation of the volume of sea water and of the volume of saturated brine which this salt would represent. Mr. Kennedy apparently interprets my remarks as opposing the idea that so large a volume of salt could have been deposited from sea water, for lie cites the area and thickness of European salt beds; but nothing was further from my intention. Given an arm of the sea, an arid climate, and evaporation under a hot sun, it is easy to conceive the formation: of almost any thickness of bedded salt. I cited the figures in connection with Harris' theory of the origin of salt domes, according to which the plugs were formed by the deposition of salt from ascending brine solutions, owing to the drop in temperature as they neared the surface. I pointed out the well-known fact that the solubility of salt is only slightly affected by temperature, and that a drop of 80° C. would cause the deposition of less than one-tenth of the salt in solution; hence that more than ten times the volume of salt actually forming the plug must have been involved. If ten times the salt actually in sight be computed in terms of brine and multiplied by 50 or 60 to take into account all of the known salt domes, the resulting figure is so enormous that I am not at all as sure as Mr. Kennedy seems to be that the porous stratified rocks of eastern Texas and western Louisiana could contain or furnish so great a volume of migratory brine.

Jan 8, 1918

By Jaime Sepúlveda

"A broad diversity of critical operating conditions may influence, in very convoluted ways, the actual performance of any given mineral processing system, to the extent that it becomes nearly impossible to empirically determine the independent, decoupled impact of each condition. An effective approach to better understand the behaviour of such complex systems is via the extensive use of sufficiently detailed and reliable mathematical models, allowing the process analyst to explore and learn about the independent effect of all relevant operational conditions. Certainly, these tasks are greatly facilitated if the appropriate software tools - such as the Moly-Cop Tools software package - are readily available. Moly-Cop Tools 3.0 is a set of 63 easy-to-use EXCEL spreadsheets designed to help process engineers characterize and evaluate the operating efficiency of any given grinding circuit, following standardized methodologies and widely accepted evaluation criteria. Moly-Cop Tools 3.0 covers conventional and semiautogenous mills, crushers, HPGR’s, rod mills and vertimills, with the corresponding mass balance and parameter estimation routines; plus complementary spreadsheets regarding Bond’s Law calculations, grinding media selection and performance evaluations and other useful utility functions. As a relevant example of application, specifically focusing on semiautogenous grinding operations, the current publication describes how the models can help us learn and rationalise critical empirical observations such as the relationship between feed particle size and optimal ball filling, the role of the rocks as autogenous grinding media as well as the impact of feed size and discharge grate opening on the mill grinding capacity. Over the last 20 years, Moly-Cop Tools has gained wide acceptance around the mineral processing community in recognition to the quality of the mathematical models contained, their simplicity of application and ease of sharing of the calculations amongst all interested parties, since it is available free-of-charge with no restrictions to make as many copies as needed."

Jan 1, 2014

By Walter Henoch

THE subject of my talk is, "What 'the college, student expects when he gets out of college." Since all of us, here tonight are engineers, I think it will be proper to limit the discussion to "What the engineering student expects when he emerges from college." It might be well to con-sider just why we chose engineering rather than some other of the many courses open to us. We certainly did not choose it because it is easy nor because we thought it would be a very pleasurable way to spend four years of our lives. The majority of students who choose this profession have a pretty well-formed idea of what an engineering course requires. They at least know that they must work to re-ceive their credits. The minority, which is no small num-ber, are soon disillusioned and as a result they either quit college entirely or they join the' Commerce School. I asked a number of my fellow students just what it was that induced them to join the engineering profession, but it was hard to obtain a clear cut statement from any of them. Some said it was due to environment; others said, they took it because they liked it, and still others said they took it because their folks wanted them to. These state-ments are rather vague and incomplete, but underlying them all is something that the fellows found hard to ex-press. I think it is this: To the layman there is something of romance and adventure about the engineering profes-sion, and this fact coupled with the reputed high salaries and high commands of certain noted engineers no doubt fired our imaginations and consequently led us to join the ranks of the engineers. No doubt many engineers have gone out into the world and have found little romance or adventure .attached to their work, but nevertheless it is an occupation of a wide diversity of interest, of constant change of scene, and of vivid and human relations.

