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By Veerendra Singh, NILAMADHABA SAHU, ARIJIT BISWAS

Manganese (Mn) ferroalloys are chiefly consumed in the steelmaking process as a deoxidizing agent or as an alloying element to produce high-quality steels. These alloys are typically produced in a submerged arc furnace using coke, fluxes and manganese ores of a specific manganese/Fe (Mn/Fe) ratio. Preferred-grade manganese ores should have Mn content greater than 46 percent and Mn/Fe content greater than 5 to produce commercial-grade, high-carbon ferromanganese [1]. Selective mining of high-grade Mn resources has created a scarcity of these resources, which is emerging as a major challenge for the manganese and steel industry. Manganese ore resources are mainly distributed in South Africa, with about 32 percent; Ukraine, with about 17 percent; Australia, with about 17 percent; and Brazil, India, China, Gabon and Mexico, with lesser amounts.

By B. R. Rakishev

Balanced development of mining works at all the benches of a working zone by use of the excavator-automobile complex is achieved by transition to the technology of the bench excavating with cross panels by the two-sub-bench scheme with the changing levels of the working platforms (is patented in Kazakhstan). In deep quarries realization of this technology will allow to reduce the current volumes of the stripping ore excavation by 10?15 % and more, to increase the opened reserves of minerals by 20?50 % to decrease volume of mining-capital works for construction of stripping excavations by 2?4 times and to provide efficient use of powerful excavator-automobile complexes.

Aug 1, 2013

By Glen C. Reed

"The Avon mica district of Latah County, Idaho, is situated on a heavily forested spur of the Thatuna Range at an average altitude of 4,000 feet. The nearest rail connection is at Avon, a small farming settlement 7 miles to the south. The main supply point is Moscow, Idaho, which is 35 miles by graveled and graded county roads from Avon.Discovery of mica-bearing pegmatites in 1885 was followed in 1888 by exploitation of several outcrop deposits. Although factual data are unavailable, past production appears to have been well in excess of $100,000. Recent production (September 191 to April 1945) amounted to 65,824.78 pounds of trimmed sheet mica and 33,294.17 pounds of punch mica having a total gross value of $434,298.The several properties of the district are variously owned by estates, individuals, and the State. When active, all wore under lease or sublease to individual operators.The deposits are in an area of pre-Cambrian schists adjacent to a small granitic outlier of the Idaho batholith. The pegmatites are nearly concordant in dip and strike to the foliation of the enclosing schists. The more persistent bodies are usually tabular, lenticular, or pegmatite. Individual dikes range in length from 10 feet or less to 300 feet.The recoverable crude mica content of the pegmatites ranges downward from a maximum of 115 pounds per ton of pegmatite mined. Material discarded in rifting and trimming operations further reduces recovery to a maximum of 5 pounds per ton -of pegmatite mined. The average recoverable sheet content of the 'recently operated .deposit is estimated at 2 pounds per ton of pegmatite -mined. Available data indicate that about 21 percent of the district's recent production of trimmed sheet mica has been ""Group A"" material suitable for exacting condenser requirements. A few of the dikes contain local concentrations of beryl.Reserves of the three known major deposits of the district are estimated at about 600 to 1,000 tons of combined ""indicated"" and ""inferred"" crude block mica. One of these deposits is inferred to contain between 150 and 450 tons of beryl.It is believed that those mica reserves are minable only under a cost-market price ratio similar to that established during 1943 and 1944. A small part of the reserves can be ""high-graded"" under the present price structure. Profitable recovery of the beryl would require a market price four to five times that paid in 1944.Preliminary examination of the Avon mica district was made by the Bureau of Mines in the summer and fall of 1942. A study of the Muscovite mine and a brief reconnaissance of the surrounding district led the Bureau to recommend that the district be explored by angle-dozer trenching."

