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By Gilbert P. Moore

The production of petroleum in Argentina during 1932 was greater than in any preceding year. The total amount produced was 13,012,-756 bbl. This figure represents a decided increase in the production of oil in Argentina and can be attributed largely to the increased activities of the Royal Dutch operations at Kilometer 27 in the Comodoro Rivadavia district and those of the Standard Oil Co. of Argentina in the Salta region. Production from the Royal Dutch wells amounted to considerably more than 10,000 bbl. per day over quite a long period during the year. Of the 13,012,756 bbl. of oil produced during 1932, the Government Agency (Y. P. F.) was responsible for some 5,578,954 bbl., while 7,433,-802 bbl. were produced by private companies. Although in 1930 the Y. P. F. produced about 55 per cent of the total production and in 1930 50 per cent of the total, in 1932 its production had fallen to about 42 1/2 per cent of the total. This decrease in the proportion produced by the Y. P. F. was not due to a decrease in its production but to the decided increase in the production of private companies, particularly those two whose names are given above. One new discovery was recorded during the year. This was a well drilled by the Y. P. F. at Cacheuta in the province of Mendoza which came in pumping 80 bbl. of 15" gravity oil from a depth of 673 meters. This well was completed at approximately the same time that the provincial government of Mendoza granted a concession to the Y. P. F. and declared the whole province as a reserve in which no concessions may be granted for a period of 10 years. Under these conditions, the federally operated Y. P. F. will have no competition in the Cacheuta area from private companies such as they have in the other three fields of the republic. The Y. P. F. wildcat west of Comodoro Rivadavia at Pampa Maria Santissima on the Rio Senguerr was completed as a dry hole during 1932. This we11 was reported as a producer in 1930 by persons returning to this country from Argentina but it appears that they had been misinformed. The site of this test has been abandoned and a new location made north of the Rio Senguerr.

Jan 1, 1933

By Nathaniel Arbiter, Brij M. Moudgil

Separating oil shale from waste material or into high- and low-grade fractions can have a major economic influence on the surface retorting process. Different shale rock components vary significantly in physical and surface chemical properties, such as specific gravity, color, and water wettability. Various techniques that use the differences in one or more of these properties have been proposed to preconcentrate oil shale before retorting. This article discusses some proposed techniques, such as gravity separation, sorting, and froth flotation. Retorting oil shale is the most common method for recovering oil from the rock. This process is accomplished either underground (in situ) or above ground (ex-situ or surface retorting). Both techniques have serious limitations, such as low recoveries in the underground method and the ability to process rich shales only in the above ground retorting method. Also, environmental problems are associated with both techniques. Some of these limitations could be overcome if feed to the retorts is of consistent grade. This could be achieved by selecting appropriate sites for in situ retorting and selective mining for surface oil shale retorting. Both of these, however, would have a serious economic impact on shale oil recovery. Obviously, little control can be exercised over preparing feed material in the in situ retorting process, but beneficiation of oil shale could definitely be advantageous in surface retorting, and should have positive influence on the overall economics of recovering oil from the shales. For example, preconcentration of mined shale could result not only in reduced costs for materials handling, but also in waste disposal costs, since waste characteristics before and after the retorting process are significantly different. Upgrading the oil shale might also allow substitution of alternate technology for retorting as well as extend the process to recover oil from marginal deposits. Also, the separated material may become a potential source for recovering other values from shale deposits. Overall, the incentives for preconcentration are to reduce costs and improve thermal efficiency of the recovery process. Oil Shale Properties Important to Preconcentration Oil shale is a carbonaceous rock assaying 5-30% organic material (kerogen) and 70-95% ash-forming minerals. Kerogen is soft, nonbrit¬tle or semi-brittle material, and generally forms a matrix between irregularly distributed mineral grains. Inorganic mineral mat¬ter-quartz, feldspar, dolomite, kaolinite, dawsonite, and nahco-

