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By A. K. Sinha

Research shows that continuous mining sections can normally produce around 950 tons per shift in a seam 5 or 6 ft. thick. The shuttle cars as well as the continuous haulage systems are able to match with the continuous miners in capacity. However, in practice both of the two systems seem to be averaging at around 450 to 500 tons per shift in production. Continuous haulage systems are preferred in thin seams particularly where the floor is soft.

Jan 1, 1981

By T. Alboussiere, O. Paireau

"We have designed an experiment to simulate natural convection in a direct-chill aluminium ingot. In this experiment we induce natural convection within a hemisphere filled with liquid-metal (mercury). The device consists of a copper hemispherical bowl closed by a copper plate. The top is maintained at superheat temperature while the boundary is cooled. While the temperature measurements were found using thermocouples, the velocities were obtained using an original technical procedure that relies on MHD properties. By applying a steady local magnetic field, we measure the velocity. The distribution of temperature and velocity can be separated into a thermally stratified core bounded by thermal wall jets. These wall jets greatly influence the degree of macrosegregation in ingots.1. Natural Convection in Aluminium CastingIt is well known that the quality of the final product in aluminium casting depends on the movements within the liquid metal such as forced or thermogravitational convection. Strong buoyancy-driven flows are known to occur in the liquid metal pool of an aluminium ingot caster. These flows have a significant influence on the metallurgical structure of the ingot through the redistribution of heat, fragmented dendrites and segregated material within the pool. Macrosegregation mechanisms depend on favourable temperature and flow distributions in the sump, and these factors will determine the pattern and degree of macrosegregation. Moreover, for many years, the importance of thermally-driven natural convection in casting has been underlined (Cole & Bolling, 1965). In this way, a precise determination of the velocity and temperature fields is required in order to improve understanding of the macrosegregation mechanisms. That is precisely the aim of this paper, which provides experimental data of the flow and the temperature distributions in a natural convection experiment. We intend, with this laboratory experiment, to determine accurately the velocity field, particularly within the thermal boundary layer where the effects of the flows may have an important role in the transport of crystals."

Jan 1, 1999

By Forrest T. Moyer

To keep the mineral and allied industries informed of trends in the causes of accidents, and to point out the need for corrective measures, the Bureau of Mines collects, analyzes, and periodically publishes health and safety statistics. This report includes data for 1963 on injury experience, with related employment statistics, at coal mines in the United States. These data are presented in tabular form under the following headings: General Injury Experience Tables 1 to 6: Current data on number of injuries, degree of injury, frequency and severity rates, average severity, and related employment and production data by work location.

Jan 1, 1966

By Forrest T. Moyer

To keep the mineral and allied industries informed of trends in the causes of accidents, and to point out the need for corrective measures, the Bureau of Mines collects, analyzes, and periodically publishes health and safety statistics. This report includes data for 1963 on injury experience, with related employment statistics, at coal mines in the United States. These data are presented in tabular form under the following headings: General Injury Experience Tables 1 to 6: Current data on number of injuries, degree of injury, frequency and severity rates, average severity, and related employment and production data by work location.

