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By Gui DeReamer

Located on the Wabash River in the central portion of the state, the city of Lafayette, IN is relatively small compared to most cit-ies that incorporate tunnels into their combined sewer overflow (CSO) systems. The original CSO plan for the city was to replace two aging lift stations that were built in the 1950s. The original con-cept was to build a new sewer using conventional cut-and-cover construction to eliminate one of the lift stations ? the Parking Lot Lift Station ? and build a new regionalized one at the location of the Pearl River Lift Station. The initial sewer design was sized just to handle the flow that would eliminate the Parking Lot Lift Station. The planning for this project was being developed as part of the city?s CSO long-term control plan (LTCP) that focused on reducing the overflows from the city?s 14 overflow points. However, during the development of the CSO LTCP, a major industrial user indicated it would be increasing its flows to the system, which required additional conveyance capacity to meet the city?s CSO LTCP. After reviewing the design concepts with city representatives during the planning phase of the project, the mayor indicated that he wanted to minimize impacts to downtown businesses; therefore, trenchless alternatives were considered. The last piece of the puzzle that drove the project away from a small diameter pipeline was discovered during a flow monitoring program to collect flow data for the CSO LTCP project. This study revealed a major restriction in the 100-year-old Pearl River Trunk Sewer, located beneath four active railroad tracks. To eliminate this flow restriction, provide more overflow storage and reduce the impact to local businesses, a new 2.7- to 3.6-m- (9- to 12-ft-) diameter tunnel sewer became the most viable solution.

Jan 1, 2011

By W. W. Adams

"During the calendar year 1925 the Bureau of Mines received from the operators of several hundred mines and quarries complete accident records covering deaths, permanent disabilities, and “lost time” temporary injuries. Some of the companies, - those that participated in the National Safety Competition-also furnished information showing the number of man-hours of work performed during the year. With these two types of information it was possible to calculate the accident-frequency and accident-severity rates of the different plants. Complete and comparable reports of accidents and Man-hours of “exposure” were received from 34 underground metal mines, 59 underground coal mines, 15 underground mines producing other nonmetallic minerals, and 102 open quarries or open-pit mines. The figures in the table appended to this report show the accident severity rates and other data for each of the 34 underground metal mines and for the group as a whole. It is proposed to publish similar figures for the other classes of mines and quarries at an early date.The 34 metal mines listed in the table are situated in Arizona, Idaho, Kansas, Michigan, Minnesota, Missouri, New Jersey, New Mexico, Oklahoma, South Dakota., and Utah; they employed 5, 925 men, an average of 174 men per mine, including employees above-ground as well as underground. With but few exceptions, 50 men was the minimum number of employees in any mine. The total period of employment or “exposure” at the 34 mines was 14,612,607 man-hours, an average of 2,466 hours per man, or the equivalent of 308 work-days of eight hours for each employee. Accidents resulted in 15 fatalities, 9 permanent disabilities, and 1,950 temporary “lost time” injuries (causing disability exceeding remainder of shift). The fatal and nonfatal injuries represented 132,575 days of disability. Thus the accident frequency rate for the group was 135 accidents per million hours and the accident severity rate was 9.073 days lost per thousand hours of 1 “exposure”. The severity rates ranged from 0.227 to 88.653, while the ·range of the accident-frequency rates was from 16. 6 to 307 .4. In compiling the accident severity rates shown in the table, all accidents caused by explosives were weighted at twice the number of days actually lost, as one of the purposes of the National Safety Competition was to stress the prevention of accidents caused by explosives."

