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The 2003 SME Annual Meeting and Exhibit will take place Feb. 24-26 in Cincinnati, OH. The meeting will feature a host of special symposia and programming geared toward the industrial minerals and aggregate industry sectors. This programming focus is particularly relevant for this annual meeting, as it will be hosted in the center of America’s transportation hub and industrial mineral center. Cincinnati, located on the Ohio River, is key to the movement of America’s supplies and construction material needs.

Jan 1, 2003

By J. Ndlovu, R. Hundermark

Keywords: Anglo Platinum, Polokwane Smelter, pyrometallurgy, metallurgy, platinum, platinum group metals, PGM, base metals, nickel, copper, sulphide smelting, electric furnace, Söderberg, electrode, matte, slag, flash drying, matte casting, crushing, granulation Anglo Platinum?s Polokwane Smelter is situated outside Polokwane, in the Limpopo Province of South Africa. Wet concentrate is received from various concentrators along the Eastern Bushveld Complex, with 60% of the total concentrate received being from UG2 reef and 40% from Merensky reef. The concentrate grade differs from 35 g/t to 110 g/t, and from this a PGM-rich nickel-copper matte is produced. The furnace matte produced is sent to the ACP converter in Rustenburg. The Anglo Platinum concentrators, from which concentrate is received, are Lebowa, Potgietersrus, Amandelbult, and Union, as well as the joint venture Modikwa. The concentrate is fed to two flash dryers that utilise coal-fired, fluidized-bed hot-gas generators to produce the hot gas that will drive off the moisture, leaving a fine bone-dry concentrate to feed to the furnace. The dry concentrate is pneumatically transferred to a storage silo before being transferred to the feed bins situated above the furnace. Lime, if required as a flux, may be transferred separately to the furnace feed bin by a pneumatic system. The concentrate and lime is fed from the feed bins onto airslides and into the furnace. The feed rate is automatically controlled, and the system is set up to optimize the power to feed ratio. The electric furnace is nominally rated at 68 MW and the power is transferred into the furnace using six 1.6 m diameter Söderberg electrodes. Concentrate is melted by energy generated when electric current passes through the electrodes and resistive slag layer. On melting, two immiscible phases form: slag and matte. Furnace matte, containing the bulk of the base metal sulphides and PGMs, is denser than slag and collects naturally at the bottom of the furnace. The furnace is constructed of a combination of refractory brick and water-cooled copper coolers. The furnace sidewalls and hearth are cooled by air drawn through the area by cooling fans. The hearth and matte zone are constructed from refractory bricks. The copper coolers reside only in the slag zone of the furnace along the entire perimeter of the furnace. One staggered row of plate coolers is installed above the waffle coolers along the perimeter of the furnace. Matte is tapped periodically through one of the two matte tapholes into 35-ton refractory-lined ladles, and cast into matte ingots on a casting machine. The cooled matte is discharged onto a concrete slab for cooling and then transferred by front-end loaders to the crushing plant. The matte is first crushed in a jaw crusher and then a Rhodax cone crusher to a size of 2 mm before it is loaded into tankers and sent to the ACP. The low-grade slag is tapped from the furnace through a water-cooled copper insert into a short water-cooled copper ?hot launder? which discharges into a granulation cold launder. The granulated slag slurry flows to three rake classifiers, from which the coarse slag is discharged onto a conveyor to dewatering silos. The dewatered slag from the silos is dumped. The water overflow from the classifiers reports to two thickeners where the water is clarified, and the thickener underflow is returned to the rake classifiers. The smelter is a zero-effluent plant and all the water from the slag silos where the slag is dewatered, as well as the storm water, is pumped to the process water dam from where it is pumped back into the plant for process water use. The off-gas from the furnace is drawn through a forced draught cooler and into a high temperature baghouse. The dust collected in the baghouse is pneumatically transferred into bins above the furnace and is fed via the concentrate airslides. The cleaned off-gas is then vented through the main stack.