Jul 1, 1946

By Richard Franchot

IN the inevitable conquest of the blast furnace by metallurgical science in the solution of the problem of how to make more and better iron or to burn less coke, or both, it is highly desirable first to find out why the efficiency of the furnace generally is no better than it is. The immediate need would seem to be a theory to account for the existing facts. Explanation for the quantitative results obtained in current practice has been strangely lacking. One reason for this may be found perhaps in the custom of think-ing of efficiency in terms of coke consumption. Before efficiency can be explained there should be a fairly exact knowledge of what it is. A statement of the pounds of coke used per ton of iron made may have value in indicating comparative results in smelting similar ores with similar cokes, but it tells very little of the efficiency of the operation. When the materials are different, one practice burning 1800 lb. may be less efficient than another using 2000 lb. A better measure is the pig and slag per unit of coke. What is needed is a specific description of efficiency as a' definite quantity having an exact meaning. Such description is afforded by expressing efficiencies in terms of the respective proportions of the fuel energy absorbed in and chargeable to the various actions known to be -taking place in the furnace. Concretely, the energy value of a pound of coke having 85 per cent carbon blown with air heated to 1200° F. is about 7650 lb.-cal. (13,770 B.t.u.). And the work done in smelting a pound of average iron from hematite, comprising reduction and melting with production of a half pound of free running slag from an equal weight of limestone,

Jan 9, 1926

By Charles T. Ford

Limestone neutralization of acid mine drainage has been proposed and examined as a method of treatment. However, limestones differ greatly in composition and structure. The reactivity of 14 selected limestones was studied, both by chemical and physical properties and by relative effectiveness in neutralizing synthetic and actual mine waters, and several guidelines for the choice of limestones are given. The particle size should be at least as small as 74 microns and preferably smaller. The most effective limestones are most nearly pure calcium carbonate. Stones which have a relatively low calcium con- tent but which contain calcite and have a large surface area are also effective neutralizers. Magnesite and dolomitic limestones are much less reactive. Disadvantages in using even the most reactive lime- stones for neutralization are indicated.

Jan 1, 1974

By D. M. Dunbar

LONG before Columbus discovered America the original inhabitants toiled in the copper workings of the Andean Cordillera. Their best diggings appear to have been at Chuquicamata, site of the huge present-day open pit owned by the Chile Exploration Co., subsidiary of Anaconda Copper Mining Co. Located in Antofagasta Province of northern Chile, Chuquicamata is in the Atacama desert, high on the western slope of the Andes at 9500-ft elevation. In the Chuquicamata area the Atacama desert merges into the Cordillera. The climate is extremely arid and precipitation, either in the form of rain or snow, is very infrequent. The temperature varies from 50° to 85°F during the day, and drops to 20° to 40° at night. The Antofagasta and Bolivia Railroad, which connects the Chilean Port of Antofagasta with La Paz, Bolivia, passes within approximately 6 miles of the plant. A branch from the main line to Chuquicamata serves the property for outgoing copper and incoming freight. The distance to Antofagasta by rail is 163 miles. Previous to the Inca conquest, this region was inhabited by the Chuco Indians, a small tribe, descendants of the Aymaras and Quechuas. The Incas, whose capital was located in Cuzco, Peru, gradually extended their empire to the south, reaching at one time to what is now central Chile. It is believed that they built crude primitive furnaces on the banks of the Rio Salado where on a small scale they smelted copper ore from Chuquicamata. At the time of the subsequent Spanish conquest, legend has it that Diego de Almagro, on his return to Cuzco in 1536 after his invasion of Chile, obtained from these furnaces metallic copper for horseshoes. During the period from 1560 to 1879 the high grade veins of the Chuquicamata deposit were worked superficially by the Spaniards and Bolivians. The treaty signed after the war of 1879 recognized the sovereignty of Chile over this territory and Chile, realizing the importance of the district, started to develop it in an orderly manner, although only high grade veins were worked. New Process Is Key Reports on the deposit were brought to the attention of Albert C. Burrage, a Boston lawyer and financier, in 1910. Chuquicamata was described as a huge deposit of ore that was not amenable to the usual treatment by concentrating and smelting since much of the copper occurred as oxides. He had been working on a process for the recovery of copper from sulphide ore which involved a sulphating roast, leaching of the roasted product, and pre-

