iron ore flotation with nontoxic reagents

Paper No. 207

IRON ORE FLOTATION WITH NONTOXIC REAGENTS K L Sandvik1 and E Larsen2,* ABSTRACT The Sydvaranger magnetite mine was shut down in 1996 due to low iron ore prices. It was reopened in 2009. The flowsheet has remained more or less as in the old plant, based upon two grinding steps followed by magnetic separation. The Sydvaranger magnetite is free of harmful elements and should therefore be sold on the high quality market. This requires a concentrate containing less than 2 % SiO2. In the old plant, this grade was achieved by amine flotation. Amine was preferred as collector because it could be used in process water from the sea. The use of amine has been restricted by the Norwegian pollution authorities during the time the mine was shut down. Test work to replace cationic with anionic flotation is therefore done. Research on activation of quartz with Ca or Mg ions prior to anionic flotation was shown by basic research in the period from 1960 to 1980 to be effective under simplified laboratory conditions. This old research has now been followed up from a practical/empirical point of view. The basic work was mainly done on quartz and iron minerals only. In this case, an actual feed containing a number of minerals common in iron ores is used, in order to demonstrate that a commercial process is possible. Calcium activation is shown to work excellently but requires too high pH to be environmentally and economically attractive. Magnesium activation is a more difficult + process, because the maximum concentration of MgOH , which was shown to be the activator, is + about 100 times lower than the alternative CaOH . Flotation success depends partly upon the ore, some amphibole minerals are more difficult to float than pure quartz. Ordinary batch flotation test work is time consuming. A South African microflotation test cell has shown promise, because work can be focused upon flotation of the more critical coarser fractions of the Sydvaranger ore. Keywords: Iron ore concentration, anionic flotation, mineral activation, microflotation, Sydvaranger

INTRODUCTION TO THE MINE AND CONCENTRATOR The banded magnetite ore body was discovered in 1902, a company was established in 1906 with first iron ore production in 1910. Sydvaranger Gruve AS has had an amazing history in a very remote yet well positioned part of northern Finnmark, Norway. Located approximately 400km north of the Artic circle, close to the Russian and Finnish border, the concentrator in the port town of Kirkenes ten kilometers by rail from the high grade deposit of Bjornevatn has been responsible to beneficiate magnetite ore up to 99.5% purity as described by Sandvik and Malvik (2001) and for it to be shipped worldwide. Using the magnetic properties of the ore and making the most out of its enviable liberation characteristics the original plant consisted of ball mills and 600mm magnetic separators. The plant was completely destroyed in World War II and reopened in 1952. In 1996, it was shut down due to market conditions. The iron ore price quadrupled with the rise of China between 2004 and 2008. Buoyed by these higher prices, Tschudi Shipping Company, who were involved in Kirkenes since the early 1990’s, purchased Sydvaranger Gruve in 2004 for its port facilities, then engaged the Australian mining community ‘which saw value where others had failed to see it’ Tschudi (2010). This engagement led to the formation of the Australian Stock Exchange listed Northern Iron who became the parent company for Sydvaranger Gruve AS which reopened again in 2009.

1. Norwegian University of Science and Technology (NTNU), NO-7491 Trondheim, Norway. Email: [email protected] 2. Norwegian University of Science and Technology (NTNU), NO-7491 Trondheim, Norway. Email: [email protected]

XXVI INTERNATIONAL MINERAL PROCESSING CONGRESS(IMPC) 2012 PROCEEDINGS / NEW DELHI, INDIA / 24 - 28 SEPTEMBER 2012

04715

SANDVIK AND LARSEN

A simplified concentration flowsheet is shown in Figure 1. The circuit had originally two stages of magnetic separation with the first after the primary milling and the second after closed circuit secondary milling.

Figure 1. Principal flowsheet of the Sydvaranger concentrator

Two major changes in processing from the old to the new company were: 1. Change from seawater to recycled fresh water as process water; including disposal of thickened tailing, this is important for flotation. 2. Introduction of a 74 µm stacker screen to treat the secondary magnetite concentrate, oversize returned to secondary mills, undersize dewatered in a tertiary magnet step before filtration.

