Search Documents

By N. F. Mescheriakov

New mechanical and pneumatic-mechanical flotation machines with axial impellers have been developed. Industrial tests show that these machines ensure lower power expenditure (on flotation) by 1.5-2.0 times as compared with reference machines (with the same or improved technological indices). Application of the newly developed axial impeller, with these advantages, is possible in any volume chambers of type OK, Dorr-Oliver, Wemko etc.

Jan 1, 1998

By Descour Jozef, Suguru Shirasagi, Takuji Yamamoto, Koji Murakami

For TBM tunneling no visual inspection of excavated ground is possible. This may hinder the high-speed excavation, particularly in complex ground conditions. The TBM Excavation Control System was developed for better prediction of geological conditions ahead of the tunnel face during excavation. The seismic reflector tracing produces volumetric images of anomalies in the ground in “near-real time.” These images are verified using logging data from pilot boring and TBM driving data, to improve reliability and precision of ground mapping. Also good correlation was obtained between the rock strength index computed from TBM data, and one derived from volumetric seismic images.

Jan 1, 2007

By Louis M. Perlman

CAN PRESENT TRENDS CONTINUE? Can population growth in the world keep going at the average rates of the last few decades, till, say, the year 2000? And beyond? At the same time, can we expect steady increases in mineral consumption per head, all over the world, as the living standards of many nations converge toward the living standards of North America? If they do converge? Does the total fund of world resources extend, with foreseeable techniques, to supplying the consequent demands for metals, non-metallic minerals and energy? Do we have the requisite economic and scientific techniques available to answer these questions? Certainly, we cannot forecast over decades with any certainty. There are many, of course, who seem to believe that by the simple fitting of trends, they can tell us something about the world and resource constraints in the year 2000, twenty-three years from now. It is significant that many studies of this sort have added sustenance to the 'doom-watching' fraternity of commentators. The reason is simple. The extrapolation of exponential trends into the distant future frequently leads to the most astounding and impossible scenarios. Another feature of exponential growth is the suddenness with which certain fixed limits are approached. It took world population two hundreds years to increase by 1/2 a billion, from 1650 to 1850. Yet some U.N. estimates suggest that the net addition to human population will be over 2 billion between now and the end of the century. Thus, the present world population would increase from a little under four thousand million today to rather over six thousand million in the year 2000.

Jan 1, 1977

By E. Gallestey, D. Groenveld, F. Kolb, M. Rozou

"Recent copper price evolution pushes modern mining industry even harder to achieve highest production with maximum efficiency. Operations are bound by numerous contractual and environmental conditions that bring a high degree of complexity to both the operators controlling the process and the plant managers assessing the circuit’s potential and setting the production goals.Advanced process control (APC) for grinding is one important tool to operate this most energy intense area in an optimal way, finding the optimal trade-offs between highest throughput, energy efficiency and product quality based on variable ore properties, production targets and management priorities.ABB’s state of the art control and optimization software platform enables user to design a controller based on model predictive control (MPC). The paper and the presentation show the technology used and the results achieved through an APC controller installed in a pilot project in Aitik (Boliden SA), Sweden. Furthermore, possibilities to interconnect grinding with other areas as stockpile management and flotation with various interconnected APC controllers will be shown. Finally, an example of the return on investment and the payback times based on the increased throughput and better product quality will be presented.INTRODUCTIONDepending on the ore characteristics and the targeted plant capacity, the design of the grinding circuit may vary significantly. Typically, the circuit consists of several mills (rod, ball, SAG, AG) in series and/or parallel with a number of classifiers and sumps at appropriate locations (see Figure 1).The main goal of the grinding section is to reduce the ore particle size to a level that enables efficient separation between gangue and precious metals in the subsiding processing step. The trade-off between quality and quantity has to be made. At the same time, it is also crucial to execute this process at the lowest possible energy and grinding media consumption."