Jan 9, 1982

By Buks Kruger

"After the milling and flotation of the platinum bearing ore, the flotation concentrate is electrically smelted in 6-in-line electric arc furnaces and the resulting furnace matte treated in Peirce Smith converters to remove the iron sulphide. The iron free (less than 1% Fe) converter matte is then granulated and sent to the Base Metals- and Precious Metals Refineries for further processing.During the smelting and converting processes two distinct gas streams are produced.The Peirce Smith converter off gas contains 4% - 9% SO2 which is treated in a conventional Sulphuric Acid Plant to produce sulphuric acid (H2SO4).The electric furnace off gas contains 0.9% SO2 and was (until September 2002) vented through a 91 meter high stack to atmosphere.Off gasses which contain less than 3% SO2 is difficult to handle and quite a few processes had been developed over the years to cope with this gas. A broad spectrum of process suppliers and technology approaches is available and the different technologies can be divided into a few main approaches i.e.Scrubbing of the off gas with an alkaline solution to produce an environmentally friendly and stable waste product which can be discharged onto a tailings dam or permanent landfill site.Physical absorption of the SO2 into a chemical solution to produce high quality chemically pure process chemicals.Catalytic reaction of SO2 with O2 to produce SO3 which is absorbed in water to produce sulphuric acid of less than 90% strength.In October 2002 Impala Platinum commissioned a Sulfacid Plant at its Smelter in Rustenburg which recovers the SO2 from the electric furnace off gasses into a weak sulphuric acid solution.The Sulfacid Plant technology is widely used in the paint pigment industry. This is, however, the first application of the technology on electric furnace off gas.The furnace off gasses contains flue dust which is removed in electrostatic precipitators. The gas exiting the precipitators normally still contains between 40 mg/Nm3 and 60 mg/Nm3 dust and passes through a quench-scrubber before entering the reactors. The quench-scrubber cools the electric furnace offgas down from ± 200 °C to 60 °C and act as a final dust removal stage."

Jan 1, 2003

By L. J. O’Connor

Mining activities in the small historic mining camp of La Colorada in the state of Sonora, Mexico can be traced back to the mid eighteenth century when gold was discovered and exploited by Jesuit missionaries. Following one hundred and twenty six years of activity, the mines were shut down due to the Mexican revolution. Very little activity took place until 1989 when a company, Minera Las Cuevas, spent US $350,000.00 in mapping and sampling and diamond and reverse circulation drilling over a period of three years. During 1991 HRC Development Corporation entered into a joint venture agreement with Minerales de Sotula, S.A. de C.V. that allowed an earn in of 70% interest. Exploraciones Eldorado, S.A. de C.V. and it's parent company, Eldorado Corporation Ltd. were subsequently incorporated by HRC to hold the La Colorada project. Property evaluation ensued and a thirty thousand tonne Run of Mine test heap was leached from August of 1992 until May of 1993. Full scale commercial production commenced in late November, 1993 and the first gold/silver doré bar was poured early in January of 1994. Pre-production costs were a total of approximately four million dollars U.S. During the fall of 1994 a buyout of the venture partners' interest gave Eldorado one hundred percent ownership of the property. The first two years of mining have utilized a local contract mining fleet and have placed nearly two million tonnes of ore on a run of mine heap. Low cost ripping and dozing mining methods have been employed in conjunction with drip irrigation leaching. An initial pilot plant carbon adsorption circuit was supplemented with a Merrill Crowe recovery circuit in June of 1994 and was completely replaced with the construction of a second Merrill Crowe circuit in July of 1995. Approximately fifty thousand troy ounces of gold and over one hundred thirty thousand ounces of silver will have been poured by year end 1995. Ongoing, on site exploration continues to increase mineable reserves in the known deposits. A feasibility study is ongoing concurrent with the operations and will provide the basis for future elevated production levels. Exploraciones Eldorado continues to explore outward from the current operations and several anomalous outlying zones may play part to the long term economic growth of the company in Mexico.

Jan 1, 1996

This paper is for mine managers. It is about maintenance of mining equipment. This is an extremely broad subject so this paper is given with the caveat that it necessarily contains generalisations, not true in all cases, but broadly correct. The job of mine managers is to be good at making money from the excavation and preliminary treatment of mineral. To do this well, they need to understand the cost and value drivers in the business. This includes understanding the drivers associated with maintenance. Maintenance is often identified as a problem at mine sites. It is frequently listed as the cause of production shortfalls when reporting. The state of maintenance within mining as an industry is generally among the lower end of the range, when compared to manufacturing, petrochemicals, etc. Why is this? Experience and review of many audits and evaluations of maintenance has shown that some patterns are consistent with good performance and some consistent with poor performance. These often centre around the time horizon of the management team and are associated with knowing the sensitivity of the business to certain maintenance drivers and the mechanisms of maintenance and their results. To summarise, some common misunderstandings can be listed as statements that are obvious and extrapolations from those, which are not necessarily correct. For example: High availability of equipment is necessary for production. Incorrect extrapolation:focus on availability will produce high production. Design is often the greatest factor in determining the reliability of a mining machine. Incorrect extrapolation: the newest and biggest machines generally have the lowest cost operations. Operation has an enormous impact on machine productivity and reliability. Incorrect extrapolation: the fastest operators are generally the best from a productive point of view. Some maintenance improves the performance and reliability of machines. Incorrect extrapolation: more maintenance will give you more reliability. Some maintenance depends on disassembly and reassembly of components. Incorrect extrapolation: assembly is basically the reverse process of disassembly. Maintenance improvement in the mining industry in the last 15 years has focused largely on planning and systems improvement. Incorrect extrapolation: most of the problems in maintenance can be traced to a lack of planning. Actions based on flawed extrapolations will steer a business suboptimally. It is therefore important that the mine manager understands these issues so that the improvement effort can be directed to the greatest benefit.