Strive to Improve Concentrate Quality Sydvaranger produces a desirable product, low in silica and exceptionally low in phosphorus (5% and 0.008% on average respectively). Improving the concentrate further will lead to a higher premium for the 2.8 Mtpa of magnetite concentrate and may lead to further opportunities to increase the amount sold to existing customers in the local market. The Metallurgy team of the mine has identified that improved magnetic separation can achieve a 4% silica, 0.005% phosphorous product. It is therefore anticipated that decreasing the magnetic separator units flux rate from ~70t/m/hr to below 38 t/m/hr through the installation of more units and decreasing the density to these magnetic separators the plant will consistently achieve less than 5% silica in the concentrate in 2012. Any further improvement will require the use of flotation. The challenge is to determine the most cost effective method of employing this practise whilst abiding by the strict environmental regulations with the tailings being discharged into the Fjord.

FLOTATION The purpose of the flotation project is to develop a process to upgrade the Sydvaranger concentrate to about 2 % silica, or 70 % Fe, in an environmentally acceptable way. A lot of work on pure minerals has been done in the past and some of the relevant literature is reviewed here. One solution is direct magnetite flotation as described


04716

IRON ORE FLOTATION WITH NONTOXIC REAGENTS

by David Young (1979). Because the amount of silicates is small, it is logical to remove the impurities however. The old concept of diamine gangue flotation is not allowed in new operations. Our original plan was to work together with the reagent suppliers to develop / test more environmentally friendly biodegradable amines. After experiencing the attitude towards amines among parts of the public at Kirkenes and the environmental authorities, it was decided that anionic collectors only, should be tested. The authors` predecessor, Marcus Digre, did some early work on anionic flotation, Cooke and Digre (1949). Seawater was used in the process at that time. Anionic reagents would form insoluble salts with the metallic ions in the sea water. Because amines worked excellently, anionic reagents were not in question when flotation became a reality in 1972. There was done a fair amount of work on anionic flotation of silicates up until 1975 or so. Fuerstenau and Palmer (1976) have in “Flotation”, the Gaudin Memorial Volume given an excellent summary of the most important aspects of this research. Some of the points that are relevant for iron ore will be reviewed in the following. FROTH FLOTATION: A Century of innovation 2000, this one again with Fuerstenau et al (2007) as the principal editor, do not show many new findings in the years between the two publications that may be relevant for iron ore flotation so it is obvious that interest in this topic has vanished. The zeta potentials of typical Sydvaranger ore minerals are negative at most pH values as shown by Sandvik and Dybdahl (1979) in Figure 2 left. This is why amine flotation works so well. From the curve, the possibility for anionic flotation appears rather limited. The trick to get anionic reagents to function is to reverse the zeta potential. Because sea water was used at Sydvaranger at the time this research was done, zeta potential curves for sea water were established. 2 % sea water solution had to be used because the conductivity of pure seawater was too high for the Zetameter to function. Figure 2 right. A shift from negative to positive zeta potential can be seen around pH 10.5 for magnetite and for the major silicate minerals quartz, hornblende, and cummingtonite.

Figure 2. Zeta potentials of typical Sydvaranger minerals as a function of pH in distilled water (left) and 2 % sea water (right) -3

-4

The metal ions, notably Mg and Ca, have in diluted seawater concentrations of 10 mol/l and 2*10 mol/l. This coincides well with the explanation given by Fuerstenau and Palmer(1976) that the active ions adsorbed and + + giving the surface charge are Mg (OH) and Ca(OH) because it can be seen from Figure 3 (right) that the + + maximum concentration of Mg(OH) occurs at 10.5 and from Figure 3 (left) that the maximum of Ca(OH) is a little above pH 13.