Jan 1, 2015

By R. J. Charles

F. C. Bond: This is an outstanding paper on comminution theory and represents a considerable advance in mathematical formulation. It clears the way for a discussion that should ultimately decide whether the work index should be considered as a parameter according to the Third Theory or as a constant. In the empirical differential equation dE = —Cdx/x [1] either Corn must be treated as a variable to satisfy known crushing and grinding conditions. The author has chosen the exponent n as the variable, with C as a constant. According to the Third Theory, in the general case n is a constant equal to 3/2, and C is a variable related to the work index Wi. If x is placed equal to P (the size in microns which 80 pct of the product passes) and F is the size 80 pct of the feed passes, then C equals 5 W1, and dE = 5 -5 Widx/x = 10 Wi 10 Wi w = vP vF which is the basic equation of the Third Theory. It is apparent that this integration involves the error of treating the parameter Wi as a constant, just as the author's treatment involves the error of integrating his variable n as a constant. These errors are probably relatively unimportant, considering the small range of the variables. When more is known about the factors contained in the work index, Eq. 1 can be revised as a partial differential equation. The Third Theory exponent of 1/2, or n = 3/2 in the Charles' differential equation (Eq. l), was not chosen primarily because it is the average between the Kick and Rittinger exponents; this may be merely fortuitous, since it is illogical to define truth as the average between two errors. It was first discovered from the study of a large mass of crushing and grinding data and confirmed by laboratory testing. The general correspondence of the Third Theory equation to actual crushing and grinding results over the entire size reduction range argues strongly that a fundamental phenomenon is involved. The a posteriori theoretical explanation of the constant Third Theory exponent is that the work done in breaking rock is proportional to the length of the crack tips formed. Compressive force applied to the rock results in deformation and strain energy. When this locally exceeds the breaking strength a crack tip forms, the surrounding strain energy flows to the crack tip and extends it to split the rock, with release of the energy as heat. Under rapid impact the crack tip may form before the strain has reached equilibrium, resulting in less energy expenditure and probably a coarser product. In any case, the elastic limit of a brittle material is practically the same as its breaking strength, and the work necessary to form the crack tip is the work required to break. The crack length of a ton of broken rock cannot be measured directly. However, the particle shapes of feed and product are similar, and the crack length is considered equal to the square root of one half of the new surface area formed. It is, therefore, proportional to 1/vD, where D is the equivalent product particle diameter. This is the basis of the Third Theory equation. Many secondary factors can be imposed upon the crack length breakage. These may cause the Third Theory work index to increase or decrease somewhat as the product size becomes finer. The work index is properly a parameter rather than a constant, making it an increasingly valuable practical criterion of the work input required in crushing and grinding.

Jan 1, 1959

By K. Awuah-Offei

The objective of asphalt mix design is to produce a mix with optimal performance using available aggregate stockpiles. Often, this process does not take into consideration aggregate plant stockpile inventories or the cost of aggregate in the mix. This is due to the lack of tools that enable the lab technician or the engineer to optimize the mix for good performance while reducing the cost of aggregate and managing the stockpile inventory levels. The Bailey method provides useful guidance on aggregate gradation selection for optimal performance. This work presents a linear programming model of the asphalt mix design problem. The algorithm is implemented in MATLAB as an Asphalt Mix Design Optimization (AMIDO) program. The program is successfully validated with real-life data. The results show that AMIDO can be used to successfully produce any mix design with known Bailey method ratios. Also, RAP can be included in a design easily by making use of the Bailey method ratios and AMIDO. This work improves the state-of-the-art in asphalt mix design for dense graded mixes and could potentially be modified for other mixes.

Jan 1, 2009

The expansion of hydrometallurgical operations, declining ore grades and upwardly spiralling production increases the demand for fresh and recycled water with low levels of contaminants. The minerals industry is also increasingly being forced both by regulatory and cost pressures to reduce the amount of liquid waste they produce. This requires a strong focus on waste reduction by recycling, regeneration and reuse. However, these measures are often not enough in arid countries for a forward progress and expansion of their mining activities.   It is often the case that in the arid and semi-arid countries the only available water is either sea water or underground (brackish to hypersaline) saline waters. Brackish waters can be desalted at a reasonable cost but the desalination of highly saline waters is challenging and expensive.   There are emerging reports that sea water and in some cases hypersaline underground waters are now used by various mining industries as process water. This study explores the possibility of the use of sea water in various mineral processing and hydrometallurgical processes and briefly discusses the benefits and problems related to the use of saline waters at the mine and mill site. Several cases studies and preliminary laboratory test results are presented.