Jan 1, 2006

By D B. Donato

Gold mining cyanide-bearing solutions can pose a risk, cause significant wildlife deaths and adverse public relations and result in non-compliance with regulatory conditions. The International Cyanide Management Code for the Manufacture, Transport and Use of Cyanide in the Production of Gold (the Code) was developed as a voluntary code to encourage industry-wide improvements in managing cyanide responsibly. The Code requires design and implementation of robust monitoring programs for compliance under the following standards of practice: 4.1, 4.4 and 4.9. A robust monitoring regime can be complicated as each processing and tailing system is different. The following systems are commonly used in Australia: carbon-in-leach (CIL); carbon-in-pulp (CIP); heap leach process ponds; tailing waste disposal systems; peripheral discharge, central discharge and in-pit disposal. Some terrestrial wildlife can be denied access by fencing tailing storage facilities (TSFs) or other physical barriers, therefore limiting wildlife interaction with cyanide-bearing tailing and solutions to mainly birds and bats. While wildlife monitoring programs for processing and TSFs contain similar core attributes, it is necessary to adapt monitoring to take into account the site-specific features of each system. These core attributes are discussed in this paper, with reference made to cyanide concentration and associated methodologies. The issue is considerably complicated if tailing solutions are deemed or discovered to be toxic to wildlife. Essentially cyanides are required to be measured to gain an understanding of the concentrations to which wildlife may be exposed. This usually equates to measuring the cyanide concentration of solutions entering and leaving a tailing system. To assess the risk to wildlife presented by tailing and heap leach systems it is necessary to monitor both the hazard (cyanide concentrations) and the exposure (live wildlife species, presence and interaction). These tasks need not be arduous and a system of simple frequent monitoring has been developed and implemented for Code compliance. The wildlife monitoring developed by Donato Environmental Services (DES) and used at mining operations departs from the industry norm. It documents observer effort, live wildlife presence, habitat and behaviour in a tailing system, as well as deaths and carcasses. A greater understanding of wildlife cyanide toxicosis associated with this monitoring approach will enable mining companies to adopt a proactive approach to significantly reduce, or preferably eliminate, the impact on wildlife and assist in gaining Code compliance.

Jan 1, 2007

By H. -U. Lim, C. -H. Kim

The sixth underground pumped powerhouse cavern is under construction now in Korea. For the stability analysis of the five underground powerhouses, finite element method was used. For this analysis, in-situ rock stresses were measured by overcoring method and horizontal to vertical stress ratio was calculated 1.07-1.32 at four sites. Rock mass strength and elasticity were assumed from intact rock properties. The ratio of elasticity of rock mass was 0.16-0.55 of intact rock core elasticity. In general, elasto-plasticity with Mohr-Coulomb failure criterion was applied and the input cohesion and internal friction angle were approximately 0.12-0.22, 0.6-0.87 of intact rock strength parameters, respectively. Der sechste Untergrund-gepumpt-Powerhouse-Cavern ist jetzt im Bau in Korea. Für die vergan-gene Stabilitätsanalyse des fünf unterirdischen Powerhouses in Gebrauch wurde ein begrenztess Elementsmethode benutzt. Für die Analyse wurden In-Situ Felsenspannungen von overcoring Methode gemessen und Horizontale zu senkrechtem Spannungsverhältnis waren als 1.07-1.32 an vier Stellen kalkuliert. Felsenmassenkraft und Elastizität wurden von unversehrtem Felsenei-gentümern angenommen. Das Verhältnis der Elastizität von Felsenmasse war 0.16-0.55 von un-versehrter Felsenkernelastizität. Im allgemeinen wurde Elasto-Plasticity mit Ausfallskriterium des Mohr-Coulombs verwendet. Die Eingabebindekraft und die innere Reibungswinkel waren unge-fähr 0.12-0.22, 0.6-0.87 von unversehrtem Felsenkraftsparametern, jeweilig. En Coree, la sixieme centrale souterraine de la pompe d'accumulation dans la caverne est en construction. La methode d'element fini a ete utilisee pour analyser la stabilite de cinq dernieres centrales en service. Dans cette analyse, la reaction de la roche sur place etait mesuree par la methode d'overcoring et celle du rapport horizontal-vertical etait calculee de 1.07-1.32 en quatre places. L'intensite et l'elasticite de la masse de roche ont ete estimees de la roche intacte. Quant a l'elastic-ite, sa masse de roche etait 0.16-0.55 fois que celle de la roche intacte. En general, elasto-plasticite avec le critere de defaut Mohr-Coulomb etait utilisee et sa cohesion d'energie et son angle de friction interieure etaient supposees respectivement 0.12-0.22, 0.6-0.87 fois que la parametre de l'intensite de roche intacte.

