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By Lucas Coelho da Silva

Este trabalho tem a pretensão de problematizar a influência da corrosão, sobretudo nos trocadores de calor, e trazer a importância de um estudo prévio detalhado, ressaltando como o desempenho deste equipamento está relacionado ao material utilizado e como esse estudo pode nos abster de grandes dispêndios. Através de ensaios eletroquímicos vamos estudar o comportamento do aço Super Duplex UNS S32750 em diferentes meios corrosivos, com o intuito de avaliar se o nível de corrosão do material é adequado para as condições severas de operação impostas a um permutador de calor.

Aug 16, 2017

By José Helio Sobrinho

O aço inoxidável austenítico alto nitrogênio ISO 5832-9 tem sido promissor na manufatura de próteses ortopédicas com resistência mecânica e de corrosão superiores a do aço tradicional ASTM F-138. Pesquisas recentes têm mostrado que esta liga apresenta resultados peculiares, tais como alta resistência mecânica, resistência à corrosão e presença de partículas de segunda fase (Fase Z) na matriz. Entretanto, não se sabe como se dá a correlação mecânica-microestrutural e de taxa de corrosão em regiões de processabilidade industrial desta liga na conformação a quente de peças. No presente trabalho, ensaios sob torção a quente isotérmico contínuo e interrompidos foram realizados após tratamento térmico de solubilização a 1.200oC, por 300 s com intervalos de temperatura variando entre 1.000°C e 1.200oC com taxa de deformação variando entre 0,05 e 5 s-1, para investigar a trabalhabilidade e a taxa de corrosão sob tensão desta liga com aplicação da técnica de espectroscopia de impedância eletroquímica e caracterização microestrutural, conduzida por microscopia óptica e varredura. Os resultados mostram que a tensão de escoamento é sensível aos parâmetros de deformação. O pico de tensão diminui com o aumento da temperatura e redução da taxa de deformação com alta taxa de recuperação dinâmica com grãos alongados, gerando regiões de acúmulo de tensão. Essas regiões de instabilidade plástica apresentam grau de corrosão superior em relação às amostras como-recebidas. As regiões de alta temperatura e baixa taxa de deformação apresentam boa trabalhabilidade com microestrutura final refinada com grãos recristalizados dinamicamente.

Aug 8, 2018

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 Lee W. Saperstien

Section 709 of Public Law 95-87, "Surface Mining Control and Reclamation Act of 1977," requests the National Academy of Sciences to determine if the standards of the act are applicable to minerals other than coal. The Academy has established the Committee on Surface Mining and Reclamation to respond to this charge. The committee is, itself, organized to be responsive to the varieties of minerals, environments, mining systems, and locations that are representative of U.S. mining systems. It is not the charge of the committee to write trial legislation, but to indicate whether or not technology, extant or proposed, is available to meet the standards of the law, and at what cost. It is within this set of conditions that the committee will examine blasting.

Jan 1, 1979

By Bill Birch, Toby White, Ian Brown

costs/benefits of electronic detonators and measure their effects on primary crusher productivity. 9 blasts (3 shock tube blasts, 6 electronic blasts). Measured electricity consumption & primary crusher throughput. Change to Electronic Detonators resulted in a 7% reduction in primary electricity costs and a 15% increase in primary crusher productivity.

Feb 1, 2020

By Junho Lee, Leopoldo Bello, Mohammad Khan

Multiline ring anchor (MRA) was developed as a cost-effective solution for securing floating offshore wind turbines (FOWTs) to the seabed. Since FOWTs are placed in arrays compared to conventional oilgas platforms, the availability of multiple mooring lines attached to a single anchor can make FOWTs economically competitive. Additionally, the MRA can provide a small and light-weight anchor with high resistance to mooring line loads due to its deep embedment depth. The combined benefits of fewer and smaller anchors allow substantial reductions in anchor material, transport, and installation costs. Such cost reductions are critical to reducing foundation costs for the wind turbine, which can comprise well over 25% of the capital costs for offshore wind power development. Since it substantially reduces the required number of anchor footprints, the multiline anchor concept also permits significant reductions in seabed site characterization costs. This study evaluates how potential advantages of the MRA can reduce the capital costs and provide a reliable cost analysis of the foundation component for offshore wind farms.