Jan 1, 1952

By M. Tominaga

An industrial port complex was constructed on the west coast of Leyte Island The work was started in December 1981 and completed in May 1984. This port complex includes a 665 m long wharf and can accommodate vessels ranging from a 60,000 DWT ore carrier to a 4 000 DWT general cargo ship. It now handles materials for and products from a copper smelter with an annual capacity of 130,000 t and a fertilizer plant with an annual capacity of 900,000 t. The Wharf structures consist mainly of interlocked steel pipe piles, steel sheet piles and steel pipe Piles. Material handling equipment such as unloaders, shiploaders, etc. were installed on the wharf. Site survey, planning, engineering design and construction aspects of the port development are discussed in this paper with particular emphasis on the wharf structure.

Jan 1, 2010

By James S. Redpath

"The ways of the world be many and most of the ways be tried, you follow in the footsteps of others till you come to the great divide". These words, written by Robert Service a long time ago, come to mind when I think of our original attitude upon entering into a contract with the International Nickel Company to sink the deepest shaft in the world outside of the Union of South Africa. This shaft was to be sunk to a depth of 7,137 feet from surface, and while it was recognized by all concerned to be a somewhat unusual undertaking in our part of the world, the fact that quite a few shafts had already been sunk to this depth in South Africa probably gave us the feeling that we were following in the footsteps of others. In retrospect, because this shaft was being sunk under entirely different conditions than those existing in South Africa, we were in fact venturing into an entirely new area of endeavour without properly recognizing this fact. It has occurred to me that others considering a similar type of job could benefit from our experience, especially in an assessment of the risks involved and how they compare to those encountered in an ordinary shaft. For comparison purposes, I would define an ordinary shaft as being sunk to a depth of 5,000 ft or less, and a deep shaft as one being sunk to a depth of between 5,000 ft and 8,000 ft. The sinking of the Creighton No. 9 Shaft has been de-

Jan 1, 1997

The resources industry is unique in that the finite nature of resource deposits requires a participant to be constantly locating and developing new ones to replace those that are depleted. An effective growth strategy that enables new project developments to be brought on line in time and within budget is therefore of fundamental importance to any mineral resources company. This paper seeks to identify the role of the development project in the company's overall growth strategy and the key performance criteria that a development project must achieve in order to allow a company to meet its growth objectives.

Jan 1, 1997

By Santiago Gómez, Pablo Segarra, José A. Sanchidrián

In this paper, differences between the industry-standard Holmberg-Persson (H-P) approach and a semi-analytical vibration prediction model called full-field solution (FFS) are analized. The analysis is conducted at two levels: vibrations and rock damage.

Feb 1, 2020

The Callie Underground Mine is located in the Tanami Desert 560 km Northwest of Alice Springs in the Northern Territory and is currently operated by Newmont Australia Limited with all mining activities carried out by Henry Walker Eltin. Callie was discovered in the region in 1991 with the box cut portal commenced south of the pit in 1995 and underground ore production commenced in 1998, after completion of the profitable Callie Open Pit. Development is currently occurring at depths of over 700 m with the decline currently below the 640 Reduced Level (RL) level (surface RL of 1400) with production haulage from 600 m below the surface. The combination of auto compression, a moderate geothermal gradient of 2¦C/100 m and diesel heat transfer causes the wet bulb temperature of intake air at the bottom of the mine to be typically 5¦C higher than on the surface. Wet Bulb temperatures can be up 5¦C higher again in active areas under secondary ventilation. During the 2002/2003 wet season many activities in the lower levels of the mine were stopped due to miners experiencing adverse heat related symptoms or the temperatures exceeding the Underground Heat Management Procedure stop job criteria of 32¦C wet bulb. To improve the working environment, combat the heat related illness and improve operational efficiency, a temporary 1.8 MW refrigerated cooling plant consisting of three 600 kW modules were installed underground to provide bulk air cooling for the lower levels of the mine during the 2003/2004 wet season. Several studies have been conducted in relation to future ventilation demands at Callie and all came to the conclusion, that some mine cooling will be necessary for Callie to meet the high production and development demands. This paper will describe the ventilation network at the Callie Underground operation focusing on the installation, operation and effectiveness of the temporary underground refrigeration plant.