Sep 1, 1926

By T. J. Lutz, J. L. Carr, J. A. Yonkey, C. C. Jobes

"Since 1984, the date of the first recorded fatality involving a remote controlled continuous mining machine, 38 miners in the United States have been killed when they were struck by a continuous mining machine. Proximity detection is a valuable tool for preventing fatalities, and the use of these systems is expected to significantly improve safety. Proximity detection provides a means to prevent accidents like these. Researchers at the National Institute for Occupational Safety and Health (NIOSH) conducted a series of underground tests to evaluate the accuracy and repeatability of proximity detection systems under a number of conditions such as varying the orientation and height of the personal wearable device (PWD), moving the PWD close to the trailing cable, and activating the machine’s mining mode. At the 2014 SME Annual Meeting, results from tests on first-generation proximity detection systems were presented. This paper will provide results from the similar tests performed with second-generation systems. The results show that performance has improved. For example, the negative impact of PWD orientation and the trailing cable has been reduced. However, some inconsistencies remain. Notably, performance changes dramatically when mining mode is activated. INTRODUCTION Since 1984 there have been 38 fatalities involving striking and pinning of the operator and other workers by the continuous mining machine (CMM) [1]. In August 2011, the Mine Safety and Health Administration (MSHA) published a proposed regulation that would require proximity detection systems on all continuous mining machines except full-face machines [2]. Proximity detection systems are designed to stop machine motion to protect miners from striking/pinning hazards. Several MSHA-approved proximity detection systems are commercially available. MSHA estimates that of the 38 fatal striking and pinning accidents that have occurred since 1984 involving CMMs, the use of proximity detection could have been a preventative factor in at least 28 cases [1]. Furthermore, MSHA estimates that proximity detection could prevent 20% of all deaths throughout the industry [3]. NIOSH researchers have been involved in the development and testing of proximity detection technology since creating the HASARD system [4] and have recently conducted research intended to develop more advanced proximity detection technology for CMMs [5,6,7,8]. In addition, NIOSH has also conducted research investigating how the signals of a proximity detection system propagate [9,10] and have developed a novel method for determining the location of a personal wearable device (PWD) relative to the magnetic field generators [11,12]."

Jan 1, 2015

The Bendigo Goldfield is AustraliaÆs second largest goldfield, having produced 22 million ounces of gold between 1851 and 1954. Development of the modern decline mine, initially for underground exploration and bulk sampling, commenced in 1998. The project is targeting new reefs below and along from historic workings beneath the city of Bendigo. The major technical challenge for the project is the inability to solely use drill core gold assays to estimate the grade of reefs, due to the coarsegrained nature of gold mineralisation, known as the nugget effect. In 2002, a program of bulk sampling commenced, resulting in the development of factors for converting drill hole grades to in situ reef grades. In mid 2006 commercial production commenced following the construction of a 600 000 t/yr processing plant adjacent to the Kangaroo Flat Mine portal, with plant feed being derived from four reefs in the Sheepshead and Deborah lines of reef. In early 2007 it became clear that the initial ore reserves had been overestimated, leading to an immediate change of strategy from production to exploration. A cash conservation program was also put in place. Exploration refocused onto the more historically productive Garden Gully and New Chum lines of reef, whilst a new grade estimation tool based primarily on the visual assessment of drill core was implemented. Immediate exploration success occurred with the discovery of four reefs in the Garden Gully line in early 2007. Successful step-out drilling during the year led to the decision to undertake trial mining in Gill reef. Trial production commenced in June 2008 and includes the processing of discrete ore blocks and the introduction of selective mining techniques. This paper examines the steps that have been taken to refocus exploration and to gain confidence in new exploration results, particularly through the visual grade estimation technique and assessment of trial mining Gill reef on the Garden Gully line. In addition, project factors relating to human skills, community relationships and financial management in a dynamic external environment are also addressed.

Jan 1, 2009

By Douglas Muir, Chas. A. Chase

The Aztec mine is not widely known, by reason of its isolation and the relative insignificance of its tonnage; financially, however, it has an enviable record and geologically it is extremely interesting. The town of Baldy, Colfax County, the mine camp, at an elevation of 10,000 ft. (3048 m.) is 8 mi. west of Ute Park, the terminus of a branch of the A. T. & S. I?. Ry. The mining tract was designated by the Maxwell Land Grant Co. as the Aztec Reservation, and lies on the east slope of Mt. Baldy, the summit of which, 12,491 ft. (3807 m.), is one of the highest in the state. Rich copper float was found on the northeast slope of the mountain in 1865, and discovery work was begun the next year. At the same time, gold placers were found on Willow Creek, on the southwest side of the mountain, and the Elizabethtown settlement followed shortly. Gold washing lasted many years. In 1867, placer gold was found also on Ute Creek, the tracing of which to its source led to the locating of the Aztec mine by Lynch, Dogherty, and Fosley in the following year. The early production came from the contact of shale and sandstone or from sandstone immediately above the contact. Exact record of production is lacking, but L. C. Gratonl gives estimates ranging between $1,250,000 and $1,500,000, of which about $1,000,000 was taken out in the first four years. After the exhaustion of the early bonanza, production continued intermittently until 1909, when the Maxwell Land Grant Co, undertook systematic development, through four adits at intervals of 75 ft. vertically or approximately 100 ft. on the dip of the contact, which was thought to contain generally a quartz-pyrite vein. The two intermediate adit? almost immediately reached ore, the most southeasterly in Fig. 2. This ore, however, was not mined until 1911-12; its grade was about $20 and

