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By Jason Morrison

The Norwegian Lake Tap Method was successfully used to tap into the side of Lake Dorothy 36.5 m (120') below the lake surface. The lake tap allows water to flowinto an access tunnel from the lake lowering the water elevation below its natural out-fall. When the lake reaches a predetermined elevation below the outfall, a 1.5 m (60?) valve in the tunnel will be shut and discharge from Lake Dorothy will stop temporarily allowing construction of a dam located 2.9 km (1.8 miles) downstream at Bart Lake. The lake tap and tunnel were one phase of the $64 million Lake Dorothy Hydroelectric Project, owned and operated by Alaska Electric Light & Power Company (AEL&P), a private utility located in Juneau, AK. The Lake Dorothy portal site and tunnel are located approximately 25.7 km (16 miles) southeast of Juneau, AK, in the remote Tongass National Forest at an elevation of 701 m (2,300') above sea level in the mountains above the Taku Inlet. Thirty degree snow covered mountainsides and limited access by helicopter were just two of the many factors which challenged this unique project. The lake tap and tunnel development consisted of 250 m (820') of horizontal tun-nel, averaging 3.6 m x 3.6 m (12' × 12'), excavated using conventional drill and blast methods in highly competent Granodiorite and Tonalite rock, placement of one 3.9 m (13') long concrete plug, or bulkhead, located at the tunnel midpoint, and one 3 m (10') diameter lake tap into the side of Lake Dorothy. This paper will provide an analysis of the logistical and engineering challenges involved in the development of one of only a handful of lake taps in North America. AEL&P and contractor J. S. Redpath Corporation worked together to complete this critical project phase safely and on schedule, in a location accessible only by helicopter and plagued by constantly changing, harsh weather conditions. A strict schedule had to be maintained to ensure the timely operation of the downstream dam and power plant.

By Agne Rustan

The 9th International Symposium on Rock Fragmentation by Blasting was held in the Granada Conference Centre, Granada, Spain from the 14th - 17th September 2009. The Symposium was preceded by a 1 ½ day Workshop on “Vibrations from Blasting” organized by the initiator Dr. Alex Spathis and his colleague M. J. Noy, both employed at Orica Mining, Australia. Prof. Josè Sanchidrian at the “Universidad Politécnic de Madrid” in Madrid, Spain was the main organizer for both the Workshop and the Symposium. A record number of participants came (292) from a record number of countries (41). There were a total of 14 papers in the Vibrations from Blasting Seminar and 96 papers and 2 Keynote Addresses in the Fragblast 9 Symposium itself. The aim with this paper is to present some results from Fragblast 9 which are of interest for a large target group. One very important result from the Fragblast 9 meeting was that it was decided there that Fragblast 10 will be held in November 2012 New Delhi India.

Jan 1, 2009

By D. E. Pollock, J. A. Organiscak

Water sprays and water powered scrubbers have both been utilized in the mining and milling industry to suppress airborne dust. Unconfined water sprays operated at lower water pressures of ?689 kPa (?100 psig) can be very effective at wetting the mine product at the dust source and significantly reducing the amount of respirable dust that becomes airborne. However, unconfined waters sprays can be somewhat ineffective in actually removing airborne dust from the air. On the other hand, water powered scrubbers operating at higher water pressures of ?1,724 kPa (?250 psig) in physical enclosures or ducts have previously been demonstrated to be very effective in removing airborne dust from the air. These higher operating water pressures are uncommon in many mines and mills, so their use is limited. The National Institute for Occupational Safety and Health (NIOSH) recently investigated the performance of a lower-pressure, water powered inline series spray scrubber for removing localized airborne dust emitted at the source. Results showed noticeable improvements in airborne dust capture efficiency through the operation o f multiple inline series hollow cone spray nozzles within a round duct or pipe at the same water pressure as a single spray. Operating hollow cone spray(s) at higher water pressures noticeably improved airflow through the scrubber and yielded some additional scrubber efficiency improvements. Thus, in-line spray scrubber efficiency trade-offs were observed to be made by altering spray power components of water pressure and/or quantity (number of sprays). Results show that, on average, up to 0.23 and 0.32 m3/s (484 and 679 cu ft per min) of airflow at 0.81 and 0.69 dust capture efficiencies can be achieved with three 81° and 33°hollow cone inline series sprays, respectively, operating at 1,655 kPa (240 psig).

