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By A. Pearson, Auld. F. A., S. O. Colbeck

"To minimise mine shaft maintenance over the long-term, some potash mine shaft sinking projects have required the provision of a continuous fully welded waterproof steel liner over a substantial depth to achieve zero water ingress. This brings many challenges, including the implementation of a bottom-up construction sequence incorporating extensive waterproof welding and how to ensure the integrity of the welded liner when subject to substantial temperature fluctuations that are likely to occur with a composite steel-concrete liner placed within an artificially frozen shaft excavation. This paper highlights the key challenges and discusses multiple mitigation measures including vertical shaft movement joints and a combined structural-thermal analysis approach. INTRODUCTION Whilst the application of combined steel and concrete shaft liners using a fully welded steel liner is not new and has been used in several previous projects, most notably at Lanigan (Storck, 1968 & Roesner, 1980), recently the need to provide a continuously welded steel waterproof liner for over 800 m length of shaft for a potash mine has been considered. The need for a fully continuous welded steel liner arises from the perceived benefits of providing a fully dry shaft environment to; increase the life expectancy of the shaft; reduce routine maintenance particularly combatting salt accretion and minimising wash down practices; and improve long-term ground stability by preventing salt dissolution. Whereas previous projects have employed discontinuous sections of steel liner, when a fully continuous liner is employed longitudinal shaft strains from thermal changes result in significant deformations that must be either explicitly accommodated or resisted. Note, these thermal vertical expansion forces are more than capable of lifting the shaft collar and head frame structures should they be fixed in place. Typically, shafts are designed to resist vertical forces (collar loads and self-weight) by creating mechanical interlock with the adjacent ground. In competent rock this occurs incrementally with the construction of the shaft liner, and occasional hitch/crib structures are constructed as intermediate foundations to guarantee the transfer of vertical forces at distinct horizons. Usually this means cribs are constructed at first rock head, above the station level, and at the sump, although sometimes additional cribs are formed when entering and exiting weak strata or when a change in shaft liner construction is required."

Jan 1, 2019

The comminution circuit at Tarkwa Gold Mine was commissioned as a Single Stage SAG mill operated in closed circuit with hydrocyclones designed to process 4Mtpa. The circuit was later expanded by adding a secondary ball mill in closed circuit with hydrocyclones with a view to treat 12Mtpa. The change in the circuit configuration and assigning the low aspect ratio SAG mill for primary grinding application makes the Tarkwa experience unique in terms the optimisation strategy and operating philosophy. This paper reports the findings from a series of surveys performed on this plant. The surveys reported include data obtained when the SAG mill was operated as a single stage mill in closed circuit with hydrocyclones and also SABC mode under two different feed conditions. The differences in operating philosophy between the initial single stage SAG circuit and the current SAG/Ball mill circuit are discussed.

Jan 1, 2012

By R. E. Miller

Hoisting coal using water and air is a technology which has potential to improve the economics and safety of transporting coal. The Coal Air-Lift Hydrohoist is a prototype coal hoisting system developed by the U.S. Bureau of Mines (USBM). This paper discusses the method used to regulate the supply of water to the 50-m-high hydrohoist U-tube such that the pressure difference at the bottom of the U-tube is held constant. Maintaining a constant pressure difference results in a steady flow of slurry out of the hydrohoist.

Jan 1, 1995

By David W. Wisecarver

Kennecott Copper Corporation and the U. S. Bureau of Mines cooperated in a full-scale slope stability research project at the Kimbley Pit near Ely, Nevada. The west wall of the Kimbley Pit was mined from an over-all slope angle of 45° to an over-all slope angle of 57° and the pit was deepened nearly 50 feet. Mining was begun February 10, 1966, and completed October 30, 1966. Approximately 2.55 million tons of ore and waste were mined. The project was designed and carried out to learn more about the behavior of a pit wall as its slope angle is steepened and to obtain a possible insight into engineering parameters for use in design of future slopes. Data from this project are presented as obtained from (1) instruments installed in two adits for deformation and displacement measurements, (2) a subsidence network installed on the surface above the steepened wall, (3) pre- and post-mining observations of stress in rock, (4) sonic-velocity measurements, and (5) seismic and microseismic investigations. Changes occurred in measured quantities as the slope was steepened but stopped when mining was completed. Agreement was noted between the measured changes and predicted changes obtained from calculations using the finite element technique. Subsidence surveys and microseismic studies show the steepened pit wall to be stable.

