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By John Drobe, Robert M. Cann

"Abstract - The Minas de Oro Cu-Au skarn and replacement deposits are located in the highlands of central Honduras, 90 km north-northwest of the capital of Tegucigalpa.The deposits formed in Cretaceous volcano-sedimentary rocks following the emplacement of an early Paleocene granodiorite to dacite intrusive complex. Three types of skarn (Type IA, IB, and IC) and a low-temperature replacement mineralization (Type II) are recognized. Type IA skarn consists of massive brown-green andradite and lesser magnetite and pyroxene. Type IB skarn has magnetite + hematite as the main constituent and garnet <50%. Type IC skarn comprises >50% pyrrhotite + chalcopyrite with interstitial garnet or pyroxene. Au and Cu occur in potentially economic concentrations in all skarn types while other metals such as Ag, Pb, Zn, and Mo are locally present in significant quantities. The highly variable distribution and nature of the skarn deposits is controlled by: (1) intrusive activity, (2) composition of host rocks, (3) faulting and fracturing, and (4) attitude of host carbonates.Type II Au-Cu-As mineralization occurs in calcareous sandstone and conglomerate 2 km distal from skarns and comprises clots of Cu sulfides hosted within a zone of brecciation and quartz flooding. Garnet and/or magnetite skarn is not present and the mineralizing event appears to have been a low-temperature replacement type.The similarity of coeval Au-Cu-Fe skarns in central Guerrero, southern Mexico to Minas de Oro skarns, and the similar local stratigraphy is consistent with the generally accepted theory that the Chortis block originated from southern Mexico and during the Tertiary moved southeast to its present position."

Jan 1, 2000

By Jean-Francois Magnan

"The old conventional process to produce lithium raw material dates from 1954 (CA patent #504477A). This process produces low purity lithium carbonate (Li2CO3) only. To obtain battery-grade carbonate, another polishing process is needed. To further obtain lithium hydroxide, another chemical conversion process is needed. That is why historically, the value of lithium hydroxide has always been higher than that of the lithium carbonate precursor. With the old conventional process, the low quality lithium carbonate is precipitated from a lithium sulfate solution (Li2SO4) by adding soda ash (sodium carbonate and its impurities). The resulting reaction produces significant quantities of sodium sulfate (salt cake) as a by-product with amounts larger than the impure Li2CO3 recovered with part of the soda impurities. When this conventional process was invented, the salt cake by-product was not an issue: it was discarded in a river near the plant. This practice cannot be done today in North America. While looking to find a solution for salt cake, electrolytic salt splitting of Na2SO4 into NaOH and dilute sulfuric acid appears as a technically feasible solution to get a rid of the salt cake. However, the sodium sulfate salt splitting process has difficulties to economically compete with the chlor-alkali process, which produces millions of tons of NaOH from sodium chloride (NaCl) every year.In contrast, the process developed by Nemaska to directly produce LiOH by Li2SO4 salt splitting not only solves the environmental problems related to the conventional process to produce lithium carbonate raw material, it also produces very high purity lithium hydroxide monohydrate directly at very competitive OPEX cost in comparison with industry peers (Figure 1). It also drastically reduces the amount of reagent and the direct greenhouse gas emissions."

Jan 1, 2016

By Paul E. Mather, Matthew J. M. Krane

Anode or fire refining is the final pyrometallurgical step in the production of copper anode. Part of this process purifies liquid copper by reducing its dissolved oxygen with a natural gas-steam mixture. Until now there have been no transient models of this process coupling the injected gas jet mass transport with its chemical reactions. A transient, 1D flow, chemical reaction model was made to estimate the time-dependent behavior of anode copper composition, soot formation, and the transient gas jet composition along its height. Model estimates were made for reduction rate, soot emission rate, and exergy loss and destruction distributions for two melt temperatures and a range of steam addition rates. It was found that increases to melt temperature and steam addition rate increased reduction rate and decreased soot emission, while having little effect on exergetic performance. Estimates for reduction histories were compared to industrial data. INTRODUCTION Anode/fire refining is the final pyrometallurgical step in the production of copper anodes; it is essential for removing dissolved gases which cause porosity in cast anodes. Despite its significance, the open literature appears devoid of any transient models of the entire process; furthermore, there are very little published process data with which to validate any models made. This dual dearth of modelling work and process data inhibits studies of the tradeoffs between process and performance changes, desirable for decreasing reduction time and soot formation as well as quantifying changes in the exergy losses and destructions of the process. This work introduces a reduced-order, transient model of reduction during fire refining, examines performance trends with operating conditions, and compares model estimates with industrial data. Overview of anode refining The purpose of anode refining is to produce liquid copper which can be cast with minimal gas porosity, which disrupts the downstream electrorefining process. Dissolved sulfur is removed first, then oxygen, by reactions with different injected gases. The removal of oxygen (the reduction step) is the subject of this proceeding and is effected by the bubbling of natural gas, typically mixed with steam, through the melt from tuyeres. A typical refining run reduces dissolved oxygen from 3000 ppm to 300 ppm in 2–3 hours (Schlesinger, King, Sole, & Davenport, 2011). The refined copper is then cast as anodes. Figure 1 shows a schematic of the furnace vessel.