By Guillermo P. Selas

Perhaps this is the first time that an analysis of the industrial minerals panorama has been made in Mexico for that type of commodities. Because the non-metallic minerals lack the glamour of the metallic minerals and are usually of less unit value than metallics, a comparative study of these commodities has seldom been attempted. Sulphur, fluorite, barite and few others are an exception. Consequently, the statistics on the lowest per unit value industrial minerals, such as limestone, dolomite, gravels, sands, etc., are usually scarce and not always accurate. The following, then, will be an attempt to show a general status of the industrial minerals in Mexico with due reservations regarding volumes and values. According to Kruger (1971), this group of minerals surprisingly very little known, and called non-metallics because the lack of metallic luster, is the largest group of the mineral kingdom. It constitutes perhaps 75% of all the mineral varieties. For the most part, these minerals are compounds of common acids, sulphides, hydroxides and two elements carbon and sulphur. The list of non-metallic minerals is therefore very long. It includes the borates, carbonates, chlorides, fluorides, nitrates, phosphates, silicates, sulphates and less commonly arsenates, chromates, molybdates, sulphides, tungstates and vanadates. The latter group and many of the oxides are sought usually for their metal content.

Jan 1, 1974

By Paul Grady, Martin Froehling, Curtis Mohns, Eliab Roman

"Since the idea of processing rock to unlock the valuable minerals inside was put into practice, efforts to improve production began. Methods have been learned that allow for operation in environments and under circumstances where it would not have been possible until now. All existing operations have experienced at least some form of productivity improvement measures. The objective is simple: generate more product of better quality in less time and at lower unit cost. Since the year 2000, various modifications have been made at the Kemess Concentrator in efforts to improve productivity. The following is a discussion of these modifications and their contributions to the success of the operation.INTRODUCTIONThe Kemess South Mine, a major producer of gold and copper, is located in north-central British Columbia. Kemess was brought into production in May 1998 by Royal Oak Mines with a modern mill controlled by a fully integrated Foxboro IA DCS system and employing 134 people. On February 11, 2000 Northgate Exploration Limited (now Northgate Minerals Corporation) acquired Kemess Mine from receivership. Since acquisition in 2000, Kemess has seen many accomplishments, including the reduction of costs by 40% site wide.The mill has a nameplate capacity of approximately 56,000 tonnes per day which is processed through a Metso (Svedala) 60” x 89” Primary Gyratory Crusher and then conveyed by a 60” conveyor to a stockpile with 48 hours of total storage capacity. The nominal minus 6” rock from the stockpile is fed to two parallel grinding circuits (A & B). Each section consists of a 10.4 m F x 4.7 m SAG mill with dual drive motors (4,480 kW apiece) discharging onto a 2.4 m x 7.3 m Metso (Svedala) single deck vibrating screen to remove the coarse material for recycle to the SAG mill by a Continental 45º high angle sandwich conveyor. Screen undersize is pumped along with ball mill discharge to a cyclopac consisting of twelve Krebs 66 cm F cyclones to classify flotation feed to a size of approximately 65% minus 200 mesh. Cyclone underflow feeds a Metso (Svedala) 6.8 m F x 11.0 m ball mill, also with dual drive motors (4,480 kW apiece). The primary grinding circuit has 35,840 kW available for grinding."

Jan 1, 2005

By Levent Ozdemir, Bjorn Nilsen

For several recent TBM tunnel projects, unexpected ground conditions have been encountered, causing extra cost, considerable delay, and in some cases major claims. Undoubtedly, insufficient and/or irrelevant site investigation and testing are among the main causes of such incidents. The purpose of this paper is to review the main investigations and tests relevant for hard rock TBM projects, with particular emphasis on the considerable development that has taken place over the last few years. Also, guidelines for defining optimum scope and extent of site investigations and testing are discussed, as well as the importance of adjusting the investigations and tests according to complexity of the ground conditions and project type.

