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By Bertrand L. Johnson

Pure talc is a hydrous magnesium silicate (Mg3Si4010(OH)2). Pyrophyllite, although it resemble: talc in certain physical properties and uses, is a hydrous aluminum silicate (Al2Si4010(OH)2) - like kaolin, except that kaolin has a somewhat different composition (Al2Si205(OH)4) and structure. Soapstone can be considered essentially an impure talc grid when pulverized is used for many of the same purposes as talc. Massive, compact varieties of talc often are called steatite. PROPERTIES Talc and pyrophyllite are characterizes; by their softness (hardness on Mohs' scale, 1) their greasy feel, pearly luster, white color, and perfect basal cleavage. They have similar crystal structures, and this similarity explains the close resemblance of the two minerals in physical acid chemical properties. Talc is a good electrical insulator, resists the action of most acids and alkalies, and has an exceptional heat-retaining capacity. Its specific gravity is about 2.7.

Jan 1, 1939

By José Antonio Enríquez

El presente texto describe las características estructurales y estratigráficas de los yacimientos epitermales de baja sulfuración en el corredor estructural de San Pablo en Porculla, Cajamarca. En este sentido, el corredor se relaciona a un sistema de fallas regional que tiene rumbo general NW-SW y está segmentado por fallas transformantes de dirección NE-SW, que limitan bloques estructurales. Asimismo, sobre este corredor estructural se ubican en forma alineada una serie de yacimientos epitermales del tipo baja sulfuración. Los yacimientos están controlados por fallas de dirección NW-SE y contienen mineralización en vetas y mantos con relleno de cuarzos, calcita y baritina con importantes valores de oro y oro-plata y las estructuras menores relacionadas a las fallas transformantes. Finalmente, estos yacimientos pueden constituir un nuevo e importante objetivo de exploración por minerales de oro – plata del tipo bonanza.

Sep 12, 2005

By Silvana Costa

The quality of life in remote mining communities is a significant factor that impacts the mining industry?s ability to attract and retain high quality workers in remote areas. The rapid growth of mining development, along with the shortage of skills and training globally are reflected in an increasing need to identify the ideal conditions for employee wellbeing and investigate ways to improve quality of life of miners and their families in remote locations.

May 1, 2007

By Paul Meadows

The Buckrange reservoir, Stephens field, Ark., is approaching the economic limit of production under present operating conditions. Over 100 million barrels of oil remaining in the reservoir will not be recovered by primary producing methods. The Bureau of Mines, in cooperation with the Arkansas Oil and Gas Com-mission, made the petroleum-engineering study presented in this report to aid operators in evaluating secondary-recovery possibilities in the reservoir and in planning secondary-recovery programs if their initiation is decided. The Stephens field, in Columbia, Nevada, and Ouachita Counties, Ark., is about 12 miles northeast of Magnolia, Ark., and 15 miles west of the Smackover field. Surface elevations in the area average about 240 feet above sea level.