Jan 1, 2019

By J. Walenzyk

This report investigates the limited use of super large backhoes in surface mines around the world. The demand for super large (> 700 t) hydraulic excavators is rising and most of the machines are sold as front shovel (FS) application. Despite advantages in faster cycle times super large backhoe (BH) machines are a minority (ca. 1% of the total machine population). This investigation is based upon the comparison between the volumes available to the BH and FS in one block and before the machine needs to move. The working conditions are created with standard parameters as given by Original Equipment Manufacturers of hydraulic excavators regarding bench height, angle, and material density. The travel amount of both machine types is derived from these volumes and compared. Furthermore, the travel amount of super large BHs is compared to the travel amount of smaller mining BHs and evaluated based on attachment length and bucket size. The calculations conclude that the volume of material recoverable in a single spot is about 75% higher for FSs. This leads to an increased travel for BHs. The combination of both the increased appearance of electrical driven machines (in the class of > 700 t excavators) and the travel amount lead to a high risk of damaging the cable or the excavator. Enlarged wear on the travelling parts is also attended by the higher travelling. A direct comparison of commonly found working set ups shows that the ratio between volumes is 32% which leads to an increased travel amount. In comparison with a 250 t BH, the calculation shows that the smaller BH is able to reach nearly 50% of the volume of the super large BHs. The main reason for this is the respectively high reach on smaller machines which, combined with other factors, leads to a significantly higher number of BHs in lighter classes.

Aug 1, 2013

By S. Bantubani, I. Mc Dougall, S. Arjun, H. Joubert, R. Hansraj

The requirements determining the design of the Söderberg electrode columns used in base metal applications are different to those of other industries. There is a strong trend to position the contact shoes above the furnace roof in order to reduce the risk of water leakages directly onto the furnace bath. Tenova Pyromet has adapted its proven electrode column design for application on base metal electric furnaces based on these requirements. The design modifications implemented as well as considerations regarding the overall electrode-roof-furnace arrangement are discussed. Equally important is the management of the Söderberg electrodes to prevent over or under baking. Tenova Pyromet has developed the AutoFurn™ furnace controller to not only optimise furnace operations, but to improve electrode management. An overview of the electrode management control philosophy implemented by AutoFurn™ is provided, focusing on intelligent electrode regulation, slipping and baking. INTRODUCTION Tenova Pyromet has developed an electrode system for use in base metal furnaces, which builds on the experience that Tenova has gained during the past 30 years with electrodes in the ferroalloy and base metals industries. In base metals applications, an important safety consideration is the minimization of water within the furnace and on the furnace roof. In addition, direct contact between copper equipment such as contact shoes and pressure rings and the contents of the furnace is discouraged. In a 3 phase AC furnace, three electrode columns transfer the electrical energy to the furnace bath. Both Söderberg and prebaked carbon or UHP graphite electrodes are in use in base metal furnaces. In each column, the electrode is held in position by a hydraulic slipping and regulating device, a pressure ring and contact shoes. The electrodes can operate in a submerged-arc or brush-arc configuration, where the electrode is immersed in the slag or a short distance above the slag (Mc Dougall 2013). The electrode equipment is divided into an upper and a lower electrode assembly. The self-baking Söderberg electrode is made from lengths of steel casings that are added to the top of the electrode column and welded into place. Carbon electrode paste is added to the inside of these casings and is baked to form a solid tip in the region of the contact shoes. The working platform that surrounds the area where new electrode casings are added, is electrically isolated from the building and each electrode is isolated from the others to ensure the safety of the personnel who perform the electrode casing additions.

