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By F. Flynn, P. Downey, G. Ritson

"Bell Creek Mill which has been in continuous operation since 1987 has undergone numerous operational changes in the past three years. The addition of over 2. 5 million tons of resource at over 13 git from Hoyle Pond Mine precipitated an expansion program for the Bell Creek Mill. This paper discusses the major changes that have occurred over the past two years to increase throughput and efficiencies. In addition, the refractory ore treatment options are also discussed.IntroductionThe Bell Creek Mill is located in the Timmins Goldcamp approximately 8km NE of porcupine Ontario. (Fig I.)The mill was originally constructed by Camamax in 1987 to treat the Bell Creek ore at a design throughput of 350tpd. The flowsheet (Fig 2.) consisted of two stage crushing, grinding, thickening, leach and drum filtration with gold recovery by Merrill Crowe.As ore depleted at Bell Creek, the Mill was purchased by Falconbridge to process ore from their Hoyle Pond property. Kinross purchased the Hoyle Pond property and Bell Creek Mill from Falconbridge in 1992.The successful exploration program at Hoyle Pond has led to several expansions and process design changes, with further changes planned over the next twelve months.In 1994, the mine produced 61,865 oz. of gold at an average cost of 197US$ per oz. In 1995 production was 91, 611 oz. at a cost of 187US$ per oz. Projecting the nine months ended September 30, production for 1996 will be 155,000 oz. at a cost of 165US$ per oz.In early 1997, the CIP expansion will be complete, with a target of 175,000 oz. at a cost of 176US$ per oz for 1997. The mill throughput will have increased from 400tpd to 1500tpd between 1994 and 1997.The changes to Bell Creek since 1994 have been numerous with very little of the existing equipment still in operation. The total cost of the expansion to date is 15 million C$. This paper discusses the major projects carried out during in this time. The focus will be to highlight the justification of each project, the design and engineering, some of the initial start-up problems and solutions.Figure 3 shows a plan of the existing mill facility together with the various additions that were carried out."

Jan 1, 1997

By Michael F. Kitto

The production of sulphuric acid at base metals smelters is predominantly an involuntary activity and now comprises approximately 17% of all-forms sulphur supply in world terms. The majority of elemental sulphur is also produced involuntarily in response to environmental regulations regarding emissions from hydrocarbon processing plants. The paper examines the points of similarity and difference in the sulphur and smelter acid markets, and reviews their interactions. The reasons for the merchant acid market collapse of 1993 are analyzed, and the prospects of a recurrence are considered.

Jan 1, 1996

By James Westfield

This publication covers the work of the Division of Coal Mine Inspection, Bureau of Mines, as performed under provisions of the Federal Coal Mine Safety Act. It presents background information and data since 1952 when the Act became effective and includes up-to-date statistics for the calendar year 1962. During that year, approximately 9,704 coal mines were active in the United States, and the Bureau made 12,810 regular inspections. BACKGROUND The Federal coal-mine inspection program was established to carry out the provisions in Public Law 49, 77th Congress, (H. R. 2082) approved May 7, 1941. This act has since been amended as Title I, Advisory Powers Relating to Health and Safety Conditions in Mines, under Public Law 552, 82d Congress, approved July 16, 1952, which also included Title II, Prevention of Major Disasters in Mines. Public Law 552 is known as the Federal Coal Mine Safety Act.

Jan 1, 1963

By Gibbs R

The objective of this paper is to provide a broad qualitative perspective of the needs and costs of infrastructure associated with typical mining projects. Infrastructure may be defined as anything necessary for the development and operation of a project, which is peripheral to the basic extraction, handling and processing of material. As depicted in Figure 1, infrastructure acts as the link between the technological heart of a project and the outside world.

Jan 1, 1990

By W. E. Benton

Diatomite or diatomaceous earth is a siliceous sedimentary rock principally made up of the skeletal remains of the diatom, a microscopic, single-cell aquatic plant related to algae. The unique structure, chemical stability and physical properties are the major attributes of diatomite not met by other forms of silica. The production of diatomite was down in 1982 vs. 1981, reflecting the worldwide recession. The principal markets for diatomite products continue to be basically filtration and fillers. The outlook is for small but steady growth into the 1990s.