Jan 1, 1956

By R. J. Sheets

"What is the primary responsibility of geo-engineers in mining? Early in my career, I was asked this question by an experienced engineer. After explaining some various tasks, I was informed that the main responsibility was to monitor. Although this proved true for many cases, I came to understand that monitoring is merely one means to achieve the primary responsibility of a geo-engineer — ensure the safety of all personnel. The potential for groundwater to impact stability allows this responsibility to be shared with hydrogeologists as well. Inherently, we monitor to identify risks; we analyze designs and develop models to mitigate risks; and we are provided opportunities to explain to supervisors the geo-risks to personnel and mine plans that will result from particular courses of action. A couple years ago, I took an opportunity to share experiences from the first several years of my career as a geotechnical engineer with junior geoengineering staff within my company. The intent was to share situations that were a part of my development, so they could learn and understand their responsibility to the safety of all personnel. Similarly, the purpose of this article will be to share these thoughts with new geo-engineers in the industry in order to help prepare them for the challenges they will face.What is our purpose?A senior hydrogeologist shared an experience that occurred in their career. The experienced hydrogeologist had spent some time with the hydrogeology staff, most of them relatively new to the mining industry, to understand how they viewed, and carried out, what they perceived as their responsibility at an active mining operation. After a couple months of observing and mentoring, the senior hydrogeologist met with everyone and posed the question “What is our purpose?” I am sure that most junior geo-engineers would start explaining different tasks or projects that they must complete. The list would include monitor ground water levels, track slope movement, evaluate designs, build models, oversee drilling programs, etc. These individuals did as well their supervisor. After going around the room allowing each individual to answer, the experienced hydrogeologist provided them the real answer to the question: Safety. If you examine the root intent of nearly every responsibility geotechnical engineers and hydrogeologists have, it is to apply our professional skills to ensure the safety of our workforce. When the objective of safety from the geotechnical and hydrogeology disciplines is achieved, inherently accomplishing the mine plan should follow suit, or it should significantly minimize the impact by gravity or ground water."

Jan 9, 2018

By K. C. Dean

During investigation of the occurrence of manganese deposits in the western United States, examining engineers of the Bureau of Mines collected samples from several properties in Montana. Extensive laboratory tests were made to determine the amenability of these ores to standard ore-dressing methods of concentration to produce marketable-grade products. The Tzarena claim, one of the properties sampled, is on the western end of the Rama or Black Chief vein in the city of Butte, Silver Bow County, Mont., on the southwestern periphery of the mining district. The Black Chief 'WAR system contains one of the largest and most extensive known rhodochrosite deposits in the world. Test work was conducted on two samples of Tzarena ore. Sample 1 is representative of a vain 5 feet wide on the 150-foot level of the Tzarena shaft. Sample 2 is representative of a vein 8 feet wide in the 100-foot level of the shaft. The Minnie Jane mine, 1/2 mile northwest of the Tzarena claim, contains ore similar in mineralogical composition to Tzarena ore but differing materially in manganese content and to some extent in physical characteristics. Ore-dressing investigations were made on three samples of Minnie Jane ore. Sample 1 represented stock-piled ore, sample 2 represented ore from Minnie Jane mine having a considerable sulfide content, and sample 3 represented low-grade ore.

Jan 1, 1951

By J. J. Pfeiffer, J. W. Myers, A. A. Orning

A study was made of the feasibility of using washery refuse in the manufacture of lightweight aggregate for building materials . Preliminary caking and burning tests showed that the refuse could be converted into lightweight aggregate but that refuse passing a No. 14 U.S. Standard screen could not be used without additional processing . The aggregate was prepared by pelletizing the refuse and burning out the carbonaceous material on a chain- grate in a refractory - lined furnace . A small stationary - grate furnace was used to study the process variables involved and establish optimum ignition and combustion conditions . Maximum ignition rates were obtained at air rates of approximately 200 lb per sq ft hr , but air rates of 450 or higher were required in the burnout period to obtain temperatures high enough to sinter the ash . The data obtained from this unit were used as a guide in the operation of the chain- grate furnace . In the chain - grate furnace , satisfactory results were obtained by using primary air rates of 200 lb per sq ft hr in the ignition zone and 500 lb per sq ft hr in the burnout zone . The ignition- zone air was preheated to 500 ° F , but the air for burnout was not preheated . A total of 2-1/2 tons of aggregate was prepared in this unit and made into experimental block . A commercial development based upon the results of this study yielded a product which met the ASTM specifications C 130-42 for lightweight aggregate , and blocks made from the aggregate were of good color , durable , and light - inweight .