Jan 1, 1923

By J. Howard Boyce

INTRODUCTION The Tsumeb Mine is situated in the northern part of South West Africa, approximately 335 air miles north-east of the port of Walvis Bay on the Atlantic Coast, and 240 air miles north of Windhoek, the capital of South west Africa. The mine came into operation in the early 1900's and has operated continuously since then except during the two world warm and the depression years of the thirties, Tsumeb Corporation purchased the property in 1947, installed a new plat, and ham built a modern tom. In addition to the concentrator which began operations in 1948, a copper smelter was started in 1962 and a lead smelter in 1963, Tonnage of ore treated in the concentrator has varied through the years, reaching a maximum of 70,000 tons per month with a current mined tonnage of approximately 50,000 tons per month. Copper concentrates, lead concentrates and zinc concentrates are produced, and for a number of years the production of a germanium enriched concentrate was an important feature of the operation, Both copper and lead concentrates are smelted in Tsumeb, while the zinc concentrates are shipped overseas. Ore change, and the installation of the smelters in Tsumeb have necessitated modifications in the concentrator operation. This paper will describe current plant practice in some detail and will touch briefly on the techniques used from 1954 to

Jan 1, 1970

By P. R. A. Andrews

Areas of bubble contact were delineated for topaz and tourmaline in the presence of chloride, fluoride, sulfate, phosphate, oxalate, and citrate ions using dodecylamine hydrochloride collector. The limits of pH and the minimum collector concentration were established using the bubble pick-up technique. Bubble contact did not occur below the p.z.c. (point of zero charge), estimated to be 5.6 for topaz and 6.6 for tourmaline, in the presence of chloride ion. Flotation below the p.z.c. was, however, extended by citrate, oxalate, phosphate, sulfate, and fluoride ions. Oxalate, citrate, and fluoride ions displayed the greatest pH range of activation, followed by phosphate and sulfate ions. Chloride ion acted as an indifferent electrolyte. Topaz was selectively activated with respect to tourmaline by fluoride and phosphate ions. Topaz was activated by fluoride ion between pH 2.4 and 5.6. Tourmaline flotation in this pH range was virtually insignificant. Topaz was activated by phosphate ion between pH 4.9 and 7.4. Tourmaline flotation in this pH range was only partial. Activation appeared to be strongest when anionic protonated species were in maximum concentration. Strong activation also corresponded to very stable metal ion species thought to be formed on the surface of each mineral.

Jan 1, 1991

This is the sixth in a series of reports on properties of commercial motor fuels, made in accordance with a cooperative agreement between the Cooperative Fuel Research (C.F.R.) Committee and the Bureau of Mines of the Department of the Interior. The first report gave data on motor gasolines sold during the winter of 1935-36.3/ The present report covers 54 brands of gasoline represented by 1,480 samples plus those from the Pacific Coast, which were averaged by the Pacific Coast Technical Group before being submitted and are published exactly as they were submitted. In some areas reports were received from 8 to 10 companies on the same brands of gasoline, whereas in other areas only one company reported on a brand. Thus, it can be seen that some of the data are more representative than others. The inclusion of data from several cities within some of the districts has improved the report by reducing the number of brands reported by only one company. Several changes have been made in the districts. A new district, Memphis, Tenn., has been formed to include Arkansas and western Tennessee. The Rocky Mountain areas have been divided into three sections. Salt Lake City and Denver form the Middle Rocky Mountain district; Laramie, Wyo., Idaho, Montana, and Spokane, Wash., are included in the North Rocky Mountain district, and Arizona, New Mexico, and El Paso, Tex., have been combined to represent the South Rocky Mountain district. The Cleveland district has been expanded to include Ohio an,"- Indiana. Eastern Pennsylvania, Maryland, and northern Virginia are included in the Washington (D. C.) district. The Atlanta district includes Georgia, South Carolina; Mississippi, and northern Florida.