Jan 1, 2003

By Christopher Mark, Anthony T. Iannacchione

Falls of ground continue to be one of the most serious causes of injury to U.S. miners. Of the 256 fatal injuries that occurred in mining between 1996 and 1998, 59 (23%) were caused by falls of ground (Table 23.1). Falls of ground affect some sectors of the mining industry more severely than others. For instance, nearly 40% of the 98 coal mine fatalities between 1996 and 1998 were caused by falls of ground. Underground miners are at much greater risk than surface miners. Nearly half (45 out of 101) of underground mine fatalities were attributed to roof, rib, and face falls, while only 6% of the 155 surface fatalities were caused by falls of highwalls or slopes. The goal of this chapter is to provide guidance to safety professionals tasked with preventing ground fall injuries. This chapter combines an analysis of the Mine Safety and Health Administration's (MS HA) accident and injury data with a survey of industry "best practices" to safeguard miners from ground falls. Ultimately, this approach can help to form the basis of a sound, proactive ground control program for the mining industry. SOURCES OF DATA All the injury data examined in this study were derived from MSHA's fatal investigation reports and the MSHA accident database. Because falls of ground often result in serious injury, MSHA "Fatalgrams" and fatal investigation reports provide a useful snapshot of ground control issues in the mining industry. These reports are available to the public (on the M FHA Web site at www.msha.gov). Fatalgrams, are one-page summaries that are usually published within a month of an accident. They contain very basic information about the accident with a graphic and a short section on relevant best practices. Fatal investigation reports are the official accident investigation reports filed by MSHA personnel. These reports contain general information about the mine, a, description of the accident, physical factors involved in the accident, a conclusion, and enforcement actions, Enforcement actions typically identify citations and discuss violations to the Federal Mining Law. MSHA also maintains comprehensive statistical data on the mining industry's accident and injury record. The law requires that mines file a report on every reportable accident that occurs, containing information on the accident's location, severity, classification, activity, and nature of injury A short narrative is generally included as well. Accident reports can be searched by many of the above fields.

By P Stoker

The need for quality assurance and quality control (QA/QC) in the sampling and assaying of minerals is fundamental to the reliability and confidence of a mineral resource estimate, reserve estimate and possibly profitability of a mine. In spite of this, QA/QC practices in the mining industry are commonly applied poorly, or misunderstood. As New Leaders, we should aim to correct the poor practices and become fully informed on appropriate procedures. This paper uses worked examples that show the usefulness and importance of best practice QA/QC procedures. Fundamentals of reviewing the performance of the laboratory by examining the results of client inserted standard reference material (standards), blank material and duplicate samples are discussed. To do this, three examples from Australian data sets have been used. For each example, simple and efficient plotting techniques are suggested with subsequent evaluations of each plot. The first case study compares the results of internal standards from a laboratory with blind standards (where the expected value is unknown to the laboratory) submitted by the client. It was demonstrated over time that the laboratory was unintentionally analysing the internal standards in a different manner to the remainder of the sample batch. In this example, not understanding the cause of a high negative laboratory bias could have led to poor understanding of the depositÆs grade and in the extreme a proposed mining operation may have reached operational stage with a subeconomic grade. The second case study discusses the quality of blank material submitted by a mining operation to monitor cross-contamination within the laboratory. The blank material contained background levels of the element under investigation. This created uncertainty in the assay results, which could have led to the need for additional assays and drilling programs. Finally, the third case study demonstrates the use of the relative paired differences (RPD) technique to cross-validate the results from duplicates and repeat analyses from two laboratories. The three examples given show either laboratory or operational error. If left undetected, each example could have had serious implications for the mineral resource estimate and ultimately, the mining operation. These examples are not isolated occurrences; poor QA/QC practices are an industry wide problem exacerbated by rapid staff turnover and increased contractor positions on-site. New Leaders must be equipped with a thorough grounding in QA/QC best practices to improve its application and confidence in the results of sample analysis.