Jan 1, 1964

By Kenneth L. Cook

The five serpentine masses exposed within the northern half of Riley County, Fig. 1, constitute a major part of the few exposures of igneous rock in Kansas. Although not many subsurface data are available for this part of Riley County, six deep wells indicate that the basement complex lies between 2500 and 3000 ft below the earth's surface. A northwestward-trending ridge in the present basement surface, which extends across the northern half of Riley County, is indicated on Mettner's contour map. The serpentine masses lie in a region that was active during Pennsylvanian and Permian time along two main tectonic trends, which may, in part at least, have developed contemporaneously. One trend is northeasterly, as evidenced by the Nemaha anticline, the Abilene anticline, the Voshell anticline, and many other smaller northeastward-trending folds. The other trend is northwesterly, as evidenced by the central Kansas uplift, the Salina basin syncline, and smaller northwestward-trending folds. The area of the Abilene anticline was actively deformed during much of this time, and the present surface rocks, for the most part Permian in age, show many normal faults

Jan 5, 1955

By Willard J. Tesch

The borehole surveying system described is used to measure ground movement over and near operating mines. The magnitude of any movement is con-trolled by the ground conditions, the mining method, and the mine progress. By correlating these factors with the measured movement, equations can be determined for predicting the effect of mine progress on the surrounding ground. These equations can also be used to predict the effect of mining a similar deposit with similar ground conditions. The borehole surveying system is used in cased boreholes to measure the inclination of the hole in two perpendicular planes. Grooved casing with an inside diameter of 3 inches is grouted in the hole. Two probes are used to obtain the data necessary to calculate the hole position versus depth. One of the probes contains a gyroscope, and the other contains a biaxial force-balance [s]ervoaccelerometer. Readings from the transducers in the probe are punched on paper tape by an automatic data acquisition system as the probe is pulled up the hole at 30 feet per minute by an electric winch. Readings from the inclinometer probe containing the accelerometer are used as input data to an on-line programmable calculator as they are punched on tape. The calculator computes preliminary results of the survey being made. The paper tapes obtained during the survey runs are later used as input to a computer that computes the offsets of the drill hole versus depth in true north-south and east-west planes. These offsets are printed for each foot of depth. Changes in these offsets between the first survey and subsequent surveys are a measure of the horizontal ground movement between the surveys.

Jan 1, 1976

By O. Engler, J. Friis, S. J. Andersen, C. D. Marioara, Y. Aruga

"This work investigates the early stages of precipitation in Al-Mg-Si alloys using a combination of high angle annular dark field scanning transmission electron microscopy (HAADF-STEM), selected area electron diffraction (SAED) and density functional theory (DFT) calculations. Quenching from a high temperature (solution) treatment leaves an alloy in a high-energy unstable state, with solute and vacancies in supersaturation. Thus, already at room temperature (RT) a large driving force exists for solute atoms to cluster and order on the Al matrix positions. DFT calculations indicate that strain alleviation is important at this stage. We observe larger regions of Mg-dominated L10 ordering together with a Si-dominated ordering of smaller clusters in a particular geometry connected to the Si-network found in all the precipitates. In addition, building blocks of the main hardening phase ?'', stabilized by vacancies, are observed at higher, typical pre-aging temperatures.INTRODUCTION Al-Mg-Si (6xxx) alloys are widely used in industry due to a combination of favorable properties such as strength, ductility, corrosion resistance and lightweight. They undergo a complex production route, usually consisting of the following thermo-mechanical steps: casting, homogenization, extrusion or rolling, solution heat treatment (SHT) and artificial aging (AA). The SHT takes place at temperatures above the solvus line in the phase diagram, usually between 500°C and 570°C, followed by a rapid cooling to RT. This ensures that high amounts of solute atoms, uniformly distributed on the cubic face-centered (FCC) Al lattice, together with a large amount of vacancies, are quenched-in. The material is therefore in a super-saturated and unstable state. During the subsequent AA, at temperatures usually between 150°C and 220°C, the solute atoms diffuse and cluster, forming metastable precipitates with needle/ lath/ rod morphologies having the main growth direction along <100>Al. With a maximum alloy solute level (Mg + Si) of only 2%, these precipitates increase material strength up to three times that of pure Al, and are essential for the final material properties. To understand and make better use of this phenomenon, it is imperative to research the nature of the precipitates (such as their atomic structure), measure their key parameters (number density, size, volume fraction) and connect it to material properties (Marioara, Andersen, Zandbergen & Holmestad, 2005; Marioara et al., 2014). The following precipitation sequence is given for the Al-Mg-Si system:"