Jun 1, 2010

By J. Boldy

This paper is an attempt to place Canadian Precambrian volcanogenic deposits in perspective, by examining and compiling data regarding certain neglected fundamentals of size, grade, value, frequency distribution and spatial density. The discovery roles played by various exploration methods are compared and contrasted, and. using empirical methods, the probabilities of discovery are discussed, combined with the use of an effective 'cost-per-target' program. From these basic observations, realistic exploration judgments may be made. Using similar methodology, this format may be applied to other types of deposits. Precambrian volcanogenic deposits occur in a natural geometric progression in size. They number 110, of which 70% are pre-producers, producers or past-producers. These deposits are bonanzas and generally have a 'net pre-smelter' value of between $43 and $62 per ton. The statistically typical deposit is that which occupies the modal position, has 400,000 tons of ore, a value of $20 million and occurs with a frequency of 40%. A median deposit has 1.4 million tons of ore, a value of $70 million and occurs with a frequency of 15%. An upper-decile deposit ha s 13.3 million tons of ore. a value of $750 million and occurs with a frequency of 6.5%. The major, mid-upper-decile deposits have a minimum of 50 million tons of ore, a value of $3 billion and occur with a frequency of 4%. Five major volcanogenic deposits occur in the Precambrian and account for 54% of production. This contrasts with 55 deposits (all below the median) which account for 3.9% of production.

Jan 1, 1977

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

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

Jan 1, 2015

By B B. Du, Y N. Lu, L X. Wang, S L. Wu

The cohesive zone in the blast furnace is largely affected by the softening and melting behaviour of ferrous burden because of its vital role in the blast furnace process. The Visual High Temperature Methodology (VHTM) has been given to study the softening and melting behaviour of sinter, two kinds of lump ores and acid pellet intuitively. Before softening, both two kinds of lump ore swelled and more serious for lump ore B because of its denser structure than lump ore A. The softening and melting temperature intervals are different from one another and this phenomenon contributed to the gas permeability of ore layer in the cohesive zone since the ‘ore window’ has formed. Besides, the softening and melting behaviour of mixed burdens have been studied, where it has been found that the softening and melting characteristics of mixed burden A (consisting of sinter and lump ore A) is better than mixed burden B (consisting of sinter and lump ore B) and mixed burden C (consisting of sinter and acid pellet) despite the poorest characteristics of lump ore A. The reason is that the high temperature interactivity between lump ore A and sinter is much better than lump ore B and acid pellet. The high temperature interactivity is mainly influenced by the chemical compositions, porosity and contact condition. This paper provides a new research method for the softening and melting behaviour of iron ore and an important technical basis for the optimisation of blast furnace burden.CITATION:Wang, L X, Wu, S L, Lu, Y N and Du, B B, 2017. The softening and melting behaviour of single and mixed iron ore burdens, in Proceedings Iron Ore 2017, pp 605–610 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By Jos. Kalbhenn

PART of the plan of re-allocation evolved by the engineers of the Dominion Steel and Coal Corporation in 1932 to retard the rapid advance of the submarine workings seawards was the opening of Dominion No. 20 Colliery, in the Harbour seam, by a pair of cross-measure tunnels from the workings of an older mine working the Phelan seam, which lies 430 feet below the Harbour seam. The tunnels, 8 ft. by 13 ft. in size, were started away from the old workings of Dominion No. 2 at a point 1.5 miles beyond the shore line and were driven on a rising grade of 1 per cent for a distance of 5,335 feet, tapping the Harbour seam at a point 2 ¼ miles off-shore. For transportation and ventilation, shafts and roadways already existing in the Phelan seam were made use of during the stages of development and up to the present time. One of the two tunnels served as a main haulage road and an intake airway, and the other as a return airway and travelling road. However, as the workings of this new colliery expanded, the need for additional airways arose as was anticipated when the plans for the development were formulated. To meet this need, work was started on October 1st, 1942, to establish a new opening between Nos. 2 and 20 Collieries in the form of a vertical shaft to provide a passageway for the ventilating air current of some 250,000 cubic feet per minute on its return to the upcast shafts in No. 2 Colliery, thus making both tunnels available for intake airways. In the interest of safety, speed, and economy, the shaft was raised from the Phelan to the Harbour seam. It is 20 ft. by 12 ft. in size and, with the exception of three timbered compartments across the narrow end, is untimbered. One of the three timbered compartments housed the ventilating tubing and pipe lines, the centre one served as a cage-way for hoisting men and material during the operation of raising, and the third one contained a ladder way. The rest of the shaft was filled with debris during the period of shaft raising.