By Esher Hsu

In Taiwan, around 8,000 tons of batteries are consumed per year but only 17% of which were collected and transported to oversea for recovery. The dumping used batteries have become a potential environment problem. How to increase the collection efficiency of used batteries is an important issue for EPA in Taiwan. Besides, some treatment plants are on process to get permission for recovering used batteries. The objective of this study is to estimate the recycling cost of spent consumer battery and explore the feasibility of recovering used consumer batteries in Taiwan based upon the principle of cost-efficiency. The recycling cost including collection cost and recovery is estimated based on a survey. Results would provide some suggestions to Taiwan EPA regarding feasibility of recovering used batteries in Taiwan.

Jan 1, 2006

By C. E. Arnold

IN attacking the problems of mining and treating large disseminated copper orebodies such as those occurring in the Miami or the Ray district of Arizona, one of the vital questions to be decided is, "What is the value of the orebody per ton in place?", The importance of this question lies in the fact that its answer helps to reach a decision concerning the method of mining to be employed. Let it be assumed for the moment that milling and smelting practice, especially the latter, is more or less standardized; and has a cost incapable of much variation, while the cost of mining has a considerable range owing to the large number of mining methods by which a mine may be exploited. For instance, a method may be adopted whereby all the available ore can be extracted perfectly clean (i.e., unmixed with barren material) at a relatively high cost; or a method may be used involving the necessary dilution of a portion of the ore with waste, in order to recover that portion, thus giving a relatively poorer extraction but nevertheless a lower working cost than the former method. Naturally, between these two extremes there lie intermediate methods. In order to arrive at a solution of a mining method problem that is of any value, it is necessary that accurate and complete estimates of mining, milling, and smelting performances and costs shall be established. By way of illustration is given a, case of mining, by the system of undercutting and caving, a large orebody with an average content of 1.50 per cent. copper supposedly all in sulphide form, to which case has been applied those results in mining operations which one may fairly assume are to be expected when viewed in the light of G. R. Lehman's ore-drawing experiments, which form the subject of a paper to be presented at this meeting. It is assumed that the total. costs will be as shown in the following table:

Jan 9, 1916

By J. H. Gasper

A cost analysis for two alternative contour mining methods is performed to compare their results. The methods evaluated are: 1) Front-End Loader/Truck Haulback and 2) Scraper Haulback. Before performing these analyses, output requirements and equipment needs are found based on integrating geometric limitations with material properties. The final geometric design establishes limits on clearances, congestion problems and grades (with associated running surface materials). All calculations are kept practical, with assumptions clearly stated. The cost analysis is a unique, easily utilized compilation of several different approaches incorporating cost engineering principles. The strengths of each are welded into a cohesive entity once again aimed at practical application. The results are presented in tabular form.

Jan 1, 1992

By Ronnie Melker, Joel Webster, Daniel Belardo, Thomas Hyatt

Quality assurance of drilled shafts is highly dependent upon the construction practices used during the excavation and concrete placement process. When performed under slurry, challenges are presented in maintaining borehole stability and placing a uniform and continuous concrete column. An integral part of the construction operation is to evaluate the as-built integrity to verify the drilled shaft is capable of safely transferring the required loads to the subsurface. One such case is exemplified by a project involving the expansion of an existing public transit rail line. The expansion required the construction of a new bridge bent supported by seven 5-ft diameter drilled shafts of various lengths. The specifications for the project required both Cross-Hole Sonic Logging (CSL) and Thermal Integrity Profiling (TIP) for drilled shaft integrity evaluation. Both of these post-construction non-destructive tests (NDT) are designed to detect anomalies or the presence of regions with reduced concrete quality. However, the implementation and methodologies of performing each of these NDT methods are significantly different. Cross-Hole Sonic Logging uses ultrasonic signals emitted from a transmitter in one access tube to a receiver in another access tube. Thermal Integrity Profiling measures elevated concrete temperatures during the curing process using embedded thermal wires. Both test methods require different access details and therefore have different costs associated to the contractor or owner to perform. This paper provides a detailed comparison of these NDT methods as performed on these seven drilled shafts. The associated financial and time costs to perform each method are compared and the results are discussed in detail. INTRODUCTION Drilled shafts are a fundamental deep foundation element that can be designed to resist large bending moments as well as to transfer large compressive axial loads through weaker upper soils to competent soils or rock below. Often times when the upper soils are very soft, lateral support by casing or slurry, or a combination of both, may be used to hold back the softer soils from collapsing into the hole before concrete has been placed. Due to the blind nature of drilled shaft construction and the qualities issues that may arise during the placement process, it is essential for design engineers to specify non-destructive testing (NDT) to assess the quality of the as-built drilled shaft. CSL and TIP testing are two NDT methods designed to detect areas of lower quality concrete or soil inclusions throughout the length of the drilled shaft and determine its overall quality. Each method uses different technologies and methods to achieve the results required to determine the as-built condition of the drilled shaft.

