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The eighth annual RMCMI Mine Rescue Evaluation was held during the 1999 Institute convention at the Crested Butte Marriott Resort, Colorado. [ ] Elliott Tsosie, Mary Rowell, Janna Aune, Guy Horton, Art Saenz and Bill Balaz were pleased with the 1999 Mine Rescue Program, The Mine Rescue committee is but one of 17 volunteer committees; this one co-chaired by Bill Balaz, Elliott Tsosie and Janna Aune. There were three evaluators, Art Saenz, Western Environmental; Mary Rowell, ProCare and Guy Horton, Horton Rope Rescue. Nine teams participated and brought 89 participants, 26 spouses, nine children and five teens. The teams were BHP LaPlata, BHP Navajo, BHP San Juan, all from Farmington, NM; Bridger Coal Co., Rock Springs, WY; The Coteau Properties Co., Beulah, ND; Northwestern Resources, Jewett, TX; Powder River Coal Co., Gillette, WY; Solvay Minerals, Green River, WY and Western Energy Co., Colstrip, MT. To observe were members from Centralia Mining Co., Centralia, WA. [ ] Evaluators: Mary Rowell, Guy Horton &Art Saenz The three professional evaluators provided challenging scenarios for the Mine Rescue participants on the grounds of the resort property. Some attending convention were alarmed to see stretchers with "victims" and phoned the hotel switchboard to report the catastrophe. Team members had no prior knowledge of what the scenarios might be and derails were provided only as they arrived at the scenario location, with one exception. Seven to ten or more articipants make up each team. Some teams also ring an alternate and team trainer. For identification purposes, the "victims" wore white shirts emblazoned both front and back with a red VICTIM designation. There was a staged demonstration by three response teams who had not previously worked together. This was enacted on the convention floor following a Technical Session meeting attended by about 200 members. Three response teams from BHP's Navajo and San Juan Mines together with Coteau Properties' Freedom Mine provided a demonstration of their response skills. For never having worked together before, they presented a very professional enactment. This is described at length in the Welcome Session copy. Training sessions were held Saturday and Sunday to enable the teams to prepare for the scenarios. A swap meet was Saturday evening for teams to swap hats, shirts, belt buckles, stickers etc. Each participant received a royal blue golf shirt and ivory cap with royal blue bill. Culmination of the Mine Rescue program was a Tuesday morning Awards breakfast when teams were presented with jade acrylic plaques. [ ] Teams participating in the Mine Rescue Evaluation received trophies on Tuesday morning. The captains above represent BHP-La Plata, BHP-Navajo, BHP¬San Juan, Bridger Coal Co., The Coteau Properties Co., Powder River Coal Co., Solvay Minerals, and Western Energy Co. Sponsors who contributed financially to the 1999 Mine Rescue program were Balaz & Associates, CMC, Civil Constructors, Colstrip Electric, Farmington Fire Equipment & Safety, Fisher Safety, Flanders Ellctric Motor Service, Magnum Metals, Richard Palmtag, Rocky Mountain Electric Motors, Sedona Consulting and Western Water Consultants. [ ] Evaluator Art Saenz likes to have the upper hand and will take whatever steps necessary to get it.

Jan 1, 1999

By Robert J. Linney

IN the latter part of the year 1943, it was decided to experiment with diamond-drill blast holes in the Old Bed magnetite mine at the Mineville mines of the Republic Steel Corporation, in sections of the ore body where vertical dimensions were large and where development for sublevel bench stopes had already been done. Other possibilities of application for this drilling method were in the reclamation of old floors and pillars in the pitching folds of the Lower Old Bed ore body and in the reclamation of large vertical pillars in the " 21 " Bonanza-Joker ore body. The cross sections and views of the two ore bodies will show the conditions to which the diamond-drill method has been applied. The Lower Old Bed magnetite ore body is lenticular and dips approximately 30° toward the southwest and strikes generally southeast and northwest. An intensely folded structure, which pitches generally on an axis approximately at right angles to the strike, is continuous from the uppermost old workings down to the present bottom of the mine, and probably continues with depth. On successively lower levels this folded structure becomes greater in vertical dimensions and the two folds are closer together in a horizontal dimension. As shown in Fig. 1, this folded structure gives to parts of the ore body considerable vertical dimensions, which provide areas that were developed for sublevel bench stoping above a slushing drift, as shown in Fig. 2. It was in one of these previously developed sublevel bench stopes, No. 26 on the 1185-ft. level, that the diamond-drill blast-hole method was applied for mining the upper sublevel. The lower sublevel was underhand-bench-drilled to the finger raises of the scram drift below. The roof of this stoping area began to break and spall on the lower sublevel, causing the roof to become inclined. Along the hanging wall of the stope this reached an elevation about half way between the two sublevels. A smooth schistose seam in the hanging wall causes the ore to break away freely. Therefore, if bench mining had been carried on in the upper sublevel, a dangerous condition of poorly supported benches would have resulted. When the drillers were slashing across the sublevel to footwall and hanging wall to make a bench, spalling roof of the opening below would probably have carried away part of the bench, and would have left a dangerous place in which to continue underhand benching. A 6 by 8-ft. raise had been driven between the lower stope and the upper sublevel about 200 ft. from a manway. It was decided to use this raise for a starting cut and to arrange holes as shown in Fig. 3, so that a slot would be opened across the ore body between the footwall and hanging wall. This slot would then be used for the purpose of breaking into the ore; by blasting vertical rings of holes drilled at right angles to the axis of the pitching folds as nearly as possible. Layouts of this slot drilling and the rings adjacent are

Jan 1, 1945

By Robert J. Linney

In the latter part of the year 1943, it was decided to experiment with diamond-drill blast holes in the Old Bed magnetite mine at the Mineville mines of the Republic Steel Corporation, in sections of the ore body where vertical dimensions were large and where development for sublevel bench stopes had already been done. Other possibilities of application for this drilling method were in the reclamation of old floors and pillars in the pitching folds of the Lower Old Bed ore body and in the reclamation of large vertical pillars in the "21" Bonanza-Joker ore body. The cross sections and views of the two ore bodies will show the conditions to which the diamond-drill method has been applied. The Lower Old Bed magnetite ore body is lenticular and dips approximately 30° toward the southwest and strikes generally southeast and northwest. An intensely folded structure, which pitches generally on an axis approximately at right angles to the strike, is continuous from the uppermost old workings down to the present bottom of the mine, and probably continues with depth. On successively lower levels this folded structure becomes greater in vertical dimensions and the two folds are closer together in a horizontal dimension. As shown in Fig. I, this folded structure gives to parts of the ore body considerable vertical dimensions, which provide areas that were developed for sublevel bench stoping above a slushing drift, as shown in Fig. 2. It was in one of these previously devel-oped sublevel bench stopes, No. 26 on the 1185-ft. level, that the diamond-drill blast-hole method was applied for mining the upper sublevel. The lower sublevel was underhand-bench-drilled to the finger raises of the scram drift below. The roof of this stoping area began to break and spall on the lower sublevel, causing the roof to become inclined. Along the hanging wall of the stope this reached an elevation about half way between the two sublevels. A smooth schistose seam in the hanging wall causes the ore to break away freely. Therefore, if bench mining had been carried on in the upper sublevel, a dangerous condition of poorly supported benches would have resulted. When the drillers were slashing across the sublevel te footwall and hanging wall to make a bench, spdlling roof of the opening below would probably have carried away part of the bench, and would have left a dangerous place in which to continue underhand benching. A 6 by 8-ft. raise had been driven between the lower stope and the upper sublevel about 200 ft. from a manway. It was decided to use this raise for a starting cut and to arrange holes as shown in Fig 3, so that a slot would be opened across the ore body between the footwall and hanging wall. This slot would then be used for the purpose of breaking into the ore, by blasting vertical rings of holes drilled at right angles to the axis of the pitching folds as nearly as possible. Layouts of this slot drilling and the rings adjacent are