Jan 1, 1966

By E. M. Elkin, J. H. Schloen

All known methods of treating and recovering the various components of copper refinery slimes are discussed. The slimes treatment processes presently used by five copper refineries are described and flowsheets are given. A bibliography is appended. THE recent increase in anode copper production at the Noranda smelter placed an additional load on the Montreal East plant of Canadian Copper Refiners Limited. The consequent increase in the amount of anode slimes led to a careful consideration of the various methods of slimes treatment. A critical review of these methods is presented here in the belief that it would be of interest to copper refiners. No such review has yet been published, although a number of individual flowsheets have been described. To make the information of greater value, a questionnaire was sent to all copper refineries. While the overall response to the questionnaire was not entirely satisfactory, the foreign refiners gave much detail in their answers and expressed willingness to cooperate further if needed. Although the data below do not cover all refineries, it is thought that examples of all flowsheets currently in use are included. A comparison of these flowsheets should prove of interest. The paper is divided into two parts. One part deals with a discussion of all known methods of treating and recovering the various components of the slimes. The second describes the processes now in use at various copper refineries for the treatment of raw slimes up to the production of Doré bullion. A brief description of the two main electrolytic methods of parting Doré bullion is appended. Since the first commercial application of electrolytic refining of copper by J. B. Elkington in 1865, the treatment of anode mud or slimes has undergone many changes. The object of treating slimes has always been the recovery and separation of precious metals. With the passage of time, increasing attention has been paid to improvement in recovery and quality of byproducts, mainly selenium and tellurium. The original methods of treatment were crude, although relatively direct. For example," the slimes were fused on a dolomitic clay bed in a reverbera-tory furnace; the slag was removed and the resulting metal cupelled to silver and gold. In another process, the dried slimes were oxidized and scorified on the lead bath of a cupelling furnace. Later, these pyrometallurgical methods were modified by sulphuric acid leaching, followed by fusion with niter in a reverb eratory furnace. A preliminary oxidizing roast improved the leaching step. Straight hydro-metallurgy has not found practical application. No matter what method is employed, the result is Dorb metal—a silver-gold alloy, with or without the metals of the platinum group. Characteristics of the Slimes Anode slimes are made up of those components of the anodes which are not soluble in the electrolyte. They contain varying quantities of copper, silver, gold, sulphur, selenium, tellurium, lead, arsenic, antimony, nickel, iron, silica, etc. The color of raw slimes is grayish black and the particle size is —200 mesh. However, Baraboshkin and Gaev (Ref. 1, p. 37), describe light gray colored slimes from the electrolysis of secondary copper which are composed predominantly of sulphates and of basic salts. In anode furnace practice primary and secondary metals are not usually segregated and such slimes do not commonly occur. Copper in the slimes originates to a large extent in the cuprous oxide of the anodes and is therefore dependent on the oxygen content of the latter: Cu2O + H2SO4 + CuSO4 + H2O + Cu [I] Opinions have been expressed that some of the copper is formed by decomposition of cuprous to cupric sulphate Cu2SO4 ? CuSO4 + Cu PI The product of either reaction is very finely divided. Depending on the slimes composition, much of the copper is combined with sulphur, selenium, tellurium, etc. The balance of the copper is in the free metallic state.

Jan 1, 1951

By N Weatherstone

A consistent and reliable international approach to public reporting of a mining companyÆs main assets, its Mineral Resources and Mineral Reserves, has become of increasing importance in recent years with the rapid globalisation of the mining industry. Through the efforts of the Committee for Mineral Reserves International Reporting Standards (CRIRSCO), a committee representing national resource/reserve reporting committees in Australia, South Africa, UK/Ireland/Western Europe, Canada, USA and Chile, the reporting standards in most æwesternÆ countries use almost identical resource and reserve definitions and are generally around 90 - 95 per cent compatible. New codes based on the CRIRSCO model are being developed or considered in other countries, including Peru and the Philippines. This unified approach has provided the mining industry, regulators and investors with a common framework to report and interpret mineral resource and reserve estimates, thereby facilitating funding to the mining industry through improved investor understanding and confidence. The common reporting standards are reflected in the CRIRSCO International Template, available on CRIRSCOÆs web site, a document that is kept æliveÆ by reflecting the best of the most recently published CRIRSCO-type national standards. In a significant exception to this unified industry approach, the USA Securities and Exchange Commission (SEC) requires adherence to its own standard, only partly compatible with CRIRSCO-type standards, with adverse and costly implications for companies listed in the USA. A major industry submission was presented to SEC in April 2005 which, if accepted, would bring about considerable compatibility of CRIRSCO and SEC standards. The UN Economic Commission for Europe (UN-ECE) has, since 1992, been developing its Framework Classification (UNFC) as a single global system for harmonising all existing national and international resource/reserve reporting systems, covering both government and industry requirements and both the hydrocarbon and solid minerals sectors. While few mining companies report their resources and reserves using the UNFC, the system is accepted or is being examined as a basis for reporting by some governments, including those of Russia, China and India. The UN-ECE agreed in 1999 to incorporate CRIRSCO definitions into the UNFC for those categories of resources and reserves used for market-related reporting, but a more recent edition of the UNFC removed or changed the definitions, requiring re-engagement by CRIRSCO to ensure compatibility. Since 2005, CRIRSCO and the Society of Petroleum Engineers (SPE) have been involved in discussions with the International Accounting Standards Board (IASB), which plans to develop a new Accounting Standard for the Extractive Industries as part of its International Financial Reporting Standards. At the request of the IASB, CRIRSCO and SPE have been examining any potential for convergence of resource/reserve definitions between the mining and hydrocarbon industries. Each of these systems is described in this paper and high level comparisons made.