Jan 1, 1999

By S. Banisi, M. Poorkani, M. Farzaneh, M. Askari

"At the Sarcheshmeh copper mine 40,000 t of ore per day with an average grade of 0.9 % Cu is ground to 70% passing 75 microns. Grinding is effected in 8 parallel lines of 8 x 5 m ball mills in a closed circuit with cyclones. The make-up balls are 80-mm forged alloy steel balls and the average ball consumption is 750 gh of ore ground. In order to optimize ball size distribution inside the mill, based on the previous investigation instead of using only 80 mm make-up balls a combination of 80 and 60 mm (75% to 25% by weight) ball charge was used in one of the mills. To evaluate the performance of three ball types namely forged alloy steel balls (A), ductile cast iron (C), high chromium cast iron (D) and also a combined (80 and 60 mm) charge of forged alloy steel balls (B), four identical parallel mills were charged accordingly. These four mills were sampled during a period of six months. The amount of material finer than 75 microns of cyclone overflows was selected as a criterion for the performance evaluation. It was observed that in a mill with the combined make-up balls the amount of fine produced was 4% higher than the other mills. The consumption of ball types A, C and D were 730, 710 and 534 g/t, respectively. The results of laboratory flotation tests showed that due to an improvement in the production of fines an increase of 1.4% in the overall copper recovery is obtainable. Due to the promising results, the new make-up ball regime was implemented in all mills in the plant.INTRODUCTIONThe Sarcheshmeh copper ore body which may rank as third or fourth largest in the world contains 1 billion tonnes averaging 0.90 % copper and 0.03 % molybdenum. It is located southeast of Iran and has been processing 40,000 t/d of ore since 1982. The mine produces 100,000 tonnes of copper and 2,200 tonnes of molybednite concentrate per year. After three stages of crushing, the ore with a F80 of 12.7 mm is fed to 8 parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of the first stage of flotation is reground and the tailing constitutes the main component of the plant final tailings. A concentrate with an average grade of 31% Cu is obtained after cleaning and re-cleaning stages. Total recovery of the plant varies between 83 and 87 % depending on the operating conditions and ore type."

Jan 1, 2002

By P. C. Carrasco

Knowing the origin of the relevant and irrelevant components of the nugget effect is essential to optimize the process under study. This paper describes the causes of the nugget effect and, its dependency on the sample support, sampling density, and on the QA/QC procedures. Several examples of natural and ?human? nugget effect are given. A method to estimate the magnitude of the relevant and irrelevant components is suggested. Finally, the economic consequences of misunderstanding the nugget effect are analysed.

Jan 1, 2010

On June 20, 1972, Ocean Energy was awarded Contract No. H0220055, entitled Portable Task Luminaire System Development. Initially, this was a 12 month program consisting of three phases: Phase I -- Design of the Luminaire System and Battery Chargers, Phase II -- Fabrication of Prototype Units, Phase III -- Fabrication of Production Prototype Units and Battery Chargers. In addition to developing the Portable Task Luminaire System, however, this contract was amended a number of times in the course of the program to include the development of a Low Coal Portable Task Luminaire, a Rib Mounted Luminaire System, and a Machine Mounted Luminaire System, all based upon the same high frequency fluorescent lighting and safety technologies involved in the Portable Task Luminaire. This report covers the activities of this program as amended, and includes the development and fabrication of all of the systems. This report constitutes the final fulfillment of the contract.