Jan 1, 1944

By W. Kotze

The gold mining industry has mainly relied upon the use of highly polluting chemicals, such as mercury and cyanide to recover gold from its ores. The Coal Gold Agglomeration (CGA) process was developed some years ago and has the advantage in that gold is recovered by a procedure which has little or no negative impact on the environment. A gold ore containing liberated gold particles is contacted with coal-oil agglomerates, where by the gold is recovered into the coal/oil phase. Laboratory scale batch tests were performed on an artificial mixture gold slurry and gold recoveries of up to 85% were found under optimized conditions. By recycling the coal/oil phase, it was found that the gold loading onto the agglomerates was increased. Tests performed on an industrial ore yielded slightly lower gold recoveries, and X-ray Diffraction (XRD) analysis on the coal/oil phase showed that minerals other than gold was recovered into this phase. A comparative study was conducted whereby the CGA process was compared to mercury amalgamation. Gold recoveries obtained through amalgamation were 15% lower than by the agglomeration process, which indicates that this process can be considered favourably as an alternative to amalgamation.

Jan 1, 2000

By Robert T. MacMillan

A process has been developed in the laboratory and expanded to semi-pilot-plant scale for recovering titania, iron, and byproducts from two types of titaniferous iron ores. A reducing roast with coke and soda ash followed by grinding and magnetic separation produced powdered iron and slag from which titania and byproducts were recovered by H2SO4 decomposition. Based on the steps tested on a. semi-pilot-plant scale supplemented by the results of the laboratory investigations, enough data are available to make a balanced flow sheet. Combined data from the laboratory and semi-pilot plant indicated that 85 to 9O percent of the titania content of the ore was recoverable and that approximately 90 percent of the iron could be recovered as a metallic-iron powder. Tahawus magnetite from the Maclntyre mine at Tahawus, N. Y., was successfully treated as follows: (1) The ore was roasted at 1,050° to 1,080° C. with carbon and soda ash, thereby reducing the iron to metal sponge and retaining the titania in the slag. (2) The metallic iron was separated, after wet grinding, by magnetic or gravity means. (3) The slag fraction was air-roasted at 850° C. and leached with NaOH to recover Na3VO4. (4) The leached solids received a dilute-H2SO4 leach to remove interfering impurities, chiefly Na2O and SiO2. (5) Titania was dissolved from the leached solids of step (4) by a H2SO4-bake followed by a water leach. (6) Titania was precipitated by seeding and boiling the resulting solution. The spent liquor was used for the dilute-H2SO4 leach in step (4). The following scheme was found to be technically feasible for recovering other byproducts of the process. (1) The gelled filtrate from the dilute-acid leach step was washed and dehydrated to produce a commercially acceptable silica gel. (2) The wash liquor containing sulfates of Na2O, Al2O3, Fe2O3, and TiO2 was combined with the highly alkaline filtrate from the wet-grind and metallic iron-separation steps, precipitating iron and titanium hydroxides, which were separated for special treatment. (3) Anhydrous Na2SO4 and Al2O3?3H2O were recoverable from the alkaline filtrate from step (2) by evaporation and carbonation, respectively, and the barren liquor was recycled to the magnetic-separation step.