Jan 1, 2006

By J. H. Henderson, R. J. Wygal, J. Naar

Experimental work is reported which shows that consolidated rocks and unconsolidated porous media exhibit different imbibition flow behavior. At a given saturation the imbibition nonwetting permeabilities for a rock are smaller than the drainage permeabilities. The contrary happens for unconsolidated aggregates — imbibition nonwetting permeabilities are larger than drainage ones. A similar difference is observed for the wetting phase. Imbibition permeabilities are larger than drainage ones for a consolidated rock but smaller than drainage permeabilities for an unconsolidated medium. The results of these differences are examined for two cases. 1. Flooding Efficiency — Craig's1 scheme for the computation of production history of a five-spot water flood is shown to agree extremely well with experimental results obtained when using a system packed with glass spheres if imbibition relative permeability curves are used. 2. Alcohol-Slug Displacement — Published theory 2 on oil displacement by alcohol slugs has been questioned despite the apparent agreement between predicted and observed results. The present work suggests that, if imbibition relative permeability curves characteristic of the unconsolidated media used in the early experiments had been available to make the predictions, the inadequacy of the theory would have been immediately evident. The experimental work shows that poorly consolidated formations tend to behave like unconsolidated media. Finally, it is shown that the difference in imbibition behavior is directly related to pore-size distribution and cementation. PART 1 — THE FLOW BEHAVIOR OF UNCONSOLIDATED AGGREGATES Experiments on scaled models of field reservoirs are useful for studying new displacement processes which are incompletely understood. Even when a mathematical description is possible, the solution might be difficult and complex. An answer obtained from a scaled model is extremely valuable in such cases. A great amount of work$-5 therefore, has been devoted to the derivation of scaling laws. Similarity groups have been defined which assume (1) that the relative Permeability curves of the prototype and the model are the same whether the displacement is an imbibition or a drainage process and (2) that there is a linear relationship between the capillary pressure of the model and the prototype. For practical reasons (simplicity in the preparation of models, duration of the experiments, etc.), the porous media of laboratory models are usually unconsolidated packs of sand or glass particles. Hence, unless the capillary and flow characteristics of unconsolidated and consolidated systems are identical, the model data are applicable only to unconsolidated formations. The usefulness of scaled-model studies may then be seriouslyrestricted since most oil-bearing sands are consolidated. Perkins and Collins6 suggested the use of model and prototype curves normalized with respect to both relative permeability and saturation to improve compliance with scaling criteria. Even this technique does not give a satisfactory model-prototype match. This paper reports an observation of two-phase flow in unconsolidated sands which shows that, for most displacements, "scaling" in the strict sense of the word is not even qualitatively feasible with a sand model. It provides, however, a firm foundation for testing a theory by matching it with observed performance of laboratory-size models. EXPERIMENTAL AS a part of a basic study of packed aggregates, the relative permeability of glass-spheres and sand-grain packs was measured with capillary control. The fluids were oil and air' The spheres were packed in plastic cylinders 1-in. in diameter;

By D. G. Krige

Discussion by D.G. Krige I read the paper "The jackknife approach" with interest. There appears to be a widespread misunderstanding of the kriging variance as a measure of the error variance of the kriged grade estimate of an ore block. Virtually all mineral deposits display skew frequency distributions and are subject to the well-known proportional effect. Thus, it is usual to find in an ore body, when subdivided a really into subpopulations, that the variances of the grades (or the total sills of the variograms) for the subpopulations will be correlated with the corresponding mean grades (usually linearly with the squares of the mean grades). Thus, the variogram will be station. When this feature is evident, the relative variograms for the subpopulations (i.e., with variogram values divided by the relevant mean squared) will be similar, or they could even be accepted at an averaged model for the whole ore body with stationarity. Using this approach, the kriging variance will reflect the relative error variance that is to be multiplied by the kriged block grade estimate squared to yield the usual kriging variance for purposes of confidential limits. The objection that the kriging variance does not take into account the local grades used for a block valuation then falls away. In the example given by the authors, the global mean grade is not recorded. But the global variance, taken at the value of the total sill, is 0.005, whereas the variance of the 11 values used in the example is 0.000152. The two variances differ by a factor of nearly six. If the relative variogram approach is applicable in this case, the need for a jackknife technique will fall away. Also note that, for skew distributions, the confidence limits for ore-block estimates will be skew as well, and symmetrical normal limits will not apply. RepIy by G.S. Adisoma and M.G. Hester There are a few interesting ways to utilize the block-by-block kriging variance that the authors have encountered. One of them is to classify a particular block into a proven-probable reserve or a possible-resource category, using a somewhat subjectively determined kriging variance value. Here, the kriging variance still represents the drilling density, i.e., data configuration more than the data values. Professor Krige mentions another ingenious approach through the use of relative variograms. This will yield a block kriging variance that takes into account local (block) grade, as he correctly points out, for purposes of confidence limits. However, we are more interested in quantifying the uncertainties associated with grade estimation in a much larger scale than a block. e.g., for a mining phase or the minable reserve. The jackknife, through the block kriging shortcut, provides a simple solution to obtain a single estimation variance for a general shape. The authors' main reference discusses different ways to handle skewed distributions typical in mineral resource estimation, e.g., the weighted jackknife approach or data transformation. Also, as with any technique, we can always improve our estimate by subdividing into smaller populations that are geologically and statistically separate domains. As we have mentioned; the jackknife is but a tool in our collection of tools to solve estimation problems. The jackknife's appeals are its simplicity and versatility. It is handy when no other solution is available, or when the available solution is not practical. It is used with other methods, not to replace them altogether. ?