Jan 1, 2005

By P. Miranda, K. Le, G. Maes, R. Peters

"The Mount Milligan Mine, owned by Thompson Creek Metals Company Inc. (TCM), started operations in September of 2013. It is located between the communities of Fort St. James and Mackenzie and is the first major metal mine to open in British Columbia in 15 years. Based on reserves as of January 1, 2014, it has an estimated 22 year mine life with 2.1 billion pounds of copper and 6 million ounces of gold in reserves. Like all commissioning operations, there have been issues achieving design throughput and recovery due to flow sheet design, equipment failures, control philosophy, and reagent scheme. Efforts to improve overall mill performance include equipment modifications, reagent screening and testing, and grind size testing. Applying these changes to the mill and working closely with the mine has resulted in continuous improvement each month as Mount Milligan looks to achieve the nominal design capacity of 60,000 dry metric tonnes per day in the later part of 2015. Considerable focus is being placed into ensuring size reduction equipment is utilized effectively. Additional modifications and tests are being undertaken to steadily improve processes and alleviate restrictions in the mill. This paper will discuss the start-up and present ramp-up history of Mount Milligan, the current mill processes, and the steps taken to achieve expected production numbers.OPERATION BACKGROUNDMount Milligan is an open-pit copper and gold operation wholly owned and operated by Thompson Creek Metals. It is located approximately 155km northwest of Prince George, British Columbia, between the communities of Mackenzie and Fort St. James. The copper-gold porphyry deposit contains a Measured and Indicated Mineral Resource of 706.7 Mt at 0.18% Cu and 0.33 g/t Au, containing 2.84 billion lb. copper and 7.50 million oz. gold. Mine life is currently estimated at 22.1 years (“Feasibility Update – Mt. Milligan,” 2009). The mill is designed to process 60,000 dry metric tonnes per day and produce 400 dry metric tonnes of concentrate per day averaging 25.5% Cu and 40 ppm Au. Concentrate is delivered by truck and trailer to Mackenzie followed by rail to Vancouver for overseas shipping to various customers.Mount Milligan is considered a “zero discharge” operation meaning no process or surface water will be discharged to the environment. As well, stockpiling is very limited and is only permitted for a short period of time. This scenario presents several unique challenges. Scheduling of ore/waste production must be carefully planned as waste rock is used to build the tailings storage facility (TSF). Reagent selection is also important as persistent chemicals will have a tendency to build up in reclaim water which feeds the process water tank."

Jan 1, 2015

By Raphael França Assumpção

O aço inoxidável duplex (AID) vem sendo considerado uma excelente alternativa para aplicações que requerem alta resistência à corrosão e alta resistência mecânica. Portanto, a microestrutura e as propriedades dos aços dúplex permanecem como temas de grande interesse. Para este proposito, estudou-se o efeito da laminação a morno em um aço inoxidável duplex. O foco deste trabalho foi o AID 2205 que é um dos tipos mais comuns de AID. O objetivo deste estudo é avaliar o comportamento eletroquímico do AID 2205 laminado a morno no ambiente de íons de cloretos utilizando medidas de polarização potenciodinâmica. O aço inoxidável dúplex foi laminado a morno à 600°C para 60 e 80% de redução na espessura e submetido a testes eletroquímicos. As amostras apresentaram uma excelente resistência à corrosão em solução de NaCl 3,5%. Apesar da existência de certa similaridade entre os valores de potencial de circuito aberto e potenciais de pites obtidos para todas as amostras, o número de pites observado foi maior na amostra CF, seguida pelas amostras com 60 e 80% de redução.