04717

SANDVIK AND LARSEN

Figure 3. Logarithmic concentration diagrams for Ca and Mg after Fuerstenau and Palmer (1976) If those findings are combined, anionic flotation of those minerals should be possible with Mg or Ca activation at high pH. A problem regarding selectivity is that not only the silicate minerals changes zeta potential this way, the magnetite seems to follow a similar zeta potential curve as a function of pH and Mg and Ca activation as shown in Figure 2. Iron activation can probably be neglected, as it seems to be at a maximum at pH 8.5, Figure 4 (left) again after Fuerstenau and Palmer (1976), so iron activation above pH 10 could be ruled out. Figure 4 (right) after Morgan (1986) shows that the adsorption of oleate may tend to follow the ferrous ion concentration while flotation + + is strongest at the maximum FeOH concentration. There is a possibility that some combination of FeOH and oleic acid may be adsorbed at the bubble surfaces. Figure 4 (right) indicates that hematite and probably magnetite as well, may have some flotation properties different from the silicates. This is also the case for amine flotation. From Figure 2 it should be obvious that magnetite should float with a cationic collector. Usually it does not, and if it does, it can be depressed with an agent like dextrin. Similar reactions may occur for anionic collectors as well. Figure 5 (left) the last one from Fuestenau`s literature survey demonstrate some of points from the preceding figures, even if petroleum sulfonate was used in this investigation. Figure 5 (right) after Fuerstenau and Cummins (1967) shows flotation with lauryl acid, which is another fatty acid. What should be noticed on this figure, a point already observed by Cooke and Digre (1949) is that higher concentrations of collector or metal ion may increase the recovery at a lower pH which otherwise will give poor flotation. The optimal pH values given in Figure 3 are therefore not absolute.


04718


++

Figure 4. Logarithmic concentration diagram for Fe after Fuerstenau and Palmer (1976) (left) and adsorbtion and flotation of hematite after Morgan (1986) (right).

SAMPLES AND EQUIPMENT Initial tests were carried out to test the validity of the hypotheses sketched in the foregoing chapter. Gangue flotation as well as magnetite flotation was carried out separately in batch scale. Then flotation of primary magnetic concentrate was carried out to give an idea of what grades may be achieved. So far, oleic acid, which has been the first choice of the literature, has been used. When a promising reagent regime is found, optimizing may begin, including testing of commercial alternatives to oleic acid. Microscale flotation has been tested out as part of this project in order to make such research more efficient.

SAMPLES Figure 1 shows the sample points as stars in the flowsheet. The samples were stored under water. Final concentrates, the first taken at the concentrate filter: 66.3 % Fe and later stacker screen undersize (another final concentrate) contained at the day of sampling: 67 % Fe, 5 % SiO2, 0.3% CaO and 0.35 MgO. Stacker screen oversize, 53 % Fe. This sample gave an opportunity to work on larger amounts of the more tricky silicates. Figure 6 shows a polished section of the size fraction -150 + 106µm in transmitted light. All gangue particles can be seen to have larger or smaller specs of magnetite, which is why they end in the magnetic concentrate. For upgrading regrind is necessary, but the possibility to make a 2 % SiO2 concentrate by more grinding and magnetic separation is limited as size distributions in Figure 7 demonstrate that the grinding already is fine.


04719

SANDVIK AND LARSEN

Figure 5. Minimum flotation edges of quartz as a function of pH (conditions 1*10-4 M Sulfonate, 1*10-4 M metal ion) (left) Fuerstenau and Palmer (1976) and flotation recovery as a function of lauric acid and calcium chloride addition at pH 11,5 (right) after Fuerstenau and Cummins (1967)


04720


Figure 6. Picture of polished section showing specs of magnetite in all gangue particles


04721

SANDVIK AND LARSEN

Figure 7. Sieve analysis for flotation feeds The + 106 µm fraction of the stacker oversize 19 wt % will contain only 12 % of the iron, but 41 % of the SiO2 38 % of the Al2O3 and 39 % of the CaO in the sample. This means a considerable enrichment of gangue. It was assumed to be in the floatable range. Samples for microflotation were taken from the stacker screen oversize by screening at 54 µm. 0.5 kg of the oversize was split out wet and stored under de-ionized water in a fridge. 5 gram wet samples were taken at random for each test. This resulted naturally in some spread in the feed grade. The feed grade was averaged after the products from the tests were analyzed and the tests with calculated feeds deviating more than 20 % from the average were rejected. Primary magnetic separator tailing which contains approximately: 0.7 % Fe, 3 % Al2O3, 79 % SiO2, 4 % CaO, and 4 % MgO. Major minerals are magnetite, quartz, and different amphiboles. The samples were dried and split into smaller batches for further treatment, which included regrind to create fresh surfaces.