Jan 1, 1938

By Mack JF

A metallurgical evaluation of the Argyle Kimberlite No. 1 (AKl) orebody has been carried out using a 40 tonnes per hour pilot plant. Pilot plant results, together with laboratory investigations provided a basis for the determination of process design criteria for a major treatment plant. Various sampling programmes were carried out to determine specific aspects in the evaluation of diamond liberation. Treatment parameters such as the top crush size, recrush size and the bottom cut-off size were investigated. A description of the pilot plant is given together with the design criteria for a 3 million tonne per annum treatment plant

Jan 1, 1984

By Richard H. T. Garnett

On the western continental shelf of southern Africa diamonds derived from onshore sources are concentrated in gravels within drowned terrestrial and marine geomorphological features. First discovered offshore in the late 1950's, they are known to exist for about 1400 km along the Namibian and South African coast in water depths of up to 200 m. Shallow water operations yielded 1.5 Mct during the decade ending 1970 when De Beers acquired a majority of the offshore leases and commenced a 15 year exploration program. Deeper water and the prevailing sea conditions required the development of new, unique, marine mining equipment. The first large scale mining production of 22 kct was achieved by De Beers in 1989 with an output of 22 kct. Annual marine output from all sources now exceeds 0.6 Mct worth some US.$ 170 M, and rising to 0.7-0.8 Mct for 1998. Exploration has also been taking place in the waters of West Africa, Australia, and Indonesia. Exploration is by means of high resolution geophysics followed by sampling using a variety of drills, airlifts and seabed-mounted machines which imitate the eventual mining process. Surveying is now probing depths of 500 m. Smaller scale mining by several junior companies has traditionally involved diving and airlift systems. Increasingly, however, they are adopting different and improved techniques which to date have been employed only by De Beers Marine as contractor to Namdeb. Marine exploration and unit mining costs are high compared to those applicable to alluvial deposits onshore. By 1996 the total development costs incurred by De Beers alone reportedly had exceeded US.$ 500 M. However, the financial rewards are commen-surate because of the higher recovered grades. The operations are not without risk for the total of nearly 10 large and medium scale production vessels deployed by various organisations. The greatest shortfalls are experienced in the estimated throughput, not the grade.

Jan 1, 1998

By Clinton C. Knox

During the latter part of 1943, a magnetometric survey was conducted by the Geophysical Division of the Bureau of Mines at the Mahopac magnetite mine, Putnam County, N. Y., in an endeavor to locate a segment of the ore body that reports state was cut off by a fault. The Mahopac mine was abandoned about 1895 after producing nearly 1,000,000 tons of high-grade magnetite iron ore. The results of the investigation indicate the termination of the ore body by "pinching", or a displacement of such magnitude as to place any extension at a considerable distance from the original are body. This report describes the work done on the project and presents the data obtained.