Jan 1, 1983

By J. Sliwinski, H. Kikutani, J. Atwood, D. Larimer, J. Pyecha

"The Pogo mine start-up milling targets of 2005 are vastly different than the targets of today. Improvements have been driven by unintentional design oversight, continual circuit optimization or optimization due to changing mineralogy. The more significant process improvements to date have occurred in the gravity, flotation and leach/CIP unit operations. Details of the significant improvements are presented along with the impacts to the process.INTRODUCTION Deposit Overview Pogo mine is located along the Goodpaster River, southeast of Fairbanks, Alaska. This deposit sits in the center of the Tintina Gold Province (Figure 1), a region of similar Intrusion Related Gold deposits in Interior Alaska, the Yukon and British Columbia Canada (McCoy et al., 1997). The Pogo deposit is hosted in a rock sequence of amphibolite-grade gneisses intruded by mid-Cretaceous granitic plutons and dikes, which in turn, were cut by and host the auriferous quartz veins of the Pogo deposit.The Pogo deposit is comprised of three distinct zones of auriferous quartz veins spatially adjacent to each other. These are the Liese Zone, North Zone, and the East Deep Zone. The Liese Zone consists of three stacked, massive quartz veins, designated the L1, L2, and L3. Vein thickness averages 3.0 m, with variations from 0.3–18 m thick. The North Zone mineralization is comprised of four veins with more consistent thickness, between 1 and 1.5 m. The East Deep deposit lies northeast of the Liese Zone ore body. It is separated by a post mineralization, quartz diorite pluton. The East Deep Zone is comprised of two stacked massive quartz veins with similar orientation as the Liese Zone."

Jan 1, 2017

By E. A. Roux

As the first mine-to-metal zinc operation utilising atmospheric leaching, solvent extraction and electrowinning, Skorpion Zinc faces unique challenges and opportunities. One of these challenges is found in the solvent extraction circuit, in which di-ethylhexyl phosphoric acid (D2EHPA) is used as the organic extractant. Although D2EHPA is an excellent extractant for zinc, it suffers from the co-extraction of ferric iron, which is produced during leaching and by the corrosion of the grinding media. This situation is partially mitigated by iron precipitation in the neutralisation step prior to solvent extraction. Unfortunately, some ferric iron still reports to the solvent extraction process. This leads to a slow build up of iron in the organic phase. Chemical stripping is used to maintain the Fe concentration at an acceptable level. In this paper, the efficiency of the iron removal process is discussed.

Jan 1, 2006

By C. R. McKee

In-situ mining of hard rock deposits will only become successful as we understand the engineering parameters involved and the design limitations they impose. As recognized by Wadsworth1 the most critical area is the enhancement of permeability in a hard rock mass. Along with creating permeability, the appropriate flow and fragment size conditions must be generated to insure economic rates of mineral recovery. In this article our interest is confined to true in-situ processes in which boreholes are drilled from the ground surface or a mine drift to deeper strata. We assume that fractures and permeability enhancement will be due to explosives detonated in boreholes. Other techniques such as hydraulic fracturing do not produce the necessary size distribution to promote economic extract ion rates unless the medium is already highly fractured and interconnected. Notable exceptions where hydraulic fracturing is applicable are many oil, gas and salt deposits. For many in-situ processes, an array of explosive charges offers the opportunity of creating a distribution of many fractures not achievable by other techniques. Such a situation would be suitable, for example, to in-situ mine hard rock gold, uranium, copper or oil shale deposits and may enhance oil and gas recovery in certain situations. The purpose of our article is to review and extend the explosive permeability theory of McKee and Hanson 3,4,5 and to apply its results to derive design parameters suitable for in-situ mining.

Jan 1, 1981

By J. Kortnik

Mining has always been, and still is, an activity which is vital for the development of mankind. This is because a predominant proportion of industrial products and energy is produced from raw materials that are excavated using mining methods. The need for environmental protection and the close proximity of urban settlements and mines in Slovenia have necessitated ever stricter requirements for the excavation of raw materials and the introduction of more environmentally-friendly mining methods. This is achieved by implementing a closed eco-technological cycle of raw material excavation within the scope of mining methods. In surface mines, this means prompt reclamation of the degraded surfaces, and in underground mines the use of methods with minimized impact on the surface and on underground water. In underground mines, mining methods with backfilling installed according to the multi-barrier disposal system are the only ones that allow the possibility of implementing a closed eco-techno-logical cycle of excavating raw materials. Various industrial waste materials are also included in the backfilling material in order to improve its strength properties. Backfilling consisting of several different waste materials is produced in the form of a composite. Before backfilling consisting of waste material composites can even begin to be used in mines, an environmental impact assessment of its suitability for mines needs to be made. In order to make such an assessment, a pressure leaching cell and an open diffusion cell were made, i.e. a pressure leachate test and a sorption test were developed. The two cells and the developed tests will enable the study of the influence of backfilling on the environment if installed in an underground mine. The pressure leachate test performed with the use of the pressure leaching cell simulates leaching from a consolidated backfill during the seeping of underground water into the backfilling material under pressure. The sorption test performed in an open diffusion cell simulates the flow of eluate leached from the backfilling material through the rock/geotechnical barrier. The two tests were developed in such a manner that they are performed either in succession immediately after one another, but they can also be performed separately at a certain time interval. The paper presents an assessment of the suitability of backfilling made of different waste material composites. Keywords: environmental impact assessment, multi-barrier disposal system, waste material composites, pressure leachate test, sorption test, pressure leaching cell, open diffusion cell, eco-technological cycle, mining methods with backfilling.