Batch equipment and procedure Grinding was done in a steel batch mill ø 205x100 mm with 3.5 kg balls. A Denver flotation machine with a 2 l cylindrical stainless steel cell was used. -4

Reagents: The values of 10 mol/l from the literature were translated into grams per ton. For CaCl2 and MgCl2 this is about 30 g/t and oleic acid was stage added from 100 g/t. pH should be in the range 11.7-12.4 for Ca activation and around 10.5 for Mg activation. pH regulator: NaOH; 0.5 molar 20 g/l. pH decreased with time and sodium hydroxide had to be added stepwise during every test. Activator: CaCl2 or MgCl2 0.5 % solution. Depressant: Dextrin 1 from Lyckebo 0.5 % sol Collector: Na-oleate 0.5 % solution


04722


Frother: MIBC 1 drop, later Flotanol C Agitation: 2200 rpm with 1100ml water then flotation at 1500 rpm Flotation at 1200 rpm to barren froth.

Microflotation equipment and procedure A microflotation procedure was developed to have a method for rapid screening of reagents on the coarse fractions only, which represent the separation problem. For this work, the equipment described by Bradshaw and O’Connor (1996) was available. The flotation cell is shown in Figure 8. The cell can be regarded as an automated and enlarged Hallimond tube. The auxiliary equipment including a pump for circulation of the pulp makes this more flotation system more versatile. The reagents used were the same as for the batch tests except that no frother was used. Conditioning of a 5 g sample was done separately in a glass container with 300 ml fluid.

Figure 8. Microflotation test cell After conditioning, 100 ml of deionized water was added to the dispersion in order to make a total volume of 400 ml. The pH was registered with a Metrohm 744 pH meter immediately after conditioning. After adding, the pulp to the microflotation cell system, aeration, and flotation was carried out for 10 minutes. Air flow was 10 ml/min, and the pump speed was 1700 rpm. Two products, the float and the non-float (sink), were left for settling for a minimum of 30 minutes, prior to decanting and drying at 85C overnight. After registering sample weight, chemical analysis was performed with a Thermo Scientific Niton XL3t XRF Analyzer. The analysis was performed on powder samples prepared by pouring the products into polyethylene sample cups, where the powder is coated with a thin film foil. Rapid analysis is necessary to get the full advantage of a speedier flotation setup.


04723

SANDVIK AND LARSEN

TESTWORK To establish approximate flotation conditions, tests were carried out separately upon gangue and concentrate, then work to clean concentrate began.

Flotation of primary tailing The tailing from the primary separators contains a collection of all gangue minerals in the ore. 750 g was ground for 25 min. Some of the softer gangue minerals were concentrated in this stream and was deslimed in a plexigass cylinder. Removal of the top liquid resulted in a split around 25 µm, which gave around 500 grams flotation feed. The results are summarized in Figure 9 (right).

Figure 9. Determination of basic flotation properties of magnetite (left) and gangue (right) SV 6. The amount of activator CaCl2 has been set at 30g/t to conform to literature values and pH rose to 12,2, very good flotation. SV 9. This test was run under the same conditions as SV 6 except that depressant was used. Dextrin is shown to depress silicates to some degree. Extra introduction of CaCl2 in the last flotation stages improves flotation somewhat. SV12. MgCl2 was used as an activator instead of CaCl2. Dextrin is still used as depressant. pH above 10.4.. Initial flotation was much better than with CaCl2. No extra activator was added in the end of the test so the last steps gave a performance approaching SV6. A examination of the products in the microscope showed that uncolored minerals, (quartz?) floated best. In the material not floated, larger uncolored minerals constituted about 70 % while dark minerals, (amphiboles?) made up 30 %. In addition, a few pinkish minerals and a few clear green minerals could be seen.

Flotation of final magnetic concentrate The purpose of the initial tests was to find the response of magnetite to the different reagent regimes. Weight recovery results for those tests are shown in Figure 9 (left) and can be compared to the gangue flotation.


04724


SV 7 Calcium chloride was used as an activator. No depressant. The test is run on same conditions as SV 6 on primary tailing. Magnetite floated very well. This confirms with Figure 2, which shows the change of zeta potentials in sea water. SV 8 This test is run at the same conditions as tailing SV 9. The test demonstrates that addition of dextrin has a very depressing effect upon magnetite flotation. 69.8 % Fe and 2.5 % silica. SV 13 This test is run the same way as SV 12 on tailing. Magnesium chloride is used as an activator. The test demonstrates again the influence of dextrin.

Flotation of stacksizer over and undersize Further tests were conducted upon material from the stacksizers. Results from the most important tests and conditions are summarized in the following and in Figure 10.