Jan 1, 2006

By J. F. Barkley

Field tests were made on the rate of corrosion of 30 materials used in air preheaters under regular operating conditions, and the buildup of deposits was observed. Twenty-three materials were tested with coal as the fuel on an underfeed stoker and seven with oil, including a short period with pulverized coal as fuel, The plant equipment is described, and pertinent plant operating data are tabulated. Figures 17 and 19 (pp. 24 and 27) give the results of corrosion tests on the uncoated metal test specimens. The information obtained from these tests follows the general pattern described in Bureau of Mines Report of Investigations 4996, Corrosion and Deposits in Regenerative Air Preheaters. The deposits were similar, and the corrosion rates of Cor-Ten and Mayari-R steel (previously found to be essentially the same at the Harding Street Station using coal) continued to show little difference at the Marion Generating Station using principally oil. The one high-alloy steel tested, type 321, showed a much higher corrosion rate than the low-alloy steel, Cor-Ten.

Jan 1, 1959

By Sanjeeva Latchireddi

"Deteriorating ore grade and high demand of metals has tremendously influenced the unit capacity of SAG mills. Pebble circuits have been added to the existing SAG circuits to increase their unit capacity. To make the pebble circuits justifiable, the grate openings are increased almost to 4.5 inches to draw large quantity of pebbles. The inherent transportation problems associated with SAG mills in closed circuit with pebble fraction are more complicated than the single stage SAG mill operation, because the coarse pebbles behave significantly different to that of slurry in pulp lifters. In addition, operation of mills at higher speed, to take advantage of precise design of high release angled shell lifters significantly reduces the discharge efficiency of the conventional pulp dischargers. The presence of a sizable quantity of pebbles in the pulp lifter limits the space available and reduces the flow gradient through the grate thus increases the load inside the mill and hence the mill power-draw. This paper analyses all the problems associated in discharge of slurry and pebbles in SAG mills with pebble circuits. The Outokumpu patented Turbo Pulp Lifter (TPLTM) aims at eliminating the inherent material transport problems to ensure good grinding conditions. Recent installation of TPLTM in a 26-ft SAG mill has proven these claims by reducing the specific power consumption while increasing the mill capacity.INTRODUCTIONThe conventional wisdom is to look at the semi-autogenous grinding (SAG) mill as just another grinding mill in which ball charge, mill filling, and mill speed are varied to attain maximum throughput. Liner design for the SAG was also approached more from a wear perspective rather than a metallurgical and hydraulic-flow perspective. The extensive research work carried out by Latchireddi and Morrell (1997,2003a and 2003b) has demonstrated to think of the SAG as more of a pump in which the grates and pulp discharger end (“impeller”) were designed to efficiently discharge the product sized particles smaller than the grate size with minimal recirculation out of the mill.The rock load in the mill is essentially depends on ore characteristics and the discharge rate of broken particles. The discharge rate of the product material depends on how efficiently the discharge pump (grate and pulp lifters) is operating. Similar to the impeller design affecting the pump capacity, the pulp lifter design affects the discharge capacity (or mill throughput) of the SAG/AG grinding mills. The general arrangement of AG/SAG mills discharge system is shown in Figure 1."

Jan 1, 2007

By M Sherman

The challenge facing the engineering provider is to calculate throughput rates when faced with disparate specific energies calculated by different testing methodologies. Typically, an ownerÆs team would expect that the throughput calculations be checked using more than one method, therefore it is inevitable that there will be differences, based on which methodology is used. This engineering challenge is exacerbated by the continuing evolution of contemporary circuit configurations such as two stage pebble crushing with or without high pressure grinding rolls (HPGRs) as the second stage, or second stage pebble crusher product reporting to the ball milling circuit, etc.This paper is a companion to the paper first presented at SAG 2011, which showed that the Bond Indices are an excellent tool for estimating the specific energy requirements for semi-autogenous grinding (SAG) mills in single-stage SAG milling (SSS) circuits, SAG and ball mill circuits (SAB) or SAG and ball mill circuits with pebble crushers (SABC). This paper presents how the Bond Indices have been used to calculate ball mill specific energies, and therefore throughput rates, for SAG mill based circuits. Once the ball milling specific energy is known, tower mill and vertimill energy requirements can also be calculated. It is noteworthy that the method presented in this paper does not require the use of any unpublished correction factors: the proposed method uses the Bond parameters and equations published by Rowland in 1982.CITATION:Sherman, M, 2013. 21st century bond, in Proceedings Iron Ore 2013 , pp 309-312 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 12, 2013