Jan 1, 2019

By E. Demedde

Carbon anodes in aluminium reduction cells are continuously consumed during electrolysis. In prebake cells the anodes are replaced at regular intervals. The spent anodes, called butts, are often allowed to cool in air while emitting fluorides into the building from the hot alumina cover and the adhering bath. In modem smelters spent anodes are placed inside enclosures and fluoride emissions during cooling are effectively reduced. Fluoride emissions from spent anodes were estimated by considering the kinetics of fluoride formation, a heat transfer analysis of the cooling anode butt and the ventilation flow from a cavity. Fluoride emissions from large anode butts cooling inside enclosures were compared with those cooling in air. Good agreement with published experimental values was obtained for enclosures having different leakage opemngs, when a hydrofluorination reaction of the spent anodes was included.

Jan 1, 2006

By Marcel M. Djivre

The paper will outline the destructive test methods and test procedures including the evaluation criteria currently used at the Materials Testing Laboratory (MTL), Sudbury, Ontario, Canada, a laboratory approved by the Ontario Minister of Labour for the purpose of testing or examining mine shaft ropes or other hoisting appliances. For the evaluation of the mechanical properties of rope samples used in "ultra-deep" single-lift shafts, additional discard criteria used in Ontario, Quebec and South Africa shall be presented. Most of the methods and test procedures currently used at MTL can also be found in the current CSA G4-00 Standard, Steel Wire Rope for General Purpose and for Mine Hoisting and Mine Haulage, published in November 2000. In this latest or sixth edition of CSA Standard G4-00, a new Clause 14 has been added outlining the requirements for in-service destructive testing of steel wire ropes for mine hoisting applications. The CSA G4-00 Standard is also intended for the manufacturing of new ropes, and includes requirements for raw materials and construction features for round strand and flattened strand rope, and locked coil rope in bright and galvanized finish. To-date the Ontario Mining Regulations require only that destructive test of new ropes be performed in accordance with CSA Standard G4-00, but does not require conformity to the Standard, and on the other hand the inservice wire ropes are required to be tested in accordance with CSA G4 and thus require conformity. The author recommends that the Ontario Mining Regulations (and regulations of other mining jurisdictions) be improved in order to ensure quality control in the manufacturing of wire ropes destined for mines in Ontario (and Canada). Also, the author recommends the revision of the CSA G4-00 Standard by the technical committee in order to address manufacturer?s concerns with respect to compliance, and harmonize the Standard with Internationally recognized Standards.

Apr 1, 2005

By I. Silin, M. Stapelmann, H. Wotruba, K. M. Hahn, C. Gerold, J. Katzmarzyk

New deposits are often characterized as low-grade and complex ore bodies. To date, processing of fine grained ore deposits implies complex processing routes with high energy consumption. For that reason, the investigation of dry grinding, e.g. with vertical roller mills (VRM), is an approach to save energy reserves. It is already proven that dry grinding can influence subsequent processing steps. In terms of flotation, the surface characteristics of the minerals are crucial for a successful flotation process, e.g. the sufficient adhesion of mineral particles and air bubbles. Wet grinding results in different surface properties of the product entering the conditioning stage than dry grinding, which has direct influences on the subsequent flotation process. To investigate the consequences for flotation, processing of a copper sulfide ore containing chalcopyrite and bornite was investigated, including grinding and rougher flotation. The effects of wet tumbling mill and dry VRM grinding on the degree of liberation of valuable minerals are characterized using mineral liberation analysis by scanning electron microscopy. The obtained flotation results, as well as the recorded flotation parameters such as dissolved oxygen, oxidation reduction potential and conductivity are presented and discussed. INTRODUCTION Copper production is mainly connected to porphyry copper deposits as main source for copper sulfides such as chalcopyrite and bornite. State-of-the-art processing of these deposits is already established worldwide for decades, combining crushing, fine grinding and subsequent flotation with rougher, cleaner and scavenger stages (del Villar, Desbiens, & Maldonado, Bouchard, 2010). While the demand for copper is increasing, copper grades of the deposits are decreasing and the remaining ore bodies have more complex mineralogy and finer particle size distribution. This implies processing routes with high water and – due to the liberation of fine grained valuable minerals – high energy consumption by fine grinding. One of the main energy consumers in processing plants is the grinding section, which accounts for more than 50% of the energy consumption in processing plants and even 4% of the worldwide energy consumption (Jeswiet, & Szekeres, 2016). Therefore, dry fine grinding alternatives are applied to enhance the overall energy consumption of mineral processing plants: high pressure grinding rolls (HPGR) and vertical roller mills (VRM).