Jan 1, 1946

By D. L. Katz

The Oklahoma City Wilcox sand, discovered on March 26, 1930, has produced 394 million barrels of crude oil and 819 billion cubic feet of natural gas as of July I, 1941. The 100,000-bbl. wells, pressures of 2600 lb. per sq. in., and flows of "Wilcox" sand caused exciting moments in the early life of the field. Today the reservoir is substantially at atmospheric pressure and produccs 75,000 bbl. of crude oil daily through 466 pumping wells. Early estimates of the initial crude-oil content of the "Wilcox" sand1 indicated that about 600 million barrels of crude oil would remain after the pumping wells ceased to produce economically. The Oklahoma City Wilcox Secondary Recovery Association was organized in March 1939, to study the possibilities of recovering a portion of this oil. Only a very brief summary of the information available can be presented in a paper of this nature. Some simplifications and generalizations are necessary to give the over-all picture in SO short a space and statements made may not apply to a specific well or small area. Since the area under consideration by the Association is that north of the south line of section 22-11-3, some information will be presented for areas A, B, C, 1, 2, 3 and part of 4 of Fig. 2, and some will be for the entire Wilcox zone west of the fault. This paper is essentially an abstract of an engineering report made by A. D. Small, B. T. Murphree, M. C. Sons and the author for the Oklahoma City Wilcox Secondary Recovery Association on Sept. 5, 1940. Some information is taken from a report to the Association by C. V. Sidwell on March I, 1940, and some data are from a recent report by A. D. Small. It gives a brief history of the Wilcox zone, estimates of the initial oil contents of the reservoir, present conditions in the reservoir, and probable results from several programs of "secondary recovery." History of Wilcox Zone The Wilcox sand lies on the flank of the general structure, as shown by cross section of Fig. I2,3,4 and areal map of Fig. 2. The sand is of Ordovician age and is composed of well assorted, rounded sand grains with no apparent cementing material. The loose texture of the sand is shown by the enormous quantities of sand production from the wells and from the soft cores. The productive sand varies from zero thickness on the east to 220 ft. maximum and to zero thickness on the west because of water. The development of the pool4 as dependent upon drilling restrictions within the city limits, which accounts for the spacing of drilling operations listed in Table I by years. In 1931, areas I through 7 of Fig. 2 were open for drilling with the exception of some west-edge properties, and all wells drilled before Feb. I, 1935, were in this portion of the field. Nearly all wells drilled from 1933 to 1037 were in areas A, B and C. Since 1939, the newly drilled wells have gradually enlarged the producing area toward the west in the northern portion.

Jan 1, 1942

By D. L. Katz

The Oklahoma City Wilcox sand, discovered on March 26, 1930, has produced 394 million barrels of crude oil and 819 billion cubic feet of natural gas as of July I, 1941. The 100,000-bbl. wells, pressures of 2600 lb. per sq. in., and flows of "Wilcox" sand caused exciting moments in the early life of the field. Today the reservoir is substantially at atmospheric pressure and produccs 75,000 bbl. of crude oil daily through 466 pumping wells. Early estimates of the initial crude-oil content of the "Wilcox" sand1 indicated that about 600 million barrels of crude oil would remain after the pumping wells ceased to produce economically. The Oklahoma City Wilcox Secondary Recovery Association was organized in March 1939, to study the possibilities of recovering a portion of this oil. Only a very brief summary of the information available can be presented in a paper of this nature. Some simplifications and generalizations are necessary to give the over-all picture in SO short a space and statements made may not apply to a specific well or small area. Since the area under consideration by the Association is that north of the south line of section 22-11-3, some information will be presented for areas A, B, C, 1, 2, 3 and part of 4 of Fig. 2, and some will be for the entire Wilcox zone west of the fault. This paper is essentially an abstract of an engineering report made by A. D. Small, B. T. Murphree, M. C. Sons and the author for the Oklahoma City Wilcox Secondary Recovery Association on Sept. 5, 1940. Some information is taken from a report to the Association by C. V. Sidwell on March I, 1940, and some data are from a recent report by A. D. Small. It gives a brief history of the Wilcox zone, estimates of the initial oil contents of the reservoir, present conditions in the reservoir, and probable results from several programs of "secondary recovery." History of Wilcox Zone The Wilcox sand lies on the flank of the general structure, as shown by cross section of Fig. I2,3,4 and areal map of Fig. 2. The sand is of Ordovician age and is composed of well assorted, rounded sand grains with no apparent cementing material. The loose texture of the sand is shown by the enormous quantities of sand production from the wells and from the soft cores. The productive sand varies from zero thickness on the east to 220 ft. maximum and to zero thickness on the west because of water. The development of the pool4 as dependent upon drilling restrictions within the city limits, which accounts for the spacing of drilling operations listed in Table I by years. In 1931, areas I through 7 of Fig. 2 were open for drilling with the exception of some west-edge properties, and all wells drilled before Feb. I, 1935, were in this portion of the field. Nearly all wells drilled from 1933 to 1037 were in areas A, B and C. Since 1939, the newly drilled wells have gradually enlarged the producing area toward the west in the northern portion.

