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By B. Pesic

Manganese is about twelfth in the list of the most abundant elements in the earth's crust (Reidies, 1981). More than 150 known minerals contain manganese as an essential constituent (Mellor, 1932). In general, ores containing at least 35%Mn are classified as manganese ores, ores with 10-35%Mn are called ferroginous manganese ores, and ores containing 5-10%Mn are known as manganiferrous ores. One very important class of manganiferrous ores are manganiferrous-silver ores. They occur in the oxidized portions of the silver deposits of the Western United States. Silver is very often associated with manganese oxides which in some ore deposits are dispersed through calcite, causing the black color of the host rock, which is known as black calcite. Black calcites can occur in carbonate rocks (10-1500oz Ag/ton) and in volcanic rocks (less than 1.0oz Ag/ton).

Jan 1, 1986

"Twin Creeks is a gold mining complex that is part of Newmont’s Nevada operations. The mining complex is composed of two openpits, Mega and Vista, external ore sources and a set of existing stockpiles, all of which provide ore for an autoclave, an oxide mill and a heap leach. Mega pit provides sulfide ore for the autoclave, whereas Vista pit provides oxide ore for the oxide mill and the heap leach. Both pits operate the same mining equipment and their extraction sequence must, therefore, be optimized simultaneously. Stringent blending requirements are associated with the operation of the autoclave for the sulfide ore. Strategic blending optimization on a large scale has brought significant value to the operation by increasing synergies. This paper presents the implementation of a stochastic optimization framework for the long-term production planning at Twin Creeks that simultaneously optimizes mining, stockpiling, blending and processing decision variables. The uncertainty and variability associated with the different sources of material are incorporated in the optimization model by means of stochastic simulations. The stochastic solution generated shows substantial benefits by increasing forecasted recoverable gold, leading to increased cash flows, while reducing the risk of not achieving forecasts by increasing the probabilities of meeting production and blending targets.IntroductionMine operations consist of multiple components such as mineral deposits, stockpiles, external sources, mills and autoclaves. These components represent a mineral value chain, also known in the literature as a mining complex (Montiel and Dimitrakopoulos 2015; Goodfellow and Dimitrakopoulos 2016). The material extracted from the mines flows through the different components and is transformed into sellable products. This flow of material in a mineral value chain involves complex, nonlinear transformations that render the use of industry-standard optimizers difficult. Furthermore, the optimization of a mining complex can be formulated as a mixed-integer, nonlinear optimization program with millions of integer variables representing mining, blending, stockpiling and processing decisions, prohibiting the use of conventional nonlinear optimization approaches. Previous methods that consider the joint optimization of multiple components of the mineral value chain are available (Hoerger, Seymour and Hoffman, 1999; Chanda, 2005; Stone et al., 2005; Wooller, 2005; Whittle, 2007, 2010; Bodon et al., 2010). Although these methods add value by increasing the synergy between the different components of the mining complex, they rely on simplifying assumptions that constrain their efficacy, while ignoring a most critical aspect. The latter aspect includes the geological uncertainty and actual variability of the materials in the mineral deposits mined and are supplied to the related mineral value chain, similarly to conventional life-of-mine planning methods used for individual mines (Dimitrakopoulos, Farrelly and Godoy, 2002; Godoy and Dimitrakopoulos, 2011; Spleit and Dimitrakopoulos, 2017). Additional simplifying assumptions may include: the consideration of a limited number of components, the combination of different methods, the aggregation of variables, fixed production schedules the definition of policies prior to optimization, such as cutoff- grade definition or destination of mining blocks, the linearization of nonlinear transformation functions and/or the disregard for the uncertainty of several critical parameters (Whittle, 2007). The simultaneous optimization of multiple components of a mining complex was introduced by Newmont. Hoerger, Seymour and Hoffman (1999) implemented a mixed-integer programming formulation to optimize Newmont’s Nevada operations. The model is based on the work of Urbaez and Dagdelen (1999) and simultaneously optimizes the extraction of materials from multiple deposits and the delivery of ore to multiple plants, as well as as"