Jan 1, 2018

By Jose R. Hernandez-Aguilar

"This paper revisits data from previous studies (Hernandez-Aguilar, 2010, 2011) to better understand the link between metallurgical performance and bubble size (Db) in the collection (pulp) zone of full-scale flotation systems. To obtain conclusive evidence exclusively related to the collection zone, the experiments were designed in those studies to control both the bubble generation method and the influence of the froth zone on the overall metallurgical performance in industrial and pilot-scale flotation columns. The results showed that generating small (1 mm) bubbles had a significantly higher beneficial effect on the recovery of coarse (> 75 µm) particles compared to the effect on fine (< 16 µm) particles. It was observed that, under certain practical conditions, the recovery of liberated coarse particles was close to zero, but could be recovered almost completely by reducing bubble size. The evidence revealed an apparent inconsistency of the so-called “k–PSb” model: at constant froth recovery, increasing Sb did not always result in higher recovery, as the model predicts. Overall, the evidence showed that it was not the Sb but the bubble size that proved to be a more reliable link to metallurgical performance. Only under highly suboptimal bubble size conditions (Db = 3 mm) did the froth zone seem to play a role, possibly reflecting an effect of pulp-zone bubble size on froth stability. However, for the most part, both the bubble generation method and the froth zone played inconsequential roles. An empirical model that better describes the kinetics of particle collection is proposed.INTRODUCTIONStudies showing the link between metallurgical performance and the interactions of bubbles and particles in the collection (pulp) zone of industrial flotation systems are scarce. A common belief among a number of mineral processing practitioners considers that “small” bubbles are favourable to recover “small” particles and that, conversely, “larger” bubbles are more efficient at floating “coarse” particles. However, perceptions like this are often substantiated by anecdotal evidence rather than by rigorous data analysis. What is the optimum bubble size for a particular case, how operating conditions can be adjusted to control a target bubble size, and whether the bubble size needs to be changed in the first place to optimize the operation of a flotation plant are questions that usually remain without solid answers.The studies reported by Hernandez-Aguilar (2010, 2011) were, in part, performed to test the effect of bubble size on flotation response and, thus, help address some of the above questions. The focus of the studies was to optimize the metallurgical grade-recovery performance of the Cu-Mo separation plant at Teck’s Highland Valley Copper mine (HVC). The outcome was compelling and significantly valuable: consolidated production data showed that in 2008 – the year when the recommendations from the studies started to be implemented – the Mo plant recovery was 89%, and gradually increased to 92% and 94% in 2009 and 2010, respectively, along with an increase in Mo concentrate grade. (Note: these production figures were not included in the publications quoted above; however, they were shown during an oral presentation at the 43rd Annual Meeting of the Canadian Mineral Processors in 2011.)"

Jan 1, 2016

By Diómedes Sarmiento

Centromin is a Mining Company owned by the peruvian government. It&apos;s considered as one of the major mining and metallurgical center in Latin America. The main operations of the Company are: Mining, Concentration, Smelting and Refining of several metals (Figure 1). At present, dentromin operates with six mining units: Cerro de Pasco, Morococha, Casapalca, San Cristobal, Yauricocha and Cobriza; which are located in the Central part of the country between 2 100 m and up to 4 800 m above sea level. The installed capacity of concentration all together ranges about 14 000 tons per day. There&apos;s also a metallurgical plant in La Oroya with a production capacity of 220 000 tons per year of refined metal. The production of the Company is diversified in twenty two (22) final pro- ducts, whose major productions are : copper, lead, zinc, silver, gold, tungg ten and bismuth. The minor products are: arsenic, antimony, cadmiun, teluriun, sulphuric acid, indium, selenium, etc.