Jan 1, 1968

MORE than 2500 members of the AIME from all sections of the U. S., Canada, Central and South America are expected to attend the 175th General Meeting scheduled for Los Angeles Feb. 16 to 21, 1953. The Southern California Section will be host for the meeting. It will be the first time the Annual Meeting has been held in Los Angeles. In keeping with the historic nature of the event, headquarters will be the fabulous Statler Hotel, recently opened in downtown Los Angeles. Meeting plans are under the direction of Henry T. Mudd, General Chairman, assisted by members of the Southern California Section, headed by Basil P. Kantzer. The annual business meeting is scheduled to be held Tuesday, February 17, at the Statler Hotel, at 4 pm. Minerals Beneficiation Div. has planned a very important program, with 10 technical sessions, the famous Scotch Breakfast, divisional luncheon, and participation in the Mining Branch and All-Institute activities. The MBD business meeting is set for the morning of Feb. 16, 1953.

Jan 1, 1952

By Roberto Fustos, Gonzalo Nelis, Víctor Sanhueza, René Gómez

Hang-ups are a challenging operational problem in block caving mines. They interrupt ore flow and imply hang-up removal activities, which take time, increase operational costs, and involve safety risks for workers. In block caving, the number of hangup events of mechanical arches that appear in the drawbell could be decreased if the drawbell geometry is adequately designed based on the expected rock fragmentation. Here, we compile and analyze hang-ups reported in several block caving mines to evaluate the impact of drawbell geometry, rock size, and vertical load on hang-up frequency at the mine scale. Results show that the ratios between drawbell length and d80 (dL/d80), between drawbell width and d80 (dW/d80), between drawbell volume and rock volume (VDB/VP), and vertical stress are critical parameters that define hang-up frequency. These findings confirm previous experimental studies, where a decrease in hang-up events is associated with an increase in these ratios. Additionally, these outcomes are used in a support vector machine algorithm to define categories of hang-up frequency. The results of this classification algorithm are presented in diagrams, which can be used to estimate the hang-up frequency in block caving projects.

Oct 15, 2022

The range of environments which control the level of ground support required vary greatly in terms of rock type, rock structure and the mining method used. At the operating mines visited during this review the level of ground support was such that incidence of rock-fall- related accidents were rare. It was also found that methods existed to support most of the conditions encountered. The main problem areas were weak rock, major geological structures, areas of heavy jointing, high development density and areas affected by high mining-induced stresses and blasting. With stope support the level of success was much lower than that achieved in development; the larger spans and more rapid changes in geometry of open stopes compared with cut and fill stoping (where cables were originally used) led to significant problems. Before a useable design method for cable support of stope walls and pillars can be obtained the precise failure mode of the walls and pillars, and the interaction of the support and the rock mass must be investigated.

Jan 1, 1982

By Stephen Krickovic

THERE is today no proven continuous mining machine that can be used under all the varying conditions found in most bituminous coal mines. During the last five years, however, both the machines and methods of using them have been improved. Coal operators can now exercise a fair degree of selectivity, except those who must struggle with thick and hard partings and operate in seams less than 40 in. thick. It must be realized, then, that more than one type of continuous miner may be used advantageously in the same coal mine. Although this appears to violate the standardization generally sought after, the gain can more than offset added maintenance and, to a lesser degree, operating problems. It is equally important to realize that most coal mines will require one or more conventional loader setups in the plan of total mechanization. Seam variations, adverse physical conditions, and special designs of main headings for haulage and ventilation, to name only three reasons, may render 100 pct use of continuous miners uneconomical. Of course, this situation may change. In one mine, prior to 1954, machines that ripped coal from the face achieved only average results, as compared to the impressive performance of boring-type units in another coal mine. A boring-type unit was therefore selected for the first mine in the belief that the chance of success would be greater if it were used in a combined cutting and wedging operation, cutting a full face in its forward movement. The miner was located in a part of the mine where a special projection could be readily developed, a solid area scheduled for mining according to the overall progress plan. All openings surrounding the particular area were carefully investigated for the presence of sulfur balls. Fig. 1 (opposite) shows the representative seam section, containing the several pyritic bands 1/4 to 1/2 in. thick; a characteristic Pittsburgh seam parting averaging 5 to 6 in. of slate, coal, and slate; and the occasional py-rite lenses 2 to 3 in. thick. Under normal conventional mining methods, the top 10 or 12 in. and the bottom 4 in. are left in place to offer roof protection