Dec 1, 2018

By Dong Hoon Han

The simultaneous sulfation by sulhr dioxide and oxygen of CaO and MgO in calcined CMA (calcium magnesium acetate) were investigated. he experiments were carried out in the temperature range from 700 to 830 "C and at the SO2 concentrations of 0.3 to 1 mol %. The oxygen concentration was kept at 5 %. Each sample was calcined completely before the temperature was lowered to the sulfation temperature. According to the ranges of partial pressure of sulhr dioxide and temperature, there exist three regions for the sulfation reaction: region I in which neither oxide is sulfated, region I1 in which only CaO is sulfated, and region 111 in which both oxides are sulfated. It was shown that the MgO in calcined CMA is sulfated at a rate comparable to that of CaO by forming a mixed sulfate phase. Thus, unlike MgO contained in other desulhrization materials, the MgO content in calcined CMA can be fully utilized for flue gas desulhrization.

Jan 1, 1999

By N. Arbiter

Fracture under low velocity free-fall and double impact, and under slow compression have been-investigated. The pattern of breakage and the size distribution of resulting fragments for sand-cement and glass spheres have been determined. Photoelasticity methods were used to simulate the stress distributions in free-fall impact in order to explain the observed patterns of breakage. Oblique fracture planes, occurring only in free-fall impact, develop along lines which coincide with the trajectories of principal compression as determined by mathematical analysis Breakage efficiencies for different modes of fracture were compared for both types of spheres. For the same specimen and loading system, static loading and low velocity dynamic loading induce geometrically similar stress fields resulting in reasonably similar fracture patterns and shapes of fragments.

Jan 1, 1968

By E Marambio, R Brändle, G Fischer, G von Rickenbach, L Burgos, R Luis Fonseca

Ground support systems must provide safe and effective designs for underground excavations under high stress conditions. These systems must be capable to resist dynamic impacts and yielding during the loading process. In this context dynamic testing of the reinforcement and retaining elements that compose the ground support system are required to study and improve the behaviour of these elements under dynamic load events. During the last years, Geobrugg has been working on the improvement of retaining products by testing them in a large-scale impact test facility located at Walenstadt, Switzerland. The test facility is composed of a double level platform with a square-shaped pyramidal trunk geometry, in the upper level housing a loading mass that drop from a height up to 5m. The loading mass is guided by one central steel pipe, and the impact occurs in the sample to be tested located at the lower level in a slab with an area of 3.6m × 3.6m where the ground support system is installed. During the last years, this innovative facility has been used to test several configurations of ground support systems. The results of these tests have enabled to improve the understating of the behaviour of ground support systems under dynamic loads. In this manuscript, the arrangement, measurement, results, and the preliminary analysis of a single impact dynamic test performed during 2021 of a ground support system with a high-tensile steel wire mesh MINAX 80-4.6 is presented.

Apr 26, 2022

By T Roberts, D Faulkner, G Watt

Traditional mechanised resin rock bolting in hard rock underground mines requires the installation of resin capsules into a pre-drilled hole prior to the rock bolt being spun into a hole to mix the resin and set the bolt in place. In poor rockmass conditions, especially at depth or in highly altered faults and shears, the condition of the holes often deteriorates as the poor ground around it unravels and falls into the hole. This results in partially or fully blocked holes that either damages the resin capsule when they are being inserted into the holes or prevent it entirely until the hole is continuously cleaned of debris. This deterioration of the hole condition, and continuous need to clean and re-drill holes, also increases the diameter of the hole which in turn affects the continuously mechanically coupled anchorage mechanism of the rock bolt to the rockmass as there is simply not enough resin in the hole to fill the annulus between the bolt and the rock. This results in a large increase to critical heading cycle time and, more importantly, results in the rock reinforcement not performing as per design specification and increases the risk of instability. To overcome the quality control issues with resin bolt installation in poor ground, a single pass drill, install, inject and tension self-drilling resin bolt application was designed and implemented for use in such conditions. This system allows the operator to inject a two part polyester resin through a self-drilled hollow bolt to mix and fully encapsulate the rock reinforcement. The self-drilling bolt allows for post injection of the hole after the bolt has been installed into the rock and gives the operator manual or automated control on the amount of resin used in the hole to ensure complete encapsulation and continuously mechanically coupled anchorage to the rock. This allows the reinforcement to work as designed and provide the Geotechnical Engineer extra quality control confidence in these conditions. The implementation of this single pass resin bolting system resulted in significant decreases to the resin bolting cycle time and rock bolt quality control failures. In this paper the design, implementation and embedment into the development cycle will be discussed and the in situ quality control checks used to validate the increase in rock bolt installation quality will be shared. CITATION:Watt, G, Roberts, T and Faulkner, D, 2018. Single pass drill, install and inject self-drilling resin bolt application in poor ground, in Proceedings The Fourth Australasian Ground Control in Mining Conference (AusRock), pp 323–343 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 30, 2018