Jan 1, 2009

By Felix B. Shay

THE Agua Fria mine is in eastern Honduras at Lat. 140° 06' N., Long. 86° 36’ W. Danli, the seat of the local municipal government, is the nearest town. The property is approximately 44 miles east of Tegucigalpa, the terminus of a truck road from the Pacific coast. The mine lies at an altitude of 3300 ft. with surrounding elevations as high as 6000 feet. For 20 years the oxidized croppings of the Agua Fria vein had been worked intermittently when in 1927 an attempt was made to open the property. A shaft was sunk 300 ft. and considerable ore was blocked out, but before a mill could be installed financial difficulties arose and the mine reverted to the original owners. Erle P. Halliburton, of Los Angeles, purchased the property in 1934. At that time a careful consideration of all factors involved showed that air transport was the only practical method of moving the equipment and machinery necessary to commence production. The other alternative was by muleback, which required sectionalizing all equipment to a maximum load of 200 lb., and during the rainy season this type of freighting was subject to serious delays. It was decided to purchase a trimotor Ford powered by three Pratt and Whitney Wasp B motors developing 400 hp. each. There were several reasons for the choice: (1) this type of plane was available at a reasonable price, owing to its replacement by faster models on most of the airlines in the United States; (2) the ship was capable of carrying a pay load of 3000 lb. over the required distance; (3) since this airplane had been widely used there would be no difficulty in securing competent personnel or in purchasing spare parts. It was necessary to rebuild the plane for cargo service. The windows were removed and the fuselage was entirely covered. A hatch was cut in the top of the fuselage to receive heavy crane loads and pieces of freight that were too large or too long to go through the rear door. The seats were removed and the interior of the plane was lined with Duralumin sheets. A heavy-duty landing gear was installed and the floor was reinforced. The nearest site for an airport was two miles north of the mine. The necessary grading was done by a small track-type tractor with a wheeled grader, which was brought in overland. Six weeks was required for bring-

Jan 1, 1940

By John E. Lens

Driving bridge foundation piles into glacio-fluvial sands and gravels and glacial tills brought surprisingly longer than predicted driving lengths on several recent projects in Vermont. In some instances, piles drove more than 40 feet past refusal depths predicted from Standard Penetration Resistances from borings terminating in glacial till. Overall, piles developed substantially less skin friction than predicted using common capacity formulas and drove as much as 80 feet longer than predicted in the moderate to dense gravelly and cobbly soils. These piles were driven for one to five span highway bridges within narrow valleys with bedrock ranging from 80 to over 200 feet deep. Piles were for both conventional and integral abutment bridges. H-piles were used for a variety of reasons including lower unit cost, better availability, and better ability to drive through cobbly soils in upper strata. Dynamic testing was used to evaluate capacity, including skin friction and end bearing capacity. In order to develop better pile length and capacity predictions for future projects, VTrans and one of their geotechnical consultants are collaborating to develop criteria for improved pile capacity predictions in these river valleys. We are doing this by evaluating driving and dynamic testing data from four projects in comparison with the Federal Highway Administration (FHWA) DRIVEN pile design software, and methods from Myerhof (1976) and Tomlinson (1981). Our overall objective is to provide a basis for choosing appropriate pile types and reliably estimating pile lengths for projects in this terrain. Bridge foundation design projects in other mountainous terrain share similar design and construction challenges to those encountered in these Vermont projects. These include: difficult exploration drill rig access requiring smaller rigs with limited capability to drill through cobbles and boulders deep enough to reach bedrock; limited experience with displacement piles as alternatives to H-piles in this geology; and relatively greater variability in soil stratigraphy between foundations at these sites as a consequence of the relatively high energy depositional environment in mountainous terrain. This paper presents the case history summaries of the pile driving, results of the ?back analysis? capacity evaluations, examples of solutions employed on these projects to address the lower capacities achieved, and resulting pile capacity design guidelines and prediction criteria for these geologic conditions based on data from these projects.

Jan 1, 2007

By Ken Deter

I have some general comments on optimizing profitability based on working in heap leach operations in tropical environments. Many of my planned comments this morning have been stated earlier in the Conference. I get the opportunity to say only a few words, and follow up on some specific comments and concerns. In normal operations, stack the ore higher, where possible. The surface area of the stacked ore controls the total solution volume that can be sprayed on the pad at a specific time period. With higher lifts, solution will contact more ore particles as it percolate through the lift, and therefore the solution grade flowing through the base will be increased, whether it is freshly stacked or almost completely leached ore. This also allows more tonnage to be stacked per lift on the pad, decreasing conveyor movement and downtime. Increasing from 6 to 8 meter (20 to 26 feet) lifts have worked successfully, and some mines are using 10 meter (33 feet) lifts in 10,000-15,000 dmtpd plants. The same concept will work with smaller plants, but initial 6 meter lifts should be used until the start-up period has shown the leaching is working as planned. As the stacked ore lifts get higher, the solution flowing through the highest lift is diluted by solution retained in the lower ones. Eventually no real increase is noted from putting fresh ore in service, and gold production rate suffers. To address the dilution problem, an HDPE liner should be installed between each 2 to 4 lifts, depending on each site specific lift height, recovery, % slump in the lift, recovery rates, and grade treated, among others. The surface of the previous lift should be leveled and sloped to control the effluent draining from the planned lift, then the liner is placed, including berms and discharge pipes. The intermediate HDPE liner does not need to be completely sealed as would be the case with the base liner, and can accept some solution penetrating into the lower lifts. A one-mm (40-mil) thickness liner can be used. This intermediate liner allows the upper lift(s) higher-grade effluent to report to the pad discharge point. This will maintain higher gold solution grades going to the ponds. The slopes of the previous lifts should not be covered with HDPE in this concept, and precipitation falling on the lower lift slopes and benches can be directed to barren or surge ponds.