Jan 1, 1979

By W. G. Stevenson

The centennial celebration of the AIME provides an opportunity to review some aspects of developments in mineral exploration since 1871 and to reflect on the changes that have occurred. During this period seven Factors have fundamentally altered the procedures, techniques and philosophy of mineral exploration. Although most of these developed gradually, and unheralded, they have drastically altered the art of exploration and discovery. 1. ADVANCES IN MINING AND MILLING The development of huge shovels, draglines, trucks and track-mounted drills, to replace the wheelbarrow and single Jack has drastically expanded the rate of production for both underground and surface mining. The development of selective flotation has had a far-reaching affect on mineral exploration and has made it possible to extract metals from the metallurgically complex deposits discovered years before. These developments have reduced the costs of mining and treatment, and have transformed material into ore that was considered waste one hundred years ago. These advances have permitted exploration for progressively lower grade and larger mineral deposits.

Jan 1, 1971

By Joseph D. Davis, G. S. Scott, Palmer B. Place

In the course of investigations made by the Bureau of Mines in cooperation with the Trent Process Corporation, into the properties of mixtures of oil and coal, a comparative study was ma.de of the 'behavior of these mixtures and their cons ti tuents on destructive distillation. Mixtures of a petroleum oil and bitminous coal were first distilled under carefully controlled conditions, and then charges of the eame oil and coal were distilled separately under similar conditions. The quantity and quality of the distillation products were determined, and. resul. ts were compared. This comparison furnished a basis tor judging whether grea, ... er values in products can be obtained by distillation of oil and coal in mixtures than· b:, their separate distillation. In making the mixtures, the method used in t~e Trent process for cleaning coal was followed. The paste or plastic emulsion of oil and coal obtained is called "amalgam".••• When finely pulverized wet coal is mixed with approximately one-third its weight of. oil and agitated with a large excass of waj;er, the coal and oil agglomer-. ate to form a paste or amalgam. A large portion of the extraneous ash original.17 in the coal remains suspended in the water, with which it can be decanted from the amalgam. Ir the latter is now thoroughly drained and kneaded after the manner ot working butter, most of the excess water enclosed can be removed. This, .in brief, constitutes the Trent process for cleaning coal, The effectiveness of the process for cleaning coal has been studied••••. by the Bureau, and the results of the study show, in general, a high ash reduction an:l good recovery of combus ti bl e f'rom coals of medi ,mi and hi~ rg,nk. With coals of tht lowest rank, such as li@!Rlites, the process is only moderately auccessf'ul, owing to the difficulty attending the preparation of coherent amalgams, Oil losses to wate~ and refuse are small, and apparently it would be practical to rotate the same oil in the cleaning process if some efficient method were devised for recovering it from the amalgam, so a distillation process naturally suggested itself.

Dec 1, 1921

By W. J. Kroll

This investigation was undertaken with the idea of studying a number of unconventional methods for making chromium to determine whether production cost could be lowered and to discuss some of the more recently described processes. These comprise carbon reduction of the oxide in vacuum, nitride dissociation, fusion electrolysis of the chloride, amalgam method, iodide dissociation, calcium reduction of the oxide, hydrogen reduction of the chloride, and reduction of this compound with zinc and with magnesium. Preference is given this latter method, which is shown in a flow sheet. Special attention is given purification of the chloride and separation of the vacuum at about 900° C. A reduction method for the elimination of traces of oxide contained in pure chromium by pure hydrogen obtained from the dissociation of zirconium hydride is described. The various methods are compared in a table in the appendix. Aluminothermic and silicothermic methods for the reduction of chromium oxide and the aqueous electrolysis have been disregarded in this investigation.