By D D. Deacon

A brief description is given of the requirements which No. 4 Shaft has been designed to meet and of the methods by which two-stage winding of rock has been accommodated in a single shaft sunk to a depth of 7,421 ft. This is followed by a consideration of the factors which led to the decision to sink the shaft as rapidly as possible. Introduction T HE No. 4 Shaft of Hartebeest-fontein Gold Mining Company, Ltd., has been sited to serve the deep southern area of the mine, some 1,500 claims in extent, between the elevations of 6,250 and 7,300 ft. below surface. From this area, some two-thirds of the mine production will ultimately be obtained. Above the 6,-250-ft. elevation, an existing shaft hoisting system, situated adjacent to the reduction plant, is of adequate capacity to cater for all ore hoisted to surface for milling. No. 4 Shaft will therefore be required to provide downcast ventilation of 1,000,000 c.f.m., together with surface hoisting facilities for men, material and development rock and underground hoisting

Jan 1, 1965

By E. C. Keeley

THE property of Buffalo Ankerite Gold Mines, Limited, is in Delora township, Porcupine area, Ontario. The southwestern part of the property is ground formerly owned by Marbuan Gold Mines, Limited, which was acquired by Buffalo Ankerite in January, 1936. Until 1938, the Ankerite (or North) mine and the Marbuan (or South) mine continued as separate operations, each developed through a surface shaft and winze to the 1050 level and each having its own mill. This meant high hoisting costs, which, in 1936 and 1937, amounted to 36 cents per ton. In 1938, the underground workings of the two mines were connected on the 1050 level, and ore from both mines was hoisted at the main, or No. 2, shaft of the North mine, location of which is shown in Figure l. This shaft was discontinued at a depth of 1150 feet because, owing to the soft nature of the talc (serpentine) rock in which it was sunk, it had to be concreted throughout, and also because of its great distance--1,100 feet and 1,400 feet, respectively-from the North and South orebodies.

Jan 1, 1945

By The Staff

"A general description of the equipment, personnel and work sequence •being used in the sinking of a circular, concrete lined, 14 foot diameter shaft which at present is at a depth of 2,484 feet. Equipping, lining and sinking of the shaft is being carried out concurrently and waste is re-moved using a mechanical mucker.INTRODUCTIONTHE BUCHANS UNIT of the American Smelting and Refining Company operates a group of mines at Buchans, Newfoundland. Buchans is in the centre of the Province (Lat. 48°47'N.; Long. 56°50'W.), five miles north of Red Indian lake at an elevation of 900 feet and is accessible by 37 miles <Jf private railway which connects with the Canadian National railway at Millertown Junction, and by 47 miles of gravel highway, which connects with the Trans-'Canada Highway at Badger."

Jan 1, 1959

By H. A. Kee

In the mining industry, as in other business, one of the more satisfactory results of sincere but cheerful effort lies in the proper completion of undertakings and problems, which, to be successful, must continually hold the interest of all concerned in the work. In most undertakings, genuine interest and satisfaction is largely dependent on the proper appraisal of opportunities presented, and the opportunity to associate ourselves with sinking operations at McIntyre central shaft, from the surface to a depth of 4,133 feet, was genuinely appreciated. This opportunity was made possible by the previous persistent and successful leadership of Mr. J. P. Bickell, President, with the assistance of his associates, and of the general manager, Mr. R.J. Ennis, in the development of the McIntyre Porcupine mine, and I beg to state that any success in this sinking operation was very largely due to the congenial co-operation and splendid workmanship of Captain Martin C. Horne, his shift bosses, leaders, and all those who worked with us . Our slogan throughout was " Safety First for Pleasant and Efficient Workmanship," and this general policy made it possible to sink and timber 2.,123 feet during the first year without a serious accident.