Jan 1, 1942

By P. D. Rao, H. B. Charmbury, D. R. Mitchell

As a result of a changing market pattern for anthracite, there is now an increasing demand for steam size and a decreasing demand for domestic sizes. To help the producer meet this new demand, the authors present the technical aspects of the problem of crushing domestic sizes to steam fineness and size. The market pattern for anthracite coal has changed considerably in recent years. The demand for domestic sizes has been gradually decreasing and the demand for the steam sizes has been increasing. The companies however aim for a maximum production of domestic sizes, since these sizes command premium prices. To meet the market demands, particularly in certain periods of the year, further size reduction of the coarse clean coal is necessary. The present practice of multistage crushing in toothed rolls leads to complicated flow arrangements and is high in cost. Data presented herein indicate that crushers with high reduction ratios could be used with a resulting simplification of crushing flows and an appreciable reduction of capital and operating costs. The different types of crushing equipment investigated included the rotary-type breaker, the jaw crusher, the impactor-type crusher, and the Aerofall mill. CURRENT PRACTICE The standard sizes of anthracite are presented in Table I and market specifications require that any given size be maintained with only a limited amount of oversize and undersize material. The 1958 prices of the various sizes also are presented, so it may be seen that size reduction is only performed when dictated by market demands. At the present time the general practice of crushing anthracite consists of removing the top size and crushing it in a toothed double roll crusher to the next largest size. Consequently, to reduce the domestic egg, stove, and nut sizes to steam sizes of pea, buck, rice, and barley, the original material must be screened to produce the three top sizes and each size must be crushed in three separate rolls to the next largest size. Each crushed product must then be rescreened to remove the oversize for re crush ing. Thus a typical operation would consist of scalping the egg size from a mixture, crushing it to stove size, and rescreening the product. This same cycle would be repeated for the stove and nut sizes requiring three screens and three crushers as well as conveying equipment to handle the crushed products. This results in an expensive series of operations to make a product which sells for a lower price. If the three top sizes are crushed in one operation to the smaller sizes then the yield of steam sizes is low and the production of Nos. 4 and 5 sizes is large. These sizes sell for lower prices than the steam sizes. A typical operation is presented in Fig. 1. From the figure it may be seen that the original cleaned anthracite contains 67.4 pet domestic sizes, 26.4 pct steam sizes, and 6.2 pct fines. After multistage crushing of the domestic to steam sizes the composition of the product is 0.6 pct domestic sizes, 89.6 pct steam sizes, and 9.8 pct fines. From a technical standpoint these results are reasonable but the handling and crushing costs are

Jan 1, 1961

Profiled on these pages are the eight RMCMI 1994-95 Scholarship Award Winners. Six of the eight winners were present as guests of the Institute at the June convention in Crested Butte, Colorado. Each received recognition at the Institute Awards Luncheon Monday. June 26, 1995. Each state vice president served as the host for the collegiate winner. The first and last winners - Arizona and Wyoming - were absent. Besides receiving a plaque from the Scholarship Chairman, Alan Czarnowsky, each winner receives $750 tuition credit for the coming academic year, as well as transportation, hotel lodging and meals at convention as guests of the Institute. This latter portion is paid by each state vice president. Because this is a continuing award, the winners from the previous year also receive a second $750 for tuition credit in their senior year. A resident from each of the eight RMCMI member states is selected from among Board approved institutions. Criteria for these awards to be made in the spring annually is that the recipient (1) Is a full time sophomore student who will use the tuition credit during the junior year. (2) Be both a U.S. citizen and a resident of the member state. (3) Maintain good academic standing, be pursuing a degree in a mining related field and have an expressed interest in Western coal as a possible career path. The scholarship program includes curricula other than in the mining departments, but in engineering disciplines, such as geology, mineral processing and metallurgy. Contributions to this worthy project are tax deductible to each individual RMCMI member or his firm. Through annual donations of $100 or more, the Patrons of the Institute fund the scholarship program. A listing of current donors is found elsewhere in this book. Donations to this worthwhile program are payable to RMCMI Scholarship and may be mailed to the RMCMI Office. 3000 Youngfield Street, Suite 324, Lakewood, Colorado 80215. Profiles of the 1995 Scholarship Winners follow. [ ] ARIZONA WINNER Brad Carroll University of Arizona - Tucson, Arizona Brad Carroll's hometown is Tucson, Arizona but his summer work is in Oriskany Falls, New York where he is a quality control technician, truck driver and equipment operator for the Eastern Rock Products. Brad served as vice president and treasurer for the SMWAIME student chapter at the University of Arizona for the past three years He was assistant and mine foreman at the San Xavier Mining Laboratory near Tucson. Upon graduation Brad plans to find a job in a southwest mine. He is a totally self supporting student. [ ] COLORADO WINNER Zane Znamenacek Colorado School of Mines - Golden, Colorado Zane Znamenacek hails from a small town in Colorado - Alamosa - where he 4 was a 4.3 GPA in high school. As a student at Colorado School of Mines, he has many hobbies to ease the tension of education. Foremost among these is skiing, mountain biking, billiards and spring rugby. He hopes to further mining technologies and contribute to the industry by working in it full time. In 1997 he will pursue whatever opportunities are available to graduating mining engineers

Jan 1, 1995

By E. P. Youngman

Before the World War the use of zirconium and its compounds was largely experimental and confined almost wholly to Germany and Austria . Occasionally small lots of zirconium ore had been produced in the United States ( 1,000 pounds in 1869 , 26 tons in 1883 , and 3,000 pounds in 1903 ) and used (with monazite ) in the manufacture of various lighting devices , especially incandescent gas mantles ; but it was not until 1906 that the discovery of the natural zirconium oxide in large quantities near Sao Paulo , Brazil , first gave promise that a zirconium industry eventually might be developed . Edward Rietz , who worked the Brazilian deposits , interested German chemists in extending the use of zirconium , and by 1911 trade journals were listing German firms as manufacturers of compounds or products derived therefrom . For a time zirconia was one of the various oxides used in the Nernst lamp . It was used also for crucibles and hearths as refractory material ; in the manufacture of chemical utensils ; in enamels ; in the place of bismuth subnitrate for defining Röntgenray pictures of the stomach ; as a pigment ; in medicine ; and in a number of other ways . The Germans were credited also with producing a remarkable zirconium steel , which , it was claimed , was much superior to other alloy steels . During the World War the United States Government investigated the possibilities of zirconium , as well as those of other steel -hardening elements , and a large automobile manufacturer actually purchased a substantial quantity of ore with the purpose of employing zirconium in automobile steels . Immediately after the armistice the Government ceased its investigations ; but despite the difficulties experienced by various investigators in reaching concordant results , interest in the use of zirconium as a steel -hardening agent , as well as in other uses of the element and its compounds , continued . In 1918 the Foote Mineral Co. , which had been interested in zirconium for several years , was exploiting the deposit in Brazil formerly worked by the Germans and was supplying zirconia in the form of ore , brick , and cement for refractory and other purposes . About the same time the Electro Metallurgical Co. , which before the entrance of the United States into the war was experimenting with zirconium alloys , supplied in tonnage quantities the demand for these alloys . In 1919 it was reported that American manufacturers , after several years of research, were producing in commercial quantities pure zirconium oxide , which was finding a place in the manufacture of small refractory articles , scorifying dishes , crucibles , tubes , and one- piece furnace linings . In 1920 the lowering of the price of zirconium metal powder ( because of improved metallurgical methods ) promised the production of the metal on a semicommercial scale . In 1924 the Titanium Alloy Manufacturing Co. introduced a new opacifying agent , " Opax , " made from zircon . A year or two later some of the large pottery manufacturers became interested in the use of zirconium oxide as an opacifying agent , the increased production of titanium pigments having resulted in the production of zirconium as a by-product at a price that could compete with tin oxide . In 1928 it was reported that the industrial use of zirconium oxide was increasing steadily in the manufacture of brick , crucibles , and other refractory products , such as high- temperature cements . However , the reluctance of producers to install additional and more suitable