Jan 5, 1928

By C. M. Ball

The very considerable progress made during the past three years in the crushing and concentration of ores, lends special interest at the present time to any means of a more efficient character than such as are already generally known and in use, for the rapid and regular reduction of large quantities of the harder and heavier grades of mineral-bearing rock. Such means are of exceptional importance in connection with the preparation of iron-ore for separation by the magnetic process, since in this operation the capability of dealing with large volumes of material and of securing by ready means of adjustment, a granulation of the material under treatment to any required mesh or fineness, is of the greatest possible value. With these, however, must be coupled the further considerations of the amount of power expended for the work done, and the wear and tear, represented in the cost for renewal of parts of the crushing machinery, per ton of material reduced to a given degree of fineness. In general terms it may be stated that the operation of a crushing train, capable of reducing an ore-rock of average density and hardness, from the condition of mine lumps to pass 1/16-inch screenopenings at the rate of 20 tons per hour, should not call for an expenditure of more than 125 horse-power, and that the wear and tear of the crushing machinery, all told, should not exceed 2 cents per ton. In the design of a crushing-train for an ore-dressing plant, there are, moreover, considerations of great importance which relate to the simplicity or complexity of construction and arrangement, the wear and tear of few or many elements, the easy ability or difficulty of maintaining the working adjustment of the various elements of the train, and the number of hands required for the operation of the system. With these points in view, it is thought that the Cummings ore-

Jan 1, 1893

By Roswell Johnson

THE principle of sliding or graduation in royalties is accomplished either by the block, period, cumulative, or class method. The block method calls for a very low royalty rate on all oil produced up to. a certain amount. This amount would normally be that which is believed to suffice to repay the cost of the well. A higher rate would be charged after the well is making a profit. This method, common in many industries, is peculiarly inadaptable to any industry such as this where the productivity per day of the units gradually and inevitably decreases. This, for reasons given later, is because it forces an early abandonment of the well and so offends against conservation. In the period method, the royalty rate is changed by some definite amount when the well produces less than a specified quantity per day or other given unit of time. While simple to explain, it operates badly because a great deal is made to depend upon a small reduction in production, whereas the reduction as a matter of fact declines rather erratically owing to the exigency of lease management, connecting of tankage, and vagaries of the gaugers. These disadvantages are obviated by the cumulative method. In this method, all the production less than a certain amount per week has one rate of royalty, or none at all, and the oil produced in excess of that amount pays an additional rate of royalty. In the long run the calculation is probably no greater than with the period method, for simple tables will be published and printed in the oil papers that can be cut out, enabling quick readings to be made. Even without the table the calculation is easy with the few royalty rates, two or three, which would be used. In this method, the amount paid in royalty declines gradually, rather than suddenly as in the period method.

Jan 6, 1915

By O. Guyot, A. Broussaud, M. Revalor, G. Legrand, D. Kok, E. Roux

"The paper describes the advanced grinding control system implemented at Newmont Ahafo gold mine, Ghana in 2009 and 2010. It shows how the different components of the Advanced Process Control (APC) system cooperate for increasing plant performance.The SABC circuit processes hard rock from different pits mixed with a variable proportion of oxides. The control strategy is designed to address specific disturbances, including start/stops of pebble crushers, unavailability of oxides, changes in rock size distribution resulting from segregation in the stockpile, and changes in ore type and hardness.The system consists of an expert system per se with extensive use of fuzzy logic, supplemented by an online dynamic model of the process used as a soft sensor, a VisioRock™ rock size measurement system over SAG mill feed before addition of recycled pebbles, and an AudioMill™ system to analyze the SAG mill sound and detect impacts on liners. All these components are embedded in the same OCS© software platform, which, among other benefits, makes it easy to support the entire system remotely. The paper details the physical constraints and the variables that are manipulated. It describes the control strategy and the contribution of each component of the system. It also describes the project implementation and the performance of the system. In addition to a strong involvement and support of management, the very good acceptance of the operators has been facilitated by the remote access system that made it possible to immediately modify the system each time it was needed during the baby sitting phase.The system made it possible to increase gold production significantly through increases in tonnage and recovery thanks to a better control of the grind size through a stabilization of the secondary ball milling circuit and an improvement of the operating conditions of the classifying cyclones."