Jan 1, 1928

By L. Weaver

THE Tennessee Copper Co.&apos;s mines are in the southeast corner of the state of Tennessee, Polk Co., in the well-known Ducktown copper basin. Their new circular production shaft will eventually be the only ore hoisting shaft, and is located for handling ore from Boyd, Calloway, Mary, and Cherokee mines. A circular, rather than a rectangular, shaft was decided upon because: 1-The circular shaft section is stronger when lined with concrete. Reinforcing steel was not required, thus reducing lining cost by one- quarter. 2-Water sealing was desired; it was thought it could best be done with a concrete lining followed by pressure grouting to eliminate water seeps. 3-The shaft labor saving equipment could best be used in a large opening, so the round section was made for economy, sinking, mucking, concreting and water sealing. Less shaft steel is used than could have been possible with steel sets. There are no posts required; center beams and Tee plate assembly placed at 10 ft hold the 100 lb steel guide rails.

Jan 1, 1950

By L. Weaver

THE Tennessee Copper Co.&apos;s mines are in the southeast corner of the state of Tennessee. Polk Co., in the well-known Ducktown copper basin. Their new circular production shaft will eventually be the only ore hoisting shaft, and is located for handling ore from Boyd, Calloway, Mary, and Cherokee mines. A circular, rather than a rectangular, shaft was decided upon because: 1-The circular shaft section is stronger when lined with concrete. Reinforcing steel was not required, thus reducing lining cost by one- quarter. 2-Water sealing was desired; it was thought it could best be done with a concrete lining followed by pressure grouting to eliminate water seeps. 3-The shaft labor saving equipment could best be used in a large opening, so the round section was made for economy, sinking, mucking, concreting and water sealing. Less shaft steel is used than could have been possible with steel sets. There are no posts required; center beams and Tee plate assembly placed at 10 ft hold the 100 lb steel guide rails. The shaft is 12 ft 7 1/2 in. ID of the concrete walls. This is large enough for two 10-ton hoisting skips and a ladder compartment. One of the skip compartments is for hauling equipment. The grouting equipment was a 2%xlO-in. Gardner-Denver grout pump, air powered and developing 1000 lb pressure. A grout mixer, also air driven, fed the mixed solution into a tank coupled to the pump suction. Grout materials used were mostly water and Portland cement. Some Bentonite was used to seal porous ground near the surface.

Jan 11, 1950

By C. W. Davis

AN inclined shaft is usually more difficult to sink than a vertical shaft, yet it is often chosen because of the subsequent saving in cross-cuts required to reach an inclined ore body. This paper briefly discusses the factors which led to the choice of an incline and then describes the operations involved in the preparatory work and in the sinking. These operations included a programme of mining out the underground facilities and installing the structures required for a hoist-room, cable-raise, waste-pass, ore pass, two sets of skip dumps, and a collar station equipped with counterbalanced drop-rails. The shaft was sunk 1,000 feet and four stations were cut, each with loading pockets. As the sinking progressed, the rock changed from hard porphyries and dense cherts to blocky limestones and earthy rubble. These conditions called for a change in methods and it is to the red it of the shaft-sinking crew that they overcame the new difficulties as the shaft progressed, while maintaining a satisfactory monthly advance. The section mined out was nine feet six inches by seventeen feet overall. This gave room for three compartments: two skip compartments four feet three inches by five feet nine inches inside the timber, to handle two-ton skips, and a manway three feet six inches by five feet nine inches. The sinking began on December 16th, 1941, and was completed on February 15th, 1943.

Jan 1, 1945

By Westwater, J. S.