Jun 1, 1931

By Jack Bedder

"Editor’s note: We would like to thank those who make this annual issue possible. A special thank you goes out to the industrial minerals annual review editor, to the Industrial Minerals & Aggregates Division Technical Committee chair and vice chairs, and to the authors of the individual commodity profiles.Jim Norman, member SME, of Tetra Tech, is the industrial minerals annual review editor.Ebrahim K. Tarshizi, member SME, of Michigan Tech, is chair of the Industrial Minerals &Aggregates Division’s Technical Committee.The technical committee vice chairs include: • Hyunjung Kim, member SME, of Chonbuk National University, agricultural materials.• Xihui Yin, member SME, of Kemira, ceramics and refractory materials.• Nikhil Gupta, member SME, of Virginia Tech, chemical raw materials.• Jasson Giddens and Swadhin Saurabh, members SME, of Oil Dri Corp of America and Mill Creek Engineering, respectively, chemical raw materials.• Snehamoy Chatterjee, member SME, of the Michigan Tech.• LeDeana Roberts and Shubham Verma, members SME, of Imerys and U.S. Oil Sands, respectively, fillers and pigments.• Thomas Hilderbrand and Kaiwu Huang, members SME, of Thiele Kaolin Co. and Virginia Tech, respectively, sorbents, filters and process aids.• Rhaghav Dubey and Edgar Blanco, members SME, of Outotec and FL Smidth, respectively, specialty mineralsAntimony generally occurs along with gold, lead, copper and silver. There are more than 100 antimony minerals, although the sulfide mineral stibnite is the main one.Antimony enters the supply chain either as a result of mining (primary production) or recycling (secondary production). China has by far the largest antimony resources and is, as a result, the world’s center for antimony mine production. In 2016, it accounted for more than three quarters of the global mine supply out of a world production figure of 130 kt (143,000 st), according to the U.S. Geological Survey. The biggest, and most high-grade deposits are found in southern China, with the largest mines located in Hunan, Guangxi and Yunnan provinces. Tajikistan, Russia, Australia and Bolivia are the largest mine producers outside China. A new antimony project in Oman made progress in 2016, producing first metal in April 2016 in test facilities, and then ordered three furnaces for a 26 kt/a (28,600 stpy) antimony metal and oxide smelter that is due on stream in late 2017.Antimony ingot production is not as widespread as is antimony mine production, owing to the high levels of pollution that are associated with antimony smelting. It is for this reason that smelting has been largely confined to China and other Asian countries, where environmental legislation is less stringent.Antimony oxide production is undertaken on four continents with China being the biggest producer. There is also a considerable amount of oxide production in Europe — which relies on imported feedstock, mostly from China. Antimony also enters the supply chain through recycling. Lead containing antimony is produced in a number of countries, mainly in secondary lead smelters. Typically, a blast furnace charge containing used or discarded battery plates, type metal or bearing metal is reduced to lead bullion, which is refined in reverberatory furnaces and melting pots. In the main, the antimony recovered is consumed in metallurgical applications. China was the biggest secondary antimony producer in 2016, and accounted for 25 retardants and lead-acid batteries. Flame retardants accounted for roughly 50 percent of consumption in 2016 with batteries representing about 25 percent. Other end uses include plastics and heat stabilizers, ceramics and glass and a variety of metallurgical applications. Demand has been sluggish since the global financial crisis, mostly because key markets failed to recover strongly from the downturn, particularly in Europe."

Jan 7, 2017

By N. W. Bliss

"Bauxite has been the almost universal ore of aluminum for about 100 years, with most of the world's reserves oc¬curring in the developing countries. As a result of fiscal and political activities prompted in part by the energy crisis, the formation of the IBA combined with increasing freight and labour costs, these reserves are no longer as competitive or as secure as they once were. This has given fresh impetus to the development of processes for alumina production that utilize raw material other than bauxite. Canada has no bauxite and it is prudent to look for alter¬native raw materials within our own boundaries suitable for one of these processes.Of the three igneous rocks that have been used as a source of alumina, only anorthosite is considered as a potential source in Canada. The Al content would be extracted by a sinter process, which currently requires a raw material with a. minimum of 28% Al2O2. Thus, about 4 tons of anorthosite and 10 to 12 tons of limestone are needed to produce one ton of alumina. There are two petrologically distinct types of anorthosite, one containing the plagioclase andesine-labradorite and around 26% Al2O3, and the other containing upwards of 30% Al2O3, in by¬townite-anorthite; for our purposes they may be con¬sidered as low- and high-alumina anorthosite respectively. The high-alumina anorthosites are somewhat rare; the low-alumina anorthosites are much more common, partic¬ularly in Eastern Canada. Geologically speaking, there is no shortage of anorthosite in Canada. The reserves are im¬mense, although their low Al2O3 content coupled with con¬siderations of location, access and proximity to limestone could be constraints to their development. Fly ash from coal-fired thermal power stations and nepheline syenite can also be used in a sinter process, but neither is likely to be utilized in Canada in the immediate future.The H+ process utilizes an acid attack to leach Al from a raw material with at least 20% Al2O3. One of the best raw materials is kaolinite-bearing clay; not only is it high in Al, but most of it can be extracted. With the possible exception of the Whitemud clay in Saskatchewan, Canada is apparently deficient in clays with sufficient reserves to support an alumina plant utilizing the H+ process. Low-alumina clays with illite and montmorillonite may also prove to be a suitable raw material if advantageously sited with respect to industrial centers. The average shale con¬tains only 15% Al2O33, much of which may be contained in minerals, such as feldspar, not amenable to acid leaching. If shale is used as a source of alumina, it will be because it contains some other valuable commodity, such as oil or uranium, the extraction of which leaves the spent shale upgraded in extractable Al2O3 and lowers the over-all costs to a point where it is economic, Thus, rejects from coal washing plants are most commonly considered as a source of alumina; they are already mined, are available in large and ever increasing amounts, and may contain significant amounts of combustible material which upgrades the Al2O3, content of the resulting ash. Coal rejects from existing and developing coalfields in Alberta and British Columbia are a possible source of raw material, but present data indicate they may contain insufficient Al2O3."