By Mani Manimaran, Inna Dymov, Mariam Melashvili, Chris Fleming, Joe O’Day

This paper presents experimental evidence that supports the use of bromine/bromide as a potential alternative to conventional cyanidation for gold leaching. The study evaluated a brominebased lixiviant, Stabilized Bromine, with considerably lower vapour pressure than liquid bromine. Using a rotating device, the rate of gold dissolution was determined at various concentrations of this reagent. The reactivity of pure pyrite, arsenopyrite and chalcopyrite minerals was also evaluated in stabilized bromine reagent. Furthermore, bromine leaching of gold in a number of gold bearing ores was undertaken. Results showed that significantly higher gold recovery can be achieved with bromine than with cyanide when leaching gold encapsulated in sulphides, but bromine consumption was high owing to simultaneous oxidation of the sulphides. The greatest promise for the bromine leach process was with oxidized gold ores. The recovery of gold from oxide ores with bromine was comparable to that achievable with cyanide and bromine consumption was reasonable when leaching was conducted at a near neutral pH of ~6. In the case of oxidized gold ores containing copper mineralization, it is possible that bromine consumption may be lower than cyanide consumption, since there is evidence to suggest that bromine is less reactive than cyanide with copper minerals.

Jan 1, 2014

By John Kuyt, Zili Li, Michael Mooney, Michele Mangione

"INTRODUCTIONTwin bored tunnels are commonly used for a variety of road and tunnel projects throughout the world. A major component of these twin tunnel projects is the construction of connecting cross passages that provide emergency access between tunnels and space for tunnel service equipment. Cross passages can present significant geotechnical risk and technical challenges to a project – support systems must be considered for redistributing loads around the mainline tunnel openings and supporting new loading resulting from excavation activities. Additionally, cross passages pose risks to construction schedule and cost because of their sequencing at the end of construction, so mitigation of schedule overruns can be critical. By observing field instrumentation data records from completed projects, a betterunderstanding of the force development in each temporary structure can be achieved with indication of which structures are the most critical to support ground loads.This paper discusses the data obtained during the Brisbane Airport Link construction project, a road tunnel located in Brisbane, Australia. Strain gauge instrumentation data from the precast segmental lining and propped opening steel sets is presented with relation to the construction sequence and local geology. Interpretations and conclusions based on the observed trends are proposed, with several key mechanisms affecting the loading and unloading identified.Brisbane Airport Link Project OverviewThe Brisbane Airport Link project in Brisbane, Australia, connects central Brisbane to the airport located northeast of the downtown with over 7km of twin-bore road tunnels, two lanes each, with several caverns, interchanges, and access ramps constructed along the alignment to connect to other key routes in the city. The project was executed under a Public-Private Partnership (P3) scheme, funded and operated by the consortium BrisConnections. The main contractors were a joint venture of John Holland and Theiss, while the design was carried out by Arup and Parsons Brinckerhoff.Construction of the tunnels was conducted using a combination of 17 roadheaders and two 12.5m double-shield Herrenknecht TBMs [1]. The TBMs were used in EPB-type closed mode for the initial portion of their drive, starting at the northeast end of the alignment. Tunneling started in soft ground, weathered rock, and mixed face conditions initially, progressing to competent sedimentary rocks as the drive continued. Support was accomplished using a 12.1 m diameter steel fiber reinforced concrete (SFRC) segmental lining installed behind the TBMs in a 9 segments + 1 key arrangement. The innovative design of these segments is detailed in the paper by Harding and Francis [2]. Ancillary roads and underground interchange caverns were constructed by sequential excavation techniques using roadheaders, accounting for nearly half of the total underground excavations."

Jan 1, 2016

By Tony Peglas, Harry Asche, Eva Wood

The North-South Bypass Tunnel is a new $AUD1.5 billion underground traffic route in Queensland, Australia, a project of the Brisbane City Council. The tunnel is approximately 5.2 km long, bypassing the Central Business District, and is adjacent to and below major buildings and services. Road tunnel alignment issues include the complex southern connections to the Pacific Motorway and the design of ramps to the east and the north. A review of road tunnel alignment criteria and accident data for Australian and overseas tunnels, similar to the proposed NSBT, are discussed. The key findings of this review lead to a rationalization of the traditional tunnel cross sectional width allocation. The road alignment design elements adopted for the NSBT are also discussed. The NSBT route crosses beneath the Brisbane River. Geotechnical investigations and hydrogeological modelling are described. The crossing has a fundamental impact on the alignment and the tunnelling methods. The tunnelling methods to be adopted and the tunnel support are described. The geotechnical investigations and hydrogeology analysis culminate in the selection of a single shielded TBM for the reference design. These outcomes are related to complex issues of hydrogeology and contaminated groundwater. These are discussed in detail.