Jan 1, 1999

By R Williams, P Godwin

In a time of changing rainfall patterns and in the light of recent tailings dam failures, there is closer environmental scrutiny for proposed tailings dam construction or expansion. Therefore the elimination of the traditional tailings dam could provide both environmental and cost benefits to an operation. One method to dewater tails is using pressure filtration to achieve lower moistures than what is currently available using other dewatering technologies. Further benefits include a higher rate of filtrate recovery, reuse of waste material and the potential for dry stacking of the filtered tailings cake, thereby minimising both risk and cost. This paper chronicles the replacement of two existing belt filter presses with a single McLanahan recessed plate filter press to dewater a tailings stream at a Bacchus Marsh site. It covers the slurry testing, equipment selection, installation, commissioning and operation of the filter press to achieve the following objectives:• removing the current tailings dam and slimes cells from the existing process to minimise failure risk and reduce the cost of a dam upgrade, as well as operational and maintenance costs• free up the surface area which was previously occupied by tailings dams and slimes cells• recover additional water to be returned to the process• reduce flocculent usage• reduce truck movements and handling.The project’s owner also expects to reduce carbon emissions by over 3000 t over the lifetime of operation as well as substantially reducing operating expenditure. By recognising the carbon reduction benefits, the project owner was able to successfully achieve partial funding through the Voluntary Waste Industry Protocol.An unanticipated, but as yet unquantifiable, bonus for the project was to seek an offtake agreement for the tailings thereby allowing the tailings to be considered a valuable source of beneficiated clay suitable for the construction industry.CITATION:Godwin, P and Williams, R, 2017. Pressure filtered tailings – environmental and cost benefits, in Proceedings MetPlant 2017, pp 324–334 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 11, 2017

By H. S. Martin

The Phosphate Recovery Corporation operates three flotation plants, NOS. 1 and 2 concentrators about three miles northeast of Mulberry, Florida, and No. 3 plant at Wales, Tennessee. These plants represent the first commercial phosphate flotation operation and are also believed to be the largest nonmetallic flotation operation in the world at the present time. The details of this paper will be confined to plant No. 2, this being the newest and largest concentrator in Florida, and the flow sheet being practically the same as at plant No. 1. The No. 3 plant has been operated but a short time and has been shut down since December, 1932, on account of the decrease in phosphate rock production. The phosphate washers usually are located centrally on a known deposit so that about two hundred acres of ground can be mined economically to each washer location. The washers are moved to new areas when the distance becomes too great for economical pumping. The flotation plants are built either adjacent to a washer, so that current debris may be utilized, or on the site of a former washer where there is sufficient accumulated debris available for mining, and the flotation plants must also be moved from time to time. Most of the water used in the Florida flotation plants is recovered after passing through a circulating system where most of the fine SUSpended matter is settled out, but some make-up water is obtained from deep wells near the plant. Power is obtained from the Tampa Electric Co. through the distribution facilities of the International Agricultural Corporation. Flotation concentrate is hauled from the plant by trolley over the International Agricultural Corporation's trolley system, to the drying plant at Prairie, one mile northwest of Mulberry. Ore Treated In the pebble phosphate district of Florida, the phosphate mineral, or collophane, is usually found in horizontal beds of varying thickness

Jan 1, 1934

By Nora M. Hutt, Kevin A. Morin

There is currently some pessimism over mixing of net-acid-neutralizing (NAN) and net-acid-generating (NAG) rock to obtain indefinite near-neutral seepage from mined-rock piles. This is due in large part to the lack of well- documented, successful cases of blending and layering, and to the knowledge of unsuccessful attempts. Past work often defined a potentially successful mix in terms of the ratio of net-acid-neutralizing rock to net-acid-generating rock. For example, one case study indicated the bulk pile-wide ratio was approximately 3: 1. Since this exceeded a common ratio used in acid-base accounting, no acidic drainage was predicted. However, acidic drainage appeared within two years. A large part of the ambiguity in mixing arises from attempts to predict success based on a ratio. This paper looks in detail at simple discrete-zone mixing, where at least one truckload of NAG rock is placed in a particular location with no further rehandling or dozing. The lack of acidic drainage in this scenario is dependent on the intervening distance between the discrete NAG zones, along flowpaths occupied by NAN rock. Calculations show that the bulk mixing ratio of a mined-rock pile is not an accurate indication of whether near-neutral drainage can be maintained. For example, a pile with a bulk ratio around 300: 1 could still release some acidic drainage. Calculations also show that the minimum separation distance of the NAG zones (L) can be so great, even with typical NAG and NAN rock, that the mixed pile could safely contain only a few percent of NAG rock, which would not be sufficient for many minesites with NAG rock. The comparison to a more elaborate, computer-based statistical model shows that the simple discrete-zone approach will provide a reliable yes-no indication of whether any acidic drainage will leave a pile. However, predictions of whether combined downstream drainages will be acidic depend on numerous site-specific physical and geochemical factors within and below a pile that cannot be modelled in a general way, like the number of flowpaths and the nature of their intersections.