Jan 1, 1952

By Paul M. Tyler

Graphite occurs in many places in the United States, but previous to 1914 the domestic production amounted to only 15 to 20 per cent of the natural graphite consumed in this country. Including artificial, graphite, the proportion was about 25 per cent. During the World War the production of both natural and artificial graphite increased until, in 1918 it, contributed more than one-half of the available supply, and since 1919 the proportion has not fallen below 40 per cent. Except during the war period, however, this apparently increasing independence has not been due to any large increase in production from domestic mines. It simply reflects a larger output of artificial graphite and reduced imports of Ceylon lung and chip, grades which have never been produced in substantial quantity in the United States. DOMESTIC DEPOSITS3 American graphite deposits, though numerous and often large, are characteristically low-grade. They constitute an abundant source of potential supply, but in normal times they have proved relatively costly to work, and even at the same price the various products have never been able wholly to displace certain imported qualities, notably crucible grades from Ceylon and pencil graphite from Mexico.

Jan 1, 1929

By Org de Wet

This paper summarises the strategy for selection between a two module coal plant incorporating large, but less equipment, and a three module coal plant with more but smaller equipment. The two options were evaluated in terms of three main categories: ? Size ? size of the equipment and the effect on the process efficiency as well as the difference in physical dimensions of the plant to accommodate the larger equipment; ? Cost ? capital savings between a two and a three module plant, processing the same ROM (run-of-mine) tons per hour (t/hr); and ? Practicality ? in terms of constructability of the plant, safety of using larger equipment and maintainability of the equipment. Decisions that were taken to reduce the risk associated with size, cost and practicality on the chosen option of a two module plant will be highlighted.

Jan 1, 2006

By Walter Murphy

For a number of years the Bureau of Mines has been studying the properties of crude petroleums produced in the United States and the Western Hemisphere. A number of reports have been published giving the analyses of crudes from the different field and regions as made by the Bureau of Mines Hempel method of analysis.4/ Several of these reports include analyses off crude oils from the Rocky Mountain region.5/ The search for new fields and new producing horizons in old fields in the Rocky Mountain region has been quite intensive during the last few years, and a number of discoveries have been made.

Jan 1, 1937

By P Luo, J E. Danoczi, P Nakutnyy

Sensor-based sorting methods to remove waste or for upgrading ore prior to downstream metallurgical processing is becoming more prevalent for mining operations. X-ray transmission (XRT) sorters are a popular choice for mineral processing plants since they can sort a wide variety of minerals. XRT sorters pass high intensity X-rays through particles to generate images of varying greyscale, based on their effective atomic density. The ore sorter then mechanically separates the particles using algorithms based on the images. A popular test currently used to evaluate the amenability of the ore to XRT sorting is to pass the particles through the sorter and obtain the greyscale images showing mineralogical differences. This test provides information on the mineralogical heterogeneity between ore particles for their separation. However, it does not provide an understanding of the actual mineralogical composition of the ore or information pertaining to which other minerals may respond in a similar manner, nor does it provide liberation information. This paper describes a test that combines high resolution X-ray micro Computer Tomography (CT) combined with QEMScan to obtain information about amenability of ore to XRT sorting and to obtain valuable information that can streamline metallurgical testwork programs. This test also provides information relating to particle thickness and associated X-ray attenuation coefficients, desired mineral presence and mineralogical information relating to associated minerals and clays to assist with developing sorting algorithms. Combining the CT and QEMScan can address many limitations of the current method to determine whether XRT sorting technology is appropriate for particular ores. Mineralogy information from QEMScan can identify incidences of; mineral encapsulation, problematic clays, detrimental minerals as well as calibrate the greyscale values of CT data. Additional 3D CT imaging carried out on cores provides spatial information of the minerals within the ore and mineral grades.