Jan 1, 2005

By Gerry Houlahan

The east span of the 3.5 km long SF-Oakland Bay Bridge is supported by different types of foundations along its span. Starting from the west, its first 460 m section comprising twin concrete box-girders is supported by sands and rock on Yerba Buena Island. A 620 m long steel suspension span then bridges a 25 m deep channel that is formed in sediments overlying rock. Twin concrete box girder structures, each 2.4 km long, complete the stretch over the remainder of San Francisco Bay to the engineered earth fill landing at its east end. For most of its length, the Skyway bridges water 3 to 10 m deep over soft surficial Young Bay Mud. The foundation types include spread footings with hold-downs, steel and concrete piles that are set in drilled rock sockets up to 3 m in diameter and 2.5 m diameter steel tubular piles driven to penetrations of up to 100 m. The design and construction-related considerations are presented. The different foundation types, influences of the site geology and earthquake demands on the substructures are discussed. The seismic motions of the bridge superstructures and substructures, as well as the depth-varying soil motions, influenced the foundation design. Together with performance-based design criteria, including post-earthquake footing displacement limits, these seismic demands influenced various aspects of the foundation design, including the use of battered and vertical piles.

Jan 1, 2005

By T. C. Holmes

"Dome mine, in the Porcupine mining camp of northern Ontario, is one of the important gold mines of Canada, its production of gold from the commencement in 1910 to the end of 1946 having a value of $149,000,000.GENERAL GEOLOGYThe oldest rocks in the mine are greenstones of Keewatin age(l). Those to the south are known as the southern greenstones, but little is known of their structure. Those in the northwest part of Fig. 1 and the north part of Fig. 2 are known as the northern greenstones or greenstone 'nose'. They face north, dip steeply north, and end to the east against an unconformity which pitches northeast about 20°. The volcanics are overlain unconformably by Timiskaming .conglomerate and slate. All these rocks are cut by porphyry. Closely associated with the porphyry is a group of rocks that have suffered extreme alteration. Diabase dykes intrude all the other rocks and the ore. The principal massive rocks in the mine are two adjacent flows known asthe andesite (south) and dacite (Fig. 1). They are composed chiefly ofmassive, medium-grained lava. Pillows, flow lines, and other typically volcanicfeatures are found along the tops of both flows and at other places in theandesite. These parts nearly everywhere show a slight schistosity panllelwith the flows except in the orebodies, where they are massive."

Jan 1, 1954

By X. Wu, M. G. Karfakis, K. Y. Haramy, B. T. Brady

A quantitative evaluation of damage caused by ground motions associated with strong roof caving is presented. The strong roofs are modeled as either cantilevering or bridging beams over continuous elastic foundations that are subject to an exponentially distributed vertical stress. Major factors influencing roof caveability and strain-energy buildup are identified and investigated. A methodology for the evaluation of strain-energy accumulation caused by, an overhanging roof is presented. The potential for mine tremors is determined in terms of the local Richter magnitude (ML), which depends on the seismic energy transformed from the stored strain energy that is released when the roof fails. A particle-velocity model is applied to estimate the damage area associated with roof collapse. A case study from a Utah coal mine shows that estimations of bump severity and the damage area agree well with field observations.

Jan 1, 1997

By G. Halverson, K. Blower

"Giant Yellowknife Mines Limited (Giant) has been producing gold since 1948. Located at Yellowknife, Northwest Territories, Canada, see Figure 1, the mine processes approximately 1,200 short tons per day of 0.23 ounce gold per ton of ore.The majority of the gold in the ore is sub-microscopic and associated within the mineral arsenopyrite. This association complicates the extraction method since straight cyanidation will only recover 30% of the gold. The operations at Giant involve four essential steps, with an overall recovery of 87%. The steps shown in Figure 2 are crushing and grinding, flotation, roasting and cyanidation. These processes result in a number of effluent waste streams characterized in Table 1. These streams are combined and sent to a tailings storage pond for subsequent treatment prior to discharge to the environment.Giant is located just north of the 62nd parallel where winter temperatures drop to -40°C for extended periods. This entails storing effluent in dams over the seven month winter period for treatment during the relatively short summer season.The treatment method currently employed at Giant is an ""alkaline chlorination-arsenic precipitation-flocculation"" process. The pond water is subjected to three phases of treatment prior to being discharged to the environment as shown in Figure 3."