Jan 1, 1997

By F. G. Stone

A strain anneal process has been developed for the production of iron tensile test specimens, the gauge length consisting substantially of a single crystal. Since the method has been quite successful and the results of general interest, the process will be described in detail. The essential feature of the method is the continuous, slow increase of furnace temperature during the grain growth annealing treatment. Briefly, the procedure is as follows: I. Heat machined bars to 950°C in vacuum, hold I hr, furnace cool. 2. Strain to 3.5-3.8 pct elongation (axial loading). 3. Anneal, continuously raising furnace temperature from 450 to goo°C at a rate of so0 per day. 4. Hold at 900°C for 100 hr, furnace cool. Tensile test bars with a 1/4- X I-in. gauge section have been machined from 35 in. Armco iron plate and from 3/8 in. rounds of Puron, a high purity laboratory iron. The axes of the test bars were parallel to the rolling direction. During the machining operation the bars were polished in the lathe with No. 600 emery paper, and this polished surface proved satisfactory for subsequent macrographic examination. A micrographic examination of the materials showed that both were high in FeO inclusions, those of the Puron being smaller but much more numerous. An initial vacuum anneal of I hr at 950°C followed by furnace cooling resulted in a ferritic grain size of ASTM No. 3 for Armco and No. I for Puron. Special heat treatments which reversed these initial grain sizes produced no marked improvement in single crystal production after subsequent treatments. The strain for subsequent grain growth was determined by measuring the decrease in diameter of the tensile bar during elongation. The change in diameter could be measured with a micrometer to +0.0001 in., corresponding to an accuracy of 0.08 pct for the elongation. The accuracy of the diameter measurement is also dependent on grain size because of surface roughening during elongation, and for this reason, a No. 3 grain size is preferable. Optimum grain growth was obtained for strains between 3.5 and 3.8 pct elongation. Once this critical range had been cstablished, all tests were held within it, and thus only meager data were obtained on grain growth following strains outside this range. In general, grain size decreased with increasing strain to about 5.5 pct elongation. Grain growth was occasionally ohserved with less than 3.5 pct elongation although this may have been the result of error in measurement. The program control device for the grain growth anneal consisted of a I rpm motor and worm gear drive, operating the dial shaft of a pyrometer controller through a pulley. Dial speeds of 33 and 50°C per day were tried, but the results at the slower speed were not sufficiently better to warrant the increased furnace time. The furnace tube was evacuated at intervals during the run and kept sealed the remainder of the time. When the controller reached 900°C, it was stopped and the temperature held for 100 hr. This eliminated most of the numerous minute crystallites of about 0.5 mm dimension usually visible on the etched surface of the single crystal. Since recrystallization occurs at the A3 point, the top temperature of 900°C provided a 10"