Jan 1, 2003

By V. A. Nieberlein

The purpose of this investigation is to determine the reserves of columbium (niobium) minerals in Arkansas and to develop metallurgical methods for concentrating the minerals and extracting the metal. The columbium resources of central Arkansas were explored by drilling and trenching, and a number of samples were taken. On the basis of this exploration the known reserves of the metal were estimated to be over 78,000,000 pounds. The reserves could possibly be extended by additional exploration to 150,000,000 pounds.

Jan 1, 1954

By Wujian jun

In the process of wire rod production, the rolled wire products must be cooled from the red-hot state to the normal temperature state. How to cool the wire rod after rolling is one of the key links of product quality control in the whole wire rod production process. In this paper, combined with the closed-loop transformation project of controlled cooling temperature of Xuanhuan iron and steel group co., LTD., the closed-loop related controlled cooling equipment (including water cooler, valve table, etc.), nozzle layout and process layout of water tank, and closed-loop software control system of controlled cooling temperature of high-line are analyzed and designed. Closed-loop system design USES temperature of cooling water cooling and water cooling two temperature closed loop regulating circuit, through temperature feedback and feedforward to achieve the uniformity of temperature adjustment, in order to prevent occur due to temperature closed loop control function with the phenomenon of a steel before and after the temperature difference is bigger, system Settings have "absolute temperature regulation" and "relative temperature regulation" selection mode. The system has a self-learning mode, which can modify the temperature feed forward formula. In addition, this paper takes the hard wire 82B rod as the test object, and studies the process parameters and the comprehensive mechanical properties of the rod before and after the water cooling system reform. The research results show that the temperature closed loop system, can ensure the rolled piece temperature control precision of plus or minus 10℃, enhances the stability of wire rod cleaner performance. It can reduce water consumption by about 186m3/h and improve the utilization rate of cooling water.

Oct 1, 2019

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 Russell C., Fleming

FOK, many years the officials of the Magma Copper Co. mine at Superior, Ariz., have had to contend with adverse conditions underground in the form of high rock temperatures, hot water, and high relative humidity. These conditions in the mine have been such that working the mine would have been impossible were it not for the constantly expanding and improved sys- tem of ventilation to produce bearable conditions in the working places. Results that have already been achieved make an interesting study of what can be done along those lines. From the time the mine was first opened in 1910 it has been hot. An idea of the conditions that have persistently been encountered can be obtained from parts of the mine that are not being worked at present or that do not yet have sufficient ventilation. Virgin rock temperatures as determined in drill holes at the faces range from 104° to 116°*. An observation made at the face of No. 5 West footwall drift on the 2800-ft. level.

Jan 1, 1930

By P. Walford, C. Mealin, D. Good

"The geology of the Ore Fault Ni-Cu deposit has been reinterpreted in the light of recent studies of the geology of the Neoarchean Bird River Sill Complex and the occurrence of economic Ni-Cu and Zn-Cu-Ag mineralization. The Ni-Cu sulfides are hosted by the lower of two parallel NNW-trending and moderately W-dipping mafic-ultramafic intrusions related to the Bird River Sill Complex. The bod¬ies intruded a bimodal volcanic suite consisting of weakly deformed mafic flows and felsic pyroclastic rocks. The mafic volcanic rocks are geochemically related to the MORB-type rocks of the Lamprey Falls Formation, and the felsic volcanic rocks are related to arc-type rocks of the Peterson Creek Formation. The original pyroxene- and olivine-dominated mineralogy of the sills has been replaced by variable ser¬pentine-amphibole-talc-carbonate assemblages.The Ni-Cu sulfide assemblage exhibits textures and geochemical signatures typical of orthomagmatic sulfide mineralization. The sulfide minerals are associated with cumulate layers of magnetic ferrochromite that range in thickness from a few to tens of centimeters. A later event of Zn-Cu-Ag sulfide mineralization associated with quartz veining and garnet-chlorite alteration along a fault that strikes north, dips vertically, and cuts all rock types. The mineralized fault is cut by the NE-trending Peterson Creek shear zone. Where the Ni-Cu and Zn-Cu-Ag mineralized zones intersect, remnant magmatic features such as amphibolite or cumulate ferrochromite bands occur within quartz veins or local chlorite alteration. The mixed zones contain a very unusual polymetallic assemblage of Ni, Cu, Zn, Ag, and platinum group elements (PGE)."