Jan 1, 1958

By D Gelling

Mineral sands dry plants consist of a large number of repetitive unit operations with significant transport delays, in the form of stream pipes, conveyors and elevators. As part of the P255 mineral sands sponsorship and using the JKMRC developed conductor/non-conductor (CNC) gauges, control loops have been implemented at the multi-duty dry plant of Cable Sands Limited at Bunbury. This paper describes the loops for kV control on a rougher high tension roll machines (HTR). The options available are product grade control and ratio control using the JKMRC CNC gauges, which measure the percent conductors in a stream. The purpose of control is to stabilise the circuit operation against disturbances such as atmospheric and feed disturbances.

Jan 1, 2001

By N. A. Duke, R. L. Carpenter

Abstract - The West Meliadine area is underlain by structurally interleaved panels of mafic and minor ultramafic metavolcanic rocks and metasedimentary rocks that occur along the northern margin of the Neoarchean Rankin Inlet greenstone belt. Three structural and metamorphic domains are recognized: (1) the easterly Wesmeg domain; (2) the central Barracuda-Ridge domain; and (3) the westerly Peter Lake domain. The Wesmeg domain is characterized by a series of southeast-trending, north-dipping, foliation-parallel panels of greenschist facies mafic metavolcanic and metasedimentary rocks. The Barracuda-Ridge domain is comprised of greenschist to amphibolite facies mafic metavolcanic rocks that define an east-northeast-trending structural grain. The Peter Lake domain consists of amphibolite facies mafic metavolcanic rocks and minor metasedimentary rocks intruded by a monzonite pluton. West Meliadine hosts the economically significant Wesmeg gold deposits, as well as other important gold showings across the Barracuda-Ridge and Peter Lake domains. The geological setting of the Wesmeg gold deposits resembles that of a break or fault zone. The Pyke Break is a major geophysical discontinuity (>65-km strike length) and is the first-order structural control on gold mineralization at West Meliadine. It is several kilometers wide and characterized by polyphase deformation and shear zone development accompanied by lode-gold mineralization. In general, gold concentration is related to quartz and iron-carbonate veining, iron sulfides (mainly arsenopyrite and pyrrhotite), and accompanying silicate alteration minerals that overprint favorable chemical and structural traps late in the history of deformation.

Jan 1, 2004

By S. McCalpine, W. Muir, P. J. Downey

"Golden Bear Operating Company Ltd. (GBOC) was a joint venture between Chevron Minerals Ltd. and North American Metals Corp. that was formed for the development of the Golden Bear Project. The mine is located at Muddy Lake, Western British Columbia, approximately 80 km northwest of Telegraph Creek and 135 km west of Dease Lake (see Figure 1). Road access to the mine was developed by Golden Bear and consists of a radio controlled 155 km gravel road originating at highway 114.A feasibility study was completed in 1987 and project approval was finalized in 1988 with Golden Bear issuing a ""Request For Proposal"" for detailed engineering in May 1988. Golden Bear also commenced negotiating with vendors at this time for the supply of the two principal process equipment packages: the dry grinding and classification system and the whole ore roaster package.Process plant engineering and plant construction was completed in October 1989 with plant startup commencing in November 1989. Two months into the commissioning phase it became apparent that the plant, as designed, would not achieve the anticipated production targets.Kilborn Engineering Pacific were subsequently contracted by GBOC to review and modify the plant design, assist the operating personnel with plant commissioning and startup and provide operator training.This Paper describes some of the problems encountered during the startup and the plant modifications incorporated to achieve design parameters."

Jan 1, 1992

By Muhammad Hashim Rasa, Ali Aras

Recovering valuable metals from tailings not only provides a source of income but also helps conserve natural resources, thus reducing the environmental impact of mining activities. This study aimed to recover cobalt using tailings of the Küre flotation plant, which were determined to contain 0.17% Co. Flotation tests were conducted on samples taken from the tailings to investigate the effects of pH and pH conditioning time, as well as types and amounts of depressant, collector, and frother on the grade and recovery of Co in rougher flotation tests. Optimal flotation conditions were determined to be pH 6, a pH conditioning time of 7 min, and dosages of 50 g/t, 100 g/t, and 20 g/t of sodium hexametaphosphate, X-231, and DF250, respectively. According to the rougher flotation test results, a concentrate with a Co grade of 0.46% and a Co recovery of 84.15% was obtained. In the concentrate obtained using cleaner flotation tests, the Co grade increased to 0.49% and Co recovery increased to 90.56%. These results show that froth flotation, carried out with the appropriate reagents, is a successful method for recovering Co by enriching the flotation plant tailings.