Jan 1, 2019

By T. Gonzales, M. Maleki, M. Caçador, B. Climenhaga, H. Ghorbani

The campaign life of furnaces in the copper making industry is greatly affected by damage caused by normal operation cycling and abnormal conditions. Proper monitoring and assessment of the existing conditions and safe remaining life of the vessels and their refractory lining is key to prevent an unplanned early replacement. This paper presents an industry-proven refractory management program as well as a typical fitness-for-service assessment of a rotary vessel shell in order to extend the campaign life of copper making furnaces and enhance safety. The furnace refractory management system involves a comprehensive audit and monitoring program which has been successfully applied to several copper flash smelting furnaces over the past 15 years. The fitness for service assessment provides a methodology to estimate the remaining life of the steel components by evaluating the existing damage and future damage rate based on the historical survey data. The above practices can provide the operator with cost savings by extending the vessel campaign life in lieu of implementing more rigorous refractory lining and shell audits. INTRODUCTION Primary copper smelting is generally divided into two categories of flash smelting and open bath smelting. There are many different types of vessels that can be used to convert the copper concentrate to blister copper including, flash, Mitsubishi process, TSL (ISA/Ausmelt), SKS Bottom Blowing, Anode and Teniente/Noranda furnaces, reactors, and Peirce Smith converters. Each vessel has its own complex and intense environment which requires careful refractory and steel design to withstand the severe thermomechanical and thermochemical attacks within its campaign life. The campaign life of furnaces is greatly affected by damage caused by normal operation cycling and abnormal conditions. Proper assessment of the existing conditions and safe remaining life of the vessels and/or refractory lining is key to prevent an unplanned early replacement. This paper presents an industry-proven refractory management program as well as a fitness-for- service assessment methodology to estimate the remaining campaign life of copper making furnaces resulting in cost savings and enhanced safety of the operating vessels.

Jan 1, 2019

By G. Ladouceur

"Northgate Patino Mines produces a copper concentrate assaying 22% Cu. Moisture content of the product averages 12.5% H20 when dewatering with conventional vacuum disk filters. Recent smelter contracts reflect higher energy costs and increased moisutre penalty charges. This situation forced Northgate to investigate possible alternatives such as rotary dryers and pressure filtration. After laboratory testing, it was found that pressure filtration could economically meet the smelter&apos;s requirements. Pilot plant testing was then undertaken and confirmed the laboratory test results. Based on these findings, a model I.S.F. 1500 x 28/22 I.R. Lasta filter press was purchased.This paper describes the Copper Rand dewatering flowsheet the operation of the press and the economics of the process.IntroductionNorthgate Patino Inc. operates 2 und erground copper mines and a 3300 ton per day concentrator located in Chibougamau, Quebec, some 425 miles north of Montreal and 300 miles northeast of Noranda, Quebec.Copper concentrate is produced at a rate of 250 tons per day and trucked to a C.N. siding 10 miles from the mill site and then railed to the Noranda Smelter.Until early 1981, moisture content of the concentrate at 12.5070 was acceptable since smelter penalty charges were relatively low ($0.10 per ton per unit over 10070 moisutre). In January 1981, a new contract was signed in which the penalty was raised to $1.10 per ton per unit over 8070 . This and the increasing truck and rail costs justified the change-over from vacuum to pressure filtration . The Lasta I.S .F. 1500 x 28/22 was then purchased and installed, and operation of the plant started in December 1981."

Jan 1, 1987

By L. E. Gutiérrez

A new filtration model based on porous media transport theory is presented. It predicts capacity and final moisture content in the filtered cake as a function of design and operational variables at the plant. FilterSim® is the computer code where this new model is programmed. Application of FilterSim® to ten-month operation of an industrial copper concentrate filtration plant is shown on a daily basis and using a single set of internal parameters for the whole period. Average error is less than 5%, both in capacity and moisture estimation. FilterSim® abilities to optimize industrial filtration plants as well as to estimate future budgets are illustrated with several real examples.