Jan 1, 1948

By K. V. Raj, M. McNinch, A. L. Miller, J. Wilson, D. A. Parks

Diesel particulate matter (DPM) has been shown to contribute to various adverse health effects on underground miners. In order to reduce worker exposures, it is critical to measure the levels of DPM in the active work settings and in real time. However, the 5040 method, developed by the National Institute for Occupational Safety and Health (NIOSH), which quantifies the mass of elemental carbon (EC) and organic carbon (OC) in the samples, is based on full-shift samples, and can take days or even weeks for results. Use of real-time monitors could significantly reduce the risk of DPM exposure to miners, and available monitors have been shown to quantify DPM, mainly using the EC and other surrogates. The NIOSH Spokane Mining Research Division is conducting research aimed at developing a field-portable DPM monitor that is able to quantify both EC and OC. One part of this work is to understand the effect of OC on currently available real-time DPM monitors. This paper will present the results from experiments that were designed to observe the effect of OC on real time monitors and compare the results with the NIOSH 5040 method. BACKGROUND Diesel engines remain ubiquitous in mining environments and, as such, are one of the primary sources of ultrafine- and nano-aerosols in the mining environment in the form of diesel particulate matter (DPM). As a result of their small size, DPM has longer residence times in a mine atmosphere than the typically larger mechanically generated dust [1]. In addition, the small size enables DPM to penetrate deeply into the respiratory system [1], and exposure to DPM has been shown to contribute to various adverse health outcomes of the pulmonary system [2] and cardiovascular system [3]. Diesel exhaust has been classified as a carcinogen to humans (Group 1) by the International Agency for Research on Cancer [4]. Due to mounting concern about adverse health outcomes, extensive efforts are being made to reduce workers’ exposure to DPM, with the high exposures that often occur in underground mines being of particular concern [5]. New technologies are being developed to reduce engine emissions and thereby airborne DPM levels, including advanced engine technologies, a host of aftertreatment systems [6], and a variety of alternative fuels. The latter have been shown to markedly decrease respirable DPM and other hazardous components of diesel emissions from mining vehicles [7], yet high exposures persist, especially in underground mines [8]. In order to reduce worker exposures, it is critical to measure the levels of DPM in active work settings. In mine settings, exposures to DPM are measured by collecting air samples on filters (usually full-shift samples) and sending them to a lab for analysis using the Mine Safety and Health Administration (MSHA) P13 method, a variant of the National Institute for Occupational Safety and Health (NIOSH) 5040 method. The 5040 method is designed to quantify the mass of elemental carbon (EC) and organic carbon (OC) on the samples [9], which is then used to determine the concentration in the sampled air. MSHA has set a limit for exposure, which is based on the concentration of total carbon (TC), or the sum of both EC and OC. The allowable TC in the air (time-weighted average over the shift) is 160 μg/m3 of TC [10].

Jan 1, 2019

By D. D. Dunlop, J. R. Welker

The growing interest in the use of CO, in crude oil recovery increases the need for data on the effect of CO, on hydrocarbon physical properties. Data are presented on the solubility of CO, in various dead oils, the swelling changes in CO2-oil solutions and the effect of CO, on dead oil viscosity. This latter property shows the most pronounced effect, with viscosity reductions up to 98 per cent of the uncarbonated viscosities. An empirical method of estimating the viscosity of carbonated oils is presented. The apparatus and procedures used are described in sufficient detail to allow others to make similar studies. INTRODUCTION The effect of dissolved carbon dioxide on the swelling and viscosity reduction of specific hydrocarbon oils has been observed and recorded by a number of investigators.'- me object of this paper is to offer a means of predicting these effects for crude oils free from natural gas, using the dead state viscosity and gravity of the crude oils. The CO, solubility and swelling of numerous crude oils were determined in a visual cell at various pressure levels. The viscosity of the oils carbonated to various pressure levels was then determined by measuring the pressure drop across a capillary tube. From these data, the physical properties were correlated empirically. The resulting correlations allow the prediction of CO, solubility, swelling and viscosity reduction if the dead state gravity and viscosity of the oils are known. SOLUBILITY AND SWELLING MEASUREMENT EQUIPMENT AND PROCEDURE A high pressure visual cell was installed in a constant temperature cabinet. A test gauge was attached at the top of the cell for pressure measurement, and a line was run through the cabinet wall to a wet test meter which was used for volumetric measurement of the gas. The first step in making a test run was to put the oil in the cell up to a level about half to two-thirds of the total volume. This required about 50 to 65 ml of oil. carbon dioxide was then bubbled up through the oil for a time during which the pressure of CO2 in the cell was kept above 800 psia. Saturation of the oil with CO2 at this pressure and ambient temperature was confirmed by slowly bleeding CO2 through a valve to the atmosphere. If the oil was completely saturated with CO2, bubbles of gas would form in the oil at the first small decrease in pressure. If the oil was under-saturated, no bubbles formed until the pressure was decreased to the saturation pressure existing in the oil. If this saturation pressure was lower than that desired, more CO2 was bubbled through the oil until the desired level was reached. After saturation at ambient temperature was completed, the cabinet temperature was adjusted to the desired level and the cell was allowed to reach temperature equilibrium. After temperature equilibrium was reached, the pressure was again decreased slightly, and the oil again checked for full CO2 saturation at the cell pressure. The pressure now had changed because of the difference in solubility of the CO, in the oil at higher temperatures and the expansion of CO2 as the temperature increased. The outlet tube from the cell was then connected to the wet test meter and the CO2 was allowed to flow slowly out of the cell and through the wet test meter at ambient temperature and pressure. The water in the wet test meter had previously been saturated with CO2 at ambient temperature and pressure by allowing CO2 to flow continuously through it lor a period of several hours. The gas flow was stopped at several pressures during the run and the cell was allowed to come to equilibrium; this made possible the measurement of solubility and swelling data at the intermediate pressures. The volume of the oil in the cell was recorded at each of the equilibrium pressures in order to obtain swelling data. DATA AND RESULTS The solubility of CO2 in the oil was calculated by the relationship V — V, where R. = solubility of CO, in crude oil, cubic feet of CO, measured at 60F and 1.0 atm/ bbl of dead state oil at the temperature under which solubility was measured, V, = volume of gas released from the cell between the saturation pressure and zero pressure, corrected to 60F and 1.0 atm, cu ft, V, = volume of CO: contained in the gas space above the oil, corrected to 60F and 1.0 atm, cu ft, and V, = volume of the dead oil in the cell in bbl at the temperature of the run. The volumetric data of Sage and Lacey' were used to calculate V., from the volume of CO2 at high pressures. The swelling factor was calculated as where V, is the volume of the C0,-