Nov 10, 2020

This paper presents an initial investigation into the viability of co-kriging the grades of multiple mining cuts in a two dimensional, thin reef scenario. Deposits with gradational mineralized zones, in the perpendicular sense, are not as well served by the use of reef accumulation and thickness as deposits characterized by a sharp distinction between ore and waste rock. South African Bushveld Complex platinum deposits and Witwatersrand gold deposits are examples of this distinction. The calculation of an optimal mining cut is of paramount importance in the mining of Bushveld platinum. Two-dimensional kriging of a constant stoping width mining cut is an effective method for resource calculations once a cut is decided, but this begs the question of how to decide on the cut. Comparison of multiple cuts is used as a way of achieving this decision on many of Anglo Platinum’s operations. Although the kriging of each cut is valid in isolation, occasional negative values occur when incremental grades between overlapping cuts are calculated. Negative incremental grades cannot reflect reality. They are thought to be related to the inability of the method to include the strong continuity in the perpendicular direction demonstrated by the relevant variograms. Co-kriging of related variables has been investigated as a potential means tore impose this inherent order relationship. Various co-kriging options have been investigated. The method has been tested on a real data set from an established platinum operation. One of three types of anomaly has been investigated so far. The results show that co-kriging the mining cut grade as the primary variable, with just one secondary variable, the reef thickness, eliminates most occurrences. Severe anomalies remaining are arguably indicative of population mixing and might be more effectively treated by rezoning the data. Aim—To investigate the viability of co-kriging the grades of multiple mining cuts as a method for resolving anomalies in the estimation of such cuts on Bushveld platinum deposits, with applicability to other thin reef type deposits with a gradational not sharp cut-off between ore and waste in the direction perpendicular to the reef. Keywords: co-kriging, cross variograms, ordinary kriging, platinum, two-dimensional, grade profile

Jan 1, 2003

By Terrence M. Carroll, Chu E. Ho, Dimitrios Niarchos

"Construction of the new extension to Terminal 4 at John F Kennedy International Airport in Queens, New York involved the application of tapered piles for foundation support. Initial design based on past load testing experience had indicated that a pile depth of 65 feet was necessary to achieve a capacity of 394 tons. However, piles with a single length of 62 feet were ordered and installed, which led to marginal performance of the piles. This paper presents the technical considerations for mitigating the situation based on a systematic review of three preliminary pile load tests and dynamic capacity measurements on the shorter piles. Consideration was given to the long term gain in capacity of tapered piles due to set up in the Glacial Sand. The observational approach to the evaluation of piling performance adopted on this project resulted in salvaging many piles which failed to comply with the specified driving criteria. It was acknowledged that a more prudent course of action was for the Contractor to have ordered more appropriate pile lengths that allowed for potential variation in soil conditions across the site. The 4.6% saving in pile material did not warrant the effort that was required to mitigate the 39% of installed piles that did not meet the specified driving criterion.INTRODUCTIONAn expansion project at John F. Kennedy International Airport in Queens, New York involved the Phase 2 extension of Concourse B at Terminal 4. The site consists of 4 to 5 feet of pavement over about 10 feet of hydraulically placed sandy fill. Below this is a layer of marsh deposit, comprising of soft organic clay and peat, up to about 11 feet thick. Beneath the marsh deposits is a deep layer of glacial sand which extends to well over 100 feet below ground. Figure 1 shows the typical subsurface profile at the site. The new structures were supported on 223 steel tapered piles. The piles consisted of a straight steel pipe 18 inches in diameter and 0.375 inches thick, with a 25 feet long tapered bottom section (diameter varying from 18 inches down to 8 inches at the pile toe). The piles were first driven to the founding depth and subsequently in-filled with concrete. The concrete strength (fc) specified was 4,500 psi and the yield strength of the steel piles (fy) was 50 ksi. Initial pile design, based on earlier experience with static pile load testing for Phase 1 development at the same terminal, indicated that a pile length of 65 feet was necessary. However, piles with a single length of 62 feet were ordered and installed by the Contractor, leading to marginal performance of the piles. This paper presents the technical considerations for the adoption of alternative pile driving criteria, based on a systematic review of the preliminary static load test results for the present site and PDA measurements obtained during pile installation, in order to accommodate the capacity of the shorter piles."

Jan 1, 2017

By Carlos Alberto Sanchéz

El presente trabajo describe la realización del monitoreo de zonas revegetadas de cierre de minas utilizando técnicas de teledetección espacial. El estudio evaluó las áreas iniciales del programa de revegetación final de la minera Yanacocha. Las áreas de estudio son Maqui Maqui y San José, se localizan en la zona este de operaciones y se encuentran en cierre desde el año 200 y son restauradas dentro del programa de revegetación de áreas finales de la empresa. Como parte de la medición del éxito del programa se realiza el monitoreo de las mismas utilizando estas nuevas tecnologías. Por ello, es aplicable como parte de la conducción de los programas de revegetación de los planes de cierre de minas.

Sep 12, 2005