Jan 1, 2009

By F. J. Plaul

The dominating technologies for steelmaking are the basic oxygen furnace (BOF or LD process, Linz-Donawitz) and the electric-arc furnace (EAF). The main iron input materials for both are liquid iron as hot metal, or solid iron as pig iron, DRI (direct reduced iron), HBI (hot briquetted iron) and scrap. Hot metal, pig iron (i.e., solidified hot metal) and DRI/HBI are virgin iron materials, which have to be produced from iron ore by so-called ironmaking technologies. The family of ironmaking technologies includes three process routes: blast furnace, smelting reduction and direct reduction. Driven by steadily increasing costs of raw materials in the last two decades, the sector has seen a number of new developments in ironmaking technologies, developments based on fluidized-bed technology. The main advantage of a fluidized-bed technology is that fine iron ore can be directly charged to the process without prior treatment; it does away with agglomeration and its concomitant cost, a step practised in blast-furnace, COREX® and MIDREX® processes. With Posco of South Korea, Siemens VAI Metals Technologies has successfully developed the FINEX® process, a smelting reduction based on the direct use of iron ore fines to produce hot metal. The key technology is the four-stage, bubbling fluidized-bed-reactor system, in which fine iron ore is reduced to DRI fines in a countercurrent flow with a reducing gas generated by coal gasification. Beside surveying the current state of the art, this paper discusses the technological principles of smelting-reduction and direct-reduction processes. The status of FINEX® and the outlook for further developments are described. Crucial to the successful development of the new ironmaking technologies for the direct use of fine ore was the scaling up of the fluidized-bed reactor system, which demonstrated new design features.

Jan 1, 2009

By Ling Yang, John Henderson

Self-diffusion coefficients of copper in molten copper have been measured by the capillary reservoir method in the temperature range 1140o to 1260°C. The results can be represented by the equation D = [(1.46 * 0.01) x 10-3) exp [(-9710 * 710)/RT] cm2/sec. The energy of activation agrees within the limit of experimental uncertainty with that for viscous flow. The relationship between the diffusivity and coefficient of viscosity is best described by the equation D = kT/4pnr cm2/sec. MEASUREMENT was made of the self-diffusion coefficient of copper in molten copper in the temperature range 1413º to 1533ºK, using the capillary reservoir technique and CU64 as the radioactive tracer.The procedures for filling the capillaries, making the diffusion run and evaluating Dcu from the counting results were the same as described previously.1,2 All experimental operations involving the melt were carried out in a purified argon atmosphere. A graphite crucible, enclosed in a McDanel tube, was used as a container for the melt because of its satisfactory nature as a refractory material and the low solubility of carbon in copper. Graphite was also used for the capillary material. A McDanel thermocouple sheath, to which graphite radiation shields were affixed, was used to hold the capillaries. By moving this sheath through a rubber seal at the top of the McDanel tube, the capillaries could be lowered into or raised out of the melt. A Globar furnace with a constant temperature zone (± 2°C) of 2 in. was used to heat the assembly. The temperature of the furnace was the equilibrium value corresponding to a given setting of the transformer. Temperature was measured with a Pt-10 pct Rh/Pt thermocouple and varied during a run