By K. N. Han

Coal is becoming increasingly important as a major energy source in the face of the world's depleting energy mainly based on oil reserves. Coal is by far the most abundant fossil fuel both in the United States and abroad. Currently, because of increased mine mechanization, the fineness of preparation plant feed is increasing. Further, in the future as the grade of coal diminishes and the sulfur content increases, upgrading preparation of coal is very important. Because the density of coal is less than that of rock materials that occur as impurities, coal is separated from such waste largely by gravity-controlled processes for particle size between 5 inches and 28 mesh. If the size of coal particles is below 28 mesh, froth flotation is a good method for upgrading fine coal. Previously, numerous investigators have tried to upgrade coal values by froth flotation(1-16) and have achieved moderate success. However, the recovery and grade of coals in the concentrates are not always satisfactory, especially when the particle size is small. In general, when the size of coal is below 100 mesh, the above methods suffer from poor grade and low recovery. The reason for this is that as size decreases, the effects of gravity diminish, and other forces take effect.

Jan 1, 1985

(Under this heading will be published notes sent to the Secretary of the Institute by members or other persons introduced by members.) No. 504.-Member, technical graduate, draft exempt, who has been employed as research chemist in cyanidation and flotation for past 5 years, desires position with large company as metallurgist or assistant mill superintendent. Especially experienced in flotation of gold and silver ores, but am also familiar with flotation of lead-zinc ores. Foreign service not desired. Will be at liberty after Dec. 1, 1918. No. 505.-Member, graduate mining engineer, with 8 years' practical experience, including experience as superintendent. Desires position as engineer, foreman or superintendent of mine. No. 509.-Petroleum geologist and engineer, member, now engaged in consulting work, with five years' experience in Mid-Continent and Eastern fields, desires position with responsible concern either in South or South American countries. No. 510.- Chemical engineer, member, technical graduate, at present officer in Chemical Warfare Service of the United States Army, desires position in the East. Several-years' experience in design, construction, and operation of chemical and metallurgical plants. Minimum salary, $200. Available Jan. 1, 1919. No. 511.-Mining engineer, 32 years old, single, desires position. Speaks Spanish and French. Recently manager of large copper mines. Experience in steam shovel and underground mining. No. 512.-Mining engineer, technical graduate, member, 32 years old, single, 10 years' experience in United States, Mexico and Central America in mining, prospecting, milling, chemical research and business experience. Speaks Spanish. Open for engagement as superintendent or assistant. Can give references as to ability and integrity.