Jan 1, 2012

By Perry G. Harrison

The first autlientic description of an iron bath for the deposition of iron is probably that of Bottger in 1846, who used a bath containing ferrous sulfate and ammonium chloride. In 1861, Kramer deposited iron from ferrous chloride solution. Electrolytic iron was produced in 1869 by Bietz, who used it for making some magnetic tests. In the same year, Klein is. said to have worked out his process. Jacobil in 1869 described the results obtained in electroplating with iron on copper for making dies and plates for bank notes, by Klein's process. A letter from Klein to Jacobi describes the bath more fully. It consisted of FeS04 + (NH4)2SO4. Lenz2 described some properties of electrolytically precipitated iron. He used Klein's bath, FeS0, + MgSO4 + MgCO3, to neutralize and obtained fine-grained and hard iron. His conclusions were as follows: 1. Electrolytic iron and copper contain gases, notably hydrogen. 2. The volume of gas absorbed by iron varies within wide limits, but it is easy for iron to take up very considerable quantities of gas. 3. The absorption of gas is principally in the first formed layers of iron. 4. When iron is heated, gas begins to come off below 100°, principally hydrogen. 5. Heated to redness, reduced iron oxidizes in water, in part at least at the expense of the oxygen of the water, decomposing it and setting free hydrogen, which is partly or wholly absorbed. Siemens,3 in 1889, was the first to propose a general process embodying a step for leaching a sulfide mineral with ferric chloride or sulfate and a step for regenerating the leach and depositing iron on a cathode in a diaphragm cell. So far as is known, Siemens did not carry out his patent

Jan 1, 1930

By Ivor Hawkes

A wide variety of borehole gages have been developed for measuring rock stress. They all operate on the principle that stress changes around a borehole result in deformation that can be measured by the gage and interpreted in terms of the stress change either by calculation or calibration. In general terms, gages that comply with the borehole deformations are referred to as “deformation meters” and those that restrain the deformations by reason of their rigidity are referred to as “stress meters.” The main advantages claimed for stress meters are: their “stress” sensitivity is much less dependent on the rock modulus than that of deformation meters, and they are clamped firmly in the borehole. The use of each of the borehole gages developed for stress determinations in rock in the main has been confined to the laboratory responsible for its development. To be generally accepted a gage must meet the following specifications : 1) Gages to be left in place to measure stress changes must be relatively cheap, and if possible recoverable, so that large numbers can be used to obtain an overall picture of the changing conditions. This, of course, is not necessarily the case for recoverable gages designed for absolute rock stress measurements where usually only one or two gages are involved. 2) For use in the adverse environmental conditions encountered underground, it is essential that gages be robust, stable, and unaffected by high humidity, corrosive water, or dust. 3) Highly qualified technicians are rarely employed specifically for rock mechanics studies in mines or on civil engineering sites and there- fore gages designed for measuring stress changes must be easy to set and

Jan 1, 1972

By Harry Power

THE relative merits of air and natural gas as propulsive agents in pressure drive operations have been discussed for a number of years. When air or gas is introduced into the sand, various factors lead the operator to the selection of what he believes to be the proper propelling medium; but when the question of relative drive effect is raised, several physical and chemical premises must be taken into consideration in arriving at a conclusion as to the superiority of one medium over the other. Dow and Calkins1 and Beecher and Parkhurst2 have contributed fundamental experimental work on the effect of dissolved gas in lowering the viscosity and surface tension of crude oil. Herold3 and Tickell4 have indicated that whatever type of reservoir control obtains, gas plays other important parts in the movement of oil towards the well, namely: 1. The propulsive and expansive force of the gas. 2. The "Jamin effect," or resistance to the movement of oil in a porous media due to the formation of gas bubbles. When quantitative studies are made, parallel conditions must be selected whereby we may say definitely whether air is superior to gas as a driving medium, or vice versa. Examples of similar conditions are not easy to find in the field. However, in the laboratory we can make determinations which may lead to general conclusions when applied to the field. The- experimental and research work described here was started at the laboratories of the University of California in January, 1927, and completed in private laboratories at Tulsa, Okla., in February, 1928. The apparatus used was designed especially for the work.