Jan 1, 2007

By Nosa O. Egiebor, Kafui Nyavor

"The effectiveness of sulphuric acid, nitric acid, and hydrogen peroxide pressure oxidation on the cyanide extractability of gold from a double-refractory concentrate is presented. Whileover90% of the sulphide content could be oxidized by either of the two acid pretreatments, only 65% was achieved by the peroxide oxidation. Carbon oxidation by both acid and peroxide pressure oxidation were less than 15%. Although the peroxide pressure oxidation pretreatment resulted in much lower sulphur and carbon oxidation, it gave pretreated residues with decreased levels of preg-robbing activities and higher gold recoveries during subsequent cyanidation, thus indicating some deactivation ofthe preg-robbing carbon during pretreatment with hydrogen peroxide. The refractoriness due to carbon preg-robbing activity was found to contribute more to the refractory behaviour of the concentrate than the occlusion of gold particles by the sulphide mineralization.IntroductionGold ores and concentrates are classified as refractory when a significant portion of the gold content cannot be extracted efficiently by conventional cyanidation, even after fme grinding (Weir and Berezowsky, 1984); (Demopoulos and Papangelakis, 1987). The refractory behaviour of precious metal ores and concentrates is generally attributed to two main factors: The presence of carbonaceous materials which could cause preg-robbing activity during cyanidation (Zaitseva et al., 1973); (Nice, 1971); Scheiner et al., 1971); (Eisele et al., 1983), or occlude gold-bearing minerals (Wells and Mullens, 1973); (Gasparrini, 1983); (Chen and Reddby, 1990), and the occurrence of sulphide minerals which can occlude gold, thus reducing the access of lixiviants (Schwartz, 1944); (Weir et al., 1986); (papangelakis and Demopoulos, 1990); (Michaelis, 1989); (Schweigart, 1965). When the refractoriness is due to a combination of these two factors, the ore or concentrate can be referred to as being ""double-refractory"". Because of the rarity of gold ores exhibiting this double-refractory phenomenon, only a few studies have been carried out on their suitability for pretreatment by existing techniques."

Jan 1, 1992

By Robert B. Parsons

"The new Ontario Liberal government introduced its first budget in October. The changes to The Mining Tax Act proposed in the budget are mean t 10 simplify the current system without any loss of revenue to the government. It goes without saying that some operators will benefit from the changes while others will pay more taxes after 1985.With regard to ongoing matters at the federal level, the proposed carve-out legislation and the issue of northern allowances are being discussed with government officials, and a small victory has recently been scored with regard 10 the laxation of certain share transactions of prospectors and grubstakers. A report on these issues is provided herein.Ontario BudgetOn October 24, 1985, Ontario Finance Minister Robert F. Nixon tabled a budget proposing several significant changes to both the Ontario Mining Tax Act and the province’s income tax legislation.Mining Tax ActA flat rate of tax and reduced processing allowances highlight proposed changes to The Mining Tax Act. The proposed changes, with a comparison to the existing system, are set out in Table I.The most significant change is the replacement of the graduated tax rate schedule with a flat 20010 tax rate. This is a welcome change for most mine operators who viewed the high rate of 30% on profit over $20 million as counter-productive.However, the flat rate of ta x may not be welcome to all operators. Generally speaking, operators with less than $ 10 million annual profit are better off under the graduated rate system, while operators with annual profit exceeding $10 million will benefit from the 20010 tax rate."

Jan 1, 1986

By Patrice Nadeau, Yuejun Zhang, Dan Chen, Liming Huang, Claude Gagnon

"The traditional Denver flotation cell is generally suitable to process material ranging approximately from 5 µm to 300 µm. For floating over 300 µm material, its efficiency is significantly getting lower mainly due to the difficulty to suspend coarse particles uniformly, which decreases the chances of particles to attach to bubbles and increases the chances of detachment.RTFT (Rio Tinto Fer et Titane) owns one of the world’s largest hemo-ilmenite deposits. Considering the whole processing technology, the ore is only crushed to -1180 µm. The crushed ore contains about 0.32% S and most of the sulfur is removed by rotary kilns in the ore preparation plant. In order to achieve a similar S reduction, without emitting SO2 into the air, an alternative technology by flotation was pursued.Laboratory and pilot scale-up tests showed that traditional Denver flotation cell could perfectly remove the sulfur from the material sizing 0-300 µm. For the coarser particle sizes, two different models of equipment were tested to improve flotation efficiency: HF (HydroFloat Separator) from Eriez and CLF from BGRIMM (Beijing General Research Institute of Mining and Metallurgy), China. The D6” HF, which is a device that combines flotation and gravity separation, demonstrated its ability to recover those very coarse particles. As compared to the Denver flotation cell, the modified 25 L CLF coarse particle flotation cell achieved better results at higher pulp solids percentage and with a shorter flowsheet."