Jan 1, 1949

By Philip C. Good

The Bureau of Mines investigated the feasibility of utilizing low-grade phosphatic shale as a feed for phosphoric acid production. Present practice in the phosphate mining operations of southeastern Idaho is to remove and either stockpile or waste low-grade phosphatic shales as the high-grade phosphate beds are selectively mined. Acid digestion studies were performed on uncalcined shale and on shale calcined to remove free carbon to determine phosphoric acid quality, yield, and amenability to filtration. The untreated shale, containing 21.4 pct P205, was digested with sulfuric acid in a single-tank reactor to yield phosporic acid and gypsum residue. Although 96 pct P205 recovery was attained in digesting uncalcined shale, only the calcined product gave satisfactory filtration rates. Calcined shale yielded phosphoric acid recovery of about 93 pct of the contained P205. It was higher in impurities than commercially prepared acid derived from high-grade western phosphate rock. An acid filtration rate of 745 lb P205/ft2/day was attained. This value is lower than the rate of 900 lb P205/ft that is generally considered to be acceptable in industrial production of phosphoric acid from high-grade phosphate rock.

Jan 1, 1981

By J. J. Dowd

1. The Pittsburgh bed is the most important bed in Ohio County, and virtually the entire production of coal in the county comes from this bed. 2. Known measured and indicated reserves in all beds, based on a minimum thickness of 14 inches and 1,800 tons per acre-foot of-coal in place, are estimated at 365 million short tons as of January 1, 1954. Of this total, 356 million short tons is in beds 28 inches and more thick. Areas in each bed were omitted from the estimate because available data relative to the bed characteristics are too meager to make an estimate that conforms with the definitions of measured and indicated coal adopted for this study. Should future drilling or development prove reserves in these areas, such reserves should be added to the total estimated reserves.

Jan 1, 1955

AlbioN S. Howe, San Francisco, Cal. (communication to the Secretary*):—In discussing this paper, 'I mould suggest that. the dip for iron pipe be liquid asphaltum, instead of coal-tar and pitch. Asphaltum is superior to coal-tar, and the variety

Jan 1, 1905

By A. C. Johnson

The T. L. shaft of the Eureka Corp., Ltd., near Adams Hill, Eureka, Nev., was sunk in 1954 and 1955 to permit exploration and development of the lead-zinc-silver ore that had been indicated by drilling. This paper describes the sinking of the 4-compartment, vertical shaft to a depth of 1,127 feet and includes pertinent details of equipment, methods, and costs. A shaft crew of 10 men, supplemented by 13 hoistmen, topmen, carpenter, and shop crewmen, sank and timbered the 1,127 feet of shaft in 630 workdays. Total labor per foot of shaft was 107.95 man-hours. Supplies per foot of shaft included 20 pounds of explosives, 1.61 pounds of drill steel, and 200 board-feet of timber. Power consumption, including that required for pumping, averaged 3,726 kw.-hr. per foot.

Jan 1, 1958

By Douglas Bartley, Kevin Hachmeister, Rich Barry

The use of explosive energy as a tool in mining and construction is the most effective way to move large amounts of rock that is otherwise unable to be efficiently excavated. Since the early days of explosive engineering utilizing black powder initiated by cap and fuse to the present day diversity of specialty products and technologies the goal has always been to "Break and Move" as much rock for the least expense possible. It has primarily been the competitive nature of our economy that has driven the advancement of technological innovations and practice. However, the concept of removing rock is not specific to the mining and construction industry. Since the dawn of creation, man has used rock as a medium to express his creativity in the form of sculpture. Even though the process of sculpture requires the removal of rock, the final work is defined by the rock that remains intact. This typically necessitates the extreme care of the sculptor to painstakingly chip away the rock allowing the image to emerge from the original mass of stone. The famous sculptor Michelangelo once said, "I saw an angel in the marble and carved until I set him free."

Jan 1, 2004

By Hiroshi Tabuchi, Akihiko Takahashi

"High strength aluminum matrix composites were prepared and studied using as reinforcement newly developed nearly mono-dispersed a-alumina single crystal particles each of which has a regular polyhedral shape. The composites were formed by centrifugal casting using aluminum casting alloys (Al-Si and Al-Cu-Mg) as the matrices, followed by heat treatment. The volume fraction of the alumina particles was as high as 60%. The same matrix composites reinforced by conventional crushed fused alumina particles were similarly prepared as the reference. The bending strength and tensile strength of the Al-Cu-Mg matrix composites reinforced by the novel alumina particles of 10 µm in average diameter are as high as 1020 MPa and 560 MPa, respectively, which are higher than those of the conventional alumina composites. Furthermore, the bending strength of the novel alumina composites degraded less than that of the conventional alumina composites after thermal fatigue tests. The number of cracks in the matrix seemed to influence the thermal fatigue property because many cracks were observed in the conventional alumina composites, whereas only a few cracks were observed in the novel alumina composites. This was confirmed by SEM observation of the sample composites before and after the thermal fatigue tests.IntroductionAluminum matrix composites strengthened by dispersed reinforcements are of interest for various applications due to their improved properties, such as high specific strength and stiffness, wear resistance, etc .. A large number of investigations have been made concerning fabrication processes and properties of these composites. With regard to the fabrication processes, powder metallurgy has been well-known as a solid phase process, and recently many liquid phase processes such as a compocasting, a squeeze casting, the Lanxide process, and the Osprey process have been studied and developed.(1)"