Jan 1, 1919

By R. Maddin, C. H. Mathewson, W. R. Hibbard

According to conventional crystal mechanics, a face-centered cubic single crystal slips on the system {111} <110> which receives the highest shear stress in terms of the stated orientation of the stress axis of the crystal.l However, in brass, the slip process is never as simple as this conventional analysis indicates. There is always a second cooperating plane which contains the same slip direction as the primary plane.2&apos;3 There is still another cooperating slip system commonly called the conjugate system3 previously observed to function with considerable lag only after the rotation due to the action of the first plane had produced a symmetrical condition of equal shear stress on the two systems. In the 1943 Campbell Lecture,4 a slip mechanism was outlined in which the {111) planes moved step-wise in two adjacent <112> directions integrating into the <110> slip direction. "... the slipping process would be visualized as a succession of steps 30" to the right and left of the resultant, microscopically observed, slip direction; a first step setting up the twinned configuration in the single spacing involved and an ensuing slip returning the atoms to the untwinned configuration while completing a fully integrated <110> movement."4 This process is shown diagrammatically in Fig 1 which shows two close packed layers of atoms at a slip site. A movement of the full circles through the "valleys" into the adjacent sites would produce a twin configuration with respect to the open circles. The term, "twin fault," seems most appropriate to define this faulted condition of the lattice. Annealing twins in face-centered cubic metals may arise from preformed nucleii which could be described as twin faults produced by the first movement of this two stage slip process. However, it is clear that this kind of slip does not take place so as to embrace a measurable number of adjacent planes because careful X ray examination has not conclusively shown the existence of ponderable twin lamellae (as in the case of Neumann bands in the body-centered ferrite). In view of the imponderable character of these hypothetical twin faults, the search for evidence of their existence seems restricted to indirect methods. The most obvious experimental technique serving this purpose is the determination by micrographic and X ray methods of the alignment of the composition planes of annealing twins with the former slip planes in axially strained and recrystallized single crystals. In previous attempts of this sort not all of the annealing twins could be traced to slip planes known to be operative in the conventional sense and hence evidence appeared negative or inconclusive. With the discovery of additional slip systems2&apos;3 operating in conjunction with&apos; the major system of highest resolved shear stress in axially strained brass it became apparent that the simple technique outlined above might be re-employed with more gratifying results. If, on annealing single crystals of alpha brass previously strained in tension, twins are found with their composition planes parallel to all three operative slip planes and with no participation of the fourth plane there is good evidence in favor of the existence of twin faults and of the two stage slip process. Experimental Procedure and Resuits Single crystals of alpha brass of the 70-30 composition were grown by solidification from the molten state.

Jan 1, 1950

Laboratory and pilot plant investigations aimed at recovering gold as an electroformed sheet from gold eluate and gold leach solutions and slurries are discussed. The electrolytic test work undertaken makes use of high surface area electrodes, in particular nickel-chrome foam metal and activated carbon, as well as ultrasonic activation. The process involves initial gold electrowinning (or adsorption onto activated carbon), followed by gold electroforming, both operations being carried out in a specially designed reactor. The pilot plant work mainly involved the treatment of gold eluate solutions from a South African gold mine containing 600 mg/L Au: some tests were also carried out on gold leach solutions containing 6 mg/L Au. In the former case, very efficient gold extraction onto the foam metal cathode was achieved using electrolysis combined with ultrasonics. The polarity of the gold-loaded cathode was then reversed and, using a replacement electroforming electrolyte, gold was electroformed in situ onto a polished stainless steel mandrel, from which gold sheet was subsequently removed. In the case of leach solutions, pilot plant tests confirmed laboratory findings that in the presence of calcium (resulting from the lime used during the leaching process) no electrowinning is possible. Laboratory tests showed, however, that when sodium (in the form of NaOH or Na2CO3) is added to the leach solution, electrowinning can indeed take place. Alternatively, if a special activated carbon electrode is used (without current), the non-ionic calcium aurocyanide complex is adsorbed on to the activated carbon. When a gold-loaded activated carbon electrode is made anodic in a gold electroforming electrolyte, gold electroforming is feasible, although the efficiency is very low. The possible implications of these developments for gold recovery are further examined.