Jan 1, 1974

By L. Adler

Longwall caving, one of the most economical and attractive mining methods, is yet one of the most difficult and hazardous.1 This dualism is inherent in a method which manipulates the mine supports themselves to aid in excavation. The difficulty is further compounded by the lack of an analytical method to predict and control results. This paper is intended to meet this need by providing a first approximation to the mechanics involved. The analysis will be limited to materials which are linear-elastic, homogeneous and isotropic. The rock structure is assumed to consist of beds of uniform thickness with no interbed bonding.2 Consequently, the mine roof can be likened to a beam, in which the action characteristic of longwall caving is the controlled deflection of the barrier supports. This deflection causes load to be intentionally transferred from these barrier props to the working face.3 Two of the requirements of this method, shared in common with other mining methods, are the provision of (1) adequate and (2) safe working space. The third requirement is the one characteristic of longwall caving and is somewhat in conflict with the second; (3) the development of near-failure static loads at the face. Requirement (1) is handled by spacing the barrier props sufficiently far from the working face for efficient operation. Requirement (2) limits the static stresses in the structure to predetermined allowable values. This means that it is not desirable to induce a maximum load on the working face as requirement (3) may imply. Such a condition could produce collapse in the whole structure. It is the purpose of the undercut to limit the depth, and thereby control the failure of the face. Therefore, requirement (3) really means the development of an optimum, not maximum, load which, in conjunction with undercutting, produces failure at the working face, with minimum powder consumption. Stress analysis requires an exact delineation of the structure and loads. The structure will be defined as the immediate roof and its supports. Since the roof is severed at the caving line, the supports reduce to the barrier props and the working face, as shown in Fig. 1. The mechanics of failure within the working face itself are complex and beyond the scope of this study, since the face represents a typical mine invert structure.4 However, this lack can be made good by the judicious collection of field data. The primary load on the structure is its own weight or body load which can be considered uniformly distributed over its upper surface. Additional loading occurs due to interference of overbeds with the structure. Wherever this occurs, the effect can be approximated by a load uniformly distributed on the upper surface of the structure.5 When this interference is intermittent, as with a much thicker overbed, the elastic curve for the overbed is superimposed on that for the structure and the extent of interference noted, as Fig. 2 shows. However, for simplicity, this study will ignore such intermittent loads without impairing the generality of approach. Consequently, the free-body diagram is that of a uniformly loaded cantilever beam with an inter-mediate support, as shown in Fig. 3. For convenience of analysis, the diagram will be divided just to the

Jan 1, 1961

By William Seymour, L. D. Schmidt, J. W. Myers, W. T. Reid

"This investigation was conducted by the Bureau of Mines as part of a program of the Solid Fuels Administration to encourage the utilization of anthracite fines, the objective being to determine the practicability of using anthracite fines in the manufacture of packagedThe experimental loss of packaged fuel covered by this report were made at a commercial plant using the Glenn Smith process and ordinarily making packaged fuel from Pocahontas screenings. These lots included several made from straight antimacite mines and others made from various blends of anthracite fines with four different bituminous coals. In all cases an asphalt binder was used. Determinations of the physical properties and the cobustion characteristics o the various samples of packaged fuel were made by the. Bureau of Mines at the Central Experiment Station.The physical tests indicated that a packaged fuel of sufficient strength for ordinary handling can be made from straight anthracite coal or from a wide variety of blends of anthracite fixes with bitunimous coal, using asphaltic binder. However the combustion tests indicated that packaged fuel made from anthracite alone tended to crumble excessively during burning and that as much as 30 to 50 percent of a caking coal may be required In the mixture to produce a block having satisfactory combustion characteristics.The authors wish to acknowledge the. splendid cooperation of Messrs. H. W. Kelly, J. L. Mooney and C. H. Stoss of Columbus, Ohio, who were chiefly responsible for obtaining the materials and making all the arrangements necessary for the production of the packaged fuel for testing.Thanks are due to the officials of Carnegie Institute of Technology for the use of their Olsen. Testing Machine and to H. M. Cooper and W. A. Selvig of the Bureau of Mines for analyses and tests."