Jan 1, 2016

By E. Sigal

Overview ?Intrinsik has completed several wide area HHRA and ERA studies of mining and smelting communities ?Several ?Lessons Learned? ?Key areas related to: ?Study Design ?Data Quality ?Bioaccessibility ?Speciation ?Weight-of-Evidence Approaches ?Advances in Ecological Risk Assessment Techniques

May 1, 2009

By Avi Nanda, Robert Brandt, Alex Piatkiewicz, David M. Jones

In 2010 the United States Environmental Protection Agency (UESPA) revised the National Ambient Air Quality Standards (NAAQS) by lowering the permissible value to a 1-hour average SO2 concentration of 75 ppb. In 2012, the USEPA designated the Miami area as non-attaining for the new SO2 NAAQS. A number of options were studied to assist in bringing the Miami smelter into attainment with the new standard. The chosen solution included: Installation of a large tertiary hooding system over the converter aisle; Hooding of the anode vessels and treatment of the anode gas in a baghouse; Installation of a large 2,325,212 Nm3/hr caustic scrubber to treat the captured converter fugitives and anode off-gases; Installation of a secondary ESP for the tapping fugitive off-gas treatment system; and the continuous scrubbing of all acid plant tail stack gases. The modifications have reduced smelter SO2 emissions by 85%, achieving a world class SO2 stack emissions rate of 0.45 kg of SO2/tonne Cu. INTRODUCTION The Freeport McMoRan Miami smelter has been previously described in literature (Steinborn, Pasca, Sallee, Binegar, Bhappu). For the sake of brevity a brief description is given here. The Miami smelter is located 129 km east of Phoenix, Arizona near the town of Miami, Arizona. The smelter was established in 1915 and has undergone a number of process upgrades and modernizations over the years. The present smelter consists of an Isasmelt furnace, an electric settling furnace, four Hoboken converters, two anode refining vessels & one casting wheel; and one double contact sulfuric acid plant. Concentrates are railed & trucked to Miami from four Freeport McMoRan mines located in Arizona & New Mexico. The wet concentrates, typically containing 10% H2O with a Cu/S ratio of 0.85, are smelted in an Isasmelt furnace to a matte grade of 58%. The entire output from the Isasmelt is laundered into an Elkem electric furnace (originally installed in 1972 as the primary smelting furnace) for separation into matte and slag. Slag containing 0.85% Cu is skimmed from the electric furnace and discarded. Molten matte, containing 60% Cu, is tapped from the electric furnace and processed in one of four Hoboken converters to produce blister copper. Two anode refining vessels further refine the blister copper before the final product is cast on a single, 30 mold, casting wheel. All of the anodes produced at Miami are shipped to the Freeport McMoRan electrolytic refinery in El Paso Texas for refining into cathode copper. Smelter off-gases from the Isasmelt, electric furnace and Hoboken converters are combined and treated in a double contact acid plant to produce sulfuric acid. Acid produced at Miami is internally consumed at Freeport McMoRan mines in the region.