Jan 1, 1928

By Gene E. McClelland

4.1 INTRODUCTION Metallurgical testing is required on any ore, waste or tailing material before amenability to heap leaching treatment can be accurately determined. The amount of testing necessary to determine and optimize heap leach amenability and conditions for any feed material is determined by the ore type, grade, size of deposit, and the overall financial commitment a particular company is willing to make. It is very easy to either "under-test" or "over-test" a particular deposit or feed material. A good heap leach testing program will provide the necessary design information without "bogging" down in too much detail. The three basic stages of ore testing include: • Preliminary--bottle roll tests and small column percolation leach tests; • Detailed--larger column tests, different feed sizes, and feed size optimization; and • Pilot-Scale--Column or field heap. The following subchapters describe the stages of heap leach testing necessary to put a viable property into production. 4.2 PRELIMINARY METALLURGICAL TESTING The two types of preliminary tests that can be undertaken include bottle roll or cyanide agitation tests and small column leach tests. Each of these is described below, along with information about what the test results provide in terms of furthering knowledge of ore qualities. 4.2.1 Bottle Roll Tests The first step in a heap leach testing program is preliminary bottle roll tests at a relatively coarse feed size to determine the ore's leachability. Bottle roll tests are usually conducted on reverse circulation drill cuttings samples (= 1/4 inch (6 mm) in size). Bottle roll tests provide initial information about precious metal recovery, recovery rate, and reagent requirements. The recovery results are to be considered maximums because attrition grinding creates fines and liberates recoverable values that would not be liberated from the material during a static leach. Bottle roll tests can be conducted on up to 25 lb (11 kg) ore charges of up to two inches (50 mm) in size. However, the coarser the feed the more attrition grinding occurs during rolling.

Jan 1, 1988

By D Pepper

Despite the limited literature regarding the pressure filtration of mine tailings (compared to the large number of papers published each year on paste thickening of tailings) the number of mines dry stacking filtered tailings is approximately equal to the number of mines storing tailings in paste form. Main drivers for filtering tailings are: water recovery, future rehabilitation costs for tailings storage area and permitting requirements. In arid climates, the cost and scarcity of water justify the capital and operating costs associated with pressure filters. Further justification for filtered tailings is the potential for progressive reclamation of the dry stack, which can proceed earlier and faster than the reclamation of a tailings dam.Filtered tailings must be transported to the storage area by trucks or conveyors. For dry stacking, 70 - 80 per cent solids content (by mass) is typically acceptable, however the primary goal is to provide a filter cake that is consolidated and dry enough to prevent bridging and build-up on the materials handling equipment. It has been observed that high average solids contents (~70 per cent) can be achieved while the centre of the filter cakes is still wet and sticky, presenting challenges for the transportation of the solids. Appropriate controls must be in place to ensure consistently easy to handle solids are discharged from the filter.FLSmidth has installed two fine coal tailings dewatering systems in the United States. One of these systems incorporates a Deep Cone Thickener« upstream of the filter press to minimise the amount of water that must be filtered, while the other tailings filter press is fed by a high rate thickener. The operation and performance of these filters is discussed and compared to the lab testing used to size the full-scale equipment.CITATION:Pepper, D, 2012. Operation of filter presses for tailings dewatering, in Proceedings 11th AusIMM Mill OperatorsÆ Conference, pp 241-246 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 29, 2012