Jan 1, 1976

By Enver Alan

This Research & Development (R&D) study was conducted in the field of applied blasting engineering in Ethiopia, within the AKH Railway Project, which was built in Girena, within the boundaries of the province of Mersa, in the Basalt Quarry at km 346, Ashangi Formation, occured in Eocene geological time in the aphanitic trachibasalt. Under the scope of the study, which has been performed to determine the effects of the artificial discontinuities at different depths and the effects of artificial discontinuities through structural damage risk, three test areas, at the same location and same formation, were considered. Six numbers of artificial discontinuities, each 10 m (32, 81 ft) transverse, were created by using pre-splitting and drilling-blasting technology. The number of sequential artificial discontinuities on each test area is two and they have the same length. The length of artificial discontinuities are; 6,80 m (22,31 ft) at the first test area, 13,60 m (44,62 ft) at the second test area and 17,00 m (55,77 ft) at the third test area respectively. The distance between sequential artificial discontinuities is 8 m (26,25ft) on each test area. Thirty-one shots were executed in order to produce vibration (10 shots on the first test area, 11 shots on the second area, 10 shots on the third test area). One hundred and twenty-three seismic recordings were obtained during the execution of the shots by using 4 seismographs. The closest distance between the shot points and the nearest measuring station is 5,30 m (17,39 ft). During the shots, the rock was cracked and there was no fragmentation and/or fly rock. At the first test area, measured PPV values are between 25,40-127,00 mm/s (1-5 inch/s) before the first artificial discontinuity and less than 50,80 mm/s (2 inch/s) after the second artificial discontinuity. At the second test area, measured PPV values are between 76,20-177,80 mm/s (3-7 inch/s) before the first artificial discontinuity and less than 76,20 mm/s (3 inch/s) after the second artificial discontinuity. At the third test area, measured PPV values are between 50,80-279,40 mm/s (2-11 inch/s) before the first artificial discontinuity and less than 50,80 mm/s (2 inch/s) after the second artificial discontinuity. Peak reduction rates in Peak Particle Velocity (PPV) values determined by measuring as follows: In the first test area 95,77% (106,00 mm/s to 4,48 mm/s) (4,17 in/s to 0,18 in/s), in the second test area 79,77% (176,00 mm/s to 35,60 mm/s) (6,93 in/s to 1,40 in/s), in the third test area 95,85% (from 205,00 mm/s to 8,55 mm/s) (from 8,07 in/s to 0,34 in/s). The arithmetic mean of the differences of PPV values (decrease rate) is 82,66% in the first test area, 69,37% in the second test area, and 92,03% in the third test area. The highest sound value is 138,50 dB and is defined as “noise”. When the measured parameter values are examined it was determined on the USBM norm graphs that these vibrations carry a risk of structural damage in the region before the first artificial discontinuity. While passing through artificial discontinuities the vibrations progress by losing energy and the vibrations passing through two artificial discontinuities do not carry structural damage risk under current conditions.

By Syd S. Peng

According to accident statistics (l, 2), roof fall accidents occur quite frequently at the intersections of underground? openings and account for up to 30% of the roof fall fatalities. But this figure will certainly go up considerably if roof falls not associated with any fatal accident are included in the analysis. Unfortunately, nonfatal roof falls are seldom documented. In the underground coal mines, there are two types of entry intersections. One is the three-way (V type) and the other is the four-way (T type) intersections. A three-way intersection occurs whenever the mainline track haulageway turns at an angle into a section haulageway. The four-way intersections by far the most common type in the U.S. under-ground coal mines are formed by the intersections of two entries which are oriented perpendicular to each other or by the intersections of entries and crosscuts. However, crosscuts are frequently driven at an angle to the entry to facilitate machine movements. The four-way intersections formed by this method of crosscut development are angled. At the intersection of two underground openings, the diagonal roof span is wider than the width of each individual opening. This is one of the major reasons why roof falls occur more likely at the intersections than in the entry between pillars. Consequently, special intersection support plans are usually employed (Figs. 1 and 2) for intersections. The most popular practices are to decrease roof bolt spacing 1 ft. less than or to increase roof bolt length 1 to 2 ft. longer than those employed in the entry between pillars (3). The increase in bolt length or decrease in bolt spacing can be applied to all of the bolts in the intersection or employed only for the strategically located bolts depending on the individual mine practices. These rather generalized practices seems to be adequate for some areas but in many cases inadequate for others. Occasionally, the neighboring two lines of chain pillars are staggered such that the crosscuts in one row of chain pillars face the centers of the chain pillars in the neigh-boring rows. This layout forms T-type three-way intersections which are also inadequate for most weak roofs (3, 4). A study was therefore initiated to develop special techniques for intersection supports. The first phase of the study was the analysis of the stress fields for the underground entry intersections. A three-dimensional stress analysis was necessary for this purpose. Unfortunately, the three-dimensional problems of complicated boundary conditions are beyond the existing analytical methods. The three-dimensional finite element method (FEM) was therefore employed in this study. In the analysis, square pillars of 40 feet wide and an entry height of 8 feet were assumed while entry widths varied at 14, 20 and 26 feet. This model was selected to simulate the mine where roof falls data were collected. The second phase of the study was the collection of field data. The sizes of the roof falls including heights, lengths and widths were collected and analyzed. In the third phase study, the results of the first and second phases were compared and discussed. The analysis of field data revealed an arching zone within which roof falls occurred. The arching zone above an entry obtained empirically was compared with the calculated stress contours. Assuming that the bolts must be anchored in the rock strata outside the arching zone, the preferable systems of bolting were proposed. Also the amount of pretension required to inhibit sliding between layers was discussed.

Jan 1, 1978

By Kira Mizell, James R. Hein, Tracy Conrad

"The Ninetyeast Ridge (NER) extends roughly N-S as a topographic feature for 5000 km, from 31° south to 10° north (Fig. 1) where it becomes buried beneath the Bengal Fan at its northern end. These two geographic characteristics make the NER an ideal natural laboratory for determining the influence of latitude and terrestrial input to ferromanganese (Fe-Mn) crust formation and composition. In addition to study of geographic controls on crust composition, the NER is the largest feature in the Indian Ocean that may host crusts of economic interest.Crusts were collected from 9 of 33 dredge sites from 1.4° north to 26° south latitudes during the 2007 R.V. Roger Revelle Knox06 cruise. Nineteen samples were analyzed for 69 elements each. Nine bulk crusts and 10 layers from four crusts were analyzed. Three of the crusts sub-sampled for layers did not have the bulk crust analyzed and the weighted means for those layers were calculated to obtain bulk compositions. One sample was left out of the statistical analyses because it is phosphatized and chemically quite different from the other samples (Table 1).In order to determine changes in composition with latitude, Pearson product-moment correlation coefficients were calculated for the chemical dataset and latitude of dredge sites, as well as water depth. At the 95-99% confidence level, the following elements increase in crusts moving north along the NER: Mg, Ti, S, Cd, Ni, Tl, Mn, Ba, Co, Cu, In, Li, and Zn; and the following elements decrease northward: Fe, Fe/Mn, P, As, Sc, the rare earth elements plus yttrium (REY), Bi, Ge, and Pb. These correlations indicate that the iron oxyhydroxide component and its associated elements are more prevalent to the south, and the manganese-oxide phase and associated elements increase toward the north end of the NER. This change occurs between dredges D20 and D14 (Fig. 1), with crusts from D20 and south having Fe/Mn !1.0 and those north of D20 ""1.2, with the exception of the phosphatized crust. Crust composition also changes with water depth: Th, Ge, Pb, and the middle rare earth elements (REEs) increase with increasing water depth. Thorium commonly increases in deeper water crusts, which explains the higher Th concentrations in Atlantic and Indian ocean crusts in general compared with Pacific crusts (Fig. 2 and Hein et al., 2012), including the samples analyzed here with a mean water depth of 3584 m. However, water depth does not explain the latitudinal changes because the mean water depths of samples along the northern and southern segments of the NER are roughly comparable. Surprisingly, Si and Al, representing the aluminosilicate/detrital phase, do not correlate with latitude; that is as the ridge approaches and is buried by the Bengal Fan the Si and Al contents do not increase. However, overall, Si and Al have higher concentrations in NER crusts than in central Pacific crusts. Ba can be used as a proxy for primary productivity and statistically does increase to the north along the ridge in the direction of the equatorial zone of high biological productivity. Correlations of Mg, Li, and Zn with latitude may also reflect increased primary productivity at lower latitudes."