May 17, 2023

By Xiqing Wu, Tao Yue

"Magnetic seeding co-agglomeration, i.e. adding magnetic seeds, organic depressant (such as the starch), and pre-magnetization for co-agglomeration, was applied in the reverse flotation for hematite ore slimes. Samples of the hematite ore slimes and the pure hematite fines were investigated by the co-agglomeration, respectively. Flotation tests suggest that the depressing effect of starch in the flotation was further enhanced with the addition of magnetic seeds and the pre-magnetization. Following the laser size analysis and flotation tests, TOC adsorption analysis, VSM, surface charge measurement, AFM, FTIR study, and SEM imaging have been utilized to understand the depressing mechanism. Based on the results it was demonstrated that with the addition of magnetic seeds and the low field pre-magnetization, the adsorption of starch onto hematite minerals was strengthened through the magnetic seeding co-agglomeration. The co-agglomeration improved the flotation performance by introducing the magnetic interactions and acting the bridging adsorption of starch between fine hematite and magnetic seeds in a suspension, and then resulting in an intensified coverage of the starch onto hematite particles and positive action in the coagglomeration under the pre-magnetization.INTRODUCTIONFlotation is undoubtedly the most important and versatile mineral processing technique, and both its use and application are continually being expanded to treat greater tonnages and to cover new area. However, its influence was limited to a relatively narrow particle size range of 5–300 µm (Wills and Napier-Munn, 2006). The main reason behind this limitation lies in that fine particles have high specific surface area, so need relatively high dosage of reagents (collectors or depressants), and exhibit poor flotation performance while intermediate or properly-sized particles produce high flotation kinetics and need low consumption of reagents. In order to increase the recovery of fine particles, quite a few of flotation techniques have been investigated. Most of them are based on particle aggregation, namely, by enlarging the apparent particle size, such as carrier flotation (Valderrama and Rubio, 1998; Zhu et al., 2009), shear flocculation (Duzyola and Ozkan, 2010; Yin et al., 2011), selective polymer flocculation (Rubio et al., 2007; Song et al., 2012; Ng et al., 2015), oil agglomeration (Laskowski and Lopez- Vladivieso, 2004; Forbes, 2011), temperature responsive polymer flocculation flotation (Forbes et al, 2011), and etc. However, these techniques above-mentioned are mainly studied for the direct flotation of fine minerals, and most of these techniques have not found applications for iron ore slimes especially recovered by the reverse cationic or anionic flotation."

Jan 1, 2018

By None None

This month, Mining, Metallurgy & Exploration (MME) is launching with a special issue on mineral and metallurgical processing in honor of the 90th birthday of Prof. Emeritus Douglas W. Fuerstenau. Headed by a special review article by Prof. Fuerstenau, this special issue contains 22 papers by the top experts in the field.