Jan 1, 1998

By D. K. Weatherley, T. P. Oladele, L. B. Bbosa

Comminution is a critical stage of mineral processing which aims to reduce the size of ore particles through breakage, consequently increasing the likelihood of liberation of valuable minerals. Several factors, such as pre-existing cracks, mineralogical composition and particle size, are known to affect ore breakage behaviour. While investigating the contributing role of each factor is essential to improve understanding of breakage, isolating individual factors by experiment is typically impractical. Numerical techniques such as the Bonded Particle Model-Discrete Element Method (BPM-DEM) have been developed as a means of studying closely controlled breakage conditions. This study evaluates this technique in modelling the effect of pre-existing cracks breakage behaviour for synthetic ore specimens during impact breakage in a short impact load cell (SILC). Synthetic ore was modelled as a collection of discrete entities connected by brittle-elastic beams. Model parameters were calibrated against macroscopic material properties of a conventional homogenous synthetic ore. The extent of pre-existing cracks within specimens was prescribed by varying discontinuities between the connecting entities. An approach with the insertion of seed points was used to model the mineralogy of the specimens. Results showed that pre-existing cracks resulted in a decrease in the fracture force as the crack density increased. Additionally, the irregularity of fracture patterns within rock specimens after impact became more apparent as the crack density increased. The findings were used to emphasise some of the attributable causes of the variation observed between ore hardness characterisation tests and in situ ore behaviour. Keywords: Element Method; Bonded Particle Model, synthetic rock sample; rock fracture, preexisting cracks

Jan 1, 2020

By S G. Gatehouse

Extensive gold mining activity û dominantly alluvial û has taken place in the Sofala-Wattle Flat district since the 1850s.This work will have located most, if not all, of the visible sources for the gold in current stream systems.Also, there are many gold bearing elevated Tertiary alluvial terraces in the district which may confuse possible gold signals from primary hard rock sources in the streams.Associated element signatures eg As and other chalcophiles, will discriminate non-alluvial signatures but will not necessarily distinguish non gold bearing mineralisation.Both trace elements and conventional gold analyses will dominantly re-identify known occurrences. Conventional BCL sampling is considered misleading as experience from the Oberon area to the south of Sofala demonstrated that BCL sampling was very sensitive û perhaps too sensitive.All occurrences û major or minor û of gold will be detected (Gatehouse, 1992). To improve on BCL sampling, a small, -120# sample with selective duplication was recommended to screen coarse-grained gold anomalism and a program of sampling covering an area thought to be prospective was enacted.Samples returning anomalous Au were analysed in duplicate and if the results demonstrated reproducibility, indicating fine gold only, the sample site was assigned a high priority.It was found that most of the reproducible samples clustered in one area û Spring Gully. The results of this sampling program led to immediate identification and, ultimately, the successful delineation of the Spring Gully gold deposit near Wattle Flat.