May 1, 1946

By Thomas C. Anderson

Tiedback walls are now recognized and accepted as a good solution for a variety of structural problems: ? Retaining Walls and Bridge Abutments in cut situations ? Landslide stabilization ? Repair/support of existing walls Acceptable techniques have been developed for anchoring in cohesive and granular soils and in any kind of rock. Methods have been developed for adequately protecting tieback tendons against corrosion and for testing every tieback to verify its long-term load-carrying capacity. However, the major problem that the industry faces today is how do we get constructable and economical tiedback walls contracted for and built with minimal claims. This paper will explore the issue of "constructability" through a look at specific problems from actual projects. From a research report from the Construction Industry Institute, Austin, Texas (1986), constructability is defined "as the optimum use of construction knowledge and experience in planning, design, procurement and field/operations to achieve overall project objectives." Maximum benefits occur when people with construction knowledge and experience become involved at the very beginning of the project." The design and installation of tiebacks has been developed primarily by specialty contractors; each contractor has evolved its own methods of performing the work, and consequently, many of the techniques are patented and proprietary. The real question is, how do we make effective use of this knowledge and experience during the design phase of the project?

Jan 1, 1991

By L C. Taylor, H U. Lesiste

This paper describes a method for estimating the density distribution in the soil cap as a function of distance from its top at the time of impact with the target. It provides an estimate of the density distribution in the soil on the center line of the soil cap at the time of target impact for a single depth of burial of the charge and two heights of target and shows that, for the heights of target and the depth of burial considered, the thickness of the soil cap first increases from the in situ value and then decreases. The distribution of soil density along the centerline of the soil cap changes significantly with distance from its original position. These results can be used for validating hydro-code computations and the soil models used in the computations.

Jan 1, 2011

The acid plant consists of the following units: 1. Glover tower. 6. Mills-Packard type chambers. 3. Reaction towers. 6. Gay-Lussac towers.The sulphur dioxide bearing gases, used in the manufacture of the acid, are obtained from the roasting of zinc concentrate. Six Herreshoff type furnaces are available, each with an effective hearth area of 520 sq. ft.The concentrate contains approximately 30% sulphur, of which about 72 is ultimately recovered as acid. The oxides of nitrogen are added to the gases in the requisite proportions partly from nitrous vitriol and from nitric acid in the Glover tower, and partly from nitrous vitriol in some of the chambers and connecting flues.The acid leaving the Glover tower is cooled by the ordinary water-cooling coils, while the cooling of the chamber gases is effected by circulating water over the outer surface of the chambers, and by spraying nitrous vitriol and weak acid into them.Flow of gases throughout the plant is regulated and controlled by means of regulus metal fans, fitted with variable speed gears.

Jan 1, 1934

By T. Ogasawara, J. P. Barbosa, A. J. Monhemius, F. T. Da Silvia

A thermodynamic analysis of the sulphation of rare-earth phosphates and the water-leaching of rare-earth sulphates was used to guide experimental work on the sulphation and leaching of concentrates from Pitinga. It was concluded that the sulphuric acid digestion of xenotime is technically feasible and could be used industrially to produce yttrium and other rare-earth oxides. Yttrium extractions of more than 90% were obtained by digestion at 503 K for 4 h at an acid:concentrate ratio of 2:1 with particle size -0.053 mm. The apparent activation energy at temperatures less than 473 K, as determined from the Arrhenius plot, was 11.0 kJ/mol, indicating that the reaction is controlled by the diffusion of the reactants or products through the sulphate layer

Jun 19, 1905

By Nicolae Dobritoiu

The literature in this field and the practical activity record that a method, respectively the technology used for an underground mining work have many components which, in order to be applied, demand that for each of the parameters that describe the environment of the mining work and that influence the chosing of that method and that technology of execution, must have values in an interval well determined. There are situations when one or many parameters of the set of parameters that describe the environment of the mining work, can modify its values, and, as a result, one or more components of the methods or technology cannot be applied anymore, making necessary a replacement of their values. The information offered by this graphical-analytical model allow the determination of optimum solutions for the process of digging of the underground mining work and thus it will result the incresing of the technical-economical efficiency of the digging process and, also the increasing of the quality of the underground mining work.

Oct 1, 2009

By I. Yildirim

Free energy of interaction between particles in water can be calculated if one knows the values of appropriate interfacial tensions, which in turn can be determined from the surface free energies of the particles involved. The latter can be determined using the Van Oss-Chaudhury-Good (OCG) equation. This method requires the measurement of contact angles of well-defined liquids on the surfaces of the particles of interest. In the present work, the contact angles were determined using heat of wetting, capillary rise, and thin-layer wicking methods. The results are compared, and the utility of the surface free energy data obtained from the contact angles are discussed.