By Paul R. Bluekamp

The Mather Mine property is composed of a 5.2 sq km (2 sq mile) area within the Cities of Ishpeming and Negaunee which are located in the Upper Peninsula of Michigan. Production in this mine started in 1943; it ran continuously until its closing in 1979, having produced slightly over 55 million tons of iron ore. This mine was a joint venture between several steel companies and The Cleveland- Cliffs Iron Company (CCI), with CCI being the fee holder and operator. For the first 17 years of operation, the ore was shipped in its natural state directly from the mine to the steel mill. By 1960, iron ore pellets were on the market and they proved so superior to the natural soft Mather ore that the latter became difficult to sell except for the coarser portions. It was decided to develop a pelletizing process for the Mather ore, and this was accomplished by 1965 when the first Mather pellets were produced. From this date, all but the coarse fraction of the Mather ore was shipped as pellets. The geological setting is that of a large east-west trending synclinorium which plunges to the west. The lowest member of this trough-like structure is, for the most part, a quartzite-like graywacke, the upper 20 m of which grades into a softer, fine grained slate. Lying conformably on this graywacke footwall member is an iron- formation member which is over 1,300 m in thickness. This iron-formation is cow posed of thin alternate bands of iron oxides and chert and is intruded by a number of diorite sills - some up to 122 rn (400 ft) in thickness. The north limb of this synclinorium dips at approximately 45' and bottoms out at about 1,000 m from surface in the central part of the mine. There are two sets of faults, many of which are intruded by diorite dikes, which trend east-west and southeast-northwest. Displacements are varied, reaching a maximum displacement of 243 m (800 ft). The ore is found lying directly on the slatey footwall and its position is largely controlled by the faults and dikes, with the bulk of the ore being on the upper side of these structures. The ore is composed of soft earthy hematite and martite with vertical thicknesses up to 122 m (400 ft) although the average thickness would be closer to 46 m (150 ft). The ore averaged 60% iron and 7% silica on a dried analysis. The Mather Mine is located on the north limb of the syncline and was worked from two shafts, the deepest of which. was 1,09 7 m. These two shafts are about 2 km apart and serviced a total of 8 working levels between them during the life of the mine. The level spacing was about 61 m (200 ft). The main haulageways were driven parallel to the ore/footwall contact in the hard cow petent graywacke wherever possible. On the lower levels, deeper into the footwall, naked development was common. This material graded into roof bolting ground towards the upper stratigraphic portion. As drifting progressed further into the upper stratigraphic portion of the footwall, progressively stronger steel sets had to be used. From the main haulageway, cross-cuts were turned into the orebody on 61 m (200 ft) centers and extended as far as needed to recover the ore available to that particular cross-cut. The main haulageways and cross-cuts were driven 3 m (10 ft) high, 3 m (10 ft) wide at the top and 4.6 m (15 ft) wide at the bottom. This configuration would accomodate a set composed of a 2.7 m (9 ft) cap on top of two 2.7 m (9 ft) legs angled out at 18'. A sill plate was used under each leg to prevent its sinking into the ore below. The type of steel used was dependent on the expected weight to be experienced. Sets were placed on 1.63 m (5 ft 4 in.) centers; however, in extremely heavy areas, it was sometimes necessary to install sets on 0.8 m (2 ft 8 in.) centers. On the top two levels (5th and 6th), the ore was considerably hard-er and was mined by sub-level stoping and long hole drilling. Very light steel sets were used in the main haulageways and cross-cuts and timber was used in the production drifts. Some con- crete production drifts were installed on 6th level, but proved to be uneconomical. However, as mining reached greater depths, the ore became softer and more massive, reaching its maximum vertical heights on 11th and 12th levels. On levels 7 through 10, yielding steel sets were used extensively in the slusher drifts. While they were satisfactory on 7th and 8th levels, their success diminished with depth and they were

Jan 1, 1981

By Catia Casagrande

As principais operações unitárias de um processo de pelotização são: moagem, filtragem, aglomeração e queima. Das etapas citadas, a moagem, como início do processo, é uma etapa crucial para as demais operações. Foi baseado neste fato, que o presente estudo foi realizado, tendo como objetivo principal, avaliar o comportamento das misturas de minério de ferro com diferentes índices de moabilidade e definir os possíveis ajustes dos parâmetros de processo de moagem em função dos diferentes perfis mineralógicos. Foram usadas três misturas de diferentes níveis de índices de moabilidade: baixo, médio e alto, a fim de observar as respostas nas variáveis operacionais de possível atuação na rotina operacional. Para isso, uma matriz de experimento foi realizada, com oito testes para cada mistura de minério de ferro, variando os parâmetros de moagem de mínimo a máximo. As variáveis respostas foram às características físicas das misturas, superfície específica e granulometria, onde foram medidas no sistema automatizado PSI. Através dos resultados dos testes, foram calculados modelos matemáticos de superfície específica e granulometria das misturas em função dos parâmetros operacionais. Os modelos mostraram satisfatórios e aplicáveis, apresentando correlação entre as variáveis dependentes e independentes.