Jan 1, 2019

By Robert B. Parsons

"New Brunswick Finance Minister John Baxter pulled something out of a hat for the mining industry in April, but there is a general concensus in the industry that he should put II back where it came from.The industry has never been in the mood f or progressive rate structures, but that is exactly what New Brunswick &apos;s 1984 budget offers to this depressed sector of the economy.IntroductionThe Province of New Brunswick stunned the Canadian mining community in April with the announcement that the province&apos; s 16% mining tax rate would be replaced with a progressive rate structure effective January 1, 1984 (Table I). The proposed change, contained in the bud get papers tabled by Finance Minister John Baxter on April 17, caught the industry totally by surprise. The move to a progressive rate structure was completely unexpected for a variety of reasons:• Generally speaking, relations between the province and the industry have been relatively good, yet the industry was apparently not consulted with respect to the proposed measure and had no indication that a change of th is nature was coming. This is in stark contrast to the mood which prevailed during the major revamping of the province&apos; s mining tax regime in 1977: at that time, the industry was invited to contribute in advance to the province&apos;s plan for a new tax structure.• The new structure would appear to make th e province less competitive in terms of mining tax rates with Ontario , Quebec, and other provinces, despite indications in the past that New Brunswick is very cognizant of this competitive factor in formulating mineral policy.• A joint report in November 1978 by federal and provincial officials to federal and provincial finance and resource ministers, which was careful not to directly criticize Ontario and Quebec&apos;s progressive mining tax rate structures, recommended that provincial "" tax regimes be examined to ensure the inclusion of appropriate mechanisms to avoid the excessive taxation of cyclical type profit "" , and that ""the federal and provincial governments consider the concern over progressive tax regimes expressed by the industry and examine mining and royalty tax regimes and the income tax regime with a view to alleviating the effect of such regime on the aggregate marginal tax rate"". In this context, any move to a progressive rate structure would be viewed by many observers as retrograde."

Jan 1, 1984

By I. A. Cameron

In 1979, John Peacey and William (Bill) Davenport published The Iron Blast Furnace Theory and Practice billed as "An up to date and comprehensive treatment of blast furnaces from fundamental principles to modern operations" (back cover). As a 1979 graduate of McGill University’s Metallurgical Engineering program, I belonged to the last "beta" student group which completed the prescribed blast furnace course work before The Iron Blast Furnace Theory and Practice was published. As my career developed, I become an international blast furnace consultant and have worked with blast furnace operators and raw material suppliers on 5 continents. The blast furnace is the dominate iron making process responsible for 70% of global steel production. Blast furnace engineers have implemented many improvements by manipulating the blast furnace process metallurgy using the principles presented by Peacey and Davenport to reduce coke consumption and increase oxygen usage and blast furnace productivity. This paper briefly summarizes some key principles presented by Peacey and Davenport (1979), describes process improvements made by blast furnace operators over the 35 years since The Iron Blast Furnace Theory and Practice was published and outlines important future developments in blast furnace technology.

Jan 1, 2014

By Robert B. Parsons

"This month, Tax Notes focuses on two features which can potentially ""dress up"" a new issue of common shares to make the shares more attractive to investors. Attaching warrants to a share issue is a relatively common practice, and recent years have seen an imaginative variety of different kinds of warrants, some of which can be subject to somewhat different tax treatment . A less common financing strategy which certainly deserves some consideration involves the possibility of utilizing ""Share Purchase Tax Credits "" in conjunction with a new share issue.IntroductionPrevious subjects covered in Tax Notes have dealt with flow-through shares and other tax angles which can make equity issues more appealing to potential investor s. Two other techniques which can add zest to a new stock issue involve the utilization of warrants and ""share purchase tax credits"" (or SPTCs). This article takes a brief look at the better known subject of warrants and the lesser known matter of SPTCs.Of course, this paper presents only a relatively general summary of the income tax implications of warrants and SPTCs, and is not intended to be a definitive commentary on all of the detailed and complex tax issues that might be relevant in the circumstances."

Jan 1, 1985

By H. Tokuichi, H. Ohta, K. Sato

A physical separation technique of tantalum capacitors from waste print circuit boards is in a development phase in Japan. A tantalum capacitor consists of an exterior coating of resin (epoxy resin) mixed with silica fillers and a sintered tantalum compact, the latter of which is impregnated with manganese dioxide (MnO2). In this study, we developed a process to recover tantalum oxide from tantalum capacitors by roasting, fluidized-bed separation, and hydrometallurgical treatments. Our aim was to recover tantalum oxide with high purity from waste tantalum capacitors. In this process, the exterior resin was selectively decomposed by roasting without oxidation and pulverization of the sintered tantalum compact at a proper temperature. Furthermore, the decomposed resin with silica filler on the surface of the sintered tantalum compact was mostly removed by fluidized-bed separation. The surface of the sintered tantalum compact was cleaned by ultrasonic irradiation in an acid solution to remove the MnO2 paste as well as residual Ag paste and silica filler. The impregnated MnO2 in the sintered tantalum compact was leached as Mn(II) ions using acid solutions via the reducing effect of tantalum metal to recover purified tantalum oxide.

Jan 1, 2014