Sep 1, 2014

By G. C. Smith

Process control applications in the minerals processing industry have grown extensively over the years due to the many benefits derived, both financially and operation-wise. Although control theory is widely known, conventional control does not address many of the problems posed by minerals processing circuits. Some of these problems include interactions between process units, large dead times, few and sometimes unreliable measurements of process variables, non-linear systems, disturbances entering the process, processes with varying dynamics, and many more. For the past 26 years Mintek has been active in the field of providing specific, customized solutions for the minerals processing industry. This paper focuses specifically on applications in milling and flotation control. The underlying control principles have been applied to a wide range of other processes including submerged arc furnaces. Advanced control application is implemented on a common advanced control platform, called the PlantStar. Much innovation has been brought to the fore in terms of adapting theoretical techniques and developing new methods to resolve specific milling and flotation circuit issues by using process control. The general approach adopted by Mintek when implementing advanced control is to enforce stabilization of the process as a solid foundation for optimization strategies to be built upon. In doing this, a robust control solution is presented, and the end result is flow and level stability, better product quality regulation, increase in throughput and a stable feed to downstream processes. The MillStar advanced milling control system ensures that the milling circuit is at all times easy to operate while achieving optimum product quality and throughput. Stabilization should include product size regulation and sump level control. Another important aspect of stabilization control is the handling of milling circuit constraints without introducing cyclical behaviour and with minimal disturbance to the final product quality. Process constraints may include factors such as process water limitations, cyclone feed density, and cyclone underflow angle. An optimization strategy should be developed to increase the mill throughput without affecting product quality or introducing mill overloads. The relationship between solids feed and power draw is examined and manipulated for optimization. The interactive nature of flotation circuits and the presence of recirculating streams tend to make satisfactory control almost impossible with conventional techniques such as PI (Proportional plus Integral) or PID (Proportional, Integral and Derivative) control. The FloatStar Level Stabilizer simultaneously considers all the flotation levels in a circuit by use of a multivariable algorithm, and takes compensating action in anticipation of disturbances. The FloatStar Flow Optimizer is an optimization strategy that controls the concentrate flow rate between various sections of a flotation circuit (e.g. Rougher ?Cleaner; Cleaner ?Recleaner). By controlling the concentrate flow between the sections, the mass pull, circulating loads and residence times of the circuit can be stabilized, thereby enhancing the performance of the flotation process. Optimizing grade and recovery are coupled with the dynamic optimization of circulating flows. The FloatStar Grade-Recovery Optimizer will maximize the grade in a section, whilst ensuring that recovery or any other process specific limits remain within predefined limits. With successful applications on several milling and flotation plants worldwide processing amongst others PGMs, gold, copper and nickel, Mintek has shown that it is a leader in milling and flotation control innovation

Jan 1, 2004

By Theresa Y. Schulz

It is interesting to study the noise standards that have been promulgated in the US over the last decade or so. These regulations are likely to have long-lasting impact. The Occupational Safety and Health Administration (OSHA) Hearing Conservation Amendment (March 1983) continues to have influence not only in the workplace but in the debate over new regulations. Both the Mine Safety and Health Administration (MSHA) and Federal Railroad Administration (FRA) regulatory preamble documents state the desire to be consistent with OSHA. An examination of Table 1, which compares the major components of the three regulations and the 1998 NIOSH “best practices” criteria, depicts the extent to which that intent is met. There has been some regulatory activity in the last decade which may give some hope for evolution and updating based on the wealth of science that has occurred during the last quarter of a century since the OSHA regulation was enacted. However there has also been some “back-sliding” toward more lenient standards. The MSHA noise standard made regulatory progress in September 2000 by emphasizing engineering and administrative controls, followed by personal protective equipment, in the hierarchy of noise intervention. MSHA’s requirement for technician certification (today only available from CAOHC) strengthened the training requirements for audiometric testing in hearing conservation programs and MSHA also added the requirement of dual hearing protection at 105 dB TWA. There were many subtle differences between OSHA and MSHA based on comments and a desire to clarify some of the vague aspects of the OSHA noise standard, and meet the needs of the regulated mining industry. One example pertains to the ceiling for exposures. OSHA says, “no exposures >115 dBA,” which is interpreted to mean no unprotected exposures above that level, giving credit for the assumed effectiveness of hearing conservation programs, hearing protection devices and administrative and engineering controls. MSHA specified that a “P” code1 violation be issued for any protected or unprotected exposures >115 dBA. The Federal Railroad Administration (FRA) noise standard for railroad operating employees, which went The Modern Evolution of Hearing Conservation Regulationsinto effect February 26, 2007, was expressly based on the OSHA standard but also uses MSHA for comparison. The preamble states that the FRA defers to OSHA as the “primary regulator of noise in the workplace,” but also acknowledges the need for some departure from the OSHA regulation (FRA Preamble II.B). As an example, FRA requires testing at 8000 Hz “because it will allow employers to identify hearing loss sooner.” The FRA rejected MSHA’s hierarchy of noise controls in favor of requiring specific engineering interventions and focusing on appropriate hearing protection which would still allow communication and audibility, and identification of excessive noise through employee-filed “excessive noise reports.”2 Where OSHA has no specific mandate requiring employees to take advantage of the employer-paid audiogram, it has been traditionally a condition of employment and is generally accepted that OSHA-covered workers require an annual audiogram. MSHA addressed this issue in its preamble; however, they made no significant change. MSHA employers are required to offer annual audiograms but MSHA stopped short of requiring employees to comply with annual audiometric testing. The MSHA preamble does allow that mine operators can also make audiometric monitoring a condition of employment. FRA requires employees to