By M. R. Khalesi, S. Arfania

"Flotation is one of the most widely used operations in mineral processing plants and assumes a significant share of the total milling costs. The purpose of this paper is to introduce a new set of capital and operating cost models for major flotation machines based on the application of single (SRA) and multiple regression analysis (MRA). Thirty-seven major flotation machines were analysed for this purpose. Depending on the machinery type, different technical variables such as diameter, required air flow rate, required floor space, cell volume, required air pressure, and power were considered as predictor variables, individually (in SRA) or simultaneously (in MRA). Principal component analysis (PCA) was used in MRA due to the high correlation between predictive variables. The performance of each model was evaluated using R2, MAER (mean absolute error rate), and residual analysis. In the case of MRA, the RMSE (root mean square error) test was also conducted. Maximum obtained MAER of 13.5% and minimum R2 of 0.86 indicated that these models could be applied as credible tools in estimation of capital and operating costs of flotation machines for design and feasibility studies.IntroductionMineral processing is a vital part of mining projects and mainly involves comminution, sizing, concentration, extractive metallurgical processes, and dewatering. Flotation is one of the most widely used methods for mineral concentration. Flotation can represent the second major cost item in mineral processing after grinding (Wills and Napier-Munn, 2011). Accordingly, it is a main concern of mining project managers to select and optimize flotation circuits in order to decrease costs and increase productivity. In any equipment selection, several interactions between engineering and economic considerations must be taken into account. Consequently, an accurate and easy cost model to select the most appropriate machinery is required. Moreover, cost models could be used in flow sheet simulations applied in design and optimization. Models of unit operations built into the simulators could be improved by linking the equipment cost models (Khalesi et al., 2015)."

Jan 1, 2017

By F. L. Wolf

THERE is a tendency on the part of practical foundrymen to accept with reluctance the results of tests on sands, binders, and. such materials made in the chemical laboratory alone. They feel that such tests frequently emphasize and measure properties that are of minor importance and fail to take account of characteristics that are important in determining the value of the material in actual practice. Experience seems to indicate that the most thorough and conclusive, test for any new foundry material is an actual run in the foundry alongside of a standard material of known value. Such a test, however, involves certain difficulties and necessitates certain precautions. It is not always easy to obtain a fair trial for a new material. We all have certain prejudices, which unconsciously affect our judgment to a marked degree. A practical test frequently involves so many workmen, each of whom reacts in a different manner to the test, that; the personal equation becomes very, complex and the true results of the test are obscured. There are times, too, when a thorough experimental trial of a material would interfere too much with the rush of work in a department or would hazard the quality, of the product to such an extent as to make it inadvisable. Under such conditions a laboratory test is advisable before the material is sent into the foundry. These tests should be made as practical as possible and so closely related to foundry practice as to give a true index to the value of a, product for the purpose desired. Core sands and molding sands are as numerous "as the sands of the sea" and vary as much in quality. Old sources of supply are continually failing and new ones must be opened, so that we must continually search for new and better grades of sand. During the past year, the Ohio Brass Co. has developed tests for sands', cores, and binders which bear a close relation to the particular use to which these materials are put in

Jan 9, 1920

By G. E. McElroy, A. S. Richardson

"The Bureau of Mines, in cooperation with the Anaconda Copper Mining Co., has conducted a series of experiments an air measurement in fan-pipe installa¬tions in a Butte mine during the fall and winter of 1922-3. The chief purpose was to determine the friction factors for different types and sizes 3f fan piping, but much attention was devoted to methods of air measurement, not only with respect to the methods used for friction factor tests, but also to com¬parison with other methods.The air flow was measured simultaneously at two points 380.6 feet apart in six different fan-pipe installations, each over 500 feet in overall length, in a rock cross cut adjacent to an air shaft,The air pressure was regulated by two sinoco multivane fans, driven by 7 ½ h.p. motors, the fans being connected in tandem so that one discharged into the intake of the other through a short connector and the latter discharged into the line through a straight connection equipped with a honeycomb of 4-inch squares or S radial vanes to prevent spiral flow of air, At each point where it was desired to measure the air flow, measuring sections equipped with piezometer rings pierced for easy insertion of pitot tubes, thermometers, etc. Piezometer static pressures were used in all the tests as standard, and dynamic or total pressures only were obtained with the pitot tubes, except for purposes of comparison.The tubes had special guide blocks, designed especially for underground work. To insure greater accuracy two tubes were used simultaneously, in order to obtain the velocity pressure at the center of the section simultaneously with that nearer the circumference, so that the mean velocity would be determined as accurately as possible. The tubes were held in proper place by means of flat guide boards. Velocity pressures were measured in manometers."