Jan 1, 1962

By Arun Kumar Singh

Wide variation in the nature of the coal seams and the surrounding rock mass of the Indian coalfields and very little information about the associated stress fields were observed to be the major hurdles in the straight forward application of a continuous miner (CM) based mechanised depillaring approach. Due to various techno-economic reasons, a considerably large number of coal seams have been developed and are standing on pillars in India. The development of pillars is done as per Indian coal mines regulation, which results in pillars of relatively large size. In fact, the large size of the pillars provides stiff natural support to control caving of the massive overlying strata. However, the size of these pillars is not conducive for the single pass pillar stripping technique and, therefore, there is a need for pillar splitting before pillar extraction/retreating. Splitting of the pillar may reduce the competency/stiffness of the pillar, which creates a chance of overriding during final extraction below a massive roof stratum. This paper first discusses the existing coal mining scenario in India followed by a brief presentation of geo-technical details of pillar extraction at the three sites where a CM has been introduced for depillaring/pillar retreating in the country to date. Based on past rock mechanics experiences with conventional depillaring faces, stability issues of the pillars, fenders, and stumps in and around the depillaring face are discussed. This may prove to be an important factor in the success of future CM faces. It is recognized that the application of the modern strata movement monitoring techniques may improve operation safety by providing valuable qualitative and quantitative information that can be used to improve the design of modern concepts like roof bolt based breaker line support during pillaring.

Jan 1, 2009

By Run-bo Wang, Zhe Wang, Fan-yi Meng, Jian-Liang Zhang

The influence of semicharcoal substitution for coke breeze on sintering performance was investigated by pot-grate sintering test. At a fixed fuel ratio, sinter productivity and strength decreased with the substitution increasing from 0 to 100% due to better combustibility. To enhance the utilization efficiency of 100% semicharcoal, sintering parameters (moisture and suction pressure) was adjusted. With the moisture increasing, sinter performance of 100% semicharcoal increased because of the suppression of semicharcoal incomplete combustion. As the moisture further increasing to 9.0%, sinter quality gave rise to a decrease trend. The results also showed that the increase of suction pressure was benefit for the increase of heat transfer rate, leading to the improvement of heat utilization energy. Measures to enhance magnetite oxidation during sintering are discussed based on current understanding of the fundamentals that control the sintering process.

Jan 1, 2015

By W. L. Connell, O&apos

The properties of some heavy liquids which can be used in sink-float separations are discussed with major emphasis being given to the halogenated hydrocarbons. Data are presented on density, viscosity, thermal characteristics and the effects of moisture and temperature on the corrosiveness of the liquids towards various materials of construction. The toxicity of these materials, the hazards involved in handling them, and the procedures for handling them safely are also discussed. True liquids have certain advantages over colloidal suspensions for sink-float separation of minerals which makes their use especially advantageous when close separations, fine particle sizes or water-soluble minerals are to be treated. For best separating results, the medium should have an easily controlled, uniform, stable density. Here, true liquids have the advantages of allowing closer control of density and of not requiring agitation to maintain a uniform den- sity in the equipment. Low viscosities are desired in a separating liquid so that rapid separations of fine material can be made. Many liquids have lower viscosities than the colloidal suspensions at operating gravities of 1.4 to 5.0. For economic reasons and to prevent contamination of the product, the separating medium must be easily and completely removed from the ore. Normally, true liquids must be more completely removed than the cheaper suspensions and thus cause more problems in this area. The fact that the medium should not react with the minerals, air, water or the common materials of construction limits the choice of liquids which can be used, but many are available which are sufficiently stable. A more serious objection to the use of true liquids is that many of them present a toxic hazard to operating personnel and users of the product. Most true liquids are toxic to some extent, but modem knowledge of toxicity allows us to recognize the hazards inherent in handling them and enables us to design around them. Three types of true liquids are available for this use: 1) Halogenated hydrocarbons cover the gravity range from 1.2 to 3.3. They are generally more volatile and less viscous than the other liquids.

Jan 1, 1963

By D. A. Wilson

Several promising systems for selectively oxidizing the iron in an iron-tin alloy commonly called "hardhead" were investigated in a research project conducted to devise a method for recovering tin from this troublesome alloy. A satisfactory oxidation method was developed and tested successfully on melts of 2,000 grams. Molten hardhead (near 1,080° C.) was blown with air to oxidize iron and leave tin in the reduced state. Recovery of available tin was 96 to 98 percent. Tin products contained 2 to 0.1 percent iron which was reduced to 0.06 percent by filtration at 250° C. Iron oxide formed during oxidation was absorbed by a fluid flux cover composed of 43 percent silica, 17 percent burned limestone, and 40 percent fused borax. Less successful but nevertheless promising results were obtained with mixtures of steam and hydrogen which were sufficiently selective at temperatures between 600° and 950° C. but did not result in complete oxidation of iron. Carbon dioxide-carbon monoxide mixtures also oxidized iron selectively but incompletely. Both stern and carbon dioxide systems produced sufficiently encouraging results to additional research on these systems.

Jan 1, 1961