Jan 1, 1987

By D. C. Oliver

Paper written on project work carried out in partial fulfilment of B.Sc. (Eng.) (Metallurgy and Materials) degree Corrosion in exhaust systems can occur on both the interior and exterior surfaces. Chloride-containing de-icing salts on the roads in colder countries greatly increases the rate of corrosion on the exterior surfaces. This paper investigates the relative exterior corrosion resistance of three alloys?two ferritic stainless steels( AISI Types 409 and 441) and an aluminized mild steel. The exhaust system is split into two categories?the hot end and the cold end. A test for each of these ends was designed?the standard ASTM G85test was modified to create a simulated driving cycle. The cold end test samples were exposed to ten driving cycles, and the hot end test samples were exposed to five driving cycles. A driving cycle constitutes the car being started from cold, driven until the operating temperatures are reached, and switched off and allowed to cool. Pitting corrosion is the primary corrosion mechanism occurring at the cold end of the exhaust system. The aluminized mild steel was the most resistant alloy in this test, and Type 409 the least. However, once the protective aluminium layer degrades, corrosion will occur rapidly. At the hot end of the exhaust system, spalling of the surface oxide layer is the primary corrosion mechanism occurring. Type 441 was the most resistant in this test, and the aluminized mild steel the least resistant. Potentiodynamic tests were performed on the alloys?the results agreed with the results obtained in the exposure tests.

Jan 1, 2003

Sluiceboxes can provide a much higher concentration ratio than most other gravity concentrators (up to 50 000:1) at very high overall placer gold recoveries (greater than 99 per cent). They are also very reliable, inexpensive and simple to operate. This combination is very difficult to beat and explains why the sluicebox is still the most important placer gold concentrator in northwestern Canada, Alaska and the rest of the world. A sluicebox is a rectangular fl ume containing riffles on matting, through which a dilute (<12 per cent) slurry of water and alluvial gravel flows. Sluiceboxes are actually centrifugal concentrators with settling velocity playing a minor role in gold recovery. Due to their higher density, gold particles tend to segregate to the bottom of the slurry flow where they form a streamline that is diverted by a low pressure zone into a riffle. Under ideal conditions, this ribbon of slurry will be overturned as it flows down the rear of the following riffle and will continue fl owing in a circular path to form a vortex. At the bottom of this vortex, centrifugal and gravitational forces combine to drive gold particles into or beside the matting. To remove the gold concentrates, sluiceboxes are shut down and the riffles and matting are taken apart and cleaned. The slurry velocity provides the energy that powers the vortex. If the velocity of the slurry is reduced through overloading with solids, insufficient water flows or shallow gradients it may not sustain a full size vortex. If the riffles are too close, too far apart, too tall, or if there is not enough energy available to the vortex, the vortex will not be formed properly and gold recovery will be reduced. If the slurry velocity is too high, extreme turbulence and the resulting scouring will also cause gold losses. Testing sluiceboxes with conventional sampling and evaluation techniques is very costly, time consuming and problematic. Most placer gold ores contain a very small number of gold particles in a large volume of pay gravels. Sluiceboxes can lose coarse gold particles. The presence or absence of one of these particles in a tailings sample can lead to high unpredictable errors (nugget effect) even when large sample volumes are processed with care. Every time conventional samples are upgraded, additional errors are introduced due to the inefficiency of recovery equipment. Nuclear tracer tests are more accurate, faster, cheaper and safer than conventional sampling. In 1989 through 2002, the recovery efficiency of dozens of sluicing systems was determined by mixing radioactive gold particles (tracers) into the feed streams of dozens of placer mines in the Yukon Territory, British Columbia, Alaska and Guyana. Four distinct sizes of nuclear tracers were used and their recovery was related to the design and operational characteristics of the individual sluiceboxes and their pay gravels. At clean up, scintillometers were used to detect the very low level X-ray and gamma ray radiation emitted by these tracers and locate them in the sluice runs. The effectiveness of any given section of riffles was easily diagnosed by the presence or absence of tracers. At every mine the gravel feed rates, water flows, equipment dimensions and riffle performance were measured. After the gold tracers were removed from the fi nal concentrate and counted, the concentrate was sieved and weighed. These gold recovery, weight and sieve data were used to calculate the quantities and size distributions of the gold particles in the original pay gravels and those lost in the tailings. The expired tracers were stored in a lead lined container until their radioactivity was near background levels (about two months). The standard errors from these radiotracer tests were estimated from binomial probability theory with the equation: SE = {(n + p + q) ^ 0.5}/n. Where n was the total number of radiotracers added, p is the proportion recovered and q was the proportion lost. Each overall recovery estimate would be within one standard error of the true recovery value (14 times out of 20) and almost always within two standard errors of the true recovery (19 times out of 20). The maximum standard error with 100 tracers was five per cent and occurred when the recovery approached 50 per cent. With higher recoveries, standard errors usually ranged from one to two per cent.