Figure 10. 10 Flotation of stacksizer under and overflow by calcium and magnesium activation. Cumulative silica content as a function of cumulative iron distribution. SV18 Undersize. Mg activation. pH 10.6. Depression with dextrine. 70.2% Fe, 2.6 % SiO 2 and 87.5 % recovery. SV 19 Same sample, Ca activation, pH 12. Depression with dextrine. 71.8 % Fe, 0.6 % SiO2 and 80.7 % recovery. SV 25 Oversize. This sample was reground and deslimed. Ca activation. The desliming product is represented in the figure as the first point on the flotation curves. In spite of a very different feed grade fom the undersize more or less the same final concentrate grade is achieved. The reagent consumption is also similar. 70.7% Fe, 1.6 % SiO 2 and 79.9 % recovery.


04725

SANDVIK AND LARSEN

Microflotation The coarser particles of stacker oversize assumed to be a perfect material for microflotation test work, because the problematic minerals and middlings are concentrated here. This offers an opportunity to screen many variables like commercial fatty acids and petroleum sulfonates on the problematic fast. The reagents were the same as in the batch tests. The tests so far has been focused on getting the procedure up and going concentrating on the important parameter of pH. The tests shown in Figure 11 have been run with approximately 750 g/t of dextrin, 1500 g/t of MgCl2 and 1000 g/t of oleic acid. Average feed grade was 41.9 % Fe. The amount of feed at each test varied from 3.8 g to 5.1 g. This is the reason reagent consumptions are approximate. The figure shows Fe grade and recovery in the sink. There is considerable spread in the results, indicating that the operating procedure has to be refined. The figure indicates anyway that flotation with Mg activation below 10.8 gives little upgrading of the feed in the coarse fractions. The upgrading takes place at expense of iron recovery, which should be expected. Microscopic examination revealed that coarse green minerals (hornblende?) had a tendency to remain in the nonfloat, while the float fraction was dominated by somewhat finer quartz.

Figure 11. Fe grade and recovery as function of pH by microflotation

DISCUSSION Most work on Ca and Mg activation cited in the literature has been focused upon quartz flotation. Silicate removal from iron ore is a much more complex process dealing with a number of silicates, at times changing from day to day. This investigation is carried on activation of silicates with calcium or magnesium ions and subsequent anionic flotation, under more “real life” mineral mixtures. So far, oleic acid has been the only collector tested. Activation of magnetite can be counteracted by the use of general depressants which depress magnetite stronger than most of the silicates. Balancing of depressants and activators is critical and variations in the surface area of the respective minerals from sample to sample make it difficult to hit the right dosage. Calcium appeared to be the most + -4 + promising activator because the maximum concentration of CaOH is 8*10 mol/l while the maximum MgOH -5 concentration only is 10 mol/l. This makes balancing the pulp chemistry more delicate. This is probably why calcium activation has been favored in the literature. The initial work was therefore concentrated upon Ca. A look at Figures 3 and 5 as well as test results reveals that for Ca activation to take place pH has to be above 12. Calcium activation of the stacker undersize has so far given the best test result (SV19) so far with 71.8 % Fe, 0.6 % SiO2 and 80.7 % recovery. At 90 %, iron recovery the SiO2 content was 1%.


04726


A pH above 12 does not appear to be very high, but considering that this requires at least 5 kg NaOH per ton it will at present production amount to 14 000 tons of NaOH. This could add to the environmental challenge and the recycled water will have a high pH as well causing special emphasis on the working environment. Those considerations lead to the conclusion that Mg activation was the only possible route. pH will then be around 10.5 + to ensure a sufficient supply of MgOH , requiring about 0.6 kg of NaOH. Magnesium activation test work has been performed upon primary tailing, final concentrate and stacker oversize and has been shown to work. Results to date are a little inferior to Ca activation. The major problem is that it appears to be more difficult to float some of the green minerals (hornblende?).The Mg activation system seems more affected by slime because of the low + concentration of MgOH . Flotation recoveries are not impressive compared to recoveries by magnetic separation. The losses are mainly caused by liberated fine magnetite ending in the scavenger concentrates. Those products have either to be refloated or selectively flocculated to bring recovery up. An extra complication is the fact that recycling of water causes Ca and Mg to be leached from the silicate minerals making directly recycled water unsuited for flotation. A similar problem is reported by Potapova et al (2009) for apatite flotation from Swedish iron ores. So far, the solution at Sydvaranger appears to be use of fresh water exclusively for the flotation circuit. Microflotation was taken up as part of this study. It gives a possibility to make a rapid survey of a number of parameters. As we know that most of the silicates are in the coarser fraction, which is small, the test work can be concentrated where we find the non liberated gangue. So far, the tests have confirmed that large green minerals are hard to float. Higher concentrate grades with higher pH demonstrate probably that more green minerals are floated at a higher pH.