Jan 1, 2019

By M. K. Buder

Over the years, electric utilities have been faced with declining steam-coal quality, rising coal prices, and increasingly rigid and costly environmental standards. As a result, many utilities are beginning to show an interest in coal cleaning to upgrade coal by improving its heating value and reducing its sulfur content. Among the primary questions that must be answered for these electric utilities is whether the cost of coal cleaning can be offset by the potential cost savings in coal transportation and utilization in power generation. In search of at least a preliminary answer to this question, the Electric Power Research Institute (EPRI) engaged Bechtel to perform a case-by-case economic study to examine the cost effects of using cleaned versus raw coal to fuel new electric power plants. The results of two of the seven cases studied which involved western coals - one from Montana and another from Utah - are reported in this paper. Coal cleaning appeared to be a cost effective option for the Utah coal, but only marginally economical for the Montana coal. These cost results are expected to change to favor coal cleaning to a greater extent with the application of the final revised New Source Performance Standards regulating SO2 emissions just issued by the EPA. The study was based on the regulation originally proposed by the EPA under the 1977 Clean Air Act Amendments.

Jan 1, 1979

By Karl-Michael Zett, Marcus Kirschen, Ashraf Hanna

"Bottom purging systems are well established practice in refining metallurgy and an increasingly important technology for the EAF process in steel industry. Direct bottom purging provides efficient mixing of the entire melt volume, increases turbulence intensity in the melt, and improves mass exchange of alloy additions or oxygen. Continuously rising columns of dispersed gas bubbles helps to increase removal of volatile components carbon monoxide and hydrogen avoiding boiling retardation. Direct purging (DPP) systems provide cost benefits especially for EAF processes where metallurgical constraints have to be considered, e.g. for alloyed and stainless steel production, foundries. Case studies are presented.IntroductionCurrent EAF Technology and Production Constraints in Steel IndustryApproxinlately 30 % of steel, 446 Mio. tonnes in 2011, was melted in the electric arc furnace (EAF) with increasing tendency due to its large flexibility in production volume and raw materials. With recent increase of prices for ferrous raw materials the requirements for production of high quality steels from low quality scrap, direct reduced iron (DRI) and blends of ferrous raw material in various qualities increased. Maximum yield of ferrous raw material, oxygen, carbon, and alloys and minimum energy costs are of high importance. At modem high productivity levels even small process improvements provide considerable cost savings. Such improvements are caused, e.g., by efficiency increase of mass and energy transfer in the EAF. Optimization of flow pattern in the steel bath is therefore important for efficient melting of scrap and DRI and high homogeneity of the melt [1-4]."