Aug 1, 2010

By Liao Jianfeng

The Baosteel Zhanjiang 5,050m3 blast furnaces have been built with highly independent integration and innovation technologies and represented the state of the art of the world’s huge BFs. The proper of BF1 and BF2 is designed with long-campaign technology; both BFs share the same set of stockhouse, water and PCI systems; the BCQS serial-hopper no-bell top, TRT and blower are all local equipment as applied; four top-combustion stoves systems are matched; the BFG is cleaned by fully dry process; and the automation control system for the main process is self integrated. The BFs have been started up and performing with sound indicators. Their smooth proceedings symbolize China’s huge blast furnace ironmaking technology ascending a new high.

Oct 30, 2017

By John D. Sullivan, Alfred P. Towne

"In copper percolation-leaching plants the ore is usually crushed to a maximum size of about three-eighths inch. Ordinarily an ore is crushed as finely as feasible since the rate of extraction of copper is increased with decreasing size of particles owing to the greater surface exposed and. the shorter path required for solutions to come into contact with the minerals to be dissolved. However, greater degree of crushing also increases the amount of fines. When ores are charged into leaching vats the coarse and fine material tend to segregate, resulting in channeling of solutions and leaving the areas where fines predominate only partly leached. An improvement in the method of charging the ore into vats that would minimize the deleterious effects of the segregations of fines would increase the percentage extraction of copper.The Inspiration Consolidated Copper Co.4 recently installed classifiers to remove the primary slimes from the ore, before charging it into the leaching vats. This necessitates a separate agitation treatment for the slimes.Recent publications5 of the bureau have shown the advantages of agglomeration followed by trickle leaching to extract copper from concentrator tailings. It was shown that when equal parts of ore crushed to about 1/4-inch size and concentrator tailings were agglomerated with approximately 10 per cent of moisture, either in the form of water or leaching solution, and the mass was charred into a leaching vat the ore could be treated successfully by open-drainage trickle leaching. Even with 30 or 35 per cent of minus 200-mesh material present the ore could be leached, if it was not flooded with solution.The results obtained in agglomerating ores and slimes suggested the possible advantages to be gained by wetting the crushed ore for standard percolation-leaching plants before charging it into the vats. Wetting the ore whould make the fine material and slime cohere to the larger pieces of ore thus tend to prevent segregation when the ore is charged into the vats. The bureau, at its southwest Experiment Station and in cooperation with the Department of Mining and University of Arizona, Tucson Ariz., therefore made a study to work: out A way in which slimes could be handled in percolation leaching. The study showed that the method of charging ore into the leaching vats governed the distribution the fines and consequently the rate extraction of copper. The results of the study reported in this paper."