Aug 1, 2018

By Maria Stella Vasconcellos Limaverde

A utilização de moinhos verticais na indústria mineral vem ganhando destaque nas últimas décadas, quando granulometrias mais finas passaram a ser exigidas, principalmente por tais equipamentos apresentarem maior eficiência energética para moagem fina e remoagem em comparação aos moinhos tubulares tradicionais. O dimensionamento destes equipamentos ainda é fortemente embasado por experiências em projetos similares, metodologias proprietárias dos fornecedores de equipamentos e plantas piloto, criando a necessidade do desenvolvimento de ensaios laboratoriais abertos e confiáveis para guiar tais trabalhos. Outro problema observado é a quantidade de material necessária para a realização dos ensaios, da ordem de ao menos 1 kg, o que pode ser um impedimento nos estágios inicias de um projeto, quando pouca massa de material proveniente de furos de sonda está disponível. Este trabalho apresenta o desenvolvimento de um ensaio laboratorial com moinho de jarro para dimensionamento de moinhos verticais segundo uma abordagem energética que utilize uma pequena massa, de aproximadamente 500 g. O ensaio proposto mostrou-se de maior praticidade em relação em relação ao ensaio com jarro maior, por utilizar aproximadamente metade da massa necessária e também apresentar boa aderência aos resultados industriais.

Aug 16, 2017

The Hyde-Macraes Shear Zone (HMSZ) is low angle, duplex thrust system that has developed in a graphitic pelite schist within the Otago schist belt, New Zealand . At Round Hill lateral ramps have developed parallel to the movement direction between+ the footwall and hangingwall shears. These structures exhibit normal oblique-slip displacement which has been an important mechanism for creating space in an otherwise compressional system. Gold-bearing quartz veins were emplaced along the boundary shears and lateral ramps during thrusting. Gold within the lateral ramps is coarse (up to 300 microns) however hanging wall cataclasite development in the boundary shears after the gold was deposited has reduced the grainzise of most of the gold to less than one micron. This bimodal gold distribution has metallurgical implications with respect to the maximisation of gold recovery at Macraes.

Jan 1, 1993

By Steferson Ferreira de Oliveira

A máquina convencional de usinagem por descargas elétricas (EDM) é utilizada no processo de nitretação por descargas elétricas (NDE), pois permite enriquecer a superfície de um substrato de aço SAE 52100 com nitretos em meio a uma solução dielétrica composta por água deionizada e ureia. O objetivo do presente trabalho é caracterizar as camadas formadas ao utilizar uma solução dielétrica (30g / l) como fonte de nitrogênio e comparar as propriedades obtidas quando se utiliza eletrodo de cobre eletrolítico. As caracterizações das amostras no processo foram por meio das técnicas de microscopia óptica (OM), microscopia eletrônica de varredura (MEV) e microdureza Vickers (HV), revelando a formação de uma camada uniforme e endurecida de nitretos, recoberta por uma camada branca (Zona Refundida), como comprovado por difração de Raio-X (DRX) a formação de compostos com nitrogênio. As camadas formadas no eletrodo peça temperado e revenido foram mais espessas do que as que formaram no eletrodo peça esferoidizado. Para avaliar o desempenho do processo de usinagem, foram utilizados parâmetros como taxa de remoção de material (TRM), taxa de desgaste do eletrodo ferramenta (TD) e desgaste volumétrico (DVR).

Oct 30, 2017

By Henrique de Araújo Resende

Em função do aumento constante da produtividade nos laminadores e linhas de processo de usinas siderúrgicas e de alumínio, está cada vez mais evidente a necessidade de aumentar as campanhas de produção dos mandris de bobinadeiras e desbobinadeiras, reduzir as paradas de manutenção das linhas além de minimizar ao máximo os custos com reparo e fabricação. Para isso, é importante conhecer os diversos projetos de mandris utilizados nos últimos anos, os maquinários envolvidos na fabricação e reparo desses equipamentos além de estar atualizado quanto às novas tecnologias que poderão ser aplicadas. Para possibilitar o entendimento descrito acima, este material será divido em 3 partes: Projetos SMS de mandris no Brasil; Condições de fabricação e reparo; Novas tecnologias. A SMS possui 58 mandris em operação hoje no Brasil estratificados entre os seus principais clientes, desde linhas de decapagem e laminação a quente até a saída do produto final após os laminadores a frio na galvanização e linhas de pintura e corte. Com isso, é possível estabelecer uma análise técnica do desempenho de cada equipamento e sobressalente de forma bem estruturada. Esse é o principal objetivo desse trabalho.

Aug 16, 2017