Jan 1, 1950

By F. C. Langenberg

It is the purpose of this paper to present a limited amount of experimental data obtained on the Charpy impact machine. Several concrete examples will be given showing the relation existing between the results obtained upon the Charpy impact machine and the behavior of the parts tested when in service. Although it is pointed out by G. Charpy and A. Cornu-Thenard1 that it is first necessary to determine a satisfactory working process before studying the correlation between the results of these tests and the behavior of the parts tested in service, this ideal course cannot be strictly adhered to. If a certain part repeatedly fails in service and the available tests give no indication of its cause, advantage must be taken of any possible means that will solve the problem; therefore, even though the reliability of the results obtained by the Charpy impact machine have been questioned by many, it was necessary to use this instrument without giving careful consideration to its possible sources of error. It would seem that the conclusions arrived at by Charpy -and Cornu-Thenard justify the assumption that the precision obtained by this machine is comparable to that obtained by the other more common forms of testing. The second general conclusion arrived at by them, namely, that the impact test upon notching does not offer any direct correlation to the ordinary tests, is also borne out by many hundred determinations made in the laboratories at Watertown Arsenal. This statement, however, should be qualified as follows: Other ordinary tests include the static tensile test from bars taken in a longitudinal direction and the Brinell hardness. The author agrees that no law is yet known for the computation of the energy absorbed in the rupture of bars of the same material but of different dimensions. Although Charpy and Cofnu-Thenard do not refer in their valuable article to the use of the Charpy machine in industrial practice, it should be stated that this instrument has been in constant use in the laboratory at Watertown Arsenal for several years, and is now employed for both routine and experimental testing.

Jan 1, 1920

By S. Banisi, H. R. Iran-Manesh, M. R. Shayestefar, H. Shekarchyan

"Flotation circuit of the Sarcheshmeh copper mine processes 41,000 t/d of ore at 0.9% Cu. Chalcopyrite and chalcocite are the main copper bearing minerals which are concentrated in three stages of roughing, cleaning and recleaning. Due to major changes in the reagent addition regime, make-up balls combination, and throughput to the primary mills an increase of 2.6% (from 87.2% to 89.8%) in the sulfide copper recovery was obtained during a period of six years. In order to investigate a rather large variation of monthly recovery (i.e., from 89.0 to 91.2%) of the plant after implementation of the changes, a study was undertaken to trace mineralogically possible causes of such a behavior. During a five-month period, monthly composite samples of feed and tailings of the flotation circuit were taken. At the first stage, the polished sample preparation method and the mineral counting procedure were optimized through a detailed study. Liberation degrees and types of locking of valuable minerals in both streams were determined in three size intervals of +75, +38 and –38 microns. It was found that for both recovery levels (i.e., 89% and 91.2%), there was no significant difference in the liberation degrees of valuable minerals in the flotation feed; the average liberation degree was 57.1% with a standard deviation of 3.1%. On the other hand, in the case of high recovery level (i.e., 91.2%) there was a reduction of 15% in the amount of free valuable minerals in the tailings stream. The highest reduction was found for the -38 micron size range in particular for chalcocite. It was concluded that compared to the grinding circuit the change in the flotation circuit had the significant effect on the plant recovery increase.INTRODUCTIONThe Sarcheshmeh copper ore body contains 1 billion tonnes averaging 0.90 % copper and 0.03 % molybdenum. It is located southeast of Iran and currently processes 41,000 t/d of ore. The concentrator has been in operation since 1982. Last year, the concentrator produced 130,000 tonnes of copper and 4,000 tonnes of molybednite concentrate. An expansion to the plant which will increase the throughput by 50% is under construction and will be effective in 2003.After three stages of crushing, the ore with a F80 of 12.7 mm is fed to 8 parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of the first stage of flotation is reground and the tailings constitute the main component of the plant final tailings. A concentrate with an average grade of 30.5% Cu is obtained after cleaning and re-cleaning stages. At the present time, the average sulfide copper recovery of the plant is 89.8% which varies depending on the operating conditions and ore type."