Feb 1, 1931

By Deborah McCombe

"By now, most geologists and engineers in the Canadian mining industry are familiar with National Instrument 43101 Standards of Disclosure for Mineral Projects (NI 43-101) and the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Definition Standards for Mineral Resources and Mineral Reserves, which are incorporated by reference in NI 43-101. In recent years, significant progress has also been made in establishing standard reporting practices internationally that recognize the globalization of the industry and enable investors to compare projects anywhere in the world and make informed investment decisions. This article is part of a series that will summarize the background and development of international resource and reserve reporting standards by the Committee for Mineral Reserves International Reporting Standards (CRIRSCO), highlight its current activities and provide insight into international mineral reserve and resource reporting issues. CRIRSCO was formerly the Mineral Definitions Working Group of the Council of Mining and Metallurgical Institutions (CMMI), an organization representing professional bodies in all of the major mining and mining investment countries. In 2002, CMMI disbanded and CRIRSCO continued as a volunteer committee with limited funding from those professional bodies and increasingly from other sources, such as mining companies and regulatory bodies. CRIRSCO was adopted as a Task Force of the International Council of Mining and Metals (ICMM) in 2007. The ICMM is an influential industry body representing all of the major mining companies and many junior companies through its associated members, such as the Minerals Council of Australia, the Chamber of Mines of South Africa and the Prospectors and Developers Association of Canada. CRIRSCO now has funding to pursue its objectives."

Jan 1, 2003

By Cheng Pan, Xuewei Lv, Guibao Qiu, Chenguang Bai, Meilong Hu

"Physic-chemical properties of slag at high temperature play an important role in the pyrometallurgical process of production of ferronickel alloy. It determines the operation efficiency, metal recovery ratio, energy consumption and the distribution of elements like S and P between the slag and metals. In the present work, the effect of MgO content in the slag on melting features and viscosity of the slag was investigated. The MgO content in the SiO2-CaO-MgO-Al2O3 -FeO in the quinary slag system varied from 14.5% to 22.5%. The results shown that all slag samples begin to softening at the same temperature, the softening temperature, melting temperature, flowing temperature decrease with the increase of MgO content when the MgO content is in range of 14.5 and 20.5 pct mass, and when MgO content up to 22.5 percent, the temperatures increase sharply. Besides the inflection point temperature of viscosity-temperature curve increase with the increase of MgO when the MgO content is in range 16 and 22.5 pct mass.IntroductionNickel is very important to modern society, without it, we would not be able to share many things in our daily life, such as airplanes, computers and many stainless steel appliances in our homes and workplaces1.. It is the essential element of the stainless steel making, and about 65% of the nickel consumed in the world is used to make stainless steel2., so with the increase of stainless steel, the nickel production and demand has continued to increase3.Nickel is produced from two distinct ore types: nickel sulfide deposits and oxide nickel laterite deposits. The great majority of world’s known nickel resources are contained in laterite deposits3.. The laterite ores are normal classified into two groups: the high-iron laterite ore and high-magnesia laterite ore1. Typical composition of high-magnesia laterite ore is in the range 1-3%, and these ores are suitable for pyrometallurgical process to produce ferronickel4."

Jan 1, 2012

By R M. Papini, A C. Araujo, A E. C Per

Flotation is an important process used in the concentration of iron ores. One of the challenges for process engineers is the low metallurgical recovery associated mainly with the fines fraction. In this work, the use of synthetic flocculants in iron ore reverse cationic flotation was investigated. The flocculants utilised were polyacrylamides of high molecular weight and different ionic degrees. Bench scale flotation tests were performed to compare the flotation performance with and without addition of polyacrylamides. It was observed that cationic and non-ionic polyacrylamides cause a significant increase in the metallurgical recovery. The anionic flocculant did not improve the flotation performance. An adverse effect was a slight increase in the silica content of the concentrate. Flotation with polyacrylamides was influenced by the hydrodynamic conditions of the pulp, such as agitation and the weight percentage of solids in the flocculant conditioning and flotation stages.

Jan 1, 2005

By David G. Winter

The first test pile for the carrier pier and south wharf at NAVSTA Puget Sound was driven on December 27, 1989, and the final pile was tested on February 26, 1990. Implementation and modification of the test pile program were done by the owners (WESTDIV and ROICC-Everett), the engineers (ABAM and Hart Crowser), and the contractor and his subcontractor (General Manson Joint Venture and GRL). The site is located as shown on Figure 1. SUMMARY OF THE SOIL CONDITIONS A soil profile along the pier is shown as Figure 2. The soil units can be generalized as follows: ? Upper Silts - Fill, sandy silt, and clayey silt extending from the mudline to an elevation of about -80 feet; ? Intermediate Sand - A medium dense to very dense silty sand unit which is about 30 feet thick near the shore, but thins and disappears offshore; ? Lower Silts - Sandy silts and clayey silts in a medium stiff to stiff condition; ? Deep Sand - A very dense overconsolidated sandy soil encountered at an elevation of about -145 feet.

Jan 1, 1990