Oct 30, 2017

By T J. Napier-Munn, T Perkins

The JK Drop Weight test was developed at the Julius Kruttschnitt Mineral Research Centre (JKMRC) in the early 1990s as a means of characterising ore for the purposes of AG/SAG mill and crusher modelling. It has since been used extensively worldwide in the minerals industry as a key ore characterisation test which produces parameters which are used in comminution circuit design and optimisation. The test generates three parameters which are used in the AG/SAG mill model of the JKSimMet software. The resistance of the ore to high energy (impact) breakage is quantified with the A and b parameters. The oreÆs resistance to low energy (abrasion) breakage is measured using the ta parameter. The accuracy and precision of the results of the JK drop weight test are of great importance given that the parameters are key inputs in the design of such capital intensive items as SAG mills. Plant optimisation exercises also rely on the accuracy of the parameters to assist with improving the performance of an existing circuit. A thorough analysis of the accuracy and precision of the parameters generated by the JK drop weight test has now been undertaken by JKTech Pty Ltd (JKTech). An experimental program was designed using statistical methods to plan a hierarchical set of 24 drop weight tests on homogeneous material. Each stage of the test was identified and incorporated into the experimental design so that the contribution to the overall error of each stage could be quantified, including the effect of operator. The data were analysed by nested analysis of variance. An overall analysis of the A, b and ta parameters generated from the test program has now quantified the variance of these parameters more thoroughly than ever before. Using the results for the impact parameters, an analysis has been carried out to investigate the effect of these variances on final SAG mill comminution circuit design. The results will also be used to refine the comparative testing procedures currently in use by licensees worldwide.

Jan 1, 2008

By G. H. Babcock, D. A. Knight, V. G. Medina

"Production commenced at the Golden Sunlight Mine in 1983 and at Kidston Gold Mine in 1985. These projects were both constructed under the management of the Placer Dome Inc. (Technical Services) group of Vancouver, B.C.Golden Sunlight uses a conventional rod mill/ball mill grinding circuit while Kidston uses a SAG /ball mill, and later on, SABC circuit. Since the hardness of the ore and the fineness of the primary grind in both mines are about the same, a meaningful cost comparison can be made. The comparison is made on the basis of capital and operating costs.The result of this study indicates that a significant capital cost savings can be achieved in favour of the SAG /ball mill circuit when the ore is amenable to semi-autogenous grinding. With a good maintenance program, the operating cost for the SAG /ball mill circuit could also be lower that the conventional rod mill/ball mill circuit.HISTORYGolden Sunlight MineGolden Sunlight is a 5 530 t/d gold mine located near Whitehall, a small town 60 km east of Butte, Montana. Originally optioned by PDI U.S. in 1958, the development of the property waited until 1981 when economic conditions became favorable for the project to proceed.Semiautogenous grinding tests were conducted during the final stages of the feasibility study. The hard near surface ore was not quite amenable to semi-autogenous grinding. The best result produced from a series of tests using a 1.83 m diameter mill was a semiautogenous work index of 39. This is substantially higher than the 22 Bond work index of the sample used in the testwork. A conventional SAG/ball mill circuit for this ore will be extremely inefficient, and for this reason, this circuit was rejected in favour of a conventional rod mill/ball mill circuit."

Jan 1, 1991

By A. R. Laplante

A novel methodology to estimate gold recovery by gravity is presented. The methodology makes use of a population balance model which represents gold liberation, breakage and classification behaviour and applies pre-concentration and recovery performance curves to gravity-recoverable gold to predict overall recovery. Derivation of the necessary parameters is explained and demonstrated using data from an actual gravity circuit and one projected application. The methodology makes it possible to estimate the performance of circuits with large gold circulating loads and relatively low unit recoveries (e.g. jigs), or with high recoveries on a bleed of the circulating load (e.g. Knelson concentrators).

Jan 1, 1994

By Bob de Raadt

The main problem in blasting a fly-over spanning a main road is the brief time available for blasting and the removal of the debris, in order to minimize the ensuing traffic disruption in the public highway below. The blasting plan is consequently based on an extreme "time is money" concept, in which speed and public safety are predominant factors. This paper discusses reliable blasting parameters for this type of demolition.