Jan 1, 2004

By J. E. Gould, H. J. Hand

MANY independent studies are being made on slag-metal relationships in the open-hearth furnace, and these studies cannot help but result in an ultimate improvement in the quality of open-hearth steel of both the open and killed types In maintaining quality control at the National Tube Company's plant, emphasis has been placed on an evaluation of the iron oxide content of the steel in as quantitative a manner as possible, in order that the most desirable level of oxidation might be obtained for a given type of steel, and to secure a more precise basis for final deoxidation In the past, various methods were used for estimating FeO content of the metal at tap to ensure uniform and thorough deoxidation of killed steel, avoiding over-deoxidation to minimize the amount of inclusions present and to make as sound a steel as possible For the open-type steels, such control involves obtaining the necessary level of oxidation for the most suitable action of the steel in the molds to provide good surface quality and cross section for the specific grade involved While rapid analytical methods are available for determination of FeO in the liquid metal, such methods are still not sufficiently rapid for routine control on the open-hearth floor, nor do such analyses provide information as to what causes high or long FeO values The method of sampling can still be considered a problem in view of the variability encountered in taking simultaneous duplicate tests With the accurate rapid methods now available for determining carbon in the metal, estimating slag basicity and obtaining temperatures of liquid metal, it was believed possible to determine the influence of the various furnace reactions in establishing the state of oxidation at a given time Simultaneous slag and metal tests were used in this analysis, aided by statistical methods, to determine the factors that affect the metal FeO, and also to make a quantitative evaluation of these factors Data obtained are from actual furnace operations, the metal test being of the spoon, Herty type, and taken in duplicate. Metal temperatures were taken with the open-tube bath pyrometer developed by the Research Laboratory of the United States Steel Corporation This method of taking open-hearth metal temperature was invented by Collins and Oseland,1 and was developed into its present form by L 0 Sordahl,2 of the Research Laboratory The instrument and its mode of operation have been fully described by Sordahl and Sosman in three publications 3,4,5 Slag components were obtained by chemical analysis Data were taken from four 165-net-ton furnaces during the working period of the heat Recognizing that additions of ore affect the

Jan 1, 1942

By Thomas V. Falkie

4.3-1. Introduction. DEFINITION OF PRODUCTION SYSTEMS ENGINEERING. Surface mines, like most other mining and manufacturing operations, are usually organized along functional departmental lines. Frequently a mine manager will have his responsibilities divided among production, maintenance, administrative, and engineering functions. Often these are separate departments, but various combinations and variations of these functions can be found in mining companies throughout the world. The production function is responsible for physically removing the ore from the ground, processing it, and shipping the product to its market. The production manager supervises and is responsible for all man-machine systems necessary to extract and process the ore. The maintenance function, which is sometimes included within the framework of production, has the task of insuring the maximum availability and longevity of equipment needed to produce the ore. The administrative function is responsible for accounting, personnel, industrial relations, purchasing, public relations, property, and other related staff functions. For purposes of this chapter, engineering can be described as the function which plans, designs, and supervises construction of physical plant and equipment. In this context engineering includes mechanical, civil, chemical, electrical, and mining engineering. However, another phase of engineering is to prepare plans to insure that the various man-machine components function in an optimum way towards the single most important goal of any enterprise-the maximization of return on the investors' money. Along with the planning, another duty of engineering is to audit all phases of the operation for the purpose of recommending improvements toward this same general goal. These latter two ideas form the basis for a discussion of production engineering and systems engineering. Production engineering is a catchall, problem-oriented function whose duties can include: 1) solution of day-to-day production problems, 2) short-

Jan 1, 1968

By N. M. Langdon

ON account of its greater economy and cleanliness, and the extent to which iuferior fuels can be utilized for its generation, there has been of late a rapidly increasing tendency to substitute gaseous for solid fuel for metallurgical and other industrial purposes, which has led to various improvements both in the process and apparatus for its manufacture, notably among which may IN mentioned the Tessie Gas Producer* described by Mr. A. L. Holley. A considerable experience by the author with producers of various kinds using anthracite fuel, in connection with the Taylor ore-roast ing furnace at Chester, N. J., with more or less unsatisfactory resnlts, finally led to the erection of the experimental producer described below, which combines in its construction, simplicity, dorability, and accessibility for cleaning and removing the ash and manipulating the fuel. It is adapted to the use of various kinds of fuel, and is capa ble of generating a steady and uniform supply of gas at a minimum cost. It has been in constant use day and night for about five months, and has given very satisfactory results. The producer, as shown in the accompanying sketch, consists of a cylindrieal furnace, preferably inclosed in an iron jacket or casing, having a bosh or inverted cone-shaped base as in a blast furnace. It is provided with a bell and hopper, the bell being hung to a forked lever, and having guide flanges to prevent it from swing ing, and an opening at its apex through which to insert a bar for manipulating the fuel, which when not in use is closed with a plug or stopper. Cleaning is effected through two small doors placed opposite each other at the hearth level, which in order to facilitate the operation is elevated above the floor. The doors are removable, and when closed are by their own weight and by means of bevelled lugs held tightly against their frames, forming a practically tight joint. A blast of commingled steam and air is injected into the fuel through a series of tuyeres underneath the bosh communicating with steam jet injectors. A small flue also connects the door

Jan 1, 1884