Jan 1, 2003

By A. Mezei

A process has been developed to recover cobalt, gold and bismuth from the NICO deposit in Canada. The process consists of a bulk rougher flotation followed by an optional separation by flotation to produce a bismuth concentrate and a cobalt concentrate, grading on average 12.8 % Bi and 1.8% Co, respectively. Cobalt is leached by pressure oxidation at 97% efficiency, producing a pregnant solution suitable for further processing. Before optimization, about 38% of the bismuth is recovered in the selective concentrate, wherefrom it is recovered by ferric chloride leaching followed by cementation, at overall efficiencies exceeding 98%. The gold is recovered from the cobalt and bismuth leach residues by cyanidation, at 95% efficiency. Previous work described the results of the leaching of the concentrates, recovery of gold (Part I) and the recovery of bismuth (Part 11). This work describes the results of the cobalt recovery testwork and the resulting conceptual flowsheet. The sulfur dioxide-oxygen system was the most efficient for the oxidationlprecipitation of iron and arsenic from cobalt bearing pregnant leach solutions, producing purified solutions assaying <1 mg/L As and <8 mg/L Fe. The efficiency of cobalt solvent extraction (from the resulting purified PLS containing about 1 g/L Co) ranged from 9 1 to 98%; the co-extractions ranged from 88 to 100% for copper, from 28 to 33% for magnesium, from 6 to 7% for calcium and from 96 to 98% for zinc. Sulphide precipitation processes (using either sulfur dioxidelelemental sulfur or sodium sulfide) produced residual copper concentrations in the cobalt strip solution ranging from 0.1 to 2 mg/L Cu. About 90% of the calcium and magnesium were separated by scrubbing the organic with cobalt sulfate, producing residual concentrations of 12 mg/L Ca and from 180 to 330 mg/L Mg in typical feed solutions directed to cobalt carbonate precipitation tests. Cobalt carbonate precipitation efficiencies ranged from 75 to 78% when calcium and magnesium co- precipitations were restricted from 8 to 17% and 0.3 to 0.4%, respectively. The limits required by most stringent specifications on commercial grade cobalt carbonate products were met or exceeded.

Jan 1, 2002

By David Mitchell, Daan Delport, Frik Marx, Mike Shapiro

The use of copper cooling elements to enhance refractory lining life in pyrometallurgical vessels is common in modern furnace design. Copper elements typically incorporate a base panel with cast in cooling passages and an integral extended surface (formed of pins, fins or similar) onto which suitable refractory material is cast. The metallic, conducting parts are located towards the cold face of the lining, with the refractory material interposing between the coolers and the process. The coolers are designed to arrest wear of the refractory material by freezing a layer of slag or partially reacted material onto the hot face, thereby rendering the lining inert to further attack [1]. These coolers have achieved success in new furnaces and relining of existing vessels. The freeze layer is prone to mechanical damage during operational and process instability. The coolers must be capable of extracting the additional heat flux under upset conditions to facilitate rapid self-repair of the freeze layer and prevent further refractory wear in the intermittently exposed areas. However, operational experience has shown that the high heat flux capabilities of the copper coolers have some disadvantages, especially in processes producing vapours which form corrosive compounds when condensing onto the cold surfaces of the coolers. When corrosive conditions develop, the entire composite will fail due to loss of structural integrity, and as a result of leaks from the cooling water passages. This paper describes recent developments in the approach to the design and manufacture of copper cooler panels customized to achieve specified heat fluxes and temperature profiles, thereby reducing the formation of corrosion mechanisms in applications with corrosive gaseous environments and enhancing refractory lining life. Condensation formation prevention is achieved by adjusting the geometry of the cooler and selection of suitable materials of construction.

Jan 1, 2007

By R. R. Odle

A laboratory and pilot-scale investigation of direct smelting of copper concentrates was conducted to determine the effect of concentrate analysis and the degree of roasting on the yield and quality of copper produced in a subsequent smelting step. A stoichiometric :mOdel of direct smelting is presented which predicts the? quantity of copper; slag, and white metal&apos; prOduced&apos;and the distribution of copper in these phases as a function? of the concentrate &apos;(calcine) analysis.

Jan 1, 1983