Jan 1, 1984

By H. L. Cheng, C. H. B. Leitch, C. T. Hood, A. J. Sinclair

"Epithermal Zn-Pb-Cu-Ag-Au quartz-carbonate-barite veins at the Silver Queen mine near Owen Lake, west-central British Columbia, are hosted by Late Cretaceous Tip Top Hill andesitic volcanic rocks. The veins form a rectilinear northwest-northeast pattern that differs slightly from regional structural trends. The main northwest-trending veins are oriented very close to the trend of regional northwest faults, but the crossing veins tend to be more east-north-easterly (070 degrees)compared to a regional north-easterly trend (040 degrees). The character of the veins suggests they are extensional features formed by dextral slip related to Paleogene oblique subduction, or intrusion of magma at depth, or both.In detailed exploration at the mine, the veins cannot be modelled as simple tabular sheet-like bodies. Instead, they are complex structures that divide and rejoin, forming multiple veins or stringers, or are found in shear zones as replacement features with quartzsericite- clay-carbonate-pyrite alteration haloes of variable width that make definition of vein margins and correlation from hole to hole ambiguous. In addition, individual veins within each of the main structures have en echelon characterboth along strike and down dip.Geological and assay data must be carefully studied to define the margins of these complex veins before any estimation of ore reserves may be made. Where multiple veins occur in a section, ore/gangue mineralogy, assays, character of wallrock alteration, and thickness of vein and mineralized hangingwall, median and footwall zones must be used to evaluate continuity of mineralized structures. Well-defined sections are projected into areas of increasing complexity. Once the vein boundaries are defined, metal concentrations, important metal ratios(Au/Ag, Pb/Zn, Cu/Zn), and true thicknesses of the vein provide the basic data for reliable ore reserve estimation."

Jan 1, 1991

By F. Ames Smith

Offshore oil and gas production operations have moved from water depths of about 50 feet in 1948 to the present record depth of 340 feet. Although mobile drilling rigs are widely used for exploratory drilling, most development drilling and production operations have been conducted above water from the decks of fixed, bottom-founded platforms. Increasing water depths have caused problems associated with design and construction of these platforms and with laying of pipelines, but not so much with other activities. After a platform is installed, drilling, production, and maintenance operations are similar to those onshore and are relatively unaffected by water depth. Events of 1968 indicate that offshore production may soon extend into much deeper water. These events include leasing of Santa Barbara Channel acreage in water depths over 1500 feet, and the successful drilling of a wildcat well in 1300 feet. Designs have been developed for platforms suitable for installation in water over 600 feet deep, and studies suggest they may be used in some areas to water depths of 800 to 1000 feet. Buoyant towers can probably support above-surface platforms in water 1000 feet and deeper. Although the cost of either of these platforms will be high, once installed they will permit relatively low-cost development drilling and production operations compared with estimates for subsea production systems. In some cases, savings in development drilling costs alone could pay out the platform investment.

Jan 1, 1969

By B. F. Westfall

Throughout the 1960's and 1970's world trade grew rapidly, reaching a peak of USS1.8 trillion in 1980. Since then the world recession of the early 1980's shrunk trade to about US$1.6 trillion. The reduced cash flow to developing economies made it difficult to service their huge international debts, Nowadays lenders are reluctant to grant Third World countries additional credit. The shortage of export financing, plus trade barriers set up by developed countries, is further inhibiting international trade. Therefore, countries are turning to new and creative alternatives to financing imports of consumer as well as of investment goods. Non-cash transactions, including different forms of countertrade, are rapidly growing alternatives to traditional trade financing patterns. In 1983 about 30% of international trade, or close to US$600 bill ion, were presumably done by non-cash transactions, up from barely 2% in 1976. By year 2000, half the world trade is likely to be conducted on basis of barter. As of mid-1984 some 88 nations used countertrade regularly and Governments of at least 15 Latin American countries had arranged or mandated commercial countertrade.

Jan 1, 1986