Sep 1, 1923

By Zhu Fangyi

"The article introduces the process EAF with De-P hot metal charging for Melting Stainless Steel in Baosteel stainless steel Branch. Based on the practice of production, the main effect on the Cr Loss during EAF with De-P Hot Metal Charging Melting Stainless Steel including using oxygen, the material charging and making slag are theoretically analyzed. The optimization of using oxygen can reduce the oxidation of Cr during EAF Melting Stainless Steel. The optimization of material charging, especially Si content is suitable can reduce the loss of Cr. Effective reduction can realize the lowest Cr oxide in slag. Based on the character of the process EAF with De-P hot metal charging for Melting Stainless Steel, applied the integrated control method, the Cr loss during EAF melting stainless steel decreased 43.10%. 1 INTRODUCTION EAF with De-P HM charging for melting stainless steel is a technology for stainless steel making. Practice prove that the new process can shorten tap to tap time, reduce the special power consumption, dilute the impurity element from the stainless steel scrap that is propitious to produce the high quality stainless steel, alleviate the shortage of stainless steel scrap to the production. As a new technology, the process has some technical problems and difficulties. The proportion of Cr alloy charging exceeds 20% during EAF with De-P hot metal charging melting stainless steel. As Cr alloy is difficult melting and easy oxidation, too much Cr alloy charging result in Cr loss too much and tap to tap time too long for EAF production. By applying the new technology, Baosteel stainless steel branch has increased the technology level and economy index. Now the content of Cr oxide in slag reduce from 8.59% to less than 6.0%. The Cr loss during EAF melting stainless steel decreased 43.10%. According to the practice application, the process of EAF with De-P HM melting stainless steel are theoretically analyzed. The main factors affecting the Cr loss during EAF with De-P hot metal charging melting stainless steel are discussed."

Jan 1, 2008

By J. A. Holmes

Venezuela has continued during 1929 to demonstrate that it is destined for many years to come to be one of our most important sources of crude petroleum. The producing fields of the Bolivar Coastal district along the east side of Lake Maracaibo have continued their phenomenal production, and two new fields have reached the near producing stage, while the results of deep test wells have strengthened the predictions and expectations that oil will be found and produced from many other pools and from several different geological formations. The finding of these new oil pools is not going to be easy, because the surface evidences of geological structure are scanty, and the subsurface problems are difficult, but intelligent effort backed by ample resources may be counted upon to overcome these obstacles and lead to a continued increase in the number of producing fields in Venezuela. During 1929 the total production of Venezuela was approximately 135,953,000 bbl., an increase of nearly 30,000,000 bbl., or 28 per cent., as compared with the preceding year. During 1929 oil in steel storage in Venezuela decreased from 13,000,000 to about 11,000,000 barrels. Tables 1, 2 and 3 briefly summarize Venezuelan production by fields and by years, and show the trend of production since its establishment on a commercial basis. Table 1.---Production of Crude Oil in Venezuela by Years, 1917 to 1929 All Figures in Barrels of 42 Gallons Total PRodUction Total PRoduction Year VeneZUela, Bbl. Year Venezuela, Bbl. 1917....................... 227,000 1927.................... 63,181,000 1918....................... 368,000 1928.................... 106,000,000 1919....................... 258,000 1929.................... 135,953,000* 1920..................... 548,000 Grand Total Since Ineep- 1921...................... 1,544,000 tion................... 380,791,000 1922....................... 2,448,000 Increase, 1929 over 1928.. . 29,953,000 1923..................... 4,170,000 Per cent. Increase......... 28.2 1924....................... 9,234,000 Per cent. 1929 Production 1925....................... 20,219,000 of Total............... 35.7 1926....................... 36,641,000 * Includes estimate for closing period of 1929.

Jan 1, 1930