By M. A. Reyes, J. Bickson, J. R. Srednicki

Federal laws mandate the presence of refuge alternatives (RAs) in underground coal mines and provide guidelines for harmful gas removal. After an explosion, harmful gases can follow survivors into an RA and can reach a deadly concentration. Investigators from the National Institute for Occupational Safety and Health (NIOSH) tested 12 ventilation system configurations (VSCs) to purge sulfur hexafluoride (SF6) tracer gas from a built-in-place (BIP) RA in the Experimental Mine in Bruceton, PA. Using a borehole air supply, tests were conducted with airflows of 750 and 1,000 standard cubic feet per minute (SCFM). Excluding a VSC where the air inlet was 1.98 m (78 in) from the air outlet, the results show that the time to purge the SF6 concentration from 1,000 parts per million (ppm) to 25 ppm ranged from 26 to 29 minutes for an airflow of 750 SCFM and ranged from 18 to 23 minutes for an airflow 1,000 SCFM. Mines can use this information to reduce purge time in case of a mine disaster and determine which VSC to avoid. INTRODUCTION After a fire or an explosion in an underground mine, the concentration of carbon monoxide can reach a fatal level. In a mine disaster, the main goal is to escape. However, if miners cannot escape, they have to migrate to a location to find refuge and, if possible, communication, and a borehole source for a fresh air supply. To ensure survival of the miners, Federal laws mandate the presence of a refuge alternative (RA) within 304.5 m (1,000 ft) of the nearest mining face. In addition, Code of Federal Regulations Title 30 CFR Section 7 sets the guidelines for personnel tracking, RAs, and the removal of noxious gases in underground coal mines (Mineral Resources, 30 CFR, 2014). Federal regulations define an RA as a protected, secure space with an isolated atmosphere and integrated components (Mineral Resources, 30 CFR, 2014). There are three types of RAs: portable metal RAs, portable tent-type RAs, and built-in-place (BIP) RAs. A BIP RA is constructed using masonry or other stoppings to create an isolated space in a mined-out area. Blast doors are installed in one or both of the stoppings for the miners to enter the RA. Furthermore, RAs must provide the following components: breathable air, food, water, air monitoring, and harmful gas removal. Federal regulations have emphasized the mandate for breathable air and harmful gas removal in the following words: “Purging or other effective procedures shall be provided for the airlock to dilute the carbon monoxide concentration to 25 ppm” (Mineral Resources, 30 CFR, 2014). The purpose of the mine ventilation system is to provide a sufficient quantity of air to the underground mine workings, to dilute methane and other contaminants, to maintain a suitable working environment, and to prevent accidents from happening (Cheng and Wei, 2014). In addition, the organic nature of coal amplifies risks from poisonous gases, explosions, and fires (Engström et al., 2018). If an underground coal mine ventilation system fails after an explosion, the concentration of toxic gases, most notably carbon monoxide, will increase. Carbon monoxide is dangerous because it creates tissue hypoxia by binding to hemoglobin (Özmen and Aksoy, 2015).

Jan 1, 2019

By H. K. VAN POOLLEEN, W. Hurst, H. C. Bixel

In an earlier publication* I showed the development of the instantaneous point source solution for a well producing at a constant rate at the center of a system of two radial, adjoining sands of different permeabilities. I also showed how a wellbore radius could be incorporated if this were the development required. For the purpose of that paper and the point I wished to make, this solution served. It yielded a simplifying solution of sands in series identified by Eqs. VI-26 and VI-27. (These equations are reproduced in part by Bixel and van Poollen as their Eq. 3, Page 302.) I am well aware that the simplified solutions are approximations. The facts that these Ei-functions satisfy the diffusivity equation and meet the boundary conditions both at the well and at the interface between sands, and that the earliest time considerations yield zero pressure drop, are evidence that the solutions are good approximations. This was the purpose for these formulas, which I interpret that where the permeabilities are of finite order of magnitudes in the respective media, the relations give satisfactory answers. However, from the authors' point of view there are reservations, which leads me to consider their entire generalization as illustrated in their Figs. 1 through 14. What I understand from their storage capacity, relationship Fs, which is the ratio of the product of porosity and compressibility of Sand II to Sand I, that is employed here by the authors as a correlative means of identifying one figure from another, there is nothing in the exact mathematics that supports such a term as a means of correlation. This is taken out of context, unsustained that this applies. Here the mathematics are known, where the true correlation is the diffusivity term of one sand to another. This entails permeability that refers back again to the term M in their paper for the transmis-sibility of fluid from one sand to the other that should have also been a part of this storage capacity. In Figs. 8 through 14, which purported to show the simplified equation following this correlation, there are apparent discontinuities for M equal to or less than 0.1000 that differ from the trends for greater values of M. In this connection there is no reference radius given for this cumulative pressure drop in the paper that would permit me to make these calculations. However, what is clear for M equal to 0.1000 or less we are treated to almost a closed barrier at the interface; a condition entered by the writers without qualification. This across-the-board treatment of these simplified formulas, whatever the motive of the authors, enters into a range of permeabilities that even I as the contributor would have reservations that such could apply. Under comparable conditions I would be looking at a closed reservoir or some facsimile for such. The extent to which the approximation or the exact equations pertain is not evident in the generality offered by Bixel and van Poollen in their paper. Only when they or other writers show the comparison of the pressure distributions within the formations can one be able to judge the rational limits to subscribe to the simplified formulas.

Jan 1, 1969

By J. R. Tonry

In the grinding of dry non-metallic materials such as rocks, ores, gypsum, lime, and cement clinker, it is recognized that only a relatively small portion of the energy is utilized for actual size reduction, and the major portion of the energy used is converted into heat. As a result, heat is generated inside grinding mills imparting to the ground product a higher temperature than that of the mill feed. In some cases, as for instance in grinding cement clinker and gypsum into finished cement, it may be desirable to keep this temperature increase within certain limits, to avoid any change in the properties of the product and, because consumers may specify a maximum temperature of the finished product for reasons associated, with the tulimate use of the cement. The actual proportions of heat and useful work are somewhat indefinite. Some tests have indicated that nearly all the energy is converted into heat, and. other investigators have found that they could account for "everything except the noise." Investigations sponsored by the American Institute of Mining and Metallurgical Engineers, and carried out with an adiabatic calorimeter, indicated that a surprisingly high 19 percent of the energy might actually be used for size reduc¬tion. However, for the purpose of the following discussion it will be assumed that: (Fig. 1) 10% is lost in motor and gears 10% is lost by radiation and convection from the mill body 10% is used for size reduction, and. that 70% of the energy input leaves the mill as heat in the ground. product. A typical example (Fig. 2) would be a 7x22-ft tube mill operating in open circuit, driven by a 400 hp motor, charged with 77,000 lb of grinding media and producing 70 bbl or 26,300 lb of cement per hour. Reducing this cement from a clinker size of 96 percent passing a No. 16 mesh sieve to a specific surface of 3,230 sq cm per gm (Blaine) will increase the temperature of the cement approximately 143F above the temperature of the feed, as explained later. If the tempera¬ture of the clinker therefore is 125F, the finished cement will be 268F, which in some cases may be too hot. : When cement temperatures approach a range during grinding where dehydration of the gypsum becomes a problem, a logical consideration would be some means of reducing the temperature before it attains the range conducive to gypsum dehydration. Regardless of the varying degrees of heat generated within the mill or mills during grindingto various finenesses, consideration of the maximum safe limit for any type of cement being ground, which may vary fromplant to plant, govern in refining the mill circuit to obtain a safe temperature range. All plants are not necessarily required to do the same cooling job, and may or may not use cement coolers following the grinding circuit. (Not discussed in this paper). In the past, various means have been employed for controlling this temperature increase, Such means have included water cooling of the outside of the mill shell; equipping the shell with cooling fins; injection of moisture inside the mill; ventilation through the mill; and., in cases of closed-circuit grinding, drawing cooling air through the classifier; the purpose in all such cases being to remove heat from the grinding system.