Jan 1, 2002

By A. J. Stewart, M. Guay, J. W. Critchfield, A. P. Hoddeson

The MetroWest Water Supply Tunnel project includes three main work shafts, excavated primarily in rock. Shaft L is 9.8 m (32 ft) diameter and 124.1 m (407 ft)deep. Shaft E is 7.0 by 12.8 m (23 by 42 ft) and 91.4 m (300 ft) deep. Shaft 5A is 7.9m (26 ft) diameter and 72.5 m (238 ft) deep. Construction methods are compared and contrasted. Lessons learned involve a differing site condition in soil overburden, contractor-proposed modifications, and remedial slurry wall work. Other aspects discussed include noise and vibration control, groundwater inflow, water treatment, muck handling and contract administration issues.

Jan 1, 1999

By Sandra Lúcia de Moraes

O presente trabalho visou contribuir para o melhor entendimento do mecanismo inerente ao uso de agente coloidisante/dispersante no processo de pelotização de concentrado de minério de ferro. Avaliou-se a influência dos agentes coloidisantes em comparação aos aglomerantes Peridur e bentonita. Para este fim, foram realizados ensaios de bancada visando a identificação de possíveis ações destes agentes no grau de dispersão das misturas, bem como a sua influência na liberação/aglomeração dos finos. Os resultados destes parâmetros foram confrontados com os resultados de caracterização das pelotas confeccionadas com os diferentes agentes pelos quais foi possível identificar que os dispersantes tem forte atuação sobre o conteúdo de finos do concentrado de minério de ferro. A ação principal se dá pela remoção das partículas ultrafinas que recobrem os grãos mais grosseiros presentes na amostra e as dispersam, no caso dos dispersantes, aumentando a disponibilidade do conteúdo de finos <4 µm, principalmente os finos <1 µm (coloides). No caso do aglomerante orgânico, observa-se também essa ação, no entanto, os finos presentes são aglomerados na faixa entre 26 e 4 µm.

Aug 16, 2017

By K. Gie Tan, Frank Wright

"Misty Mountain Gold Limited (MGL - TSE VSE, MGLCF - NASDAQ) of Vancouver, owns the Harmony Gold Project located in the Queen Charlotte Islands of British Columbia. The deposit contains a mineable reserve of 2.3 million ounces of gold, at a 1.2 g/t cut-off and an average grade of 2.1 1 g/t Au.Discovered in 1970 and frequently investigated by previous owners, full-scale production has not yet occurred because of the complex metallurgy of the deposit. Recent process studies that were completed in six months at Process Research Associates Ltd. (PRA) of Vancouver resulted in significant improvements to the project viability. The work involved detailed mineral processing studies aimed at achieving a marketable concentrate. An option for bio-oxidation and cyanidation to allow on-site dor6 production was also investigated.The main findings at PRA followed the front-end recovery of finely divided free gold (Au) by gravity concentration. Approximately 20% of Au reported to less than 0.2% of the mass, resulting in a grade exceeding 500g/t Au. The resulting gravity tailing provided a uniform feed for flotation evaluation. The optimization of this step allowed for a reasonable primary grind size of 80% passing (&) 50pm. Efficient slime’s rejection was made possible by flotation at a reduced pulp density, and pre-oxidized sulphides were reactivated with Na,S. The regrinding of the bulk flotation concentrate followed by up to five stages of cleaning resulted in a combined gravity + flotation product of 50g/t Au with 82% recovery in less than 4% of the feed mass.Further studies indicated that the flotation concentrate was amenable also to tank bio-oxidation. At 65% sulphide oxidation the gold recovery improved from less than 70% to 90% by cyanide extraction using CIL (carbon in leach). This compares to less than 45% gold recovery from whole-ore cyanidation in several previous studies. Heap leaching may also be considered for lower grade material. PRA’s methodical approach and consistent results dispelled the uncertainty about predicting the material behavior for the Harmony Property."

Jan 1, 2000