Jan 1, 2000

SMELTING is a thing essential to the end for which ores are sought, for without it every ore is a useless stone. This art is especially needed by those who, drawn on by hope, have mined ores in large quantities with great expense and effort. For this reason it is worth while to use one's wits and, if ordinary ways should not suffice, to consider how to find new ways of smelting the ores in order to extract the metals and to purge them of their earthiness. For this reason I shall now tell you of blast and other furnaces, and I shall also tell you, in order that you may be warned, that whoever wishes to do this well must first look at the nature and quality of the ore. You will already have had enlightenment on this from the experiments on assaying and with this as guiding star it is necessary to make ingenious structures and to provide for-the other needs as necessary, for one ore requires iron and another lead. Since I have already spoken of these in their own places on their ores, it is not necessary at present that I repeat, but here I wish to speak only of copper ore as a companion of silver and gold, and of where art is most required and where difficulties chiefly appear. For these things a machine must be built that is convenient to a good water supply in order to continue the work more easily. Its wheels should be large and easily moved, constructed in such a way that the force of the water lifts the bellows that are placed at the tuyère to direct the blast inside the furnaces that are to smelt the ores. For this, attention must first be given to the site of the machine, the quantity of water, and its fall; then to the wood for building the machine and for making charcoal; then to the stones that are used for the blast furnaces; then to the provision of good bellows that are wide, long, and full in their sides and made of cloth, for the better these are, the more they quicken the power of the charcoal

Jan 1, 1942

By [Dipl. -Ing. ] Karl Angerer

The "New Austrian Tunneling Method" is one of the most significant developments in European tunneling in the last 15 years. I hope this report will show the significance of this tunneling method, which was developed in Austria exactly 18 years ago and which is now used in many European countries. In Austria from 1948 to 1953, the Moll tunnel with a length of 11.6 km (about 7.3 miles) was constructed as a part of the Kaprun power plant. Depth of overburden varied from 1000 to 1400 m, and there were many rock bursts in calcareous-schist formations. Here for the first time it was found that a thin concrete lining would help to prevent rock bursts if it was sprayed on the rock soon after blasting. During construction of the Prutz-Imst tunnel in Austria from 1953 to 1956, the shotcrete lining was sprayed on the phyllite surface a certain distance behind the blasting face to prevent rapid rock decomposition. This method was also tried in soft-rock sections of the tunnel where steel and wood supports usually were necessary. To prevent loosening of the rock, application of the sprayed concrete was tried immediately after blasting. This method made it possible to tunnel through very soft ground. Where necessary, the shotcrete was supplemented by steel arches and mesh reinforcement. At that time (1953-1956), the machines and equipment for guniting or spraying concrete were being improved rapidly and it became possible to spray aggregates up to 25 mm in diam. Hardening accelerators brought quicker strength and better adhesion of the shotcrete-this also made it possible to use the method in wet tunnels. If the tunnels are extremely wet or flowing with water, it is necessary to seal off the wet rock or drain the water, using, for instance, the Oberhasli method. After these first successful steps with the new Austrian method, large portions of the Schwarzach power plant in Austria were designed and built. At this operation, shotcrete was used almost exclusively for tunneling work. It was on this site where, for the first time, a tunnel 120 m long with a 45-sq- m section was driven through overburden without one piece of timber