FURTHER WORK The mining operation at Sydvaranger is moving to pits with less hornblende so future testing will show improved results also with Mg activation. Work will be restricted to the final magnetic concentrate (stacker undersize) because flotation of the stacker oversize will result in a more complicated flowsheet with a separate regrind operation ahead of flotation. Water chemistry will be addressed more closely. This is work well suited for microflotation studies.

CONCLUSION The Sydvaranger mine has been restarted. The concentrate grade is at present too low to give the potential premium prices, which the purity of the ore warrants. A final upgrading of the concentrate by amine flotation, which was used in the past to remove the last and most troublesome silica was not is allowed for environmental reasons. Anionic flotation of silicates after Ca or Mg activation was proposed as a possibility, not at least because the plant now works with recycled fresh water instead of sea water which is high in calcium and magnesium. The tests carried out so far demonstrates that the original assumption, which was that anionic flotation of silicate from iron ore under practical conditions is possible, was correct. Magnetite, which also will be activated, by Ca and Mg can be depressed. The best concentrate so far contained 71.8 % Fe and 0.6 % SiO2 at 80 % iron recovery. This is obtained with calcium activation at a pH between 12.0 and 12.3, which raises a new environmental concern. The NaOH consumption will be around 6 kg/t, which is rather high. The solution is therefore to work further on the more difficult Mg activation. Recycle plant water contained too much Ca and Mg leached from the silicate minerals and could not be used for flotation, but the flotation circuit may eventually be provided with fresh water. Microflotation has been shown to be a useful tool, which will be used more in the future, because the problem of gangue middlings, which is restricted to the coarser fractions, can be studied separately.

ACKNOWLEDGEMENTS The authors wish to thank Sydvaranger Gruve A/S for support of this work and for permission to publish it.


04727

SANDVIK AND LARSEN

REFERENCES Bradshaw D.J. and O’Connor C.T.1996. Measurement of the sub-process of bubble loading in flotation. Minerals Engineering, 9: (4), pp. 443-448. Cooke, S.R.B and Digre, 1949M. Studies of the activation of quartz. Trans AIME 184:299. Fuerstenau M.C., Jameson G and Yoon R-H(Ed)2007 Froth Flotation: A century of innovation. Fuerstenau M.C. and Cummins W.F.1967 Role of basic aqueous complexes in anionic flotation of quartz. Trans AIME 238:196. Fuerstenau M. C. and Palmer B. R.1976. Chapter 7 Anionic flotation of oxides and silicates. A. M. Gaudin Memorial Volume vol 1 Ed. Fuerstenau M. C. Morgan L.J.1986 Oleate adsorption on hematite Problems and methods. Int. Jour. Min.Prosess.18:139. Potapova, E., Grahn, M., Holmgren, A. and Hedlund, J. 2010. The effect of calcium and sodium silicate on the adsorption of a model anionic flotation collector on magnetite studied by ATR-FTIR spectroscopy. Journal of Colloid and Interface Science. 345, (2010) pp.96-102. Sandvik, K.L. and Dybdahl, B.A1979. Paper 15.1. Upgrading of taconite concentrate to direct reduction th specifications by flotation. Minnesota Section of AIME 52.nd Annual Meeting and University of Minnesota. 40 Annual Mining Symposium, Duluth. Sandvik, K. L. and Malvik, 2001. A century of processing of fine grained iron ores in Norway. Int. Mining and Minerals (47):pp.3-8. Tschudi, F.H. August 2010 Ed. Down Under Up North”, Published by Tschudi Shipping Company, Norway. Young, D. 1979. Paper 5.1 Flotation of iron ores in systems with controlled dispersion: Minnesota Section of AIME th 52.nd Annual Meeting and University of Minnesota. 40 Annual Mining Symposium, Duluth.


04728