Jan 1, 2013

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION This method is a variation of cut-and-fill mining where either the wall rock or the vein is too unconsolidated to permit safe mining by conventional cut-and-fill methods. Basic development consists of driving a 2.4 x 2.7-m (8 x 9-ft) haulage heading laterally in the footwall side of the vein. Crosscuts are driven over to the vein at 91-m (300-ft) intervals. Then the vein, at each inter¬section, is mined out to a distance of about 9 m (30 ft ) above the track level. After the installation of an ore chute, scrams or intermediate drifts are driven in both directions along the vein. Mining then starts imme¬diately below the level bounding the upper limit of the stope. A cut varying between 3 and 4.2 m (10 and 14 ft) in depth and mined for a distance of 45.7 m (150 ft )along the strike is taken from the orepass in each direction (Fig. 1). In this mined-out void, the floor of the cut is equipped with 203 x 203-mm (8 x 8-in.) caps hitched into the wall with steel saddle hitches installed on 1.5-m (5-ft) centers. Double wood stringers, 152.4 x 152.4 mm (6 x 6 in.), are then installed on 1.2-m (4-ft) centers longitudinally down the length of the stope. Reinforc¬ing wire mesh is also placed, overlying each mat of timber. Directly above the stringers, stulls are installed from hanging wall to footwall to hold the timber in place during the sand-filling operation. Then a cement¬sand mixture in a ratio of about 1:10 weight is poured to a depth of about 2.4 m (8 ft). The entire cycle is then repeated down to the scram drift. Another variation of timbering is illustrated in Fig. 2. There is a vein width of approximately 5.5 m (18 ft) with the same unconsolidated ground conditions. The primary difference between handling this situation and the previously described undercut-and-fill method is that standard timbered sets are installed in each excavated stoping void rather than stulls. Once again the basic method is to work down the ore block from an upper timbered drift to a lower timbered drift. Cuts are taken from a timbered raise with a chute on the stope side. Each piece of ground excavated is essentially two sets high, 5.5 m (18 ft), by two sets wide, 5.5 m (18 ft), by the length of the cut along the strike of the vein. The timbered sets are 2.4 x 2.4 x 2.4 m (8 x 8 x 8 ft). The net result is, therefore, two sets of timber on a top floor separated by doubled-up 76-mm (3-in.) lagging from

Jan 1, 1982

By F. H. Mertens

The costs of an open-pit operation can be classified according to the following hierarchy : -overall cost of production (per ton of ore) -costs of basic operations and functions -costs of auxiliary operations -costs of machines and equipment -maintenance costs. Production cost is a measure of the open-pit mine performance and should be the final target of the management control. It is also a measure of the share of the mine in the processing stages.

Jan 1, 1990

By Jose Maria Rodriguez-Ibabe, Ronaldo Antonio Neves Marques Barbosa, Douglas Glenn Stalheim

Major costs in the production of both flat and long commodity grade structural steel products include alloy, labor and energy. Flat and long commodity grade structural steels such as ASTM A36, ASTM A527Gr50, S235, S275, S355 and other equivalent world societal standards represent over 500 million annual tons worldwide. These simple commodity grades consist of a minimum of three base common alloying elements, carbon, manganese, silicon and then can be supplemented with microalloying elements of either vanadium or niobium. Since 2016 raw material costs for two of the five alloying elements in these commodity grade structural steels, FeMn and FeV, have risen significantly and/or have become volatile making it difficult to maintain stability in profitability for these commodity grades. For steel plants producing hundreds of thousands and in some cases over a million tons annually of these common structural steel grades because of the significant alloy cost increase for Mn and V alloy additions has squeezed profitability of these grades. These grades typically represent the base loading for cost controls in most all steel plants and hence a significant cost increase or volatility in two of the five elements used for these grades having a negative effect on overall production costs. However, with a proper strategy for alloy designs working in conjunction with the mills existing processing capabilities to achieve the desired end metallurgy/mechanical properties, alloy costs and operational efficiencies can be realized.