Jan 1, 1970

This paper describes DRA?s involvement with the expansions at Tati Nickel, Phoenix Mine, that have taken an initial plant design of 3.6 Mtpa (originally built by Murray & Roberts), up to 5 Mtpa, and eventually 12 Mtpa capacity. Initially, the plant could not achieve the design name plate capacity for the first few years of operation, as a consequence of the notoriously hard and abrasive nickel ore affecting primarily the operation of the crusher circuit. In order to extend the life of mine, Tati? assessed and then redefined the mining plan, choosing to increase the Phoenix mine reserves to a lower cut-off grade. In order to increase nickel units produced, the concentrator had to therefore treat additional tonnage at lower head grades. Site investigations and field tests around the comminution circuits are outlined, indicated that the milling circuit could be increased to 650 tph, from the initial design of 470 tph, for a similar feed grade and recovery. A project was initiated to debottleneck the crusher circuit and achieve the new target throughput. (Project 5M). This was followed by a feasibility study of the impact of pre-concentration by DMS, which after promising laboratory tests, was followed up with the installation of a 200 tph ?demonstration? plant. The success of this demonstration plant provided the impetus to expand the supply to the 5 Mtpa mill from a large preconcentration stage. A gyratory crusher (one of the largest and most powerful) will receive ore directly from the open pit operation, feeding into a secondary-tertiary crushing plant followed by a modular DMS preparation circuit. Four large DMS processing modules have been installed to treat the tonnage, with high expected rejection levels. Finally a review of the DMS process capability on an ostensibly disseminated micro-gabbro nickel ore is given, which shows the remarkable impact that preconcentration can have on an operation, both in terms of profitability and a successful process ?expansion?. The performance of the DMS plant has significantly extended the life of mine, allowing the mine to significantly reduce mine cut-off grades and curb its previously selective mining practice.

Jan 1, 2009

By Travis Davidsavor, Stanley: Dong Vitton

Concrete is a common construction material used throughout the world, and with increasing demand of fast-track construction, blasting and concrete casting are occurring simultaneously at many sites where rock is being excavated for concrete placement. However, there has been significant disagreement in the construction industry as to what intensity of shock vibrations green, or fresh concrete can tolerate. Peak particle velocities are often arbitrarily chosen, or are simply based on materials such as plaster and drywall that have been more thoroughly studied in relation to damage. Previous research on green concrete has used the concrete’s static strength subjected to shockwaves at different ages as the primary damage criteria. In addition, the static strength tests have been based on curing time. However, changes in temperature or humidity can dramatically alter the hydration rate, which affects strength development of different cements. To account for temperature changes and humidity the concept of concrete maturity is commonly used and allows the researcher to see where the concrete was in terms of hydration progress. In general, static testing determines the failure stress that the concrete’s macrostructure can handle, and often failure occurs in the natural materials first or along large defects in casting techniques. Dynamic testing or high strain rate testing, on the other hand, loads materials at such a high rate that larger defects do not have time to react and instead the microstructure is more highly stressed leading to higher strength and stiffness values. This is known as the material’s strain rate sensitivity. Since dynamic testing better tests the microstructural strength, it is believed that it may provide a better way to assess damage to green concrete. Testing includes both static and dynamic compression as well as static and dynamic splitting tensile strength to assess whether damaged has occurred to the concrete. No significant statistical evidence was discovered that damage to the concrete occurred at five different “green” maturities from peak resultant velocities ranging from 10.66 in/s (271mm/s) to 4.40 in/s (112 mm/s) with average intensity of 6.7 in/s (170mm/s).

Jan 1, 2003

By Michel Ruel

Most plants do not track how their systems perform. Optimization is the easiest way to benefit from your control systems. We suggest an approach in four phases. The first phase is to assess all equipment to determine the performances for instruments, valves, variable speed drives, and other equipment. The second phase is to remediate instrumentation problems, including the ones pertaining to valves, variable speed drives, analyzers, transmitters. When equipment is in order, loops are put back in automatic mode. If a loop oscillates, use standard default values for this type of loop. The third phase consists in validating the control strategies, properly configuring the systems (control system, programs and parameters, HMI, alarms), and then tuning the loops and control strategies. The fourth phase consists in analyzing the process and expected performances in order to optimize the control loops and control strategies. If performance is not sufficient, a decision tree will be presented to determine the right approach: advanced regulatory control, model predictive control, or fuzzy logic control. Once all steps have been completed, control performance monitoring software is used to sustain the results and pinpoint equipment or loops not performing as stated. Alarm management and rationalization could also be performed after auditing the alarm system, if needed. Examples from plants will be used to demonstrate how this step-by-step method is applied and produces results. INTRODUCTION Maximizing what you already have is the best investment to remain competitive. This article aims to show the benefits of a systematic approach using modern tools to improve plant performance. We at BBA suggest dividing the work in four phases, and a subsequent phase can begin when the previous one has progressed: Assess, Remediate, Stabilize, Optimize, or add advanced control. This approach has been successfully applied to several sites. Examples and illustrations from these plants demonstrate how this step-by-step method is applied and produces results. Under the BBA method, there are four questions to answer before proceeding to the next step. For each control strategy, you should ask the following questions: Assessment: Does it work? Remediation: Do you use it? Stabilization: Is it stable? Optimization: Does it perform? Do you need advanced control?

Jan 1, 2019