The Mather mine of The Negaunee Mine Co. embraces nearly all of Sections 1 and 2, T47N, R27W. within the limits of the cities of Negaunee and Ishpeming on the Marquette iron range of Michigan&apos;s Upper Peninsula. Operations of the Negaunee Mine Co., the stock of which is jointly owned by Bethlehem Steel Co. and The Cleveland-Cliffs Iron Co., are under the direction of the latter company. The opening of the Mather mine was described by C. W. A1len. Mather "A" shaft was completed to an initial depth of 2352 ft in early 1943. The shaft capacity of over one million tons yearly is in prospect of being realized by orderly development and mining, and sinking of the second or "B" shaft then started in accordance with previous planning. Mather " B, " also a vertical shaft, was designed as a twin to Mather "A" shaft and the sinking technique developed at Mather "A" was utilized in "B" shaft operation.

Jan 1, 1949

By Martin Dirks

Nowadays the sinter cooler often becomes the bottleneck of the sinter production, with outdated equipment leading to low productivity and high operation and maintenance costs. Furthermore valuable energy is simply destroyed in the cooler, which is not compatible anymore with a cost efficient and environmentally friendly sinter production. Paul Wurth has developed and improved two types of coolers, the annular cooler and the shaft cooler which offer flexible solutions to the specific situation of each client. The classic design of the annular cooler was completely revised and is now composed of a rigid ringframe construction as opposed to the traditional design with individual sintering cars, avoiding the typical problems of centering. Additional features of the PW annular cooler are the low air leakage rate and the increased effective cooling surface due to the removal of dead areas. This leads, in addition to the lower weight, to lower operating costs and a smaller space requirement. The shaft cooler stands out with an even smaller space requirement and a lower sinter outlet temperature because of a very long residence time of the sinter inside the shaft. Its development into a pure counter flow cooler aims at maximizing its energy efficiency, as it is designed as a system with a very low demand for cooling air, high air discharge temperatures and highly reduced dust emissions. For both cooler types flexible heat recovery solutions can be offered. These solutions include the supply of hot combustion air to the ignition hood burners and the sinter strand, the production of steam for the clients steam grid or the production of electric energy.

Aug 16, 2017

By A. G. Roach

Introduction The plant described in this paper is located at Coniston, Ontario, the smelting centre of the former Mond Nickel Company. Coniston is situated some eight miles east of Sudbury on the main line of the Canadian Pacific railway and on the Toronto-Capreol branch of the Canadian National-railways. The plant is now under control of the International Nickel Company, as a result of the merger of the two companies in 1929. The smelter proper includes four 50-in. by 240-in. blast furnaces and five 13 ft. by 30 ft. converters, and has a capacity of about 65,000 tons of ore charge per month. As accessories to the smelter there are a mill, where some 10,000 tons per month of low-grade ores are concentrated, and the sintering plant, which is the subject of this paper. The present 6-unit sintering plant replaces a former 4-unit installation. Here, fine ores, concentrates, and flue-dust are sintered on Dwight-Lloyd machines, the sinter produced being charged into the blast furnaces. At the time of writing (1931), only the concentrator is operating at capacity, due to present world conditions.

Jan 1, 1932

By Perry Harrison

THE increased use of sintering for the beneficiation of iron ores and the reclaiming of flue dust creates a lively interest in sintering costs and, economics. The character of material sintered and geographical location of plants so radically alter sintering costs and economics that it is well to understand at least the basic elements which control such changes. The following cost factors are believed to be roughly approximated for large-scale Dwight-Lloyd &apos;continuously operated sinter-machine installations located anywhere. PER CRUDE TON Labor, supplies and repairs 0.15 Power 6.5¢ X kw-hr. rate ? Plant overhead and administration 0.05 Depreciation per ton annual capacity 0.05 Fuel: 6 per cent coke breeze or equivalent minimum ? 1 per cent coke breeze additional for each 5 per cent free and combined water in ore or dust ? Ignition distillate at 0.6 gal. per ton ? PER SINTER TON Royalty on Ore 0.10 Royalty on Flue Dust 0.15 PER CRUDE TON "Winter expense"; i. z., the cost of the five months enforced idle . period for plants tied in to Great Lakes transportation season 0.03 Using the foregoing data, it is possible to roughly approximate the cost of efficient sintering at any location for. any ore, viz:

Jan 1, 1932

By J. Ostwald, L. T. Matthews, C. E. Loo

Australia has large reserves of pisolitic limonite. Mining recently commenced on the Yandicoogina deposit to produce a high-grade sintering ore. The ore contains about 10 wt% combined water and simulated sintering temperature profiles indicate that progressive dehydration of ore particles occurs in the temperature range 300-700 degrees C. Studies by optical microscopy show that although the dehyrdrated ore particles contain numerous cracks, there is no change in the microstructure of the ore because topotactic transformation to haematite results in the formation of compact masses of very fine-grained haematite microcrysts. When the ore is held at around 1 200 degrees C for 1 min significant recrystallisation of these microcrysts occurs, resulting in the formation of optically visible, micrometre sized haematite crystallites. Studies to simulate the behaviour of ore particles during sintering (as nuclear particles surrounded by a standard adhering fines mixture) show that the recrystallisation process occurs much earlier at temperatures of about 900 degrees C. Further studies indicate that CaO in the adhering fines mixture enhances the recrystallisation of the ore nuclei. The results of the investigation indicate that the pseudo-relict haematite in sinters produced from ore blends that contain significant levels of pisolitic limonite consists of dehydrated goethite that has undergone significant recrystallisation and grain growth, a phenomenon that is facilitated by the presence of lime derived from the sintering melt.

Jan 12, 1992

By T. W. Cavers

The Dwight-Lloyd sintering machine was the answer to the flue-dust problem of the copper blast-furnace smelter. Then came concentration by flotation, which threatened to eliminate the blast furnace from the smelting field altogether, and again the sintering process solved the problem. As is well known, the sintering process is dual in its function; it expels sulphur from the charge and it converts a powdery material into nodules or chunks that hold together. The charge is ignited on top just as it passes over the edge of the first suction box, and the oxygen of the air that is drawn downward through the charge breaks up the sulphide and oxidizes about fifty percent of the sulphur. The heat produced by this reaction in the upper stratum of the charge ignites that lying immediately below, and, when the pallett has travelled the full length of the machine, the tire has reached the graces and the fine concentrate has changed its characteristics by being partially fused and has become a more or less solid mass. When the Granby Company was smelting its ore as mined, one small D-L machine was adequate for the treatment of the flue dust, but as soon as the policy of smelting the concentrate was adopted, it became necessary to add two large machines to the sintering plant. The three sintering units, of which two are 42 in. by 396 in. and the other 42 in . by 264 in. in hearth area, have a combined capacity of from 370 to 450 tons (dry) of charge per day. The capacity varies directly as the hearth area. The sinter produced is about 88 per cent of the dry charge, or 325 to 400 tons per day. The sintered fines returned for resintering average between 30 and 35 per cent of the charge, but this is not as great a disadvantage as might appear, since the addition of hot fines improves the burning qualities of the mixture. From 50 to 54 per cent of the sulphur is eliminated in the sintering operation.

Jan 1, 1934

By Cheng Pan, Qiugang Yuan, Xuewei Lv, Mei Liu, Enguang Guo

Land-based nickel resources include nickel sulfide and nickel laterite. With the consumption of high grade nickel sulfide, use of nickel laterite has received more and more attention. The mineralogy and sintering behavior of limonitic nickel laterite with high iron and low nickel and silica was studied to offer technical support for producing ferronickel through sintering–blast furnace route. The mineralogy results showed that the main phases in this kind of nickel laterite are goethite (FeO(OH)), gibbsite (Al(OH)3) and NiFe2O4. The TGA (Thermogravimetric Analysis) and DSC (Differential Scanning Calorimetry) revealed hydroxide minerals in addition to absorbed water. After the sintering experiments the chemical composition, phases present (XRD analysis) and physical properties of the sinter were studied. The content of FeO in the sinter increased with increasing basicity up to 1.3, and then decreased with further increase in basicity. The yield of sinter increased (from 73% to 80%) with increase in the basicity from 1.1 to 1.7, and then the yield decreased with the further increase in basicity. Mg(Ni)(Fe,Al)2O4 is the main phase in the sinter while MgCaSiO4 and Fe2SiO4 is the main binder phase when the basicity is 1.1~1.5, and MgCaSiO4, Fe2SiO4 and SFCA is the main binder phase when the basicity is 1.7~2.1.

Jan 1, 2014