Oct 1, 2019

By P. M. Tait, N. A. Warner

"The refining of thermally produced zinc is currently performed using a highly energy intensive and arguably outmoded process, distilling zinc under atmospheric pressure at temperatures of 900°C to 1000°C. In order to produce refined zinc more profitably and with less environmental impact a new process has been proposed. This process is generic to all thermal zinc smelters and aims to match the purity of zinc from current technology (99.99%-99.999%). The objective is to perform all separation processes using only the heat contained in the smelter product stream. Vacuum conditions are extensively used lowering operating temperatures and enhancing separations. As part of this process zinc is distilled under vacuum from high concentration obtaining a refined product with respect to lead content. A mathematical model has been developed from fundamental principles to describe this stage, the distillation of zinc across a temperature difference between two circulating zinc streams in open channels under vacuum. Heat for the distillation is supplied by the liquid phase, necessitating a large recycle to maintain high operating temperatures, and keep zinc concentrations high. In determining this model and the operating conditions required, relationships between temperature, pressure and composition are considered.IntroductionCurrent technology (the Refluxer) used to refine thermally produced zinc is both expensive and highly inefficient. At present 80% of the worlds zinc is produced via the electrolytic process reflecting the poorer performance of the thermal route, the zinc-lead blast furnace and Refluxer. A new process is proposed to overcome the drawbacks of current processes. These drawbacks are both economic and environmental. The Refluxer requires 6.6 GJ of energy to refine one tonne of SHG zinc 1, other disadvantages of the Refluxer stem from high operating temperature, around 900°C, and liquid zinc being highly corrosive towards ferrous metals. The Refluxer is constructed from silicon carbide sections which join together by means of flat machined faces making liquid metal leaks inevitable. The Refluxer columns are very sensitive to thermal cycling, thermal shock and mechanical vibration, which easily lead to structural failure. To avoid thermal stress on the columns the feed rate and feed temperature need to be kept constant which means the feed baths to the columns must be able to accept solid and liquid feed as the Refluxer has to operate when the smelter is shut down."

Jan 1, 1998

By G Lane

This paper discusses the factors that contribute to the cost effective design of a concentrator. Concentrator design and layout outcomes are functions of the team (engineer’s and owner’s) participating in each particular project. The benchmark for relatively modest projects was set in the 1980s and 1990s during the ‘gold boom’ when numerous cost effective plants were designed and constructed on a lump sum basis in a very competitive market. A number of factors contribute to lack of transference of the lessons learnt in gold plant design to concentrator design including established paradigms in the design and layout of concentrators, lack of experience in cost effective design, operator’s preferences for flowsheet and layout and simple lack of appreciation of the impact of plant footprint on materials quantities and resultant capital costs.Experiences with recent copper concentrator projects (both small and large) are used as case studies.

Aug 8, 2011

By Phil Wilkinson, Sam Whitaker

"The installation of large (XL) diameter monopiles offshore typically requires expensive and specialist plant, equipment and vessels. Due to the increasing size (weight, diameter and length) of the foundations, in order to mitigate installation risk, the use of large diameter drills is becoming an increasing requirement.Project developers often try to avoid drilling operations due to the time associated with the drilling process. This restricts projects in terms of location and it also overlooks the additional engineering benefits drilling can bring to monopile foundation design e.g. reduced foundation material and better fatigue design life.By comparing the lessons learned of three milestone overwater drilling projects; methods of maximising drill speed and decreasing drilling process time is established. Particular emphasis is put on pile internal relief drilling in a process known as Drive-Drill-Drive (DDD).The milestone overwater drilling projects are North Hoyle Offshore Wind Farm (OWF) (2003); Flamanville EPR cooling water outlet shaft (2008); and Gwynt-y-Mor OWF (2014).By evaluating the effect of foundation pile diameter on drilling process times, example foundation installation costs are derived and used to show that incorporating drilling as part of a process may not be cost prohibitive to projectsContinued advances in drill technology, and drilling rates of penetration, combined with an integrated approach to foundation design, challenge whether hammering only is always the optimal solution for XL monopile installation."

Jan 1, 2017

By A Misra

In pursuit of solving the development rate problem the coal industry has investigated various development systems. While all the systems may improve the development rate they may not necessarily he cost effective. This paper carries out an economic analysis by developing cost models for present and future development systems for retreating longwall faces with particular reference to Australian longwalls. A critical cost analysis of various development systems is presented by taking into consideration all the costs associated with development systems. The costs have been divided into equipment, face and panel services, belt and structure, wages, roof and rib support, ventilation, consumables, power and maintenance. The annualised cost of various equipment has been obtained on the basis of depreciation rates prescribed by the Australian Taxation Office.

Jan 11, 1996