Jan 1, 2007

By Esa Peuraniemi, Hanna-Leena Heikkilä

Boliden Harjavalta has been operating a copper flash smelter since 1949. Thus, in 2019, the smelter is celebrating its 70th anniversary of flash smelting. Slag formation is an integral part of the smelting process; currently the processing of slag is based on froth flotation at the slag concentrator. It provides high recovery of copper but restricts slag utilization; the final copper slag is deposited in a landfill. The Harjavalta smelter commissioned a new landfill area for final slag in May 2018. This paper describes the full project from permitting to commissioning of a new landfill area. The permitting process is described comprehensively, including environmental impact assessment together with the environmental and dam safety monitoring. During the environmental impact assessment, observations on rare animals protected on European level were made as well as archaeological findings. Successful completion of the permitting allowed the construction phase to start on time. After six years of intensive work, the new landfill was commissioned in May 2018. The new landfill area provides a sustainable way to store the smelter’s slag for decades to come. INTRODUCTION Slag processing Sulphide based concentrates have always formed the main raw material stream at Harjavalta’s copper smelter. During the smelting and converting steps, the iron included in the concentrate is oxidized and fluxed using a silica-based sand. Originally, the slag was processed by using an electric slag cleaning furnace after which the final slag was water granulated. Froth flotation has been utilized for slag treatment since 1966. In the beginning, the slag was cooled in open ditches. In 1984, controlled slow cooling of slag was licenced to enhance copper recovery at the flotation plant. Today, the slag from both the smelting and converting is tapped/skimmed to slag ladles and treated in the following steps: Slag ladle transportation from the smelter to the cooling area by slag hauler Slow cooling of slag in the ladle to allow mineral precipitation Unloading the slag by tilting the ladle; causing the slag to crush as it drops from the ladle Feeding the crushed slag (-500 mm) by wheel loader to a slag concentrator’s feed funnel followed by a conveyor belt transporting the slag to the concentrator plant Crushing and grinding using lump and pebble mills Froth flotation of the valuables Recycling the slag concentrate back to the smelter feed after dewatering Pumping the final copper slag to the landfill area.

Jan 1, 2019

By W. William Culver

James Douglas became known in the nineteenth century for two reasons fundamental to his mining career: (1) his experimental work with copper hydrometallurgy, and (2) his grasp of the idea that the quantity of copper ore matters more than the quality. On the first reason, there is a moment for any technology that’s critically important—when it moves from an area of scientific interest to one that companies are founded on. This paper examines the original 1869 Hunt and Douglas Copper Process in the context of science and engineering as they were in the 1860 and 1870s. It was a “humid” process, developed at Quebec’s Harvey Hill Mine. Chemically, the Hunt and Douglas vat leaching process succeeded, but not the original companies using it. On Douglas’ understanding of ore quality versus quantity, luck had it that the first company to field test the Hunt and Douglas Copper Process was in Chile. The Hunt and Douglas patent made that transition from experimental model to operational plant in 1870 when the Compañia de Minas de la Invernada was organized in Valparaíso. As a joint-stock company, Invernada spread the risk of investing in the revolutionary, but untested, Hunt and Douglas Process. The company brought Douglas to Chile as a consultant in 1871. His experience with hydrometallurgy, combined with what he learned about the copper business while in Chile, set him off on a career in metal mining that both surprised Douglas and made him wealthy.

By Hiroshi Hashimoto

For the application of sodium carbonate slag treatment to product high purity copper on an industrial scale, the counter-flow batch process that reduces the amount of utilized slag is proposed from the viewpoint of the post-treatment of slag. Based on the distribution ratios of several impurities between the slag and molten copper determined previously, the effectiveness of this process is discussed using a mathematical model and confirmed experimentally. According to the model, the slag consumption to remove Sb to a required level in case of the counter-flow batch process with 3 reactors is calculated to be around 30 % of that in the case of an ordinary batch process, where pure sodium carbonate is used as a slag at every stage.

Jan 1, 2003