Jan 1, 1958

By Lane White

The 1989 Annual Meeting and Exhibit of SME in Las Vegas, NV, February 27 - March 2, 1989, was the largest ever, drawing more than 4000 regis¬trants and 240 exhibit booths. Technical sessions included presentation of nearly 500 papers, including symposia on Biotechnology in Minerals and Metals Processing, Surplus Capacity in the International Metals Industry, and APCOM `89 - the 21st International Symposium on Application of Computers and Operations Research in the Mineral Industry. Technical sessions covered the gamut of minerals industry interests; however, the proximity of the meeting to recently surging gold development in Nevada drew special attention to gold operating practices. Several of the papers on gold processing are reviewed below. To keep pace with the strong world- wide interest in gold development, SME has scheduled three additional upcoming programs that will focus specifically on gold: World Gold '89 - Gold Forum on Technology and Practices will be held November 5 - 8, 1989 in Reno, NV, as a joint meeting and exhibit between SME and the Australasian Institute of Mining and Metallurgy. Technical sessions will encompass a full program of papers covering geology and exploration, mining, and mineral and metallurgical processing. Cooperating sponsors for this meeting include the Canadian Mineral Processors Division of CIM, The Institute of Metals, The Institution of Mining and Metallurgy, The Mining, Geological & Metallurgical Institute of India, and The South African Institute of Mining and Metallurgy. Gold '90 - Salt Lake City, UT, February 26 - March 1, 1990, will be a symposium held in conjunction with the 1990 SME Annual Meeting. Current research on the processing of refractory gold ores will generate innovative papers on process mineralogy at this symposium, with special emphasis on gold with pyrite and gold with carbon. Other sessions are planned for gold exploration and mining, gold metallurgy, gold plant design and practice, and gold in scams. Advances in Gold and Silver Processing Conference, October 1 - 3, 1990 will be held in Reno, NV. Preliminary planning calls for technical sessions on heap leaching practice, refractory ore processing, non-traditional lixiviants, bioleaching, metal recovery from solution, and the environmental aspects of processing operations. PLANT DESIGN, OPERATION Technical sessions organized by the Minerals and Metallurgical Processing Division of SME at this year's meeting in Las Vegas covered a wide range of gold operating practices. Eight papers relating to gold plant design, operation, and equipment are summarized here. Preprints of these papers are available by writing to Publications Sales, PO Box 625002, Littleton, CO 80162-5002 ($2.50 per paper for AIME members, $4 for nonmembers). Continuous elution process A. M. Stone, of Bechtel Australia Pty. Ltd., described the "Development of the Continuous Elution Process for Extraction of Gold from Carbon" (pre-print 89-75). Developed by Bechtel Pacific Corp. for Ok Tedi Mining Ltd., the Continuous Elution Process uses the same reagents as batch processes for elution of gold from carbon, while yielding significant savings in stripping vessel size and energy consumed in comparison with equivalent batch processes. The process is also simpler to operate than batch processes. Other advantages include reduced maintenance costs, higher average eluate grade, and near steady final eluate grade. Pilot testing for the Continuous Elution Process was conducted at Western Mining's Kambalda gold plant beginning in December 1982. The Continuous Elution Process is a continuous countercurrent system operating in a pressure vessel at elevated

Jan 1, 1989

By Jerold Bishop, Grant Bearss, Aleksey Egorov

"When a shock wave is released in the bottom of a drill hole filled with wet concrete slurry, the result is called an electrohydrodynamic effect, or EHDE. The energy is carried through the fluid with minimal loss because the slurry is an incompressible medium. Repetitive discharges provide a series of energy pulses that compact the soil along the pile shaft and at its base. The process is analogous to repeated applications of low-energy blast densification or to dynamic compaction. In general, the process is not depth dependent. EHDE was identified in the 18th century by an American scientist who observed that a spark between two electrodes placed under water induced a burst of pressure. In 1957, two Russian engineers formulated a process and received a patent for a method of creating a pressure wave by transforming power from an electrical discharge to a mechanical impact. They derived more than 200 applications for technical and industrial processes.In the 1970s, civil engineers at the Engineering and Building Institute in St. Petersburg, Russia, began to look at EHDE to improve the capacity of drilled piles. Early experiments showed that electric discharges in a wet concrete slurry mix could create a pressure wave of up to 26 MPa (3700 psi). Later tests demonstrated that this pressure level could be created in a drilled pile hole. The researchers devised equipment and in the early 1990s began to study pulse technology as applied to drilled piles.This application was accepted into the Russian Federation Building Code in the 1990s. The code treats “regular” drilled piles with equivalent bearing area as having roughly half the capacity of a driven pile. However, the code considers drilled piles treated with the EHDE technique a direct “one for one” substitution for driven piles and allows higher values if the capacity is established by static load testing. Hundreds of projects have been completed in Russia using this technology to value engineer drilled pile projects. In several instances, the number of drilled piles installed was fewer than half of the number originally planned for due to the capacity increase that EHDE-treated piles provided.Drilled PilesIn drilled pile construction, the EHDE process creates a mechanical pressure wave that impacts the base and sidewall of a drilled pile shaft. Soft, poorly consolidated or loose soil is compacted by successive pulses of pressure created by electrical discharges. These mechanical impulses are made by inserting a discharge probe into the wet concrete slurry. A discharge across the tip of the probe creates the pulse that travels outward to the sidewall and base of the drilled pile through the incompressible medium of the slurry. The power pack ratings for equipment presently in use are 35 and 50 kJ (kilojoule).Each pressure wave strikes the soil, causing compaction and expanding the hole. Workers apply the pulses until the size of the hole stops increasing. They watch the slurry level during pulsing. When the level quits dropping, further soil compaction in the drill hole sidewalls and base is not possible. After the pulse process is complete, workers insert a rebar cage and the pile is finished.The process does not alter the soil mineralogy. The electrical discharge is internally grounded and does not arc to the sidewalls of the drill hole. Soil is not scorched or altered in any manner; it is only compacted by the pressure wave. This technique has been used in urban areas with no effect on electrical or electronic equipment."

Jan 1, 2009